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• Windows 8.1 Will Start Encrypting Hard Drives By Default: Everything You Need to Know

  - by Chris Hoffman

  Windows 8.1 will automatically encrypt the storage on modern Windows PCs. This will help protect your files in case someone steals your laptop and tries to get at them, but it has important ramifications for data recovery. Previously, “BitLocker” was available on Professional and Enterprise editions of Windows, while “Device Encryption” was available on Windows RT and Windows Phone. Device encryption is included with all editions of Windows 8.1 — and it’s on by default. When Your Hard Drive Will Be Encrypted Windows 8.1 includes “Pervasive Device Encryption.” This works a bit differently from the standard BitLocker feature that has been included in Professional, Enterprise, and Ultimate editions of Windows for the past few versions. Before Windows 8.1 automatically enables Device Encryption, the following must be true: The Windows device “must support connected standby and meet the Windows Hardware Certification Kit (HCK) requirements for TPM and SecureBoot on ConnectedStandby systems.”  (Source) Older Windows PCs won’t support this feature, while new Windows 8.1 devices you pick up will have this feature enabled by default. When Windows 8.1 installs cleanly and the computer is prepared, device encryption is “initialized” on the system drive and other internal drives. Windows uses a clear key at this point, which is removed later when the recovery key is successfully backed up. The PC’s user must log in with a Microsoft account with administrator privileges or join the PC to a domain. If a Microsoft account is used, a recovery key will be backed up to Microsoft’s servers and encryption will be enabled. If a domain account is used, a recovery key will be backed up to Active Directory Domain Services and encryption will be enabled. If you have an older Windows computer that you’ve upgraded to Windows 8.1, it may not support Device Encryption. If you log in with a local user account, Device Encryption won’t be enabled. If you upgrade your Windows 8 device to Windows 8.1, you’ll need to enable device encryption, as it’s off by default when upgrading. Recovering An Encrypted Hard Drive Device encryption means that a thief can’t just pick up your laptop, insert a Linux live CD or Windows installer disc, and boot the alternate operating system to view your files without knowing your Windows password. It means that no one can just pull the hard drive from your device, connect the hard drive to another computer, and view the files. We’ve previously explained that your Windows password doesn’t actually secure your files. With Windows 8.1, average Windows users will finally be protected with encryption by default. However, there’s a problem — if you forget your password and are unable to log in, you’d also be unable to recover your files. This is likely why encryption is only enabled when a user logs in with a Microsoft account (or connects to a domain). Microsoft holds a recovery key, so you can gain access to your files by going through a recovery process. As long as you’re able to authenticate using your Microsoft account credentials — for example, by receiving an SMS message on the cell phone number connected to your Microsoft account — you’ll be able to recover your encrypted data. With Windows 8.1, it’s more important than ever to configure your Microsoft account’s security settings and recovery methods so you’ll be able to recover your files if you ever get locked out of your Microsoft account. Microsoft does hold the recovery key and would be capable of providing it to law enforcement if it was requested, which is certainly a legitimate concern in the age of PRISM. However, this encryption still provides protection from thieves picking up your hard drive and digging through your personal or business files. If you’re worried about a government or a determined thief who’s capable of gaining access to your Microsoft account, you’ll want to encrypt your hard drive with software that doesn’t upload a copy of your recovery key to the Internet, such as TrueCrypt. How to Disable Device Encryption There should be no real reason to disable device encryption. If nothing else, it’s a useful feature that will hopefully protect sensitive data in the real world where people — and even businesses — don’t enable encryption on their own. As encryption is only enabled on devices with the appropriate hardware and will be enabled by default, Microsoft has hopefully ensured that users won’t see noticeable slow-downs in performance. Encryption adds some overhead, but the overhead can hopefully be handled by dedicated hardware. If you’d like to enable a different encryption solution or just disable encryption entirely, you can control this yourself. To do so, open the PC settings app — swipe in from the right edge of the screen or press Windows Key + C, click the Settings icon, and select Change PC settings. Navigate to PC and devices -> PC info. At the bottom of the PC info pane, you’ll see a Device Encryption section. Select Turn Off if you want to disable device encryption, or select Turn On if you want to enable it — users upgrading from Windows 8 will have to enable it manually in this way. Note that Device Encryption can’t be disabled on Windows RT devices, such as Microsoft’s Surface RT and Surface 2. If you don’t see the Device Encryption section in this window, you’re likely using an older device that doesn’t meet the requirements and thus doesn’t support Device Encryption. For example, our Windows 8.1 virtual machine doesn’t offer Device Encryption configuration options. This is the new normal for Windows PCs, tablets, and devices in general. Where files on typical PCs were once ripe for easy access by thieves, Windows PCs are now encrypted by default and recovery keys are sent to Microsoft’s servers for safe keeping. This last part may be a bit creepy, but it’s easy to imagine average users forgetting their passwords — they’d be very upset if they lost all their files because they had to reset their passwords. It’s also an improvement over Windows PCs being completely unprotected by default.     

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• The DOS DEBUG Environment

  - by MarkPearl

  Today I thought I would go back in time and have a look at the DEBUG command that has been available since the beginning of dawn in DOS, MS-DOS and Microsoft Windows. up to today I always knew it was there, but had no clue on how to use it so for those that are interested this might be a great geek party trick to pull out when you want the awe the younger generation and want to show them what “real” programming is about. But wait, you will have to do it relatively quickly as it seems like DEBUG was finally dumped from the Windows group in Windows 7. Not to worry, pull out that Windows XP box which will get you even more geek points and you can still poke DEBUG a bit. So, for those that are interested and want to find out a bit about the history of DEBUG read the wiki link here. That all put aside, lets get our hands dirty.. How to Start DEBUG in Windows Make sure your version of Windows supports DEBUG. Open up a console window Make a directory where you want to play with debug – in my instance I called it C221 Enter the directory and type Debug You will get a response with a – as illustrated in the image below…   The commands available in DEBUG There are several commands available in DEBUG. The most common ones are A (Assemble) R (Register) T (Trace) G (Go) D (Dump or Display) U (Unassemble) E (Enter) P (Proceed) N (Name) L (Load) W (Write) H (Hexadecimal) I (Input) O (Output) Q (Quit) I am not going to cover all these commands, but what I will do is go through a few of them briefly. A is for Assemble Command (to write code) The A command translates assembly language statements into machine code. It is quite useful for writing small assembly programs. Below I have written a very basic assembly program. The code typed out is as follows mov ax,0015 mov cx,0023 sub cx,ax mov [120],al mov cl,[120]A nop R is for Register (to jump to a point in memory) The r command turns out to be one of the most frequent commands you will use in DEBUG. It allows you to view the contents of registers and to change their values. It can be used with the following combinations… R – Displays the contents of all the registers R f – Displays the flags register R register_name – Displays the contents of a specific register All three methods are illustrated in the image above T is for Trace (To execute a program step by step) The t command allows us to execute the program step by step. Before we can trace the program we need to point back to the beginning of the program. We do this by typing in r ip, which moves us back to memory point 100. We then type trace which executes the first line of code (line 100) (As shown in the image below starting from the red arrow). You can see from the above image that the register AX now contains 0015 as per our instruction mov ax,0015 You can also see that the IP points to line 0103 which has the MOV CX,0023 command If we type t again it will now execute the second line of the program which moves 23 in the cx register. Again, we can see that the line of code was executed and that the CX register now holds the value of 23. What I would like to highlight now is the section underlined in red. These are the status flags. The ones we are going to look at now are 1st (NV), 4th (PL), 5th (NZ) & 8th (NC) NV means no overflow, the alternate would be OV PL means that the sign of the previous arithmetic operation was Plus, the alternate would be NG (Negative) NZ means that the results of the previous arithmetic operation operation was Not Zero, the alternate would be ZR NC means that No final Carry resulted from the previous arithmetic operation. CY means that there was a final Carry. We could now follow this process of entering the t command until the entire program is executed line by line. G is for Go (To execute a program up to a certain line number) So we have looked at executing a program line by line, which is fine if your program is minuscule BUT totally unpractical if we have any decent sized program. A quicker way to run some lines of code is to use the G command. The ‘g’ command executes a program up to a certain specified point. It can be used in connection with the the reset IP command. You would set your initial point and then run the G command with the line you want to end on. P is for Proceed (Similar to trace but slightly more streamlined) Another command similar to trace is the proceed command. All that the p command does is if it is called and it encounters a CALL, INT or LOOP command it terminates the program execution. In the example below I modified our example program to include an int 20 at the end of it as illustrated in the image below… Then when executing the code when I encountered the int 20 command I typed the P command and the program terminated normally (illustrated below). D is for Dump (or for those more polite Display) So, we have all these assembly lines of code, but if you have ever opened up an exe or com file in a text/hex editor, it looks nothing like assembly code. The D command is a way that we can see what our code looks like in memory (or in a hex editor). If we examined the image above, we can see that Debug is storing our assembly code with each instruction following immediately after the previous one. For instance in memory address 110 we have int and 111 we have 20. If we examine the dump of memory we can see at memory point 110 CD is stored and at memory point 111 20 is stored. U is for Unassemble (or Convert Machine code to Assembly Code) So up to now we have gone through a bunch of commands, but probably one of the most useful is the U command. Let’s say we don’t understand machine code so well and so instead we want to see it in its equivalent assembly code. We can type the U command followed by the start memory point, followed by the end memory point and it will show us the assembly code equivalent of the machine code. E is for a bunch of things… The E command can be used for a bunch of things… One example is to enter data or machine code instructions directly into memory. It can also be used to display the contents of memory locations. I am not going to worry to much about it in this post. N / L / W is for Name, Load & Write So we have written out assembly code in debug, and now we want to save it to disk, or write it as a com file or load it. This is where the N, L & W command come in handy. The n command is used to give a name to the executable program file and is pretty simple to use. The w command is a bit trickier. It saves to disk all the memory between point bx and point cx so you need to specify the bx memory address and the cx memory address for it to write your code. Let’s look at an example illustrated below. You do this by calling the r command followed by the either bx or cx. We can then go to the directory where we were working and will see the new file with the name we specified. The L command is relatively simple. You would first specify the name of the file you would like to load using the N command, and then call the L command. Q is for Quit The last command that I am going to write about in this post is the Q command. Simply put, calling the Q command exits DEBUG. Commands we did not Cover Out of the standard DEBUG commands we covered A, T, G, D, U, E, P, R, N, L & W. The ones we did not cover were H, I & O – I might make mention of these in a later post, but for the basics they are not really needed. Some Useful Resources Please note this post is based on the COS2213 handouts for UNISA A Guide to DEBUG - http://mirror.href.com/thestarman/asm/debug/debug.htm#NT

