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• Developer Training – Difficult Questions and Alternative Perspective – Part 3

  - by pinaldave

  Developer Training - Importance and Significance - Part 1 Developer Training – Employee Morals and Ethics – Part 2 Developer Training – Difficult Questions and Alternative Perspective - Part 3 Developer Training – Various Options for Developer Training – Part 4 Developer Training – A Conclusive Summary- Part 5 Congratulations!  You are now a fully trained developer!  You spent hours in a classroom, watching webinars, and reading materials.  You are now more educated and more prepared than ever before.  Now what? Stay or Quit The simple answer is that you now have two options – stay where you are or move on to a new job.  Even though you might now be smarter than you have ever felt before, this can still be a tough decision to make.  You feel extra trained and ready for a promotion or a raise, but you and your employer might not see eye to eye on this issue.  The logical conclusion is to go on a job hunt, but that might not be the most ethical thing to do. Click Image to Enlarge Manager’s Perspective Click Image to Enlarge Try to see the issue from your manager’s perspective.  You feel that you have just spent a lot of time and energy getting trained, and you should be rewarded.  But they have invested their time and energy in you.  They might see the training as a way to help you complete the goals they require from you, or as a way to help you complete tasks that will ultimately end in a reward or promotion. Moral Compass As in most cases, honesty is the best policy.  Be open with your manager about your expectations, and ask them to explain their goals.  When there is open and honest communication, everyone can walk away happy.  If you’re unable to discuss with your manager for one reason or another, just try to keep the company policy in mind and follow your own moral compass.  If all else fails, and your company is unwilling to make allowances for your new value, offer to pay the company back for the training before moving on your way. Whether you stay at your old job or move on to a new one, you are still faced with the question of what you’re going to do with all your new knowledge.  If you feel comfortable, offer to train others around you who are interested in the same subject.  This can look very good on your resume, and if you are working in a team environment it is sure to help you in the long run! What Next? You can even offer to train other trainers at the company – managers, those above you, or even report back to your original trainer about how your education is helping you in the work place.  Obviously this should be completely voluntary on the trainer’s part.  Taking advice from a “newbie” may not be their favorite idea, but it could also show the company that you are open to expanding your horizons and being helpful to everyone around you. Last in Line for Opportunity Click Image to Enlarge At this time, let us address a subject related to training and what to do with it – what if you are always overlooked for training?  This can as thorny a problem as receiving training in the first place.  The best advice is to let your supervisors know that you are always open to training and very interested in certain topics.  If you are consistently passed over, be patient.  Your turn will probably come, but the company as a whole has to focus on other problems at the moment.  If you feel that there are more personal issues at play, be sure to bring this up with your supervisor in a calm and professional manager so that everything can be worked out best for both parties. You, Yourself and Your Future! If all else fails, offer to pay for training yourself.  Perhaps money problems are at the root of being passed over.  Even if there are other reasons, offering to pay your own way shows your dedication and could work out well for you in the long run.  Always remember – in life you have to go out and make your own way, you cannot always sit and wait for things to land in your lap. Reference: Pinal Dave (http://blog.sqlauthority.com) Filed under: Developer Training, PostADay, SQL, SQL Authority, SQL Query, SQL Server, SQL Tips and Tricks, T SQL, Technology

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• SQL SERVER – Fundamentals of Columnstore Index

  - by pinaldave

  There are two kind of storage in database. Row Store and Column Store. Row store does exactly as the name suggests – stores rows of data on a page – and column store stores all the data in a column on the same page. These columns are much easier to search – instead of a query searching all the data in an entire row whether the data is relevant or not, column store queries need only to search much lesser number of the columns. This means major increases in search speed and hard drive use. Additionally, the column store indexes are heavily compressed, which translates to even greater memory and faster searches. I am sure this looks very exciting and it does not mean that you convert every single index from row store to column store index. One has to understand the proper places where to use row store or column store indexes. Let us understand in this article what is the difference in Columnstore type of index. Column store indexes are run by Microsoft’s VertiPaq technology. However, all you really need to know is that this method of storing data is columns on a single page is much faster and more efficient. Creating a column store index is very easy, and you don’t have to learn new syntax to create them. You just need to specify the keyword “COLUMNSTORE” and enter the data as you normally would. Keep in mind that once you add a column store to a table, though, you cannot delete, insert or update the data – it is READ ONLY. However, since column store will be mainly used for data warehousing, this should not be a big problem. You can always use partitioning to avoid rebuilding the index. A columnstore index stores each column in a separate set of disk pages, rather than storing multiple rows per page as data traditionally has been stored. The difference between column store and row store approaches is illustrated below: In case of the row store indexes multiple pages will contain multiple rows of the columns spanning across multiple pages. In case of column store indexes multiple pages will contain multiple single columns. This will lead only the columns needed to solve a query will be fetched from disk. Additionally there is good chance that there will be redundant data in a single column which will further help to compress the data, this will have positive effect on buffer hit rate as most of the data will be in memory and due to same it will not need to be retrieved. Let us see small example of how columnstore index improves the performance of the query on a large table. As a first step let us create databaseset which is large enough to show performance impact of columnstore index. The time taken to create sample database may vary on different computer based on the resources. USE AdventureWorks GO -- Create New Table CREATE TABLE [dbo].[MySalesOrderDetail]( [SalesOrderID] [int] NOT NULL, [SalesOrderDetailID] [int] NOT NULL, [CarrierTrackingNumber] [nvarchar](25) NULL, [OrderQty] [smallint] NOT NULL, [ProductID] [int] NOT NULL, [SpecialOfferID] [int] NOT NULL, [UnitPrice] [money] NOT NULL, [UnitPriceDiscount] [money] NOT NULL, [LineTotal] [numeric](38, 6) NOT NULL, [rowguid] [uniqueidentifier] NOT NULL, [ModifiedDate] [datetime] NOT NULL ) ON [PRIMARY] GO -- Create clustered index CREATE CLUSTERED INDEX [CL_MySalesOrderDetail] ON [dbo].[MySalesOrderDetail] ( [SalesOrderDetailID]) GO -- Create Sample Data Table -- WARNING: This Query may run upto 2-10 minutes based on your systems resources INSERT INTO [dbo].[MySalesOrderDetail] SELECT S1.* FROM Sales.SalesOrderDetail S1 GO 100 Now let us do quick performance test. I have kept STATISTICS IO ON for measuring how much IO following queries take. In my test first I will run query which will use regular index. We will note the IO usage of the query. After that we will create columnstore index and will measure the IO of the same. -- Performance Test -- Comparing Regular Index with ColumnStore Index USE AdventureWorks GO SET STATISTICS IO ON GO -- Select Table with regular Index SELECT ProductID, SUM(UnitPrice) SumUnitPrice, AVG(UnitPrice) AvgUnitPrice, SUM(OrderQty) SumOrderQty, AVG(OrderQty) AvgOrderQty FROM [dbo].[MySalesOrderDetail] GROUP BY ProductID ORDER BY ProductID GO -- Table 'MySalesOrderDetail'. Scan count 1, logical reads 342261, physical reads 0, read-ahead reads 0. -- Create ColumnStore Index CREATE NONCLUSTERED COLUMNSTORE INDEX [IX_MySalesOrderDetail_ColumnStore] ON [MySalesOrderDetail] (UnitPrice, OrderQty, ProductID) GO -- Select Table with Columnstore Index SELECT ProductID, SUM(UnitPrice) SumUnitPrice, AVG(UnitPrice) AvgUnitPrice, SUM(OrderQty) SumOrderQty, AVG(OrderQty) AvgOrderQty FROM [dbo].[MySalesOrderDetail] GROUP BY ProductID ORDER BY ProductID GO It is very clear from the results that query is performance extremely fast after creating ColumnStore Index. The amount of the pages it has to read to run query is drastically reduced as the column which are needed in the query are stored in the same page and query does not have to go through every single page to read those columns. If we enable execution plan and compare we can see that column store index performance way better than regular index in this case. Let us clean up the database. -- Cleanup DROP INDEX [IX_MySalesOrderDetail_ColumnStore] ON [dbo].[MySalesOrderDetail] GO TRUNCATE TABLE dbo.MySalesOrderDetail GO DROP TABLE dbo.MySalesOrderDetail GO In future posts we will see cases where Columnstore index is not appropriate solution as well few other tricks and tips of the columnstore index. Reference: Pinal Dave (http://blog.SQLAuthority.com) Filed under: Pinal Dave, PostADay, SQL, SQL Authority, SQL Index, SQL Optimization, SQL Performance, SQL Query, SQL Scripts, SQL Server, SQL Tips and Tricks, T SQL, Technology

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• jQuery Selector Tester and Cheat Sheet

  - by SGWellens

  I've always appreciated these tools: Expresso and XPath Builder. They make designing regular expressions and XPath selectors almost fun! Did I say fun? I meant less painful. Being able to paste/load text and then interactively play with the search criteria is infinitely better than the code/compile/run/test cycle. It's faster and you get a much better feel for how the expressions work. So, I decided to make my own interactive tool to test jQuery selectors:  jQuery Selector Tester.   Here's a sneak peek: Note: There are some existing tools you may like better: http://www.woods.iki.fi/interactive-jquery-tester.html http://www.w3schools.com/jquery/trysel.asp?filename=trysel_basic&jqsel=p.intro,%23choose My tool is different: It is one page. You can save it and run it locally without a Web Server. It shows the results as a list of iterated objects instead of highlighted html. A cheat sheet is on the same page as the tester which is handy. I couldn't upload an .htm or .html file to this site so I hosted it on my personal site here: jQuery Selector Tester. Design Highlights: To make the interactive search work, I added a hidden div to the page: <!--Hidden div holds DOM elements for jQuery to search--><div id="HiddenDiv" style="display: none"></div> When ready to search, the searchable html text is copied into the hidden div…this renders the DOM tree in the hidden div: // get the html to search, insert it to the hidden divvar Html = $("#TextAreaHTML").val();$("#HiddenDiv").html(Html); When doing a search, I modify the search pattern to look only in the HiddenDiv. To do that, I put a space between the patterns.  The space is the Ancestor operator (see the Cheat Sheet): // modify search string to only search in our// hidden div and do the searchvar SearchString = "#HiddenDiv " + SearchPattern;try{    var $FoundItems = $(SearchString);}   Big Fat Stinking Faux Pas: I was about to publish this article when I made a big mistake: I tested the tool with Mozilla FireFox. It blowed up…it blowed up real good. In the past I’ve only had to target IE so this was quite a revelation. When I started to learn JavaScript, I was disgusted to see all the browser dependent code. Who wants to spend their time testing against different browsers and versions of browsers? Adding a bunch of ‘if-else’ code is a tedious and thankless task. I avoided client code as much as I could. Then jQuery came along and all was good. It was browser independent and freed us from the tedium of worrying about version N of the Acme browser. Right? Wrong! I had used outerHTML to display the selected elements. The problem is Mozilla FireFox doesn’t implement outerHTML. I replaced this: // encode the html markupvar OuterHtml = $('<div/>').text(this.outerHTML).html(); With this: // encode the html markupvar Html = $('<div>').append(this).html();var OuterHtml = $('<div/>').text(Html).html(); Another problem was that Mozilla FireFox doesn’t implement srcElement. I replaced this: var Row = e.srcElement.parentNode;  With this: var Row = e.target.parentNode; Another problem was the indexing. The browsers have different ways of indexing. I replaced this: // this cell has the search pattern  var Cell = Row.childNodes[1];   // put the pattern in the search box and search                    $("#TextSearchPattern").val(Cell.innerText);  With this: // get the correct cell and the text in the cell// place the text in the seach box and serachvar Cell = $(Row).find("TD:nth-child(2)");var CellText = Cell.text();$("#TextSearchPattern").val(CellText);   So much for the myth of browser independence. Was I overly optimistic and gullible? I don’t think so. And when I get my millions from the deposed Nigerian prince I sent money to, you’ll see that having faith is not futile. Notes: My goal was to have a single standalone file. I tried to keep the features and CSS to a minimum–adding only enough to make it useful and visually pleasing. When testing, I often thought there was a problem with the jQuery selector. Invariable it was invalid html code. If your results aren't what you expect, don't assume it's the jQuery selector pattern: The html may be invalid. To help in development and testing, I added a double-click handler to the rows in the Cheat Sheet table. If you double-click a row, the search pattern is put in the search box, a search is performed and the page is scrolled so you can see the results. I left the test html and code in the page. If you are using a CDN (non-local) version of the jQuery libraray, the designer in Visual Studio becomes extremely slow.  That's why there are two version of the library in the header and one is commented out. For reference, here is the jQuery documentation on selectors: http://api.jquery.com/category/selectors/ Here is a much more comprehensive list of CSS selectors (which jQuery uses): http://www.w3.org/TR/CSS2/selector.html I hope someone finds this useful. Steve WellensCodeProject

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• Liskov Substitution Principle and the Oft Forgot Third Wheel

  - by Stacy Vicknair

  Liskov Substitution Principle (LSP) is a principle of object oriented programming that many might be familiar with from the SOLID principles mnemonic from Uncle Bob Martin. The principle highlights the relationship between a type and its subtypes, and, according to Wikipedia, is defined by Barbara Liskov and Jeanette Wing as the following principle:   Let be a property provable about objects of type . Then should be provable for objects of type where is a subtype of .   Rectangles gonna rectangulate The iconic example of this principle is illustrated with the relationship between a rectangle and a square. Let’s say we have a class named Rectangle that had a property to set width and a property to set its height. 1: Public Class Rectangle 2: Overridable Property Width As Integer 3: Overridable Property Height As Integer 4: End Class   We all at some point here that inheritance mocks an “IS A” relationship, and by gosh we all know square IS A rectangle. So let’s make a square class that inherits from rectangle. However, squares do maintain the same length on every side, so let’s override and add that behavior. 1: Public Class Square 2: Inherits Rectangle 3:  4: Private _sideLength As Integer 5:  6: Public Overrides Property Width As Integer 7: Get 8: Return _sideLength 9: End Get 10: Set(value As Integer) 11: _sideLength = value 12: End Set 13: End Property 14:  15: Public Overrides Property Height As Integer 16: Get 17: Return _sideLength 18: End Get 19: Set(value As Integer) 20: _sideLength = value 21: End Set 22: End Property 23: End Class   Now, say we had the following test: 1: Public Sub SetHeight_DoesNotAffectWidth(rectangle As Rectangle) 2: 'arrange 3: Dim expectedWidth = 4 4: rectangle.Width = 4 5:  6: 'act 7: rectangle.Height = 7 8:  9: 'assert 10: Assert.AreEqual(expectedWidth, rectangle.Width) 11: End Sub   If we pass in a rectangle, this test passes just fine. What if we pass in a square?   This is where we see the violation of Liskov’s Principle! A square might "IS A” to a rectangle, but we have differing expectations on how a rectangle should function than how a square should! Great expectations Here’s where we pat ourselves on the back and take a victory lap around the office and tell everyone about how we understand LSP like a boss. And all is good… until we start trying to apply it to our work. If I can’t even change functionality on a simple setter without breaking the expectations on a parent class, what can I do with subtyping? Did Liskov just tell me to never touch subtyping again? The short answer: NO, SHE DIDN’T. When I first learned LSP, and from those I’ve talked with as well, I overlooked a very important but not appropriately stressed quality of the principle: our expectations. Our inclination is to want a logical catch-all, where we can easily apply this principle and wipe our hands, drop the mic and exit stage left. That’s not the case because in every different programming scenario, our expectations of the parent class or type will be different. We have to set reasonable expectations on the behaviors that we expect out of the parent, then make sure that those expectations are met by the child. Any expectations not explicitly expected of the parent aren’t expected of the child either, and don’t register as a violation of LSP that prevents implementation. You can see the flexibility mentioned in the Wikipedia article itself: A typical example that violates LSP is a Square class that derives from a Rectangle class, assuming getter and setter methods exist for both width and height. The Square class always assumes that the width is equal with the height. If a Square object is used in a context where a Rectangle is expected, unexpected behavior may occur because the dimensions of a Square cannot (or rather should not) be modified independently. This problem cannot be easily fixed: if we can modify the setter methods in the Square class so that they preserve the Square invariant (i.e., keep the dimensions equal), then these methods will weaken (violate) the postconditions for the Rectangle setters, which state that dimensions can be modified independently. Violations of LSP, like this one, may or may not be a problem in practice, depending on the postconditions or invariants that are actually expected by the code that uses classes violating LSP. Mutability is a key issue here. If Square and Rectangle had only getter methods (i.e., they were immutable objects), then no violation of LSP could occur. What this means is that the above situation with a rectangle and a square can be acceptable if we do not have the expectation for width to leave height unaffected, or vice-versa, in our application. Conclusion – the oft forgot third wheel Liskov Substitution Principle is meant to act as a guidance and warn us against unexpected behaviors. Objects can be stateful and as a result we can end up with unexpected situations if we don’t code carefully. Specifically when subclassing, make sure that the subclass meets the expectations held to its parent. Don’t let LSP think you cannot deviate from the behaviors of the parent, but understand that LSP is meant to highlight the importance of not only the parent and the child class, but also of the expectations WE set for the parent class and the necessity of meeting those expectations in order to help prevent sticky situations.   Code examples, in both VB and C# Technorati Tags: LSV,Liskov Substitution Principle,Uncle Bob,Robert Martin,Barbara Liskov,Liskov

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• ASP.NET WebAPI Security 2: Identity Architecture

  - by Your DisplayName here!

  Pedro has beaten me to the punch with a detailed post (and diagram) about the WebAPI hosting architecture. So go read his post first, then come back so we can have a closer look at what that means for security. The first important takeaway is that WebAPI is hosting independent-  currently it ships with two host integration implementations – one for ASP.NET (aka web host) and WCF (aka self host). Pedro nicely shows the integration into the web host. Self hosting is not done yet so we will mainly focus on the web hosting case and I will point out security related differences when they exist. The interesting part for security (amongst other things of course) is the HttpControllerHandler (see Pedro’s diagram) – this is where the host specific representation of an HTTP request gets converted to the WebAPI abstraction (called HttpRequestMessage). The ConvertRequest method does the following: Create a new HttpRequestMessage. Copy URI, method and headers from the HttpContext. Copies HttpContext.User to the Properties<string, object> dictionary on the HttpRequestMessage. The key used for that can be found on HttpPropertyKeys.UserPrincipalKey (which resolves to “MS_UserPrincipal”). So the consequence is that WebAPI receives whatever IPrincipal has been set by the ASP.NET pipeline (in the web hosting case). Common questions are: Are there situations where is property does not get set? Not in ASP.NET – the DefaultAuthenticationModule in the HTTP pipeline makes sure HttpContext.User (and Thread.CurrentPrincipal – more on that later) are always set. Either to some authenticated user – or to an anonymous principal. This may be different in other hosting environments (again more on that later). Why so generic? Keep in mind that WebAPI is hosting independent and may run on a host that materializes identity completely different compared to ASP.NET (or .NET in general). This gives them a way to evolve the system in the future. How does WebAPI code retrieve the current client identity? HttpRequestMessage has an extension method called GetUserPrincipal() which returns the property as an IPrincipal. A quick look at self hosting shows that the moral equivalent of HttpControllerHandler.ConvertRequest() is HttpSelfHostServer.ProcessRequestContext(). Here the principal property gets only set when the host is configured for Windows authentication (inconsisteny). Do I like that? Well – yes and no. Here are my thoughts: I like that it is very straightforward to let WebAPI inherit the client identity context of the host. This might not always be what you want – think of an ASP.NET app that consists of UI and APIs – the UI might use Forms authentication, the APIs token based authentication. So it would be good if the two parts would live in a separate security world. It makes total sense to have this generic hand off point for identity between the host and WebAPI. It also makes total sense for WebAPI plumbing code (especially handlers) to use the WebAPI specific identity abstraction. But – c’mon we are running on .NET. And the way .NET represents identity is via IPrincipal/IIdentity. That’s what every .NET developer on this planet is used to. So I would like to see a User property of type IPrincipal on ApiController. I don’t like the fact that Thread.CurrentPrincipal is not populated. T.CP is a well established pattern as a one stop shop to retrieve client identity on .NET.  That makes a lot of sense – even if the name is misleading at best. There might be existing library code you want to call from WebAPI that makes use of T.CP (e.g. PrincipalPermission, or a simple .Name or .IsInRole()). Having the client identity as an ambient property is useful for code that does not have access to the current HTTP request (for calling GetUserPrincipal()). I don’t like the fact that that the client identity conversion from host to WebAPI is inconsistent. This makes writing security plumbing code harder. I think the logic should always be: If the host has a client identity representation, copy it. If not, set an anonymous principal on the request message. Btw – please don’t annoy me with the “but T.CP is static, and static is bad for testing” chant. T.CP is a getter/setter and, in fact I find it beneficial to be able to set different security contexts in unit tests before calling in some logic. And, in case you have wondered – T.CP is indeed thread static (and the name comes from a time where a logical operation was bound to a thread – which is not true anymore). But all thread creation APIs in .NET actually copy T.CP to the new thread they create. This is the case since .NET 2.0 and is certainly an improvement compared to how Win32 does things. So to sum it up: The host plumbing copies the host client identity to WebAPI (this is not perfect yet, but will surely be improved). or in other words: The current WebAPI bits don’t ship with any authentication plumbing, but solely use whatever authentication (and thus client identity) is set up by the host. WebAPI developers can retrieve the client identity from the HttpRequestMessage. Hopefully my proposed changes around T.CP and the User property on ApiController will be added. In the next post, I will detail how to add WebAPI specific authentication support, e.g. for Basic Authentication and tokens. This includes integrating the notion of claims based identity. After that we will look at the built-in authorization bits and how to improve them as well. Stay tuned.

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• Why is my external USB hard drive sometimes completely inaccessible?

  - by Eliah Kagan

  I have an external USB hard drive, consisting of an 1 TB SATA drive in a Rosewill RX35-AT-SU SLV Aluminum 3.5" Silver USB 2.0 External Enclosure, plugged into my SONY VAIO VGN-NS310F laptop. It is plugged directly into the computer (not through a hub). The drive inside the enclosure is a 7200 rpm Western Digital, but I don't remember the exact model. I can remove the drive from the enclosure (again), if people think it's necessary to know that detail. The drive is formatted ext4. I mount it dynamically with udisks on my Lubuntu 11.10 system, usually automatically via PCManFM. (I have had Lubuntu 12.04 on this machine, and experienced all this same behavior with that too.) Every once in a while--once or twice a day--it becomes inaccessible, and difficult to unmount. Attempting to unmount it with sudo umount ... gives an error message saying the drive is in use and suggesting fuser and lsof to find out what is using it. Killing processes found to be using the drive with fuser and lsof is sometimes sufficient to let me unmount it, but usually isn't. Once the drive is unmounted or the machine is rebooted, the drive will not mount. Plugging in the drive and turning it on registers nothing on the computer. dmesg is unchanged. The drive's access light usually blinks vigorously, as though the drive is being accessed constantly. Then eventually, after I keep the drive off for a while (half an hour), I am able to mount it again. While the drive doesn't work on this machine for a while, it will work immediately on another machine running the same version of Ubuntu. Sometimes bringing it back over from the other machine seems to "fix" it. Sometimes it doesn't. The drive doesn't always stop being accessible while mounted, before becoming unmountable. Sometimes it works fine, I turn off the computer, I turn the computer back on, and I cannot mount the drive. Currently this is the only drive with which I have this problem, but I've had problems that I think are the same as this, with different drives, on different Ubuntu machines. This laptop has another external USB drive plugged into it regularly, which doesn't have this problem. Unplugging that drive before plugging in the "problem" drive doesn't fix the problem. I've opened the drive up and made sure the connections were tight in the past, and that didn't seem to help (any more than waiting the same amount of time that it took to open and close the drive, before attempting to remount it). Does anyone have any ideas about what could be causing this, what troubleshooting steps I should perform, and/or how I could fix this problem altogether? Update: I tried replacing the USB data cable (from the enclosure to the laptop), as Merlin suggested. I should've tried that long ago, since it fits the symptoms perfectly (the drive works on another machine, which would make sense because the cable would be bent at a different angle, possibly completing a circuit of frayed wires). Unfortunately, though, this did not help--I have the same problem with the new cable. I'll try to provide additional detailed information about the drive inside the enclosure, next time I'm able to get the drive working. (At the moment I don't have another machine available to attach it.) Major Update (28 June 2012) The drive seems to have deteriorated considerably. I think this is so, because I've attached it to another machine and gotten lots of errors about invalid characters, when copying files from it. I am less interested in recovering data from the drive than I am in figuring out what is wrong with it. I specifically want to figure out if the problem is the drive or the enclosure. Now, when I plug the drive into the original machine where I was having the problems, it still doesn't appear (including with sudo fdisk -l), but it is recognized by the kernel and messages are added to dmesg. Most of the message consist of errors like this, repeated many times: [ 7.707593] sd 5:0:0:0: [sdc] Unhandled sense code [ 7.707599] sd 5:0:0:0: [sdc] Result: hostbyte=invalid driverbyte=DRIVER_SENSE [ 7.707606] sd 5:0:0:0: [sdc] Sense Key : Medium Error [current] [ 7.707614] sd 5:0:0:0: [sdc] Add. Sense: Unrecovered read error [ 7.707621] sd 5:0:0:0: [sdc] CDB: Read(10): 28 00 00 00 00 00 00 00 08 00 [ 7.707636] end_request: critical target error, dev sdc, sector 0 [ 7.707641] Buffer I/O error on device sdc, logical block 0 Here are all the lines from dmesg starting with when the drive is recognized. Please note that: I'm back to running Lubuntu 12.04 on this machine (and perhaps that's a factor in better error messages). Now that the drive has been plugged into another machine and back into this one, and also now that this machine is back to running 12.04, the drive's access light doesn't blink as I had described. Looking at the drive, it would appear as though it is working normally, with low or no access. This behavior (the errors) occurs when rebooting the machine with the drive plugged in, and also when manually plugging in the drive. A few of the messages are about /dev/sdb. That drive is working fine. The bad drive is /dev/sdc. I just didn't want to edit anything out from the middle.