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• Authenticating clients in the new WCF Http stack

  - by cibrax

  About this time last year, I wrote a couple of posts about how to use the “Interceptors” from the REST starker kit for implementing several authentication mechanisms like “SAML”, “Basic Authentication” or “OAuth” in the WCF Web programming model. The things have changed a lot since then, and Glenn finally put on our hands a new version of the Web programming model that deserves some attention and I believe will help us a lot to build more Http oriented services in the .NET stack. What you can get today from wcf.codeplex.com is a preview with some cool features like Http Processors (which I already discussed here), a new and improved version of the HttpClient library, Dependency injection and better TDD support among others. However, the framework still does not support an standard way of doing client authentication on the services (This is something planned for the upcoming releases I believe). For that reason, moving the existing authentication interceptors to this new programming model was one of the things I did in the last few days. In order to make authentication simple and easy to extend,  I first came up with a model based on what I called “Authentication Interceptors”. An authentication interceptor maps to an existing Http authentication mechanism and implements the following interface, public interface IAuthenticationInterceptor{ string Scheme { get; } bool DoAuthentication(HttpRequestMessage request, HttpResponseMessage response, out IPrincipal principal);} An authentication interceptors basically needs to returns the http authentication schema that implements in the property “Scheme”, and implements the authentication mechanism in the method “DoAuthentication”. As you can see, this last method “DoAuthentication” only relies on the HttpRequestMessage and HttpResponseMessage classes, making the testing of this interceptor very simple (There is no need to do some black magic with the WCF context or messages). After this, I implemented a couple of interceptors for supporting basic authentication and brokered authentication with SAML (using WIF) in my services. The following code illustrates how the basic authentication interceptors looks like. public class BasicAuthenticationInterceptor : IAuthenticationInterceptor{ Func<UsernameAndPassword, bool> userValidation; string realm;  public BasicAuthenticationInterceptor(Func<UsernameAndPassword, bool> userValidation, string realm) { if (userValidation == null) throw new ArgumentNullException("userValidation");  if (string.IsNullOrEmpty(realm)) throw new ArgumentNullException("realm");  this.userValidation = userValidation; this.realm = realm; }  public string Scheme { get { return "Basic"; } }  public bool DoAuthentication(HttpRequestMessage request, HttpResponseMessage response, out IPrincipal principal) { string[] credentials = ExtractCredentials(request); if (credentials.Length == 0 || !AuthenticateUser(credentials[0], credentials[1])) { response.StatusCode = HttpStatusCode.Unauthorized; response.Content = new StringContent("Access denied"); response.Headers.WwwAuthenticate.Add(new AuthenticationHeaderValue("Basic", "realm=" + this.realm));  principal = null;  return false; } else { principal = new GenericPrincipal(new GenericIdentity(credentials[0]), new string[] {});  return true; } }  private string[] ExtractCredentials(HttpRequestMessage request) { if (request.Headers.Authorization != null && request.Headers.Authorization.Scheme.StartsWith("Basic")) { string encodedUserPass = request.Headers.Authorization.Parameter.Trim();  Encoding encoding = Encoding.GetEncoding("iso-8859-1"); string userPass = encoding.GetString(Convert.FromBase64String(encodedUserPass)); int separator = userPass.IndexOf(':');  string[] credentials = new string[2]; credentials[0] = userPass.Substring(0, separator); credentials[1] = userPass.Substring(separator + 1);  return credentials; }  return new string[] { }; }  private bool AuthenticateUser(string username, string password) { var usernameAndPassword = new UsernameAndPassword { Username = username, Password = password };  if (this.userValidation(usernameAndPassword)) { return true; }  return false; }} This interceptor receives in the constructor a callback in the form of a Func delegate for authenticating the user and the “realm”, which is required as part of the implementation. The rest is a general implementation of the basic authentication mechanism using standard http request and response messages. I also implemented another interceptor for authenticating a SAML token with WIF. public class SamlAuthenticationInterceptor : IAuthenticationInterceptor{ SecurityTokenHandlerCollection handlers = null;  public SamlAuthenticationInterceptor(SecurityTokenHandlerCollection handlers) { if (handlers == null) throw new ArgumentNullException("handlers");  this.handlers = handlers; }  public string Scheme { get { return "saml"; } }  public bool DoAuthentication(HttpRequestMessage request, HttpResponseMessage response, out IPrincipal principal) { SecurityToken token = ExtractCredentials(request);  if (token != null) { ClaimsIdentityCollection claims = handlers.ValidateToken(token);  principal = new ClaimsPrincipal(claims);  return true; } else { response.StatusCode = HttpStatusCode.Unauthorized; response.Content = new StringContent("Access denied");  principal = null;  return false; } }  private SecurityToken ExtractCredentials(HttpRequestMessage request) { if (request.Headers.Authorization != null && request.Headers.Authorization.Scheme == "saml") { XmlTextReader xmlReader = new XmlTextReader(new StringReader(request.Headers.Authorization.Parameter));  var col = SecurityTokenHandlerCollection.CreateDefaultSecurityTokenHandlerCollection(); SecurityToken token = col.ReadToken(xmlReader);  return token; }  return null; }}This implementation receives a “SecurityTokenHandlerCollection” instance as part of the constructor. This class is part of WIF, and basically represents a collection of token managers to know how to handle specific xml authentication tokens (SAML is one of them). I also created a set of extension methods for injecting these interceptors as part of a service route when the service is initialized. var basicAuthentication = new BasicAuthenticationInterceptor((u) => true, "ContactManager");var samlAuthentication = new SamlAuthenticationInterceptor(serviceConfiguration.SecurityTokenHandlers); // use MEF for providing instancesvar catalog = new AssemblyCatalog(typeof(Global).Assembly);var container = new CompositionContainer(catalog);var configuration = new ContactManagerConfiguration(container); RouteTable.Routes.AddServiceRoute<ContactResource>("contact", configuration, basicAuthentication, samlAuthentication);RouteTable.Routes.AddServiceRoute<ContactsResource>("contacts", configuration, basicAuthentication, samlAuthentication); In the code above, I am injecting the basic authentication and saml authentication interceptors in the “contact” and “contacts” resource implementations that come as samples in the code preview. I will use another post to discuss more in detail how the brokered authentication with SAML model works with this new WCF Http bits. The code is available to download in this location.