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• Joins in single-table queries

  - by Rob Farley

  Tables are only metadata. They don’t store data. I’ve written something about this before, but I want to take a viewpoint of this idea around the topic of joins, especially since it’s the topic for T-SQL Tuesday this month. Hosted this time by Sebastian Meine (@sqlity), who has a whole series on joins this month. Good for him – it’s a great topic. In that last post I discussed the fact that we write queries against tables, but that the engine turns it into a plan against indexes. My point wasn’t simply that a table is actually just a Clustered Index (or heap, which I consider just a special type of index), but that data access always happens against indexes – never tables – and we should be thinking about the indexes (specifically the non-clustered ones) when we write our queries. I described the scenario of looking up phone numbers, and how it never really occurs to us that there is a master list of phone numbers, because we think in terms of the useful non-clustered indexes that the phone companies provide us, but anyway – that’s not the point of this post. So a table is metadata. It stores information about the names of columns and their data types. Nullability, default values, constraints, triggers – these are all things that define the table, but the data isn’t stored in the table. The data that a table describes is stored in a heap or clustered index, but it goes further than this. All the useful data is going to live in non-clustered indexes. Remember this. It’s important. Stop thinking about tables, and start thinking about indexes. So let’s think about tables as indexes. This applies even in a world created by someone else, who doesn’t have the best indexes in mind for you. I’m sure you don’t need me to explain Covering Index bit – the fact that if you don’t have sufficient columns “included” in your index, your query plan will either have to do a Lookup, or else it’ll give up using your index and use one that does have everything it needs (even if that means scanning it). If you haven’t seen that before, drop me a line and I’ll run through it with you. Or go and read a post I did a long while ago about the maths involved in that decision. So – what I’m going to tell you is that a Lookup is a join. When I run SELECT CustomerID FROM Sales.SalesOrderHeader WHERE SalesPersonID = 285; against the AdventureWorks2012 get the following plan: I’m sure you can see the join. Don’t look in the query, it’s not there. But you should be able to see the join in the plan. It’s an Inner Join, implemented by a Nested Loop. It’s pulling data in from the Index Seek, and joining that to the results of a Key Lookup. It clearly is – the QO wouldn’t call it that if it wasn’t really one. It behaves exactly like any other Nested Loop (Inner Join) operator, pulling rows from one side and putting a request in from the other. You wouldn’t have a problem accepting it as a join if the query were slightly different, such as SELECT sod.OrderQty FROM Sales.SalesOrderHeader AS soh JOIN Sales.SalesOrderDetail as sod on sod.SalesOrderID = soh.SalesOrderID WHERE soh.SalesPersonID = 285; Amazingly similar, of course. This one is an explicit join, the first example was just as much a join, even thought you didn’t actually ask for one. You need to consider this when you’re thinking about your queries. But it gets more interesting. Consider this query: SELECT SalesOrderID FROM Sales.SalesOrderHeader WHERE SalesPersonID = 276 AND CustomerID = 29522; It doesn’t look like there’s a join here either, but look at the plan. That’s not some Lookup in action – that’s a proper Merge Join. The Query Optimizer has worked out that it can get the data it needs by looking in two separate indexes and then doing a Merge Join on the data that it gets. Both indexes used are ordered by the column that’s indexed (one on SalesPersonID, one on CustomerID), and then by the CIX key SalesOrderID. Just like when you seek in the phone book to Farley, the Farleys you have are ordered by FirstName, these seek operations return the data ordered by the next field. This order is SalesOrderID, even though you didn’t explicitly put that column in the index definition. The result is two datasets that are ordered by SalesOrderID, making them very mergeable. Another example is the simple query SELECT CustomerID FROM Sales.SalesOrderHeader WHERE SalesPersonID = 276; This one prefers a Hash Match to a standard lookup even! This isn’t just ordinary index intersection, this is something else again! Just like before, we could imagine it better with two whole tables, but we shouldn’t try to distinguish between joining two tables and joining two indexes. The Query Optimizer can see (using basic maths) that it’s worth doing these particular operations using these two less-than-ideal indexes (because of course, the best indexese would be on both columns – a composite such as (SalesPersonID, CustomerID – and it would have the SalesOrderID column as part of it as the CIX key still). You need to think like this too. Not in terms of excusing single-column indexes like the ones in AdventureWorks2012, but in terms of having a picture about how you’d like your queries to run. If you start to think about what data you need, where it’s coming from, and how it’s going to be used, then you will almost certainly write better queries. …and yes, this would include when you’re dealing with regular joins across multiples, not just against joins within single table queries.

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• JQuery and the multiple date selector

  - by David Carter

  Overview I recently needed to build a web page that would allow a user to capture some information and most importantly select multiple dates. This functionality was core to the application and hence had to be easy and quick to do. This is a public facing website so it had to be intuitive and very responsive. On the face of it it didn't seem too hard, I know enough juery to know what it is capable of and I was pretty sure that there would be some plugins that would help speed things along the way. I'm using ASP.Net MVC for this project as I really like the control that it gives you over the generated html and javascript. After years of Web Forms development it makes me feel like a web developer again and puts a smile on my face, that can only be a good thing!   The Calendar The first item that I needed on this page was a calender and I wanted the ability to: have the calendar be always visible select/deselect multiple dates at the same time bind to the select/deselect event so that I could update a seperate listing of the selected dates allow the user to move to another month and still have the calender remember any dates in the previous month I was hoping that there was a jQuery plugin that would meet my requirements and luckily there was! The jQuery datepicker does everything I want and there is quite a bit of documentation on how to use it. It makes use of a javascript date library date.js which I had not come across before but has a number of very useful date utilities that I have used elsewhere in the project. As you can see from the image there still needs to be some styling done! But there will be plenty of time for that later. The calendar clearly shows which dates the user has selected in red and i also make use of an unordered list to show the the selected dates so the user can always clearly see what has been selected even if they move to another month on the calendar. The javascript code that is responsible for listening to events on the calendar and synchronising the list look as follows: <script type="text/javascript">     $(function () {         $('.datepicker').datePicker({ inline: true, selectMultiple: true })         .bind(             'dateSelected',             function (e, selectedDate, $td, state) {                                 var dateInMillisecs = selectedDate.valueOf();                 if (state) { //adding a date                     var newDate = new Date(selectedDate);                     //insert the new item into the correct place in the list                     var listitems = $('#dateList').children('li').get();                     var liToAdd = "<li id='" + dateInMillisecs + "' >" + newDate.toString('ddd dd MMM yyyy') + "</li>";                     var targetIndex = -1;                     for (var i = 0; i < listitems.length; i++) {                         if (dateInMillisecs <= listitems[i].id) {                             targetIndex = i;                             break;                         }                     }                     if (targetIndex < 0) {                         $('#dateList').append(liToAdd);                     }                     else {                         $($('#dateList').children("li")[targetIndex]).before(liToAdd);                     }                 }                 else {//removing a date                     $('ul #' + dateInMillisecs).remove();                 }             }         )     }); When a date is selected on the calendar a function is called with a number of parameters passed to it. The ones I am particularly interested in are selectedDate and state. State tells me whether the user has selected or deselected the date passed in the selectedDate parameter. The <ul> that I am using to show the date has an id of dateList and this is what I will be adding and removing <li> items from. To make things a little more logical for the user I decided that the date should be sorted in chronological order, this means that each time a new date is selected it need to be placed in the correct position in the list. One way to do this would be just to append a new <li> to the list and then sort the whole list. However the approach I took was to get an array of all the items in the list var listitems = ('#dateList').children('li').get(); and then check the value of each item in the array against my new date and as soon as I found the case where the new date was less than the current item remember that position in the list as this is where I would insert it later. To make this work easily I decided to store a numeric representation of each date in the list in the id attribute of each <li> element. Fortunately javascript natively stores dates as the number of milliseconds since 1 Jan 1970. var dateInMillisecs = selectedDate.valueOf(); Please note that this is the value of the date in UTC! I always like to store dates in UTC as I learnt a long time ago that it saves a lot of refactoring at a later date... When I convert the dates back to their original back on the server I will need the UTC offset that was used when calculating the dates, this and how to actually serialise the dates and get them posted back will be the subject of another post.

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• Windows 7 doesn't boot after Ubuntu install

  - by Omu

  I had windows 7 installed on my pc, then I installed Ubuntu 10.10/ During the installation process I have chosen to manually set my partitions: I set a 10GB drive for ubuntu root 1GB drive for swap and for boot drive I've chosen the one used by windows 7 Now I can boot ubuntu, I have the windows 7 option in the boot list, but when I choose Windows 7, it shows me a black screen for a second and returns back to boot screen. Boot Info Script 0.55 dated February 15th, 2010 ============================= Boot Info Summary: ============================== = Windows is installed in the MBR of /dev/sda sda1: _________________________________________________________________________ File system: ntfs Boot sector type: Grub 2 Boot sector info: Grub 2 is installed in the boot sector of sda1 and looks at sector 304908237 of the same hard drive for core.img, but core.img can not be found at this location. No errors found in the Boot Parameter Block. Operating System: Windows 7 Boot files/dirs: /bootmgr /Boot/BCD /Windows/System32/winload.exe sda2: _________________________________________________________________________ File system: ntfs Boot sector type: Windows XP Boot sector info: No errors found in the Boot Parameter Block. Operating System: Boot files/dirs: sda3: _________________________________________________________________________ File system: Extended Partition Boot sector type: - Boot sector info: sda5: _________________________________________________________________________ File system: ext4 Boot sector type: - Boot sector info: Operating System: Ubuntu 10.10 Boot files/dirs: /boot/grub/grub.cfg /etc/fstab /boot/grub/core.img sda4: _________________________________________________________________________ File system: swap Boot sector type: - Boot sector info: =========================== Drive/Partition Info: ============================= Drive: sda ___________________ _____________________________________________________ Disk /dev/sda: 160.0 GB, 160041885696 bytes 255 heads, 63 sectors/track, 19457 cylinders, total 312581808 sectors Units = sectors of 1 * 512 = 512 bytes Sector size (logical/physical): 512 bytes / 512 bytes Partition Boot Start End Size Id System /dev/sda1 * 63 62,894,474 62,894,412 7 HPFS/NTFS /dev/sda2 62,894,478 291,579,749 228,685,272 7 HPFS/NTFS /dev/sda3 291,579,811 309,157,937 17,578,127 5 Extended /dev/sda5 291,579,813 309,157,937 17,578,125 83 Linux /dev/sda4 309,159,936 312,580,095 3,420,160 82 Linux swap / Solaris blkid -c /dev/null: ____________________________________________________________ Device UUID TYPE LABEL /dev/sda1 1266BB2766BB0A8D ntfs /dev/sda2 BEDBF1147C76F703 ntfs DATA /dev/sda3: PTTYPE="dos" /dev/sda4 dd38226d-c7c9-4ae5-a726-6d18d34a22e4 swap /dev/sda5 e1dafd1c-f855-406b-8f9a-f9d527c70255 ext4 /dev/sda: PTTYPE="dos" ============================ "mount | grep ^/dev output: =========================== Device Mount_Point Type Options /dev/sda5 / ext4 (rw,errors=remount-ro,commit=0) =========================== sda5/boot/grub/grub.cfg: =========================== # # DO NOT EDIT THIS FILE # # It is automatically generated by grub-mkconfig using templates # from /etc/grub.d and settings from /etc/default/grub # ### BEGIN /etc/grub.d/00_header ### if [ -s $prefix/grubenv ]; then set have_grubenv=true load_env fi set default="0" if [ "${prev_saved_entry}" ]; then set saved_entry="${prev_saved_entry}" save_env saved_entry set prev_saved_entry= save_env prev_saved_entry set boot_once=true fi function savedefault { if [ -z "${boot_once}" ]; then saved_entry="${chosen}" save_env saved_entry fi } function recordfail { set recordfail=1 if [ -n "${have_grubenv}" ]; then if [ -z "${boot_once}" ]; then save_env recordfail; fi; fi } function load_video { insmod vbe insmod vga } insmod part_msdos insmod ext2 set root='(hd0,msdos5)' search --no-floppy --fs-uuid --set e1dafd1c-f855-406b-8f9a-f9d527c70255 if loadfont /usr/share/grub/unicode.pf2 ; then set gfxmode=640x480 load_video insmod gfxterm fi terminal_output gfxterm insmod part_msdos insmod ext2 set root='(hd0,msdos5)' search --no-floppy --fs-uuid --set e1dafd1c-f855-406b-8f9a-f9d527c70255 set locale_dir=($root)/boot/grub/locale set lang=en insmod gettext if [ "${recordfail}" = 1 ]; then set timeout=-1 else set timeout=10 fi ### END /etc/grub.d/00_header ### ### BEGIN /etc/grub.d/05_debian_theme ### set menu_color_normal=white/black set menu_color_highlight=black/light-gray ### END /etc/grub.d/05_debian_theme ### ### BEGIN /etc/grub.d/10_linux ### menuentry 'Ubuntu, with Linux 2.6.35-22-generic' --class ubuntu --class gnu-linux --class gnu --class os { recordfail insmod part_msdos insmod ext2 set root='(hd0,msdos5)' search --no-floppy --fs-uuid --set e1dafd1c-f855-406b-8f9a-f9d527c70255 linux /boot/vmlinuz-2.6.35-22-generic root=UUID=e1dafd1c-f855-406b-8f9a-f9d527c70255 ro quiet splash initrd /boot/initrd.img-2.6.35-22-generic } menuentry 'Ubuntu, with Linux 2.6.35-22-generic (recovery mode)' --class ubuntu --class gnu-linux --class gnu --class os { recordfail insmod part_msdos insmod ext2 set root='(hd0,msdos5)' search --no-floppy --fs-uuid --set e1dafd1c-f855-406b-8f9a-f9d527c70255 echo 'Loading Linux 2.6.35-22-generic ...' linux /boot/vmlinuz-2.6.35-22-generic root=UUID=e1dafd1c-f855-406b-8f9a-f9d527c70255 ro single echo 'Loading initial ramdisk ...' initrd /boot/initrd.img-2.6.35-22-generic } ### END /etc/grub.d/10_linux ### ### BEGIN /etc/grub.d/20_linux_xen ### ### END /etc/grub.d/20_linux_xen ### ### BEGIN /etc/grub.d/20_memtest86+ ### menuentry "Memory test (memtest86+)" { insmod part_msdos insmod ext2 set root='(hd0,msdos5)' search --no-floppy --fs-uuid --set e1dafd1c-f855-406b-8f9a-f9d527c70255 linux16 /boot/memtest86+.bin } menuentry "Memory test (memtest86+, serial console 115200)" { insmod part_msdos insmod ext2 set root='(hd0,msdos5)' search --no-floppy --fs-uuid --set e1dafd1c-f855-406b-8f9a-f9d527c70255 linux16 /boot/memtest86+.bin console=ttyS0,115200n8 } ### END /etc/grub.d/20_memtest86+ ### ### BEGIN /etc/grub.d/30_os-prober ### menuentry "Windows 7 (loader) (on /dev/sda1)" { insmod part_msdos insmod ntfs set root='(hd0,msdos1)' search --no-floppy --fs-uuid --set 1266bb2766bb0a8d chainloader +1 } ### END /etc/grub.d/30_os-prober ### ### BEGIN /etc/grub.d/40_custom ### # This file provides an easy way to add custom menu entries. Simply type the # menu entries you want to add after this comment. Be careful not to change # the 'exec tail' line above. ### END /etc/grub.d/40_custom ### ### BEGIN /etc/grub.d/41_custom ### if [ -f $prefix/custom.cfg ]; then source $prefix/custom.cfg; fi ### END /etc/grub.d/41_custom ### =============================== sda5/etc/fstab: =============================== # /etc/fstab: static file system information. # # Use 'blkid -o value -s UUID' to print the universally unique identifier # for a device; this may be used with UUID= as a more robust way to name # devices that works even if disks are added and removed. See fstab(5). # # proc /proc proc nodev,noexec,nosuid 0 0 /dev/sda5 / ext4 errors=remount-ro 0 1 # swap was on /dev/sda4 during installation UUID=dd38226d-c7c9-4ae5-a726-6d18d34a22e4 none swap sw 0 0 =================== sda5: Location of files loaded by Grub: =================== 156.1GB: boot/grub/core.img 156.3GB: boot/grub/grub.cfg 149.9GB: boot/initrd.img-2.6.35-22-generic 156.3GB: boot/vmlinuz-2.6.35-22-generic 149.9GB: initrd.img 156.3GB: vmlinuz

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• Windows 8 Live Accounts and the actual Windows Account

  - by Rick Strahl

  As if Windows Security wasn't confusing enough, in Windows 8 we get thrown yet another curve ball with Windows Live accounts to logon. When I set up my Windows 8 machine I originally set it up with a 'real', non-live account that I always use on my Windows machines. I did this mainly so I have a matching account for resources around my home and intranet network so I could log on to network resources properly. At some point later I decided to set up Windows Live security just to see how changes things. Windows wants you to use Windows Live Windows 8 logins are required in order for the Windows RT account info to work. Not that I care - since installing Windows 8 I've maybe spent 10 minutes with Windows RT because - well it's pretty freaking sucky on the desktop. From shitty apps to mis-managed screen real estate I can't say that there's anything compelling there to date, but then I haven't looked that hard either. Anyway… I set up the Windows Live account to see if that changes things. It does - I do get all my live logins to work from Live Account so that Twitter and Facebook posts and pictures and calendars all show up on live tiles on the start screen and in the actual apps. That's nice-ish, but hardly that exciting given that all of the apps tied to those live tiles are average at best. And it would have been nice if all of this could be done without being forced into running with a Windows Live User Account - this all feels like strong-arming you into moving into Microsofts walled garden… and that's probably what it's meant to do. Who am I? The real problem to me though is that these Windows Live and raw Windows User accounts are a bit unpredictable especially when it comes to developer information about the account and which credentials to use. So for example Windows reports folder security like this: Notice it's showing my Windows Live account. Now if I go to Edit and try to add my Windows user account (rstrahl) it'll just automatically show up as the live account. On the other hand though the underlying system sees everything as my real Windows account. After I switched to a Windows Live login account and I have to login to Windows with my Live account, what do you suppose this returns?Console.WriteLine(Environment.UserName); It returns my raw Windows user account (rstrahl). All my permissions, all my actual settings and the desktop console altogether run under that account. If I look in TaskManager (or Process Explorer for me) I see: Everything running on the desktop shell with my login running under my Windows user account. I suppose it makes sense, but where is that association happening? When I switched to a Windows Live account, nowhere did I associate my real account with the Live account - it just happened. And looking through the account configuration dialogs I can't find any reference to the raw Windows account. Other than switching back I see no mention anywhere of the raw Windows account - everything refers to the Live account. Right then, clear as potato soup! So this is who you really are! The problem is that in some situations this schizophrenic account behavior gets a bit weird. Today I was running a local Web application in IIS that uses Windows Authentication - I tried to log-in with my real Windows account login because that's what I'm used to using with WINDOWS freaking Authentication through IIS. But… it failed. I checked my IIS settings, my apps login settings and I just could not for the life of me get into the site with my Windows username. That is until I finally realized that I should try using my Windows Live credentials instead. And that worked. So now in this Windows Authentication dialog I had to type in my Live ID and password, which is - just weird. Then in IIS if I look at a Trace page (or in my case my app's Status page) I see that the logged on account is - my Windows user account. What's really annoying about this is that in some places it uses the live account in other places it uses my Windows account. If I remote desktop into my Web server online - I have to use the local authentication dialog but I have to put in my real Windows credentials not the Live account. Oh yes, it's all so terribly intuitive and logical… So in summary, when you log on with a Live account you are actually mapped to an underlying Windows user. In any application if you check the user name it'll be the underlying user account (not sure what happens in a Windows RT app or even what mechanism is used there to get the user name info).  When logging on to local machine resource with user name and password you have to use your Live IDs even if the permissions on the resources are mapped to your underlying Windows account. Easy enough I suppose, but still not exactly intuitive behavior…© Rick Strahl, West Wind Technologies, 2005-2012Posted in Windows   Tweet !function(d,s,id){var js,fjs=d.getElementsByTagName(s)[0];if(!d.getElementById(id)){js=d.createElement(s);js.id=id;js.src="//platform.twitter.com/widgets.js";fjs.parentNode.insertBefore(js,fjs);}}(document,"script","twitter-wjs"); (function() { var po = document.createElement('script'); po.type = 'text/javascript'; po.async = true; po.src = 'https://apis.google.com/js/plusone.js'; var s = document.getElementsByTagName('script')[0]; s.parentNode.insertBefore(po, s); })();

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• Integration with Multiple Versions of BizTalk HL7 Accelerator Schemas

  - by Paul Petrov

  Microsoft BizTalk Accelerator for HL7 comes with multiple versions of the HL7 implementation. One of the typical integration tasks is to receive one format and transmit another. For example, system A works HL7 v2.4 messages, system B with v2.3, and system C with v2.2. The system A is exchanging messages with B and C. The logical solution is to create schemas in separate namespaces for each system and assign maps on send ports. Schematic diagram of the messaging solution is shown below:   Nothing is complex about that conceptually. On the implementation level things can get nasty though because of the elaborate nature of HL7 schemas and sheer amount of message types involved. If trying to implement maps directly in BizTalk Map Editor one would quickly get buried by thousands of links between subfields of HL7 segments. Since task is repetitive because HL7 segments are reused between message types it's natural to take advantage of such modular structure and reduce amount of work through reuse. Here's where it makes sense to switch from visual map editor to old plain XSLT. The implementation is done in three steps. First, create XSL templates to map from segments of one version to another. This can be done using BizTalk Map Editor subsequently copying and modifying generated XSL code to create one xsl:template per segment. Group all segments for format mapping in one XSL file (we call it SegmentTemplates.xsl). Here's how template for the PID segment (Patient Identification) would look like this: <xsl:template name="PID"> <PID_PatientIdentification> <xsl:if test="PID_PatientIdentification/PID_1_SetIdPatientId"> <PID_1_SetIdPid> <xsl:value-of select="PID_PatientIdentification/PID_1_SetIdPatientId/text()" /> </PID_1_SetIdPid> </xsl:if> <xsl:for-each select="PID_PatientIdentification/PID_2_PatientIdExternalId"> <PID_2_PatientId> <xsl:if test="CX_0_Id"> <CX_0_Id> <xsl:value-of select="CX_0_Id/text()" /> </CX_0_Id> </xsl:if> <xsl:if test="CX_1_CheckDigit"> <CX_1_CheckDigitSt> <xsl:value-of select="CX_1_CheckDigit/text()" /> </CX_1_CheckDigitSt> </xsl:if> <xsl:if test="CX_2_CodeIdentifyingTheCheckDigitSchemeEmployed"> <CX_2_CodeIdentifyingTheCheckDigitSchemeEmployed> <xsl:value-of select="CX_2_CodeIdentifyingTheCheckDigitSchemeEmployed/text()" /> </CX_2_CodeIdentifyingTheCheckDigitSchemeEmployed> . . . // skipped for brevity This is the most tedious and time consuming part. Templates can be created for only those segments that are used in message interchange. Once this is done the rest goes much easier. The next step is to create message type specific XSL that references (imports) segment templates XSL file. Inside this file simple call segment templates in appropriate places. For example, beginning of the mapping XSL for ADT_A01 message would look like this:   <xsl:import href="SegmentTemplates_23_to_24.xslt" />  <xsl:output omit-xml-declaration="yes" method="xml" version="1.0" />   <xsl:template match="/">    <xsl:apply-templates select="s0:ADT_A01_23_GLO_DEF" />  </xsl:template>   <xsl:template match="s0:ADT_A01_23_GLO_DEF">    <ns0:ADT_A01_24_GLO_DEF>      <xsl:call-template name="EVN" />      <xsl:call-template name="PID" />      <xsl:for-each select="PD1_PatientDemographic">        <xsl:call-template name="PD1" />      </xsl:for-each>      <xsl:call-template name="PV1" />      <xsl:for-each select="PV2_PatientVisitAdditionalInformation">        <xsl:call-template name="PV2" />      </xsl:for-each> This code simply calls segment template directly for required singular elements and in for-each loop for optional/repeating elements. And lastly, create BizTalk map (btm) that references message type specific XSL. It is essentially empty map with Custom XSL Path set to appropriate XSL: In the end, you will end up with one segment templates file that is referenced by many message type specific XSL files which in turn used by BizTalk maps. Once all segment maps are created they are widely reusable and all the rest work is very simple and clean.

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• Yet another blog about IValueConverter

  - by codingbloke

  After my previous blog on a Generic Boolean Value Converter I thought I might as well blog up another IValueConverter implementation that I use. The Generic Boolean Value Converter effectively converters an input which only has two possible values to one of two corresponding objects.  The next logical step would be to create a similar converter that can take an input which has multiple (but finite and discrete) values to one of multiple corresponding objects.  To put it more simply a Generic Enum Value Converter. Now we already have a tool that can help us in this area, the ResourceDictionary.  A simple IValueConverter implementation around it would create a StringToObjectConverter like so:- StringToObjectConverter using System; using System.Windows; using System.Windows.Data; using System.Linq; using System.Windows.Markup; namespace SilverlightApplication1 {     [ContentProperty("Items")]     public class StringToObjectConverter : IValueConverter     {         public ResourceDictionary Items { get; set; }         public string DefaultKey { get; set; }                  public StringToObjectConverter()         {             DefaultKey = "__default__";         }         public virtual object Convert(object value, Type targetType, object parameter, System.Globalization.CultureInfo culture)         {             if (value != null && Items.Contains(value.ToString()))                 return Items[value.ToString()];             else                 return Items[DefaultKey];         }         public virtual object ConvertBack(object value, Type targetType, object parameter, System.Globalization.CultureInfo culture)         {             return Items.FirstOrDefault(kvp => value.Equals(kvp.Value)).Key;         }     } } There are some things to note here.  The bulk of managing the relationship between an object instance and the related string key is handled by the Items property being an ResourceDictionary.  Also there is a catch all “__default__” key value which allows for only a subset of the possible input values to mapped to an object with the rest falling through to the default. We can then set one of these up in Xaml:-             <local:StringToObjectConverter x:Key="StatusToBrush">                 <ResourceDictionary>                     <SolidColorBrush Color="Red" x:Key="Overdue" />                     <SolidColorBrush Color="Orange" x:Key="Urgent" />                     <SolidColorBrush Color="Silver" x:Key="__default__" />                 </ResourceDictionary>             </local:StringToObjectConverter> You could well imagine that in the model being bound these key names would actually be members of an enum.  This still works due to the use of ToString in the Convert method.  Hence the only requirement for the incoming object is that it has a ToString implementation which generates a sensible string instead of simply the type name. I can’t imagine right now a scenario where this converter would be used in a TwoWay binding but there is no reason why it can’t.  I prefer to avoid leaving the ConvertBack throwing an exception if that can be be avoided.  Hence it just enumerates the KeyValuePair entries to find a value that matches and returns the key its mapped to. Ah but now my sense of balance is assaulted again.  Whilst StringToObjectConverter is quite happy to accept an enum type via the Convert method it returns a string from the ConvertBack method not the original input enum type that arrived in the Convert.  Now I could address this by complicating the ConvertBack method and examining the targetType parameter etc.  However I prefer to a different approach, deriving a new EnumToObjectConverter class instead. EnumToObjectConverter using System; namespace SilverlightApplication1 {     public class EnumToObjectConverter : StringToObjectConverter     {         public override object Convert(object value, Type targetType, object parameter, System.Globalization.CultureInfo culture)         {             string key = Enum.GetName(value.GetType(), value);             return base.Convert(key, targetType, parameter, culture);         }         public override object ConvertBack(object value, Type targetType, object parameter, System.Globalization.CultureInfo culture)         {             string key = (string)base.ConvertBack(value, typeof(String), parameter, culture);             return Enum.Parse(targetType, key, false);         }     } }   This is a more belts and braces solution with specific use of Enum.GetName and Enum.Parse.  Whilst its more explicit in that the a developer has to  choose to use it, it is only really necessary when using TwoWay binding, in OneWay binding the base StringToObjectConverter would serve just as well. The observant might note that there is actually no “Generic” aspect to this solution in the end.  The use of a ResourceDictionary eliminates the need for that.