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• Using SQL Execution Plans to discover the Swedish alphabet

  - by Rob Farley

  SQL Server is quite remarkable in a bunch of ways. In this post, I’m using the way that the Query Optimizer handles LIKE to keep it SARGable, the Execution Plans that result, Collations, and PowerShell to come up with the Swedish alphabet. SARGability is the ability to seek for items in an index according to a particular set of criteria. If you don’t have SARGability in play, you need to scan the whole index (or table if you don’t have an index). For example, I can find myself in the phonebook easily, because it’s sorted by LastName and I can find Farley in there by moving to the Fs, and so on. I can’t find everyone in my suburb easily, because the phonebook isn’t sorted that way. I can’t even find people who have six letters in their last name, because also the book is sorted by LastName, it’s not sorted by LEN(LastName). This is all stuff I’ve looked at before, including in the talk I gave at SQLBits in October 2010. If I try to find everyone who’s names start with F, I can do that using a query a bit like: SELECT LastName FROM dbo.PhoneBook WHERE LEFT(LastName,1) = 'F'; Unfortunately, the Query Optimizer doesn’t realise that all the entries that satisfy LEFT(LastName,1) = 'F' will be together, and it has to scan the whole table to find them. But if I write: SELECT LastName FROM dbo.PhoneBook WHERE LastName LIKE 'F%'; then SQL is smart enough to understand this, and performs an Index Seek instead. To see why, I look further into the plan, in particular, the properties of the Index Seek operator. The ToolTip shows me what I’m after: You’ll see that it does a Seek to find any entries that are at least F, but not yet G. There’s an extra Predicate in there (a Residual Predicate if you like), which checks that each LastName is really LIKE F% – I suppose it doesn’t consider that the Seek Predicate is quite enough – but most of the benefit is seen by its working out the Seek Predicate, filtering to just the “at least F but not yet G” section of the data. This got me curious though, particularly about where the G comes from, and whether I could leverage it to create the Swedish alphabet. I know that in the Swedish language, there are three extra letters that appear at the end of the alphabet. One of them is ä that appears in the word Västerås. It turns out that Västerås is quite hard to find in an index when you’re looking it up in a Swedish map. I talked about this briefly in my five-minute talk on Collation from SQLPASS (the one which was slightly less than serious). So by looking at the plan, I can work out what the next letter is in the alphabet of the collation used by the column. In other words, if my alphabet were Swedish, I’d be able to tell what the next letter after F is – just in case it’s not G. It turns out it is… Yes, the Swedish letter after F is G. But I worked this out by using a copy of my PhoneBook table that used the Finnish_Swedish_CI_AI collation. I couldn’t find how the Query Optimizer calculates the G, and my friend Paul White (@SQL_Kiwi) tells me that it’s frustratingly internal to the QO. He’s particularly smart, even if he is from New Zealand. To investigate further, I decided to do some PowerShell, leveraging the Get-SqlPlan function that I blogged about recently (make sure you also have the SqlServerCmdletSnapin100 snap-in added). I started by indicating that I was going to use Finnish_Swedish_CI_AI as my collation of choice, and that I’d start whichever letter cam straight after the number 9. I figure that this is a cheat’s way of guessing the first letter of the alphabet (but it doesn’t actually work in Unicode – luckily I’m using varchar not nvarchar. Actually, there are a few aspects of this code that only work using ASCII, so apologies if you were wanting to apply it to Greek, Japanese, etc). I also initialised my $alphabet variable. $collation = 'Finnish_Swedish_CI_AI'; $firstletter = '9'; $alphabet = ''; Now I created the table for my test. A single field would do, and putting a Clustered Index on it would suffice for the Seeks. Invoke-Sqlcmd -server . -data tempdb -query "create table dbo.collation_test (col varchar(10) collate $collation primary key);" Now I get into the looping. $c = $firstletter; $stillgoing = $true; while ($stillgoing) { I construct the query I want, seeking for entries which start with whatever $c has reached, and get the plan for it: $query = "select col from dbo.collation_test where col like '$($c)%';"; [xml] $pl = get-sqlplan $query "." "tempdb"; At this point, my $pl variable is a scary piece of XML, representing the execution plan. A bit of hunting through it showed me that the EndRange element contained what I was after, and that if it contained NULL, then I was done. $stillgoing = ($pl.ShowPlanXML.BatchSequence.Batch.Statements.StmtSimple.QueryPlan.RelOp.IndexScan.SeekPredicates.SeekPredicateNew.SeekKeys.EndRange -ne $null); Now I could grab the value out of it (which came with apostrophes that needed stripping), and append that to my $alphabet variable.   if ($stillgoing)   {  $c=$pl.ShowPlanXML.BatchSequence.Batch.Statements.StmtSimple.QueryPlan.RelOp.IndexScan.SeekPredicates.SeekPredicateNew.SeekKeys.EndRange.RangeExpressions.ScalarOperator.ScalarString.Replace("'","");     $alphabet += $c;   } Finally, finishing the loop, dropping the table, and showing my alphabet! } Invoke-Sqlcmd -server . -data tempdb -query "drop table dbo.collation_test;"; $alphabet; When I run all this, I see that the Swedish alphabet is ABCDEFGHIJKLMNOPQRSTUVXYZÅÄÖ, which matches what I see at Wikipedia. Interesting to see that the letters on the end are still there, even with Case Insensitivity. Turns out they’re not just “letters with accents”, they’re letters in their own right. I’m sure you gave up reading long ago, and really aren’t that fazed about the idea of doing this using PowerShell. I chose PowerShell because I’d already come up with an easy way of grabbing the estimated plan for a query, and PowerShell does allow for easy navigation of XML. I find the most interesting aspect of this as the fact that the Query Optimizer uses the next letter of the alphabet to maintain the SARGability of LIKE. I’m hoping they do something similar for a whole bunch of operations. Oh, and the fact that you know how to find stuff in the IKEA catalogue. Footnote: If you are interested in whether this works in other languages, you might want to consider the following screenshot, which shows that in principle, it should work with Japanese. It might be a bit harder to run this in PowerShell though, as I’m not sure how it translates. In Hiragana, the Japanese alphabet starts ?, ?, ?, ?, ?, ...

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• SQL Server Split() Function