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• ASP.NET Web Forms Extensibility: Control Adapters

  - by Ricardo Peres

  All ASP.NET controls from version 2.0 can be associated with a control adapter. A control adapter is a class that inherits from ControlAdapter and it has the chance to interact with the control(s) it is targeting so as to change some of its properties or alter its output. I talked about control adapters before and they really a cool feature. The ControlAdapter class exposes virtual methods for some well known lifecycle events, OnInit, OnLoad, OnPreRender and OnUnload that closely match their Control counterparts, but are fired before them. Because the control adapter has a reference to its target Control, it can cast it to its concrete class and do something with it before its lifecycle events are actually fired. The adapter is also notified before the control is rendered (BeginRender), after their children are renderes (RenderChildren) and after itself is rendered (Render): this way the adapter can modify the control’s output. Control adapters may be specified for any class inheriting from Control, including abstract classes, web server controls and even pages. You can, for example, specify a control adapter for the WebControl and UserControl classes, but, curiously, not for Control itself. When specifying a control adapter for a page, it must inherit from PageAdapter instead of ControlAdapter. The adapter for a control, if specified, can be found on the protected Adapter property, and for a page, on the PageAdapter property. The first use of control adapters that came to my attention was for changing the output of standard ASP.NET web controls so that they were more based on CSS and less on HTML tables: it was the CSS Friendly Control Adapters project, now available at http://code.google.com/p/aspnetcontroladapters/. They are interesting because you specify them in one location and they apply anywhere a control of the target type is created. Mind you, it applies to controls declared on markup as well as controls created by code with the new operator. So, how do you use control adapters? The most usual way is through a browser definition file. In it, you specify a set of control adapters and their target controls, for a given browser. This browser definition file is a XML file with extension .Browser, and can either be global (%WINDIR%\Microsoft.NET\Framework64\vXXXX\Config\Browsers) or local to the web application, in which case, it must be placed inside the App_Browsers folder at the root of the web site. It looks like this: 1: <browsers> 2: <browser refID="Default"> 3: <controlAdapters> 4: <adapter controlType="System.Web.UI.WebControls.TextBox" adapterType="MyNamespace.TextBoxAdapter, MyAssembly" /> 5: </controlAdapters> 6: </browser> 7: </browsers> A browser definition file targets a specific browser, so you can have different definitions for Chrome, IE, Firefox, Opera, as well as for specific version of each of those (like IE8, Firefox3). Alternatively, if you set the target to Default, it will apply to all. The reason to pick a specific browser and version might be, for example, in order to circumvent some limitation present in that specific version, so that on markup you don’t need to be concerned with that. Another option is through the the current Browser object of the request: 1: this.Context.Request.Browser.Adapters.Add(typeof(TextBox).FullName, typeof(TextBoxAdapter).FullName); This must go very early on the page lifecycle, for example, on the OnPreInit event, or even on Application_Start. You have to specify the full class name for both the target control and the adapter. Of course, you have to do this for every request, because it won’t be persisted. As an example, you may know that the classic TextBox control renders an HTML input tag if its TextMode is set to SingleLine and a textarea if set to MultiLine. Because the textarea has no notion of maximum length, unlike the input, something must be done in order to enforce this. Here’s a simple suggestion: 1: public class TextBoxControlAdapter : ControlAdapter 2: { 3: protected TextBox Target 4: { 5: get 6: { 7: return (this.Control as TextBox); 8: } 9: } 10:  11: protected override void OnLoad(EventArgs e) 12: { 13: if ((this.Target.MaxLength > 0) && (this.Target.TextMode == TextBoxMode.MultiLine)) 14: { 15: if (this.Target.Page.ClientScript.IsClientScriptBlockRegistered("TextBox_KeyUp") == false) 16: { 17: if (this.Target.Page.ClientScript.IsClientScriptBlockRegistered(this.Target.Page.GetType(), "TextBox_KeyUp") == false) 18: { 19: String script = String.Concat("function TextBox_KeyUp(sender) { if (sender.value.length > ", this.Target.MaxLength, ") { sender.value = sender.value.substr(0, ", this.Target.MaxLength, "); } }\n"); 20:  21: this.Target.Page.ClientScript.RegisterClientScriptBlock(this.Target.Page.GetType(), "TextBox_KeyUp", script, true); 22: } 23:  24: this.Target.Attributes["onkeyup"] = "TextBox_KeyUp(this)"; 25: } 26: } 27: 28: base.OnLoad(e); 29: } 30: } What it does is, for every TextBox control, if it is set for multi line and has a defined maximum length, it injects some JavaScript that will filter out any content that exceeds this maximum length. This will occur for any TextBox that you may have on your site, or any class that inherits from it. You can use any of the previous options to register this adapter. Stay tuned for more ASP.NET Web Forms extensibility tips!

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• MySQL Connect 9 Days Away – Optimizer Sessions

  - by Bertrand Matthelié

  72 1024x768 Normal 0 false false false EN-US X-NONE X-NONE /* Style Definitions */ table.MsoNormalTable {mso-style-name:"Table Normal"; mso-tstyle-rowband-size:0; mso-tstyle-colband-size:0; mso-style-noshow:yes; mso-style-priority:99; mso-style-qformat:yes; mso-style-parent:""; mso-padding-alt:0cm 5.4pt 0cm 5.4pt; mso-para-margin:0cm; mso-para-margin-bottom:.0001pt; mso-pagination:widow-orphan; font-size:11.0pt; font-family:"Calibri","sans-serif"; mso-ascii-font-family:Calibri; mso-ascii-theme-font:minor-latin; mso-fareast-font-family:"Times New Roman"; mso-fareast-theme-font:minor-fareast; mso-hansi-font-family:Calibri; mso-hansi-theme-font:minor-latin; mso-bidi-font-family:"Times New Roman"; mso-bidi-theme-font:minor-bidi;} Following my previous blog post focusing on InnoDB talks at MySQL Connect, let us review today the sessions focusing on the MySQL Optimizer: Saturday, 11.30 am, Room Golden Gate 6: MySQL Optimizer Overview—Olav Sanstå, Oracle The goal of MySQL optimizer is to take a SQL query as input and produce an optimal execution plan for the query. This session presents an overview of the main phases of the MySQL optimizer and the primary optimizations done to the query. These optimizations are based on a combination of logical transformations and cost-based decisions. Examples of optimization strategies the presentation covers are the main query transformations, the join optimizer, the data access selection strategies, and the range optimizer. For the cost-based optimizations, an overview of the cost model and the data used for doing the cost estimations is included. Saturday, 1.00 pm, Room Golden Gate 6: Overview of New Optimizer Features in MySQL 5.6—Manyi Lu, Oracle Many optimizer features have been added into MySQL 5.6. This session provides an introduction to these great features. Multirange read, index condition pushdown, and batched key access will yield huge performance improvements on large data volumes. Structured explain, explain for update/delete/insert, and optimizer tracing will help users analyze and speed up queries. And last but not least, the session covers subquery optimizations in Release 5.6. Saturday, 7.00 pm, Room Golden Gate 4: BoF: Query Optimizations: What Is New and What Is Coming? This BoF presents common techniques for query optimization, covers what is new in MySQL 5.6, and provides a discussion forum in which attendees can tell the MySQL optimizer team which optimizations they would like to see in the future. Sunday, 1.15 pm, Room Golden Gate 8: Query Performance Comparison of MySQL 5.5 and MySQL 5.6—Øystein Grøvlen, Oracle MySQL Release 5.6 contains several improvements in the query optimizer that create improved performance for complex queries. This presentation looks at how MySQL 5.6 improves the performance of many of the queries in the DBT-3 benchmark. Based on the observed improvements, the presentation discusses what makes the specific queries perform better in Release 5.6. It describes the relevant new optimization techniques and gives examples of the types of queries that will benefit from these techniques. Sunday, 4.15 pm, Room Golden Gate 4: Powerful EXPLAIN in MySQL 5.6—Evgeny Potemkin, Oracle The EXPLAIN command of MySQL has long been a very useful tool for understanding how MySQL will execute a query. Release 5.6 of the MySQL database offers several new additions that give more-detailed information about the query plan and make it easier to understand at the same time. This presentation gives an overview of new EXPLAIN features: structured EXPLAIN in JSON format, EXPLAIN for INSERT/UPDATE/DELETE, and optimizer tracing. Examples in the session give insights into how you can take advantage of the new features. They show how these features supplement and relate to each other and to classical EXPLAIN and how and why the MySQL server chooses a particular query plan. You can check out the full program here as well as in the September edition of the MySQL newsletter. Not registered yet? You can still save US$ 300 over the on-site fee – Register Now!

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• I have a problem with a AE1200 Cisco/Linksys Wireless-N USB adapter having stopped working after I ran the update manager in Ubuntu 12.04

  - by user69670

  Here is the problem, I use a Cisco/Linksys AE1200 wireless network adapter to connect my desktop to a public wifi internet connection. I use ndiswrapper to use the windows driver and it had been working fine for me untill I ran the update manager overnight a few days ago. When I woke up it was asking for the normal computer restart to implement the changes but after rebooting the computer, the wireless adapter did not work, the status light on the adapter did not light up even though ubuntu recognizes it is there and according to ndiswrapper the drivers are loaded and the hardware is present. the grep command is being a bitch for some unknown reason today so this will be long sorry Output from "lspci": 00:00.0 Host bridge: Advanced Micro Devices [AMD] nee ATI Radeon Xpress 200 Host Bridge (rev 01) 00:01.0 PCI bridge: Advanced Micro Devices [AMD] nee ATI RS480 PCI Bridge 00:12.0 SATA controller: Advanced Micro Devices [AMD] nee ATI SB600 Non-Raid-5 SATA 00:13.0 USB controller: Advanced Micro Devices [AMD] nee ATI SB600 USB (OHCI0) 00:13.1 USB controller: Advanced Micro Devices [AMD] nee ATI SB600 USB (OHCI1) 00:13.2 USB controller: Advanced Micro Devices [AMD] nee ATI SB600 USB (OHCI2) 00:13.3 USB controller: Advanced Micro Devices [AMD] nee ATI SB600 USB (OHCI3) 00:13.4 USB controller: Advanced Micro Devices [AMD] nee ATI SB600 USB (OHCI4) 00:13.5 USB controller: Advanced Micro Devices [AMD] nee ATI SB600 USB Controller (EHCI) 00:14.0 SMBus: Advanced Micro Devices [AMD] nee ATI SBx00 SMBus Controller (rev 13) 00:14.1 IDE interface: Advanced Micro Devices [AMD] nee ATI SB600 IDE 00:14.3 ISA bridge: Advanced Micro Devices [AMD] nee ATI SB600 PCI to LPC Bridge 00:14.4 PCI bridge: Advanced Micro Devices [AMD] nee ATI SBx00 PCI to PCI Bridge 01:05.0 VGA compatible controller: Advanced Micro Devices [AMD] nee ATI RC410 [Radeon Xpress 200] 02:02.0 Communication controller: Conexant Systems, Inc. HSF 56k Data/Fax Modem 02:03.0 Multimedia audio controller: Creative Labs CA0106 Soundblaster 02:05.0 Ethernet controller: Realtek Semiconductor Co., Ltd. RTL-8139/8139C/8139C+ (rev 10) Output from "lsusb": Bus 001 Device 001: ID 1d6b:0002 Linux Foundation 2.0 root hub Bus 002 Device 001: ID 1d6b:0001 Linux Foundation 1.1 root hub Bus 003 Device 001: ID 1d6b:0001 Linux Foundation 1.1 root hub Bus 004 Device 001: ID 1d6b:0001 Linux Foundation 1.1 root hub Bus 005 Device 001: ID 1d6b:0001 Linux Foundation 1.1 root hub Bus 006 Device 001: ID 1d6b:0001 Linux Foundation 1.1 root hub Bus 001 Device 009: ID 13b1:0039 Linksys AE1200 802.11bgn Wireless Adapter [Broadcom BCM43235] Bus 003 Device 002: ID 045e:0053 Microsoft Corp. Optical Mouse Bus 004 Device 002: ID 1043:8006 iCreate Technologies Corp. Flash Disk 32-256 MB Output from "ifconfig": eth0 Link encap:Ethernet HWaddr 00:19:21:b6:af:7c UP BROADCAST MULTICAST MTU:1500 Metric:1 RX packets:0 errors:0 dropped:0 overruns:0 frame:0 TX packets:0 errors:0 dropped:0 overruns:0 carrier:0 collisions:0 txqueuelen:1000 RX bytes:0 (0.0 B) TX bytes:0 (0.0 B) Interrupt:20 Base address:0xb400 lo Link encap:Local Loopback inet addr:127.0.0.1 Mask:255.0.0.0 inet6 addr: ::1/128 Scope:Host UP LOOPBACK RUNNING MTU:16436 Metric:1 RX packets:13232 errors:0 dropped:0 overruns:0 frame:0 TX packets:13232 errors:0 dropped:0 overruns:0 carrier:0 collisions:0 txqueuelen:0 RX bytes:1084624 (1.0 MB) TX bytes:1084624 (1.0 MB) Output from "iwconfig": lo no wireless extensions. eth0 no wireless extensions. Output from "lsmod": Module Size Used by nls_iso8859_1 12617 1 nls_cp437 12751 1 vfat 17308 1 fat 55605 1 vfat uas 17828 0 usb_storage 39646 1 nls_utf8 12493 1 udf 84366 1 crc_itu_t 12627 1 udf snd_ca0106 39279 2 snd_ac97_codec 106082 1 snd_ca0106 ac97_bus 12642 1 snd_ac97_codec snd_pcm 80845 2 snd_ca0106,snd_ac97_codec rfcomm 38139 0 snd_seq_midi 13132 0 snd_rawmidi 25424 2 snd_ca0106,snd_seq_midi bnep 17830 2 parport_pc 32114 0 bluetooth 158438 10 rfcomm,bnep ppdev 12849 0 snd_seq_midi_event 14475 1 snd_seq_midi snd_seq 51567 2 snd_seq_midi,snd_seq_midi_event snd_timer 28931 2 snd_pcm,snd_seq snd_seq_device 14172 3 snd_seq_midi,snd_rawmidi,snd_seq snd 62064 11 snd_ca0106, snd_ac97_codec,snd_pcm,snd_rawj9fe snd_ca0106,snd_ac97_codec,snd_pcm,snd_rawmidi,snd_seq,snd_timer,snd_seq_device soundcore 14635 1 snd snd_page_alloc 14108 2 snd_ca0106,snd_pcm sp5100_tco 13495 0 i2c_piix4 13093 0 radeon 733693 3 ttm 65344 1 radeon drm_kms_helper 45466 1 radeon drm 197692 5 radeon,ttm,drm_kms_helper i2c_algo_bit 13199 1 radeon mac_hid 13077 0 shpchp 32325 0 ati_agp 13242 0 lp 17455 0 parport 40930 3 parport_pc,ppdev,lp usbhid 41906 0 hid 77367 1 usbhid 8139too 23283 0 8139cp 26759 0 pata_atiixp 12999 1 Output from "sudo lshw -C network": *-network description: Ethernet interface product: RTL-8139/8139C/8139C+ vendor: Realtek Semiconductor Co., Ltd. physical id: 5 bus info: pci@0000:02:05.0 logical name: eth0 version: 10 serial: 00:19:21:b6:af:7c size: 10Mbit/s capacity: 100Mbit/s width: 32 bits clock: 33MHz capabilities: pm bus_master cap_list ethernet physical tp mii 10bt 10bt-fd 100bt 10 0bt-fd autonegotiation configuration: autonegotiation=on broadcast=yes driver=8139too driverversion=0.9.28 duplex=half latency=64 link=no maxlatency=64 mingnt=32 multicast=yes port=MII speed=10Mbit/s resources: irq:20 ioport:b400(size=256) memory:ff5fdc00-ff5fdcff Output from "iwlist scan": lo Interface doesn't support scanning. eth0 Interface doesn't support scanning. Output from "lsb_release -d": Ubuntu 12.04 LTS Output from "uname -mr": 3.2.0-24-generic-pae i686 Output from "sudo /etc/init.d/networking restart": * Running /etc/init.d/networking restart is deprecated because it may not enable again some interfaces * Reconfiguring network interfaces... [ OK ]

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• College Courses through distance learning

  - by Matt

  I realize this isn't really a programming question, but didn't really know where to post this in the stackexchange and because I am a computer science major i thought id ask here. This is pretty unique to the programmer community since my degree is about 95% programming. I have 1 semester left, but i work full time. I would like to finish up in December, but to make things easier i like to take online classes whenever I can. So, my question is does anyone know of any colleges that offer distance learning courses for computer science? I have been searching around and found a few potential classes, but not sure yet. I would like to gather some classes and see what i can get approval for. Class I need: Only need one C SC 437 Geometric Algorithms C SC 445 Algorithms C SC 473 Automata Only need one C SC 452 Operating Systems C SC 453 Compilers/Systems Software While i only need of each of the above courses i still need to take two more electives. These also have to be upper 400 level classes. So i can take multiple in each category. Some other classes I can take are: CSC 447 - Green Computing CSC 425 - Computer Networking CSC 460 - Database Design CSC 466 - Computer Security I hoping to take one or two of these courses over the summer. If not, then online over the regular semester would be ok too. Any help in helping find these classes would be awesome. Maybe you went to a college that offered distance learning. Some of these classes may be considered to be graduate courses too. Descriptions are listed below if you need. Thanks! Descriptions Computer Security This is an introductory course covering the fundamentals of computer security. In particular, the course will cover basic concepts of computer security such as threat models and security policies, and will show how these concepts apply to specific areas such as communication security, software security, operating systems security, network security, web security, and hardware-based security. Computer Networking Theory and practice of computer networks, emphasizing the principles underlying the design of network software and the role of the communications system in distributed computing. Topics include routing, flow and congestion control, end-to-end protocols, and multicast. Database Design Functions of a database system. Data modeling and logical database design. Query languages and query optimization. Efficient data storage and access. Database access through standalone and web applications. Green Computing This course covers fundamental principles of energy management faced by designers of hardware, operating systems, and data centers. We will explore basic energy management option in individual components such as CPUs, network interfaces, hard drives, memory. We will further present the energy management policies at the operating system level that consider performance vs. energy saving tradeoffs. Finally we will consider large scale data centers where energy management is done at multiple layers from individual components in the system to shutting down entries subset of machines. We will also discuss energy generation and delivery and well as cooling issues in large data centers. Compilers/Systems Software Basic concepts of compilation and related systems software. Topics include lexical analysis, parsing, semantic analysis, code generation; assemblers, loaders, linkers; debuggers. Operating Systems Concepts of modern operating systems; concurrent processes; process synchronization and communication; resource allocation; kernels; deadlock; memory management; file systems. Algorithms Introduction to the design and analysis of algorithms: basic analysis techniques (asymptotics, sums, recurrences); basic design techniques (divide and conquer, dynamic programming, greedy, amortization); acquiring an algorithm repertoire (sorting, median finding, strong components, spanning trees, shortest paths, maximum flow, string matching); and handling intractability (approximation algorithms, branch and bound). Automata Introduction to models of computation (finite automata, pushdown automata, Turing machines), representations of languages (regular expressions, context-free grammars), and the basic hierarchy of languages (regular, context-free, decidable, and undecidable languages). Geometric Algorithms The study of algorithms for geometric objects, using a computational geometry approach, with an emphasis on applications for graphics, VLSI, GIS, robotics, and sensor networks. Topics may include the representation and overlaying of maps, finding nearest neighbors, solving linear programming problems, and searching geometric databases.

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• Is this proper OO design for C++?

  - by user121917

  I recently took a software processes course and this is my first time attempting OO design on my own. I am trying to follow OO design principles and C++ conventions. I attempted and gave up on MVC for this application, but I am trying to "decouple" my classes such that they can be easily unit-tested and so that I can easily change the GUI library used and/or the target OS. At this time, I have finished designing classes but have not yet started implementing methods. The function of the software is to log all packets sent and received, and display them on the screen (like WireShark, but for one local process only). The software accomplishes this by hooking the send() and recv() functions in winsock32.dll, or some other pair of analogous functions depending on what the intended Target is. The hooks add packets to SendPacketList/RecvPacketList. The GuiLogic class starts a thread which checks for new packets. When new packets are found, it utilizes the PacketFilter class to determine the formatting for the new packet, and then sends it to MainWindow, a native win32 window (with intent to later port to Qt).1 Full size image of UML class diagram Here are my classes in skeleton/header form (this is my actual code): class PacketModel { protected: std::vector<byte> data; int id; public: PacketModel(); PacketModel(byte* data, unsigned int size); PacketModel(int id, byte* data, unsigned int size); int GetLen(); bool IsValid(); //len >= sizeof(opcode_t) opcode_t GetOpcode(); byte* GetData(); //returns &(data[0]) bool GetData(byte* outdata, int maxlen); void SetData(byte* pdata, int len); int GetId(); void SetId(int id); bool ParseData(char* instr); bool StringRepr(char* outstr); byte& operator[] (const int index); }; class SendPacket : public PacketModel { protected: byte* returnAddy; public: byte* GetReturnAddy(); void SetReturnAddy(byte* addy); }; class RecvPacket : public PacketModel { protected: byte* callAddy; public: byte* GetCallAddy(); void SetCallAddy(byte* addy); }; //problem: packets may be added to list at any time by any number of threads //solution: critical section associated with each packet list class Synch { public: void Enter(); void Leave(); }; template<class PacketType> class PacketList { private: static const int MAX_STORED_PACKETS = 1000; public: static const int DEFAULT_SHOWN_PACKETS = 100; private: vector<PacketType> list; Synch synch; //wrapper for critical section public: void AddPacket(PacketType* packet); PacketType* GetPacket(int id); int TotalPackets(); }; class SendPacketList : PacketList<SendPacket> { }; class RecvPacketList : PacketList<RecvPacket> { }; class Target //one socket { bool Send(SendPacket* packet); bool Inject(RecvPacket* packet); bool InitSendHook(SendPacketList* sendList); bool InitRecvHook(RecvPacketList* recvList); }; class FilterModel { private: opcode_t opcode; int colorID; bool bFilter; char name[41]; }; class FilterFile { private: FilterModel filter; public: void Save(); void Load(); FilterModel* GetFilter(opcode_t opcode); }; class PacketFilter { private: FilterFile filters; public: bool IsFiltered(opcode_t opcode); bool GetName(opcode_t opcode, char* namestr); //return false if name does not exist COLORREF GetColor(opcode_t opcode); //return default color if no custom color }; class GuiLogic { private: SendPacketList sendList; RecvPacketList recvList; PacketFilter packetFilter; void GetPacketRepr(PacketModel* packet); void ReadNew(); void AddToWindow(); public: void Refresh(); //called from thread void GetPacketInfo(int id); //called from MainWindow }; I'm looking for a general review of my OO design, use of UML, and use of C++ features. I especially just want to know if I'm doing anything considerably wrong. From what I've read, design review is on-topic for this site (and off-topic for the Code Review site). Any sort of feedback is greatly appreciated. Thanks for reading this.

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• Can't connect to Wired Network after installing 12.04

  - by ezz9

  I have installed 12.04 into a used HP Compaq DC 7100 CMT. Plugged in LAN cable into the computer and it says network disconnected. from what I understand on the things I've read, it's not getting the address right, maybe it's just a guess (hardware address 00:16:35:78:47:bb). I know the LAN cable is working I can get on the net with the old computer. (Old computer is using device MAC address; Auto eth0 00:11:11:E6:4F:FE). I have put this address into the newer HP and it shows last used (minutes ago) but no Internet sever not found. I tried the hardware address it says never. I feel and think this should be easy to fix. But I just don't know. Here is the info everyone has asked for, but they never say what I should do. sudo lshw -C network; rfkill list; cat /etc/network/interfaces; cat /etc/lsb-release; lspci -nn; lsusb; uname -a; ifconfig; route -n *-network description: ethernet interface product; NetXtreme BCM5751 Gigabit Ethernet PCI Express vendor: Broadcom Corporation physical id: 0 bus info: pci@0000:40:00.0 logical name: eth0 verson: 01 serial: 00:16:35:78:47:bb size: 100Mbit/s capacity: 1Gbit/s width: 64 bits clock: 33MHz capabilities: pm vpd msi pciexpress bus_master cap_list ethernet physical tp 10bt 10bt-fd 100bt 100bt-fd 1000bt 1000-fd configuration: autonegotiation=on broadcast=yes driver=tg3 driverversion= 3.121 duplex=full firmware=5751-v3.29a latency=0 link=yes multicast=yes port=twi sted pair speed=100Mbit/s resources: irq:17 memory:f0400000-f040ffff auto lo iface lo inet loopback DISTRIB_ID=Ubuntu DISTRIB_RELEASE=12.04 DISTRIB_CODENAME=precise DISTRIB_DESCRIPTION= "Ubuntu 12.04 LTS" 00:00.0 host bridge [0600]: Inter Corporation 82915G/P/GV/GL/PL/910GL Memory Con troller Hub [8086:2580] (rev 04) 00.02.0 VGA compatible controller [0300]: Intel Corporation 82915G/GV/910GL inte grated Graphics Controller [8086:2582] (rev 04) 00:1c.0 PCI bridge [0604]: Intel Corporation 82801FB/FBM/FR/FW/FRW (ICH6 Family) PCI Express Port 1 [8086:2660] (rev 03) 00:1c.1 PCI bridge [0604]: Intel Corporation 82801FB/FBM/FR/FW/FRW (ICH6 Family) PCI Express Port 2 [8086:2662] (rev 03) 00:1d.0 USB controller [0c03]: Intel Corporation 82801FB/FBM/FR/FW/FRW (ICH6 Fam ily) USB UHCI #1 [8086:2658] (rev 03) 00:1d.1 USB controller [0c03]: Intel Corporation 82801FB/FBM/FR/FW/FRW (ICH6 Fam ily) USB UHCI #2 [8086:2659] (rev 03) 00:1d.2 USB controller [0c03]: Intel Corporation 82801FB/FBM/FR/FW/FRW (ICH6 Fam ily) USB UHCI #3 [8086:265a] (rev 03) 00:1d.3 USB controller [0c03]: Intel Corporation 82801FB/FBM/FR/FW/FRW (ICH6 Fam ily) USB UHCI #4 [8086:265b] (rev 03) 00:1d.7 USB controller [0c03]: Intel Corporation 82801FB/FBM/FR/FW/FRW (ICH6 Fam ily) USB2 UHCI Controller [8086:265c] (rev 03) 00:1e.0 PCI bridge [0604]: Intel Corporation 82801 PCI bridge [8086:244e] (rev d 3) 00:1e.2 Multimedia audio controller [0401]: Intel Corporation 82801FB/FBM/FR/FW/ FRW (ICH6 Family) AC'97 Autio conrtroller [8086:266e] (rev 03) 00:1f.0 ISA bridge [0601]: Intel Corporation 82801FB/RF (ICH6/ICH6R) LPC Interfa ce Bridge [8086:2640] (rev 03) 00:1f.1 IDE interface [0101]: Intel Corporation 82801FB/FBM/FR/FW/FRW (ICH6 fami ly) IDE Controller [8086:266f] (rev 03) 00:1f.2 IDE interface [0101]: Intel Corporation 82801FB/FW (ICH6/ICH6W) SATA Con troller [8086:2651] (rev 03) 40:00.0 Ethernet controller [0200]: Broadcom Corporation NetXreme BCM5751 Gigab it Ethernet PCI Express [14e4:1677} (rev 01) Bus 001 Device 001: ID 1d6b:0002 Linux Foundation 2.0 root hub Bus 002 Device 001: ID 1d6b:0001 Linux Foundation 1.1 root hub Bus 003 Device 001: ID 1d6b:0001 Linux Foundation 1.1 root hub Bus 004 Device 001: ID 1d6b:0001 Linux Foundation 1.1 root hub Bus 005 Device 001: ID 1d6b:0001 Linux Foundation 1.1 root hub Linux bob-desktop 3.2.0-23-generic-pae #36-Ubuntu SMP Tpr 10 22:19:09 UTC 20 12 i686 i686 i386 GNU/Linux eth0 Link encap:Ethernet HWaddr 00:16:35:78:47:bb inet6 addr: fe80::216:35ff:47bb/64 Scope:link UP BROADCAST RUNNING MULTICAST MTU:1500 Metric:1 RX packets:2517 errors:0 dropped:0 overruns:0 frame:0 TX packets:164 errors:0 dropped:0 overruns:0 carrier:0 collisions:0 txqueuelen:1000 RX bytes:164508 (164.5 KB) TX bytes: 40884 (40.0 KB) Interrupt:17 lo Link encap:Local loopback inet addr:127.0.0.1 Mask:255.0.0.0 inet6 addr: ::1/128 Scope:Host UP LOOPBACK RUNNING MTU:16436 Metric:1 RX packets:3290 errors:0 dropped:0 overruns:0 frame:0 TX packets:3290 errors:0 dropped:0 overruns:0 carrier:0 collisions:0 txqueuelen: 0 RX bytes:267212 (267.2 KB) TX bytes: 267212 (267.2 KB) Kernel Ip routing table Destination Gateway Genmask Flags Metric Ref Use Iface ran this sudo dhclient eth0 no reply ran this ip addr 1: lo <LOOPBACK,UP,LOWER_UP> mtu 16436 qdisc noqueue state UNKNOWN link/loopback 00:00:00:00:00:00 brd 00:00:00:00:00:00 inet 127.0.0.1/8 scope host lo inet6 ::1/128 scope host valid_lft forever preferred_lft forever 2: eth0: <BROADCAST ,MULTICAST ,UP ,LOWER_UP> mtu 1500 qdisc mq state UP qlen 1000 link/enter 00:16:35:78:47:BB brd ff:ff:ff:ff:ff:ff inet 169.254.7.172/16 brd 169.254.255.255 scope link eth0:avahi inet6 fe80::216:35ff:fe78:47bb/64 scope link valid_lft forever preferred_lft forever Help please.