  - by HighAltitudeCoder

  Title goes here   Ever wanted a dbo.Split() function, but not had the time to debug it completely?  Let me guess - you are probably working on a stored procedure with 50 or more parameters; two or three of them are parameters of differing types, while the other 47 or so all of the same type (id1, id2, id3, id4, id5...).  Worse, you've found several other similar stored procedures with the ONLY DIFFERENCE being the number of like parameters taped to the end of the parameter list. If this is the situation you find yourself in now, you may be wondering, "why am I working with three different copies of what is basically the same stored procedure, and why am I having to maintain changes in three different places?  Can't I have one stored procedure that accomplishes the job of all three? My answer to you: YES!  Here is the Split() function I've created.    /******************************************************************************                                       Split.sql   ******************************************************************************/ /******************************************************************************   Split a delimited string into sub-components and return them as a table.   Parameter 1: Input string which is to be split into parts. Parameter 2: Delimiter which determines the split points in input string. Works with space or spaces as delimiter. Split() is apostrophe-safe.   SYNTAX: SELECT * FROM Split('Dvorak,Debussy,Chopin,Holst', ',') SELECT * FROM Split('Denver|Seattle|San Diego|New York', '|') SELECT * FROM Split('Denver is the super-awesomest city of them all.', ' ')   ******************************************************************************/ USE AdventureWorks GO   IF EXISTS       (SELECT *       FROM sysobjects       WHERE xtype = 'TF'       AND name = 'Split'       ) BEGIN       DROP FUNCTION Split END GO   CREATE FUNCTION Split (       @InputString                  VARCHAR(8000),       @Delimiter                    VARCHAR(50) )   RETURNS @Items TABLE (       Item                          VARCHAR(8000) )   AS BEGIN       IF @Delimiter = ' '       BEGIN             SET @Delimiter = ','             SET @InputString = REPLACE(@InputString, ' ', @Delimiter)       END         IF (@Delimiter IS NULL OR @Delimiter = '')             SET @Delimiter = ','   --INSERT INTO @Items VALUES (@Delimiter) -- Diagnostic --INSERT INTO @Items VALUES (@InputString) -- Diagnostic         DECLARE @Item                 VARCHAR(8000)       DECLARE @ItemList       VARCHAR(8000)       DECLARE @DelimIndex     INT         SET @ItemList = @InputString       SET @DelimIndex = CHARINDEX(@Delimiter, @ItemList, 0)       WHILE (@DelimIndex != 0)       BEGIN             SET @Item = SUBSTRING(@ItemList, 0, @DelimIndex)             INSERT INTO @Items VALUES (@Item)               -- Set @ItemList = @ItemList minus one less item             SET @ItemList = SUBSTRING(@ItemList, @DelimIndex+1, LEN(@ItemList)-@DelimIndex)             SET @DelimIndex = CHARINDEX(@Delimiter, @ItemList, 0)       END -- End WHILE         IF @Item IS NOT NULL -- At least one delimiter was encountered in @InputString       BEGIN             SET @Item = @ItemList             INSERT INTO @Items VALUES (@Item)       END         -- No delimiters were encountered in @InputString, so just return @InputString       ELSE INSERT INTO @Items VALUES (@InputString)         RETURN   END -- End Function GO   ---- Set Permissions --GRANT SELECT ON Split TO UserRole1 --GRANT SELECT ON Split TO UserRole2 --GO   The syntax is basically as follows: SELECT <fields> FROM Table 1 JOIN Table 2 ON ... JOIN Table 3 ON ... WHERE LOGICAL CONDITION A AND LOGICAL CONDITION B AND LOGICAL CONDITION C AND TABLE2.Id IN (SELECT * FROM Split(@IdList, ',')) @IdList is a parameter passed into the stored procedure, and the comma (',') is the delimiter you have chosen to split the parameter list on. You can also use it like this: SELECT <fields> FROM Table 1 JOIN Table 2 ON ... JOIN Table 3 ON ... WHERE LOGICAL CONDITION A AND LOGICAL CONDITION B AND LOGICAL CONDITION C HAVING COUNT(SELECT * FROM Split(@IdList, ',') Similarly, it can be used in other aggregate functions at run-time: SELECT MIN(SELECT * FROM Split(@IdList, ','), <fields> FROM Table 1 JOIN Table 2 ON ... JOIN Table 3 ON ... WHERE LOGICAL CONDITION A AND LOGICAL CONDITION B AND LOGICAL CONDITION C GROUP BY <fields> Now that I've (hopefully effectively) explained the benefits to using this function and implementing it in one or more of your database objects, let me warn you of a caveat that you are likely to encounter.  You may have a team member who waits until the right moment to ask you a pointed question: "Doesn't this function just do the same thing as using the IN function?  Why didn't you just use that instead?  In other words, why bother with this function?" What's happening is, one or more team members has failed to understand the reason for implementing this kind of function in the first place.  (Note: this is THE MOST IMPORTANT ASPECT OF THIS POST). Allow me to outline a few pros to implementing this function, so you may effectively parry this question.  Touche. 1) Code consolidation.  You don't have to maintain what is basically the same code and logic, but with varying numbers of the same parameter in several SQL objects.  I'm not going to go into the cons related to using this function, because the afore mentioned team member is probably more than adept at pointing these out.  Remember, the real positive contribution is ou are decreasing the liklihood that your team fails to update all (x) duplicate copies of what are basically the same stored procedure, and so on...  This is the classic downside to duplicate code.  It is a virus, and you should kill it. You might be better off rejecting your team member's question, and responding with your own: "Would you rather maintain the same logic in multiple different stored procedures, and hope that the team doesn't forget to always update all of them at the same time?".  In his head, he might be thinking "yes, I would like to maintain several different copies of the same stored procedure", although you probably will not get such a direct response.  2) Added flexibility - you can use the Split function elsewhere, and for splitting your data in different ways.  Plus, you can use any kind of delimiter you wish.  How can you know today the ways in which you might want to examine your data tomorrow?  Segue to my next point. 3) Because the function takes a delimiter parameter, you can split the data in any number of ways.  This greatly increases the utility of such a function and enables your team to work with the data in a variety of different ways in the future.  You can split on a single char, symbol, word, or group of words.  You can split on spaces.  (The list goes on... test it out). Finally, you can dynamically define the behavior of a stored procedure (or other SQL object) at run time, through the use of this function.  Rather than have several objects that accomplish almost the same thing, why not have only one instead?

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• Java EE 6 and NoSQL/MongoDB on GlassFish using JPA and EclipseLink 2.4 (TOTD #175)