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• Running a Mongo Replica Set on Azure VM Roles

  - by Elton Stoneman

  Originally posted on: http://geekswithblogs.net/EltonStoneman/archive/2013/10/15/running-a-mongo-replica-set-on-azure-vm-roles.aspxSetting up a MongoDB Replica Set with a bunch of Azure VMs is straightforward stuff. Here’s a step-by-step which gets you from 0 to fully-redundant 3-node document database in about 30 minutes (most of which will be spent waiting for VMs to fire up). First, create yourself 3 VM roles, which is the minimum number of nodes you need for high availability. You can use any OS that Mongo supports. This guide uses Windows but the only difference will be the mechanism for starting the Mongo service when the VM starts (Windows Service, daemon etc.) While the VMs are provisioning, download and install Mongo locally, so you can set up the replica set with the Mongo shell. We’ll create our replica set from scratch, doing one machine at a time (if you have a single node you want to upgrade to a replica set, it’s the same from step 3 onwards): 1. Setup Mongo Log into the first node, download mongo and unzip it to C:. Rename the folder to remove the version – so you have c:\MongoDB\bin etc. – and create a new folder for the logs, c:\MongoDB\logs. 2. Setup your data disk When you initialize a node in a replica set, Mongo pre-allocates a whole chunk of storage to use for data replication. It will use up to 5% of your data disk, so if you use a Windows VM image with a defsault 120Gb disk and host your data on C:, then Mongo will allocate 6Gb for replication. And that takes a while. Instead you can create yourself a new partition by shrinking down the C: drive in Computer Management, by say 10Gb, and then creating a new logical disk for your data from that spare 10Gb, which will be allocated as E:. Create a new folder, e:\data. 3. Start Mongo When that’s done, start a command line, point to the mongo binaries folder, install Mongo as a Windows Service, running in replica set mode, and start the service: cd c:\mongodb\bin mongod -logpath c:\mongodb\logs\mongod.log -dbpath e:\data -replSet TheReplicaSet –install net start mongodb 4. Open the ports Mongo uses port 27017 by default, so you need to allow access in the machine and in Azure. In the VM, open Windows Firewall and create a new inbound rule to allow access via port 27017. Then in the Azure Management Console for the VM role, under the Configure tab add a new rule, again to allow port 27017. 5. Initialise the replica set Start up your local mongo shell, connecting to your Azure VM, and initiate the replica set: c:\mongodb\bin\mongo sc-xyz-db1.cloudapp.net rs.initiate() This is the bit where the new node (at this point the only node) allocates its replication files, so if your data disk is large, this can take a long time (if you’re using the default C: drive with 120Gb, it may take so long that rs.initiate() never responds. If you’re sat waiting more than 20 minutes, start another instance of the mongo shell pointing to the same machine to check on it). Run rs.conf() and you should see one node configured. 6. Fix the host name for the primary – *don’t miss this one* For the first node in the replica set, Mongo on Windows doesn’t populate the full machine name. Run rs.conf() and the name of the primary is sc-xyz-db1, which isn’t accessible to the outside world. The replica set configuration needs the full DNS name of every node, so you need to manually rename it in your shell, which you can do like this: cfg = rs.conf() cfg.members[0].host = ‘sc-xyz-db1.cloudapp.net:27017’ rs.reconfig(cfg) When that returns, rs.conf() will have your full DNS name for the primary, and the other nodes will be able to connect. At this point you have a working database, so you can start adding documents, but there’s no replication yet. 7. Add more nodes For the next two VMs, follow steps 1 through to 4, which will give you a working Mongo database on each node, which you can add to the replica set from the shell with rs.add(), using the full DNS name of the new node and the port you’re using: rs.add(‘sc-xyz-db2.cloudapp.net:27017’) Run rs.status() and you’ll see your new node in STARTUP2 state, which means its initializing and replicating from the PRIMARY. Repeat for your third node: rs.add(‘sc-xyz-db3.cloudapp.net:27017’) When all nodes are finished initializing, you will have a PRIMARY and two SECONDARY nodes showing in rs.status(). Now you have high availability, so you can happily stop db1, and one of the other nodes will become the PRIMARY with no loss of data or service. Note – the process for AWS EC2 is exactly the same, but with one important difference. On the Azure Windows Server 2012 base image, the MongoDB release for 64-bit 2008R2+ works fine, but on the base 2012 AMI that release keeps failing with a UAC permission error. The standard 64-bit release is fine, but it lacks some optimizations that are in the 2008R2+ version.

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• Blog Buzz - Devoxx 2011

  - by Janice J. Heiss

  Some day I will make it to Devoxx – for now, I’m content to vicariously follow the blogs of attendees and pick up on what’s happening.  I’ve been doing more blog "fishing," looking for the best commentary on 2011 Devoxx. There’s plenty of food for thought – and the ideas are not half-baked.The bloggers are out in full, offering useful summaries and commentary on Devoxx goings-on.Constantin Partac, a Java developer and a member of Transylvania JUG, a community from Cluj-Napoca/Romania, offers an excellent summary of the Devoxx keynotes. Here’s a sample:“Oracle Opening Keynote and JDK 7, 8, and 9 Presentation•    Oracle is committed to Java and wants to provide support for it on any device.•    JSE 7 for Mac will be released next week.•    Oracle would like Java developers to be involved in JCP, to adopt a JSR and to attend local JUG meetings.•    JEE 7 will be released next year.•    JEE 7 is focused on cloud integration, some of the features are already implemented in glassfish 4 development branch.•    JSE 8 will be release in summer of 2013 due to “enterprise community request” as they can not keep the pace with an 18    month release cycle.•    The main features included in JSE8 are lambda support, project Jigsaw, new Date/Time API, project Coin++ and adding   support for sensors. JSE 9 probably will focus on some of these features:1.    self tuning JVM2.    improved native language integration3.    processing enhancement for big data4.    reification (adding runtime class type info for generic types)5.    unification of primitive and corresponding object classes6.    meta-object protocol in order to use type and methods define in other JVM languages7.    multi-tenancy8.    JVM resource management” Thanks Constantin! Ivan St. Ivanov, of SAP Labs Bulgaria, also commented on the keynotes with a different focus.  He summarizes Henrik Stahl’s look ahead to Java SE 8 and JavaFX 3.0; Cameron Purdy on Java EE and the cloud; celebrated Java Champion Josh Bloch on what’s good and bad about Java; Mark Reinhold’s quick look ahead to Java SE 9; and Brian Goetz on lambdas and default methods in Java SE 8. Here’s St. Ivanov’s account of Josh Bloch’s comments on the pluses of Java:“He started with the virtues of the platform. To name a few:    Tightly specified language primitives and evaluation order – int is always 32 bits and operations are executed always from left  to right, without compilers messing around    Dynamic linking – when you change a class, you need to recompile and rebuild just the jar that has it and not the whole application    Syntax  similarity with C/C++ – most existing developers at that time felt like at home    Object orientations – it was cool at that time as well as functional programming is today    It was statically typed language – helps in faster runtime, better IDE support, etc.    No operator overloading – well, I’m not sure why it is good. Scala has it for example and that’s why it is far better for defining DSLs. But I will not argue with Josh.”It’s worth checking out St. Ivanov’s summary of Bloch’s views on what’s not so great about Java as well. What's Coming in JAX-RS 2.0Marek Potociar, Principal Software Engineer at Oracle and currently specification lead of Java EE RESTful web services API (JAX-RS), blogged on his talk about what's coming in JAX-RS 2.0, scheduled for final release in mid-2012.  Here’s a taste:“Perhaps the most wanted addition to the JAX-RS is the Client API, that would complete the JAX-RS story, that is currently server-side only. In JAX-RS 2.0 we are adding a completely interface-based and fluent client API that blends nicely in with the existing fluent response builder pattern on the server-side. When we started with the client API, the first proposal contained around 30 classes. Thanks to the feedback from our Expert Group we managed to reduce the number of API classes to 14 (2 of them being exceptions)! The resulting is compact while at the same time we still managed to create an API that reflects the method invocation context flow (e.g. once you decide on the target URI and start setting headers on the request, your IDE will not try to offer you a URI setter in the code completion). This is a subtle but very important usability aspect of an API…” Obviously, Devoxx is a great Java conference, one that is hitting this year at a time when much is brewing in the platform and beginning to be anticipated.

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• Why enumerator structs are a really bad idea

  - by Simon Cooper

  If you've ever poked around the .NET class libraries in Reflector, I'm sure you would have noticed that the generic collection classes all have implementations of their IEnumerator as a struct rather than a class. As you will see, this design decision has some rather unfortunate side effects... As is generally known in the .NET world, mutable structs are a Very Bad Idea; and there are several other blogs around explaining this (Eric Lippert's blog post explains the problem quite well). In the BCL, the generic collection enumerators are all mutable structs, as they need to keep track of where they are in the collection. This bit me quite hard when I was coding a wrapper around a LinkedList<int>.Enumerator. It boils down to this code: sealed class EnumeratorWrapper : IEnumerator<int> { private readonly LinkedList<int>.Enumerator m_Enumerator; public EnumeratorWrapper(LinkedList<int> linkedList) { m_Enumerator = linkedList.GetEnumerator(); } public int Current { get { return m_Enumerator.Current; } } object System.Collections.IEnumerator.Current { get { return Current; } } public bool MoveNext() { return m_Enumerator.MoveNext(); } public void Reset() { ((System.Collections.IEnumerator)m_Enumerator).Reset(); } public void Dispose() { m_Enumerator.Dispose(); } } The key line here is the MoveNext method. When I initially coded this, I thought that the call to m_Enumerator.MoveNext() would alter the enumerator state in the m_Enumerator class variable and so the enumeration would proceed in an orderly fashion through the collection. However, when I ran this code it went into an infinite loop - the m_Enumerator.MoveNext() call wasn't actually changing the state in the m_Enumerator variable at all, and my code was looping forever on the first collection element. It was only after disassembling that method that I found out what was going on The MoveNext method above results in the following IL: .method public hidebysig newslot virtual final instance bool MoveNext() cil managed { .maxstack 1 .locals init ( [0] bool CS$1$0000, [1] valuetype [System]System.Collections.Generic.LinkedList`1/Enumerator CS$0$0001) L_0000: nop L_0001: ldarg.0 L_0002: ldfld valuetype [System]System.Collections.Generic.LinkedList`1/Enumerator EnumeratorWrapper::m_Enumerator L_0007: stloc.1 L_0008: ldloca.s CS$0$0001 L_000a: call instance bool [System]System.Collections.Generic.LinkedList`1/Enumerator::MoveNext() L_000f: stloc.0 L_0010: br.s L_0012 L_0012: ldloc.0 L_0013: ret } Here, the important line is 0002 - m_Enumerator is accessed using the ldfld operator, which does the following: Finds the value of a field in the object whose reference is currently on the evaluation stack. So, what the MoveNext method is doing is the following: public bool MoveNext() { LinkedList<int>.Enumerator CS$0$0001 = this.m_Enumerator; bool CS$1$0000 = CS$0$0001.MoveNext(); return CS$1$0000; } The enumerator instance being modified by the call to MoveNext is the one stored in the CS$0$0001 variable on the stack, and not the one in the EnumeratorWrapper class instance. Hence why the state of m_Enumerator wasn't getting updated. Hmm, ok. Well, why is it doing this? If you have a read of Eric Lippert's blog post about this issue, you'll notice he quotes a few sections of the C# spec. In particular, 7.5.4: ...if the field is readonly and the reference occurs outside an instance constructor of the class in which the field is declared, then the result is a value, namely the value of the field I in the object referenced by E. And my m_Enumerator field is readonly! Indeed, if I remove the readonly from the class variable then the problem goes away, and the code works as expected. The IL confirms this: .method public hidebysig newslot virtual final instance bool MoveNext() cil managed { .maxstack 1 .locals init ( [0] bool CS$1$0000) L_0000: nop L_0001: ldarg.0 L_0002: ldflda valuetype [System]System.Collections.Generic.LinkedList`1/Enumerator EnumeratorWrapper::m_Enumerator L_0007: call instance bool [System]System.Collections.Generic.LinkedList`1/Enumerator::MoveNext() L_000c: stloc.0 L_000d: br.s L_000f L_000f: ldloc.0 L_0010: ret } Notice on line 0002, instead of the ldfld we had before, we've got a ldflda, which does this: Finds the address of a field in the object whose reference is currently on the evaluation stack. Instead of loading the value, we're loading the address of the m_Enumerator field. So now the call to MoveNext modifies the enumerator stored in the class rather than on the stack, and everything works as expected. Previously, I had thought enumerator structs were an odd but interesting feature of the BCL that I had used in the past to do linked list slices. However, effects like this only underline how dangerous mutable structs are, and I'm at a loss to explain why the enumerators were implemented as structs in the first place. (interestingly, the SortedList<TKey, TValue> enumerator is a struct but is private, which makes it even more odd - the only way it can be accessed is as a boxed IEnumerator!). I would love to hear people's theories as to why the enumerators are implemented in such a fashion. And bonus points if you can explain why LinkedList<int>.Enumerator.Reset is an explicit implementation but Dispose is implicit... Note to self: never ever ever code a mutable struct.

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• Adding a Network Loopback Adapter to Windows 8

  - by Greg Low

  I have to say that I continue to be frustrated with finding out how to do things in Windows 8. Here's another one and it's recorded so it might help someone else. I've also documented what I tried so that if anyone from the product group ever reads this, they'll understand how I searched for it and might try to make it easier.I wanted to add a network loopback adapter, to have a fixed IP address to work with when using an "internal" network with Hyper-V. (The fact that I even need to do this is also painful. I don't know why Hyper-V can't make it easy to work with host system folders, etc. as easily as I can with VirtualPC, VirtualBox, etc. but that's a topic for another day).In the end, what I needed was a known IP address on the same network that my guest OS was using, via the internal network (which allows connectivity from the host OS to/from guest OS's).I started by looking in the network adapters areas but there is no "add" functionality there. Realising that this was likely to be another unexpected challenge, I resorted to searching for info on doing this. I found KB article 2777200 entitled "Installing the Microsoft Loopback Adapter in Windows 8 and Windows Server 2012". Aha, I thought that's what I'd need. It describes the symptom as "You are trying to install the Microsoft Loopback Adapter, but are unable to find it." and that certainly sounded like me. There's a certain irony in documenting that something's hard to find instead of making it easier to find. Anyway, you'd hope that in that article, they'd then provide a step by step example of how to do it, but what they supply is this: The Microsoft Loopback Adapter was renamed in Windows 8 and Windows Server 2012. The new name is "Microsoft KM-TEST Loopback Adapter". When using the Add Hardware Wizard to manually add a network adapter, choose Manufacturer "Microsoft" and choose network adapter "Microsoft KM-TEST Loopback Adapter".The trick with this of course is finding the "Add Hardware Wizard". In Control Panel -> Hardware and Sound, there are options to "Add a device" and for "Device Manager". I tried the "Add a device" wizard (seemed logical to me) but after that wizard tries it's best, it just tells you that there isn't any hardware that it thinks it needs to install. It offers a link for when you can't find what you're looking for, but that leads to a generic help page that tells you how to do things like turning on your printer.In Device Manager, I checked the options in the program menus, and nothing useful was present. I even tried right-clicking "Network adapters", hoping that would lead to an option to add one, also to no avail.So back to the search engine I went, to try to find out where the "Add Hardware Wizard" is. Turns out I was in the right place in Device Manager, but I needed to right-click the computer's name, and choose "Add Legacy Hardware". No doubt that hasn't changed location lately but it's a while since I needed to add one so I'd forgotten. Regardless, I'm left wondering why it couldn't be in the menu as well.Anyway, for a step by step list, you need to do the following:1. From Control Panel, select "Device Manager" under the "Devices and Printers" section of the "Hardware and Sound" tab.2. Right-click the name of the computer at the top of the tree, and choose "Add Legacy Hardware".3. In the "Welcome to the Add Hardware Wizard" window, click Next.4. In the "The wizard can help you install other hardware" window, choose "Install the hardware that I manually select from a list" option and click Next.5. In the "The wizard did not find any new hardware on your computer" window, click Next.6. In the "From the list below, select the type of hardware you are installing" window, select "Network Adapters" from the list, and click Next.7. In the "Select Network Adapter" window, from the Manufacturer list, choose Microsoft, then in the Network Adapter window, choose "Microsoft KM-TEST Loopback Adapter", then click Next.8. In the "The wizard is ready to install your hardware" window, click Next.9. In the "Completing the Add Hardware Wizard" window, click Finish.Then you need to continue to set the IP address, etc.10. Back in Control Panel, select the "Network and Internet" tab, click "View Network Status and Tasks".11. In the "View your basic network information and set up connections" window, click "Change adapter settings".12. Right-click the new adapter that has been added (find it in the list by checking the device name of "Microsoft KM-TEST Loopback Adapter"), and click Properties.   

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• Create a Smoother Period Close

  - by Get Proactive Customer Adoption Team

  Untitled Document Do You Use Oracle E-Business Suite Products Involved in Accounting Period Closes? We understand that closing the periods in your system at the end of an accounting period enables your company to make the right business decisions. We also know this requires prior preparation, good procedures, and quality data. To help you meet that need, Oracle E-Business Suite’s proactive support team developed the Period Close Advisor to help your organization conduct a smooth period close for its Oracle E-Business Suite 12 products. The Period Close Advisor is composed of logical steps you can follow, aligned by the business requirement flow. It will help with an orderly close of the product sub-ledgers before posting to the General Ledger. It combines recommendations and industry best practices with tips from subject matter experts for troubleshooting. You will find patches needed and references to assist you during each phase. Get to know the E-Business Suite Period Close Advisor The Period Close Advisor does more than help the users of Oracle E-Business Suite products close their period. You can use it before and throughout the period to stay on track. Proactively it assists you as you set up your company’s period close process. During the period, it helps evaluate your system’s readiness for initiating the period close procedures and prepare the system for a smooth period close experience. The Period Close Advisor gets you to answers when you have questions and gives you the latest news from us on Oracle E-Business Suite’s period close. The Period Close Advisor is the right place to start. How to Use the E-Business Suite Period Close The Period Close Advisor graphically guides you through your period close. The tabs show you the products (also called applications or sub-ledgers) covered, and the product order required for the processing to handle any dependencies between the products. Users of all the products it covers can benefit from the information it contains. Structure of the Period Close Advisor Clicking on a tab gives you the details for that particular step in the process. This includes an overview, showing how the products fit into the overall period close process, and step-by-step information on each phase needed to complete the period close for the tab. You will also find multimedia training and related resources you can access if you need more information. Once you click on any of the phases, you see guidance for that phase. This can include: Tips from the subject-matter experts—here are examples from a Cash Management specialist: “For organizations with high transaction volumes bank statements should be loaded and reconciled on a daily basis.” “The automatic reconciliation process can be set up to create miscellaneous transactions automatically.” References to useful Knowledge Base documents: Information Centers for the products and features FAQs on functionality Known Issues and patches with both the errors and their solutions How-to documents that explain in detail how to use a feature or complete a process White papers that give overview of a feature, list setup required to use the feature, etc. Links to diagnosticsthat help debug issues you may find in a process Additional information and alerts about a process or reports that can help you prevent issues from surfacing This excerpt from the “Process Transaction” phase for the Receivables product lists documents you’ll find helpful. How to Get Started with the Period Close Advisor The Period Close Advisor is a great resource that can be used both as a proactive tool (while setting up your period end procedures) and as the first document to refer to when you encounter an issue during the period close procedures! As mentioned earlier, the order of the product tabs in the Period Close Advisor gives you the recommended order of closing. The first thing to do is to ensure that you are following the prescribed order for closing the period, if you are using more than one sub-ledger. Next, review the information shared in the Evaluate and Prepare and Process Transactions phases. Make sure that you are following the recommended best practices; you have applied the recommended patches, etc. The Reconcile phase gives you the recommended steps to follow for reconciling a sub-ledger with the General Ledger. Ensure that your reconciliation procedure aligns with those steps. At any stage during the period close processing, if you encounter an issue, you can revisit the Period Close Advisor. Choose the product you have an issue with and then select the phase you are in. You will be able to review information that can help you find a solution to the issue you are facing. Stay Informed Oracle updates the Period Close Advisor as we learn of new issues and information. Bookmark the Oracle E-Business Suite Period Close Advisor [ID 335.1] and keep coming back to it for the latest information on period close

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• Using Delegates in C# (Part 1)

  - by rajbk

  This post provides a very basic introduction of delegates in C#. Part 2 of this post can be read here. A delegate is a class that is derived from System.Delegate.  It contains a list of one or more methods called an invocation list. When a delegate instance is “invoked” with the arguments as defined in the signature of the delegate, each of the methods in the invocation list gets invoked with the arguments. The code below shows example with static and instance methods respectively: Static Methods 1: using System; 2: using System.Linq; 3: using System.Collections.Generic; 4: 5: public delegate void SayName(string name); 6: 7: public class Program 8: { 9: [STAThread] 10: static void Main(string[] args) 11: { 12: SayName englishDelegate = new SayName(SayNameInEnglish); 13: SayName frenchDelegate = new SayName(SayNameInFrench); 14: SayName combinedDelegate =(SayName)Delegate.Combine(englishDelegate, frenchDelegate); 15: 16: combinedDelegate.Invoke("Tom"); 17: Console.ReadLine(); 18: } 19: 20: static void SayNameInFrench(string name) { 21: Console.WriteLine("J'ai m'appelle " + name); 22: } 23: 24: static void SayNameInEnglish(string name) { 25: Console.WriteLine("My name is " + name); 26: } 27: } We have declared a delegate of type SayName with return type of void and taking an input parameter of name of type string. On line 12, we create a new instance of this delegate which refers to a static method - SayNameInEnglish.  SayNameInEnglish has the same return type and parameter list as the delegate declaration.  Once a delegate is instantiated, the instance will always refer to the same target. Delegates are immutable. On line 13, we create a new instance of the delegate but point to a different static method. As you may recall, a delegate instance encapsulates an invocation list. You create an invocation list by combining delegates using the Delegate.Combine method (there is an easier syntax as you will see later). When two non null delegate instances are combined, their invocation lists get combined to form a new invocation list. This is done in line 14.  On line 16, we invoke the delegate with the Invoke method and pass in the required string parameter. Since the delegate has an invocation list with two entries, each of the method in the invocation list is invoked. If an unhandled exception occurs during the invocation of one of these methods, the exception gets bubbled up to the line where the invocation was made (line 16). If a delegate is null and you try to invoke it, you will get a System.NullReferenceException. We see the following output when the method is run: My name is TomJ'ai m'apelle Tom Instance Methods The code below outputs the same results as before. The only difference here is we are creating delegates that point to a target object (an instance of Translator) and instance methods which have the same signature as the delegate type. The target object can never be null. We also use the short cut syntax += to combine the delegates instead of Delegate.Combine. 1: public delegate void SayName(string name); 2: 3: public class Program 4: { 5: [STAThread] 6: static void Main(string[] args) 7: { 8: Translator translator = new Translator(); 9: SayName combinedDelegate = new SayName(translator.SayNameInEnglish); 10: combinedDelegate += new SayName(translator.SayNameInFrench); 11:  12: combinedDelegate.Invoke("Tom"); 13: Console.ReadLine(); 14: } 15: } 16: 17: public class Translator { 18: public void SayNameInFrench(string name) { 19: Console.WriteLine("J'ai m'appelle " + name); 20: } 21: 22: public void SayNameInEnglish(string name) { 23: Console.WriteLine("My name is " + name); 24: } 25: } A delegate can be removed from a combination of delegates by using the –= operator. Removing a delegate from an empty list or removing a delegate that does not exist in a non empty list will not result in an exception. Delegates are invoked synchronously using the Invoke method. We can also invoke them asynchronously using the BeginInvoke and EndInvoke methods which are compiler generated.

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• How to Achieve Real-Time Data Protection and Availabilty....For Real

  - by JoeMeeks

  There is a class of business and mission critical applications where downtime or data loss have substantial negative impact on revenue, customer service, reputation, cost, etc. Because the Oracle Database is used extensively to provide reliable performance and availability for this class of application, it also provides an integrated set of capabilities for real-time data protection and availability. Active Data Guard, depicted in the figure below, is the cornerstone for accomplishing these objectives because it provides the absolute best real-time data protection and availability for the Oracle Database. This is a bold statement, but it is supported by the facts. It isn’t so much that alternative solutions are bad, it’s just that their architectures prevent them from achieving the same levels of data protection, availability, simplicity, and asset utilization provided by Active Data Guard. Let’s explore further. Backups are the most popular method used to protect data and are an essential best practice for every database. Not surprisingly, Oracle Recovery Manager (RMAN) is one of the most commonly used features of the Oracle Database. But comparing Active Data Guard to backups is like comparing apples to motorcycles. Active Data Guard uses a hot (open read-only), synchronized copy of the production database to provide real-time data protection and HA. In contrast, a restore from backup takes time and often has many moving parts - people, processes, software and systems – that can create a level of uncertainty during an outage that critical applications can’t afford. This is why backups play a secondary role for your most critical databases by complementing real-time solutions that can provide both data protection and availability. Before Data Guard, enterprises used storage remote-mirroring for real-time data protection and availability. Remote-mirroring is a sophisticated storage technology promoted as a generic infrastructure solution that makes a simple promise – whatever is written to a primary volume will also be written to the mirrored volume at a remote site. Keeping this promise is also what causes data loss and downtime when the data written to primary volumes is corrupt – the same corruption is faithfully mirrored to the remote volume making both copies unusable. This happens because remote-mirroring is a generic process. It has no  intrinsic knowledge of Oracle data structures to enable advanced protection, nor can it perform independent Oracle validation BEFORE changes are applied to the remote copy. There is also nothing to prevent human error (e.g. a storage admin accidentally deleting critical files) from also impacting the remote mirrored copy. Remote-mirroring tricks users by creating a false impression that there are two separate copies of the Oracle Database. In truth; while remote-mirroring maintains two copies of the data on different volumes, both are part of a single closely coupled system. Not only will remote-mirroring propagate corruptions and administrative errors, but the changes applied to the mirrored volume are a result of the same Oracle code path that applied the change to the source volume. There is no isolation, either from a storage mirroring perspective or from an Oracle software perspective.  Bottom line, storage remote-mirroring lacks both the smarts and isolation level necessary to provide true data protection. Active Data Guard offers much more than storage remote-mirroring when your objective is protecting your enterprise from downtime and data loss. Like remote-mirroring, an Active Data Guard replica is an exact block for block copy of the primary. Unlike remote-mirroring, an Active Data Guard replica is NOT a tightly coupled copy of the source volumes - it is a completely independent Oracle Database. Active Data Guard’s inherent knowledge of Oracle data block and redo structures enables a separate Oracle Database using a different Oracle code path than the primary to use the full complement of Oracle data validation methods before changes are applied to the synchronized copy. These include: physical check sum, logical intra-block checking, lost write validation, and automatic block repair. The figure below illustrates the stark difference between the knowledge that remote-mirroring can discern from an Oracle data block and what Active Data Guard can discern. An Active Data Guard standby also provides a range of additional services enabled by the fact that it is a running Oracle Database - not just a mirrored copy of data files. An Active Data Guard standby database can be open read-only while it is synchronizing with the primary. This enables read-only workloads to be offloaded from the primary system and run on the active standby - boosting performance by utilizing all assets. An Active Data Guard standby can also be used to implement many types of system and database maintenance in rolling fashion. Maintenance and upgrades are first implemented on the standby while production runs unaffected at the primary. After the primary and standby are synchronized and all changes have been validated, the production workload is quickly switched to the standby. The only downtime is the time required for user connections to transfer from one system to the next. These capabilities further expand the expectations of availability offered by a data protection solution beyond what is possible to do using storage remote-mirroring. So don’t be fooled by appearances.  Storage remote-mirroring and Active Data Guard replication may look similar on the surface - but the devil is in the details. Only Active Data Guard has the smarts, the isolation, and the simplicity, to provide the best data protection and availability for the Oracle Database. Stay tuned for future blog posts that dive into the many differences between storage remote-mirroring and Active Data Guard along the dimensions of data protection, data availability, cost, asset utilization and return on investment. For additional information on Active Data Guard, see: Active Data Guard Technical White Paper Active Data Guard vs Storage Remote-Mirroring Active Data Guard Home Page on the Oracle Technology Network

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