  - by arungupta

  TOTD #166 explained how to use MongoDB in your Java EE 6 applications. The code in that tip used the APIs exposed by the MongoDB Java driver and so requires you to learn a new API. However if you are building Java EE 6 applications then you are already familiar with Java Persistence API (JPA). Eclipse Link 2.4, scheduled to release as part of Eclipse Juno, provides support for NoSQL databases by mapping a JPA entity to a document. Their wiki provides complete explanation of how the mapping is done. This Tip Of The Day (TOTD) will show how you can leverage that support in your Java EE 6 applications deployed on GlassFish 3.1.2. Before we dig into the code, here are the key concepts ... A POJO is mapped to a NoSQL data source using @NoSQL or <no-sql> element in "persistence.xml". A subset of JPQL and Criteria query are supported, based upon the underlying data store Connection properties are defined in "persistence.xml" Now, lets lets take a look at the code ... Download the latest EclipseLink 2.4 Nightly Bundle. There is a Installer, Source, and Bundle - make sure to download the Bundle link (20120410) and unzip. Download GlassFish 3.1.2 zip and unzip. Install the Eclipse Link 2.4 JARs in GlassFish Remove the following JARs from "glassfish/modules": org.eclipse.persistence.antlr.jar org.eclipse.persistence.asm.jar org.eclipse.persistence.core.jar org.eclipse.persistence.jpa.jar org.eclipse.persistence.jpa.modelgen.jar org.eclipse.persistence.moxy.jar org.eclipse.persistence.oracle.jar Add the following JARs from Eclipse Link 2.4 nightly build to "glassfish/modules": org.eclipse.persistence.antlr_3.2.0.v201107111232.jar org.eclipse.persistence.asm_3.3.1.v201107111215.jar org.eclipse.persistence.core.jpql_2.4.0.v20120407-r11132.jar org.eclipse.persistence.core_2.4.0.v20120407-r11132.jar org.eclipse.persistence.jpa.jpql_2.0.0.v20120407-r11132.jar org.eclipse.persistence.jpa.modelgen_2.4.0.v20120407-r11132.jar org.eclipse.persistence.jpa_2.4.0.v20120407-r11132.jar org.eclipse.persistence.moxy_2.4.0.v20120407-r11132.jar org.eclipse.persistence.nosql_2.4.0.v20120407-r11132.jar org.eclipse.persistence.oracle_2.4.0.v20120407-r11132.jar Start MongoDB Download latest MongoDB from here (2.0.4 as of this writing). Create the default data directory for MongoDB as: sudo mkdir -p /data/db/sudo chown `id -u` /data/db Refer to Quickstart for more details. Start MongoDB as: arungup-mac:mongodb-osx-x86_64-2.0.4 <arungup> ->./bin/mongod./bin/mongod --help for help and startup optionsMon Apr  9 12:56:02 [initandlisten] MongoDB starting : pid=3124 port=27017 dbpath=/data/db/ 64-bit host=arungup-mac.localMon Apr  9 12:56:02 [initandlisten] db version v2.0.4, pdfile version 4.5Mon Apr  9 12:56:02 [initandlisten] git version: 329f3c47fe8136c03392c8f0e548506cb21f8ebfMon Apr  9 12:56:02 [initandlisten] build info: Darwin erh2.10gen.cc 9.8.0 Darwin Kernel Version 9.8.0: Wed Jul 15 16:55:01 PDT 2009; root:xnu-1228.15.4~1/RELEASE_I386 i386 BOOST_LIB_VERSION=1_40Mon Apr  9 12:56:02 [initandlisten] options: {}Mon Apr  9 12:56:02 [initandlisten] journal dir=/data/db/journalMon Apr  9 12:56:02 [initandlisten] recover : no journal files present, no recovery neededMon Apr  9 12:56:02 [websvr] admin web console waiting for connections on port 28017Mon Apr  9 12:56:02 [initandlisten] waiting for connections on port 27017 Check out the JPA/NoSQL sample from SVN repository. The complete source code built in this TOTD can be downloaded here. Create Java EE 6 web app Create a Java EE 6 Maven web app as: mvn archetype:generate -DarchetypeGroupId=org.codehaus.mojo.archetypes -DarchetypeArtifactId=webapp-javaee6 -DgroupId=model -DartifactId=javaee-nosql -DarchetypeVersion=1.5 -DinteractiveMode=false Copy the model files from the checked out workspace to the generated project as: cd javaee-nosqlcp -r ~/code/workspaces/org.eclipse.persistence.example.jpa.nosql.mongo/src/model src/main/java Copy "persistence.xml" mkdir src/main/resources cp -r ~/code/workspaces/org.eclipse.persistence.example.jpa.nosql.mongo/src/META-INF ./src/main/resources Add the following dependencies: <dependency> <groupId>org.eclipse.persistence</groupId> <artifactId>org.eclipse.persistence.jpa</artifactId> <version>2.4.0-SNAPSHOT</version> <scope>provided</scope></dependency><dependency> <groupId>org.eclipse.persistence</groupId> <artifactId>org.eclipse.persistence.nosql</artifactId> <version>2.4.0-SNAPSHOT</version></dependency><dependency> <groupId>org.mongodb</groupId> <artifactId>mongo-java-driver</artifactId> <version>2.7.3</version></dependency> The first one is for the EclipseLink latest APIs, the second one is for EclipseLink/NoSQL support, and the last one is the MongoDB Java driver. And the following repository: <repositories> <repository> <id>EclipseLink Repo</id> <url>http://www.eclipse.org/downloads/download.php?r=1&amp;nf=1&amp;file=/rt/eclipselink/maven.repo</url> <snapshots> <enabled>true</enabled> </snapshots> </repository>  </repositories> Copy the "Test.java" to the generated project: mkdir src/main/java/examplecp -r ~/code/workspaces/org.eclipse.persistence.example.jpa.nosql.mongo/src/example/Test.java ./src/main/java/example/ This file contains the source code to CRUD the JPA entity to MongoDB. This sample is explained in detail on EclipseLink wiki. Create a new Servlet in "example" directory as: package example;import java.io.IOException;import java.io.PrintWriter;import javax.servlet.ServletException;import javax.servlet.annotation.WebServlet;import javax.servlet.http.HttpServlet;import javax.servlet.http.HttpServletRequest;import javax.servlet.http.HttpServletResponse;/** * @author Arun Gupta */@WebServlet(name = "TestServlet", urlPatterns = {"/TestServlet"})public class TestServlet extends HttpServlet { protected void processRequest(HttpServletRequest request, HttpServletResponse response) throws ServletException, IOException { response.setContentType("text/html;charset=UTF-8"); PrintWriter out = response.getWriter(); try { out.println("<html>"); out.println("<head>"); out.println("<title>Servlet TestServlet</title>"); out.println("</head>"); out.println("<body>"); out.println("<h1>Servlet TestServlet at " + request.getContextPath() + "</h1>"); try { Test.main(null); } catch (Exception ex) { ex.printStackTrace(); } out.println("</body>"); out.println("</html>"); } finally { out.close(); } } @Override protected void doGet(HttpServletRequest request, HttpServletResponse response) throws ServletException, IOException { processRequest(request, response); } @Override protected void doPost(HttpServletRequest request, HttpServletResponse response) throws ServletException, IOException { processRequest(request, response); }} Build the project and deploy it as: mvn clean packageglassfish3/bin/asadmin deploy --force=true target/javaee-nosql-1.0-SNAPSHOT.war Accessing http://localhost:8080/javaee-nosql/TestServlet shows the following messages in the server.log: connecting(EISLogin( platform=> MongoPlatform user name=> "" MongoConnectionSpec())) . . .Connected: User: Database: 2.7  Version: 2.7 . . .Executing MappedInteraction() spec => null properties => {mongo.collection=CUSTOMER, mongo.operation=INSERT} input => [DatabaseRecord( CUSTOMER._id => 4F848E2BDA0670307E2A8FA4 CUSTOMER.NAME => AMCE)]. . .Data access result: [{TOTALCOST=757.0, ORDERLINES=[{DESCRIPTION=table, LINENUMBER=1, COST=300.0}, {DESCRIPTION=balls, LINENUMBER=2, COST=5.0}, {DESCRIPTION=rackets, LINENUMBER=3, COST=15.0}, {DESCRIPTION=net, LINENUMBER=4, COST=2.0}, {DESCRIPTION=shipping, LINENUMBER=5, COST=80.0}, {DESCRIPTION=handling, LINENUMBER=6, COST=55.0},{DESCRIPTION=tax, LINENUMBER=7, COST=300.0}], SHIPPINGADDRESS=[{POSTALCODE=L5J1H7, PROVINCE=ON, COUNTRY=Canada, CITY=Ottawa,STREET=17 Jane St.}], VERSION=2, _id=4F848E2BDA0670307E2A8FA8,DESCRIPTION=Pingpong table, CUSTOMER__id=4F848E2BDA0670307E2A8FA7, BILLINGADDRESS=[{POSTALCODE=L5J1H8, PROVINCE=ON, COUNTRY=Canada, CITY=Ottawa, STREET=7 Bank St.}]}] You'll not see any output in the browser, just the output in the console. But the code can be easily modified to do so. Once again, the complete Maven project can be downloaded here. Do you want to try accessing relational and non-relational (aka NoSQL) databases in the same PU ?

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• User Guide to Dropbox Shared Folders

  - by Matthew Guay

  Dropbox is an incredibly useful tool for keeping all your files synced between your computers and the cloud.  Here we’re going to look at how you can keep all of your team on the same page with Dropbox shared folders. Creating a Shared Folder Setting up a shared folder in Dropbox is easy.  Add the files you want to share to a folder in Dropbox on your computer, then right-click in the folder, select Dropbox, and then choose Share This Folder.   Alternately, log into your Dropbox account online, click the drop-down menu beside the folder you want to share, and click Share this folder. Now, enter the email addresses of the people you want to share the folder with, and optionally enter a message explaining why you’re sharing the folder. The people you invite will receive an email inviting them to view and join the shared folder.  If they haven’t signed up for Dropbox, they can directly signup; otherwise, they can simply log into their Dropbox account and start adding or editing files. Shared folders have a slightly different icon in your Dropbox.  Notice the shared folder on the left has an icon with 2 people, while the folder on the right that is not shared, shows previews of its contents. See Your Shared Folder’s History Whenever your collaborators with your shared folders add or change files, you will see a tooltip notification telling you what changed. You can also view the changes online.  Log into your Dropbox account in your browser and select the Events tab.  This shows all changes to your Dropbox, but you can view only the changes in your shared folder by selecting its name on the left sidebar. Now you can see all recent changes to your folder, and can also see who added or removed each file.   On the bottom of the page, you can even add a comment that all the collaborators will see. If someone deleted a file you still need, you can restore it by clicking its link in this online history.  Or, you can view any deleted files by right-clicking in your Dropbox folder in Explorer.  Select Dropbox, and then click Show Deleted Files.   Get Notified When a Change is Made You’re not always in front of your computer; you’ve got a life beyond your projects, after all (at least hopefully).  If you really want to stay connected to what’s happening with your project, though, you can easily do that no matter where you are. Your shared Dropbox folder’s history page offers an RSS feed of all changes to the folder.  Click  the Subscribe to this feed hyperlink. Now, in the popup that opens, click “Copy to clipboard” so you can use this RSS feed. You can subscribe to RSS feeds through many web browsers, email clients, dedicated feed readers, and more.  In Firefox, Internet Explorer 7/8, or Opera, you can paste the feed address into your address bar and subscribe to the feed directly in your browser.   However, subscribing to the feed in a desktop application won’t help you much when you’re away from your computer.  One great option is to subscribe in the popular Google Reader.  Then you can check your feed from any browser, on any computer or mobile device. To add your Dropbox feed to Google Reader, log into Google Reader (link below), click Add a subscription on the top left, paste your RSS feed from Dropbox, and click Add.   Now you can see any changes to files or folders in Google Reader. You can even add your feed to your iGoogle homepage.  Click the Add it Now button on the right in the front page of Google Reader to add your feeds to iGoogle.   Now you can see updates on your files from your homepage.  If you’re using a different computer, just login to your Google account to see what’s happening. You can also access your Google Reader feeds from many programs and apps for most major Smartphones including iPhone, Windows Phone, and Blackberry. Receive a Tweet or Text When Changes are Made If you’re a hyper-connected individual, chances are you send and receive tweets on the go.  If so, this might be the best way for you to get notified when changes are made to your Dropbox shared folder.  To do this, first create a new Twitter account to publish your changes through.  If you don’t want the whole world to see your updates, click Settings and set your new Twitter account to Private. Once the new account is created, follow it with your normal Twitter account so you’ll see updates. Now, let’s publish our Dropbox RSS feed to Twitter.  Create an account with Twitterfeed (link below). Once your account is setup, add your feed to it.  Name your feed, and enter your Feed address from Dropbox.  Click Advanced Settings to make your feed work just like you want. In Advanced Settings, change the frequency to “Every 30 mins” to make sure you’re updated on changes as quick as possible.  You can also change other settings if you like. Click “Continue to Step 2”, and then click Twitter under the available services to add your account. Make sure your signed into your new Twitter account, and then click Authenticate Twitter. Allow the application. Now, finally, click Create Service. Whenever a change is made, you will receive a tweet via your new Twitter account.  And since you can receive tweets via text message or many mobile applications, you’ll never be very far away from your Dropbox changes!   Conclusion Dropbox shared folders are a great way to keep your whole team working together on the same files in a project.  And with these handy tricks, you can keep up with your shared files wherever you are! There are a lot of cool things you can do with Dropbox make sure to check out our posts on adding Dropbox to the Windows 7 Start menu, Accessing Dropbox files from Chrome, and Syncing your Pidgin Profile Across Multiple PCs. Links Signup or access your Dropbox account Google Reader Tweet your feed with Twitterfeed Similar Articles Productive Geek Tips How to Add and Manage Shared Folders on Windows Home ServerManage User Accounts in Windows Home ServerAdd "My Dropbox" to Your Windows 7 Start MenuComplete Guide to Networking Windows 7 with XP and VistaMoving Your Personal Data Folders in Windows Vista the Easy Way TouchFreeze Alternative in AutoHotkey The Icy Undertow Desktop Windows Home Server – Backup to LAN The Clear & Clean Desktop Use This Bookmarklet to Easily Get Albums Use AutoHotkey to Assign a Hotkey to a Specific Window Latest Software Reviews Tinyhacker Random Tips DVDFab 6 Revo Uninstaller Pro Registry Mechanic 9 for Windows PC Tools Internet Security Suite 2010 Office 2010 reviewed in depth by Ed Bott FoxClocks adds World Times in your Statusbar (Firefox) Have Fun Editing Photo Editing with Citrify Outlook Connector Upgrade Error Gadfly is a cool Twitter/Silverlight app Enable DreamScene in Windows 7

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• Fun with Aggregates

  - by Paul White

  There are interesting things to be learned from even the simplest queries.  For example, imagine you are given the task of writing a query to list AdventureWorks product names where the product has at least one entry in the transaction history table, but fewer than ten. One possible query to meet that specification is: SELECT p.Name FROM Production.Product AS p JOIN Production.TransactionHistory AS th ON p.ProductID = th.ProductID GROUP BY p.ProductID, p.Name HAVING COUNT_BIG(*) < 10; That query correctly returns 23 rows (execution plan and data sample shown below): The execution plan looks a bit different from the written form of the query: the base tables are accessed in reverse order, and the aggregation is performed before the join.  The general idea is to read all rows from the history table, compute the count of rows grouped by ProductID, merge join the results to the Product table on ProductID, and finally filter to only return rows where the count is less than ten. This ‘fully-optimized’ plan has an estimated cost of around 0.33 units.  The reason for the quote marks there is that this plan is not quite as optimal as it could be – surely it would make sense to push the Filter down past the join too?  To answer that, let’s look at some other ways to formulate this query.  This being SQL, there are any number of ways to write logically-equivalent query specifications, so we’ll just look at a couple of interesting ones.  The first query is an attempt to reverse-engineer T-SQL from the optimized query plan shown above.  It joins the result of pre-aggregating the history table to the Product table before filtering: SELECT p.Name FROM ( SELECT th.ProductID, cnt = COUNT_BIG(*) FROM Production.TransactionHistory AS th GROUP BY th.ProductID ) AS q1 JOIN Production.Product AS p ON p.ProductID = q1.ProductID WHERE q1.cnt < 10; Perhaps a little surprisingly, we get a slightly different execution plan: The results are the same (23 rows) but this time the Filter is pushed below the join!  The optimizer chooses nested loops for the join, because the cardinality estimate for rows passing the Filter is a bit low (estimate 1 versus 23 actual), though you can force a merge join with a hint and the Filter still appears below the join.  In yet another variation, the < 10 predicate can be ‘manually pushed’ by specifying it in a HAVING clause in the “q1” sub-query instead of in the WHERE clause as written above. The reason this predicate can be pushed past the join in this query form, but not in the original formulation is simply an optimizer limitation – it does make efforts (primarily during the simplification phase) to encourage logically-equivalent query specifications to produce the same execution plan, but the implementation is not completely comprehensive. Moving on to a second example, the following query specification results from phrasing the requirement as “list the products where there exists fewer than ten correlated rows in the history table”: SELECT p.Name FROM Production.Product AS p WHERE EXISTS ( SELECT * FROM Production.TransactionHistory AS th WHERE th.ProductID = p.ProductID HAVING COUNT_BIG(*) < 10 ); Unfortunately, this query produces an incorrect result (86 rows): The problem is that it lists products with no history rows, though the reasons are interesting.  The COUNT_BIG(*) in the EXISTS clause is a scalar aggregate (meaning there is no GROUP BY clause) and scalar aggregates always produce a value, even when the input is an empty set.  In the case of the COUNT aggregate, the result of aggregating the empty set is zero (the other standard aggregates produce a NULL).  To make the point really clear, let’s look at product 709, which happens to be one for which no history rows exist: -- Scalar aggregate SELECT COUNT_BIG(*) FROM Production.TransactionHistory AS th WHERE th.ProductID = 709;   -- Vector aggregate SELECT COUNT_BIG(*) FROM Production.TransactionHistory AS th WHERE th.ProductID = 709 GROUP BY th.ProductID; The estimated execution plans for these two statements are almost identical: You might expect the Stream Aggregate to have a Group By for the second statement, but this is not the case.  The query includes an equality comparison to a constant value (709), so all qualified rows are guaranteed to have the same value for ProductID and the Group By is optimized away. In fact there are some minor differences between the two plans (the first is auto-parameterized and qualifies for trivial plan, whereas the second is not auto-parameterized and requires cost-based optimization), but there is nothing to indicate that one is a scalar aggregate and the other is a vector aggregate.  This is something I would like to see exposed in show plan so I suggested it on Connect.  Anyway, the results of running the two queries show the difference at runtime: The scalar aggregate (no GROUP BY) returns a result of zero, whereas the vector aggregate (with a GROUP BY clause) returns nothing at all.  Returning to our EXISTS query, we could ‘fix’ it by changing the HAVING clause to reject rows where the scalar aggregate returns zero: SELECT p.Name FROM Production.Product AS p WHERE EXISTS ( SELECT * FROM Production.TransactionHistory AS th WHERE th.ProductID = p.ProductID HAVING COUNT_BIG(*) BETWEEN 1 AND 9 ); The query now returns the correct 23 rows: Unfortunately, the execution plan is less efficient now – it has an estimated cost of 0.78 compared to 0.33 for the earlier plans.  Let’s try adding a redundant GROUP BY instead of changing the HAVING clause: SELECT p.Name FROM Production.Product AS p WHERE EXISTS ( SELECT * FROM Production.TransactionHistory AS th WHERE th.ProductID = p.ProductID GROUP BY th.ProductID HAVING COUNT_BIG(*) < 10 ); Not only do we now get correct results (23 rows), this is the execution plan: I like to compare that plan to quantum physics: if you don’t find it shocking, you haven’t understood it properly :)  The simple addition of a redundant GROUP BY has resulted in the EXISTS form of the query being transformed into exactly the same optimal plan we found earlier.  What’s more, in SQL Server 2008 and later, we can replace the odd-looking GROUP BY with an explicit GROUP BY on the empty set: SELECT p.Name FROM Production.Product AS p WHERE EXISTS ( SELECT * FROM Production.TransactionHistory AS th WHERE th.ProductID = p.ProductID GROUP BY () HAVING COUNT_BIG(*) < 10 ); I offer that as an alternative because some people find it more intuitive (and it perhaps has more geek value too).  Whichever way you prefer, it’s rather satisfying to note that the result of the sub-query does not exist for a particular correlated value where a vector aggregate is used (the scalar COUNT aggregate always returns a value, even if zero, so it always ‘EXISTS’ regardless which ProductID is logically being evaluated). The following query forms also produce the optimal plan and correct results, so long as a vector aggregate is used (you can probably find more equivalent query forms): WHERE Clause SELECT p.Name FROM Production.Product AS p WHERE ( SELECT COUNT_BIG(*) FROM Production.TransactionHistory AS th WHERE th.ProductID = p.ProductID GROUP BY () ) < 10; APPLY SELECT p.Name FROM Production.Product AS p CROSS APPLY ( SELECT NULL FROM Production.TransactionHistory AS th WHERE th.ProductID = p.ProductID GROUP BY () HAVING COUNT_BIG(*) < 10 ) AS ca (dummy); FROM Clause SELECT q1.Name FROM ( SELECT p.Name, cnt = ( SELECT COUNT_BIG(*) FROM Production.TransactionHistory AS th WHERE th.ProductID = p.ProductID GROUP BY () ) FROM Production.Product AS p ) AS q1 WHERE q1.cnt < 10; This last example uses SUM(1) instead of COUNT and does not require a vector aggregate…you should be able to work out why :) SELECT q.Name FROM ( SELECT p.Name, cnt = ( SELECT SUM(1) FROM Production.TransactionHistory AS th WHERE th.ProductID = p.ProductID ) FROM Production.Product AS p ) AS q WHERE q.cnt < 10; The semantics of SQL aggregates are rather odd in places.  It definitely pays to get to know the rules, and to be careful to check whether your queries are using scalar or vector aggregates.  As we have seen, query plans do not show in which ‘mode’ an aggregate is running and getting it wrong can cause poor performance, wrong results, or both. © 2012 Paul White Twitter: @SQL_Kiwi email: [email protected]

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• Metro: Introduction to the WinJS ListView Control

  - by Stephen.Walther

  The goal of this blog entry is to provide a quick introduction to the ListView control – just the bare minimum that you need to know to start using the control. When building Metro style applications using JavaScript, the ListView control is the primary control that you use for displaying lists of items. For example, if you are building a product catalog app, then you can use the ListView control to display the list of products. The ListView control supports several advanced features that I plan to discuss in future blog entries. For example, you can group the items in a ListView, you can create master/details views with a ListView, and you can efficiently work with large sets of items with a ListView. In this blog entry, we’ll keep things simple and focus on displaying a list of products. There are three things that you need to do in order to display a list of items with a ListView: Create a data source Create an Item Template Declare the ListView Creating the ListView Data Source The first step is to create (or retrieve) the data that you want to display with the ListView. In most scenarios, you will want to bind a ListView to a WinJS.Binding.List object. The nice thing about the WinJS.Binding.List object is that it enables you to take a standard JavaScript array and convert the array into something that can be bound to the ListView. It doesn’t matter where the JavaScript array comes from. It could be a static array that you declare or you could retrieve the array as the result of an Ajax call to a remote server. The following JavaScript file – named products.js – contains a list of products which can be bound to a ListView. (function () { "use strict"; var products = new WinJS.Binding.List([ { name: "Milk", price: 2.44 }, { name: "Oranges", price: 1.99 }, { name: "Wine", price: 8.55 }, { name: "Apples", price: 2.44 }, { name: "Steak", price: 1.99 }, { name: "Eggs", price: 2.44 }, { name: "Mushrooms", price: 1.99 }, { name: "Yogurt", price: 2.44 }, { name: "Soup", price: 1.99 }, { name: "Cereal", price: 2.44 }, { name: "Pepsi", price: 1.99 } ]); WinJS.Namespace.define("ListViewDemos", { products: products }); })(); The products variable represents a WinJS.Binding.List object. This object is initialized with a plain-old JavaScript array which represents an array of products. To avoid polluting the global namespace, the code above uses the module pattern and exposes the products using a namespace. The list of products is exposed to the world as ListViewDemos.products. To learn more about the module pattern and namespaces in WinJS, see my earlier blog entry: http://stephenwalther.com/blog/archive/2012/02/22/metro-namespaces-and-modules.aspx Creating the ListView Item Template The ListView control does not know how to render anything. It doesn’t know how you want each list item to appear. To get the ListView control to render something useful, you must create an Item Template. Here’s what our template for rendering an individual product looks like: <div id="productTemplate" data-win-control="WinJS.Binding.Template"> <div class="product"> <span data-win-bind="innerText:name"></span> <span data-win-bind="innerText:price"></span> </div> </div> This template displays the product name and price from the data source. Normally, you will declare your template in the same file as you declare the ListView control. In our case, both the template and ListView are declared in the default.html file. To learn more about templates, see my earlier blog entry: http://stephenwalther.com/blog/archive/2012/02/27/metro-using-templates.aspx Declaring the ListView The final step is to declare the ListView control in a page. Here’s the markup for declaring a ListView: <div data-win-control="WinJS.UI.ListView" data-win-options="{ itemDataSource:ListViewDemos.products.dataSource, itemTemplate:select('#productTemplate') }"> </div> You declare a ListView by adding the data-win-control to an HTML DIV tag. The data-win-options attribute is used to set two properties of the ListView. The ListView is associated with its data source with the itemDataSource property. Notice that the data source is ListViewDemos.products.dataSource and not just ListViewDemos.products. You need to associate the ListView with the dataSoure property. The ListView is associated with its item template with the help of the itemTemplate property. The ID of the item template — #productTemplate – is used to select the template from the page. Here’s what the complete version of the default.html page looks like: <!DOCTYPE html> <html> <head> <meta charset="utf-8"> <title>ListViewDemos</title> <!-- WinJS references --> <link href="//Microsoft.WinJS.0.6/css/ui-dark.css" rel="stylesheet"> <script src="//Microsoft.WinJS.0.6/js/base.js"></script> <script src="//Microsoft.WinJS.0.6/js/ui.js"></script> <!-- ListViewDemos references --> <link href="/css/default.css" rel="stylesheet"> <script src="/js/default.js"></script> <script src="/js/products.js" type="text/javascript"></script> <style type="text/css"> .product { width: 200px; height: 100px; border: white solid 1px; } </style> </head> <body> <div id="productTemplate" data-win-control="WinJS.Binding.Template"> <div class="product"> <span data-win-bind="innerText:name"></span> <span data-win-bind="innerText:price"></span> </div> </div> <div data-win-control="WinJS.UI.ListView" data-win-options="{ itemDataSource:ListViewDemos.products.dataSource, itemTemplate:select('#productTemplate') }"> </div> </body> </html> Notice that the page above includes a reference to the products.js file: <script src=”/js/products.js” type=”text/javascript”></script> The page above also contains a Template control which contains the ListView item template. Finally, the page includes the declaration of the ListView control. Summary The goal of this blog entry was to describe the minimal set of steps which you must complete to use the WinJS ListView control to display a simple list of items. You learned how to create a data source, declare an item template, and declare a ListView control.

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• UppercuT &ndash; Custom Extensions Now With PowerShell and Ruby

  - by Robz / Fervent Coder

  Arguably, one of the most powerful features of UppercuT (UC) is the ability to extend any step of the build process with a pre, post, or replace hook. This customization is done in a separate location from the build so you can upgrade without wondering if you broke the build. There is a hook before each step of the build has run. There is a hook after. And back to power again, there is a replacement hook. If you don’t like what the step is doing and/or you want to replace it’s entire functionality, you just drop a custom replacement extension and UppercuT will perform the custom step instead. Up until recently all custom hooks had to be written in NAnt. Now they are a little sweeter because you no longer need to use NAnt to extend UC if you don’t want to. You can use PowerShell. Or Ruby.   Let that sink in for a moment. You don’t have to even need to interact with NAnt at all now. Extension Points On the wiki, all of the extension points are shown. The basic idea is that you would put whatever customization you are doing in a separate folder named build.custom. Each step Let’s take a look at all we can customize: The start point is default.build. It calls build.custom/default.pre.build if it exists, then it runs build/default.build (normal tasks) OR build.custom/default.replace.build if it exists, and finally build.custom/default.post.build if it exists. Every step below runs with the same extension points but changes on the file name it is looking for. NOTE: If you include default.replace.build, nothing else will run because everything is called from default.build.    * policyChecks.step    * versionBuilder.step NOTE: If you include build.custom/versionBuilder.replace.step, the items below will not run.      - svn.step, tfs.step, or git.step (the custom tasks for these need to go in build.custom/versioners)    * generateBuildInfo.step    * compile.step    * environmentBuilder.step    * analyze.step NOTE: If you include build.custom/analyze.replace.step, the items below will not run.      - test.step (the custom tasks for this need to go in build.custom/analyzers) NOTE: If you include build.custom/analyzers/test.replace.step, the items below will not run.        + mbunit2.step, gallio.step, or nunit.step (the custom tasks for these need to go in build.custom/analyzers)      - ncover.step (the custom tasks for this need to go in build.custom/analyzers)      - ndepend.step (the custom tasks for this need to go in build.custom/analyzers)      - moma.step (the custom tasks for this need to go in build.custom/analyzers)    * package.step NOTE: If you include build.custom/package.replace.step, the items below will not run.      - deploymentBuilder.step Customize UppercuT Builds With PowerShell UppercuT can now be extended with PowerShell (PS). To customize any extension point with PS, just add .ps1 to the end of the file name and write your custom tasks in PowerShell. If you are not signing your scripts you will need to change a setting in the UppercuT.config file. This does impose a security risk, because this allows PS to now run any PS script. This setting stays that way on ANY machine that runs the build until manually changed by someone. I’m not responsible if you mess up your machine or anyone else’s by doing this. You’ve been warned. Now that you are fully aware of any security holes you may open and are okay with that, let’s move on. Let’s create a file called default.replace.build.ps1 in the build.custom folder. Open that file in notepad and let’s add this to it: write-host "hello - I'm a custom task written in Powershell!" Now, let’s run build.bat. You could get some PSake action going here. I won’t dive into that in this post though. Customize UppercuT Builds With Ruby If you want to customize any extension point with Ruby, just add .rb to the end of the file name and write your custom tasks in Ruby.  Let’s write a custom ruby task for UC. If you were thinking it would be the same as the one we just wrote for PS, you’d be right! In the build.custom folder, lets create a file called default.replace.build.rb. Open that file in notepad and let’s put this in there: puts "I'm a custom ruby task!" Now, let’s run build.bat again. That’s chunky bacon. UppercuT and Albacore.NET Just for fun, I wanted to see if I could replace the compile.step with a Rake task. Not just any rake task, Albacore’s msbuild task. Albacore is a suite of rake tasks brought about by Derick Bailey to make building .NET with Rake easier. It has quite a bit of support with developers that are using Rake to build code. In my build.custom folder, I drop a compile.replace.step.rb. I also put in a separate file that will contain my Albacore rake task and I call that compile.rb. What are the contents of compile.replace.step.rb? rake = 'rake' arguments= '-f ' + Dir.pwd + '/../build.custom/compile.rb' #puts "Calling #{rake} " + arguments system("#{rake} " + arguments) Since the custom extensions call ruby, we have to shell back out and call rake. That’s what we are doing here. We also realize that ruby is called from the build folder, so we need to back out and dive into the build.custom folder to find the file that is technically next to us. What are the contents of compile.rb? require 'rubygems' require 'fileutils' require 'albacore' task :default => [:compile] puts "Using Ruby to compile UppercuT with Albacore Tasks" desc 'Compile the source' msbuild :compile do |msb| msb.properties = { :configuration => :Release, :outputpath => '../../build_output/UppercuT' } msb.targets [:clean, :build] msb.verbosity = "quiet" msb.path_to_command = 'c:/Windows/Microsoft.NET/Framework/v3.5/MSBuild.exe' msb.solution = '../uppercut.sln' end We are using the msbuild task here. We change the output path to the build_output/UppercuT folder. The output path has “../../” because this is based on every project. We could grab the current directory and then point the task specifically to a folder if we have projects that are at different levels. We want the verbosity to be quiet so we set that as well. So what kind of output do you get for this? Let’s run build.bat custom_tasks_replace:      [echo] Running custom tasks instead of normal tasks if C:\code\uppercut\build\..\build.custom\compile.replace.step exists.      [exec] (in C:/code/uppercut/build)      [exec] Using Ruby to compile UppercuT with Albacore Tasks      [exec] Microsoft (R) Build Engine Version 3.5.30729.4926      [exec] [Microsoft .NET Framework, Version 2.0.50727.4927]      [exec] Copyright (C) Microsoft Corporation 2007. All rights reserved. If you think this is awesome, you’d be right!   With this knowledge you shall build.

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• Master Data Management and Cloud Computing

  - by david.butler(at)oracle.com

  Cloud Computing is all the rage these days. There are many reasons why this is so. But like its predecessor, Service Oriented Architecture, it can fall on hard times if the underlying data is left unmanaged. Master Data Management is the perfect Cloud companion. It can materially increase the chances for successful Cloud initiatives. In this blog, I'll review the nature of the Cloud and show how MDM fits in.   Here's the National Institute of Standards and Technology Cloud definition: •          Cloud computing is a model for enabling convenient, on-demand network access to a shared pool of configurable computing resources that can be rapidly provisioned and released with minimal management effort or service provider interaction.   Cloud architectures have three main layers: applications or Software as a Service (SaaS), Platforms as a Service (PaaS), and Infrastructure as a Service (IaaS). SaaS generally refers to applications that are delivered to end-users over the Internet. Oracle CRM On Demand is an example of a SaaS application. Today there are hundreds of SaaS providers covering a wide variety of applications including Salesforce.com, Workday, and Netsuite. Oracle MDM applications are located in this layer of Oracle's On Demand enterprise Cloud platform. We call it Master Data as a Service (MDaaS). PaaS generally refers to an application deployment platform delivered as a service. They are often built on a grid computing architecture and include database and middleware. Oracle Fusion Middleware is in this category and includes the SOA and Data Integration products used to connect SaaS applications including MDM. Finally, IaaS generally refers to computing hardware (servers, storage and network) delivered as a service.  This typically includes the associated software as well: operating systems, virtualization, clustering, etc.    Cloud Computing benefits are compelling for a large number of organizations. These include significant cost savings, increased flexibility, and fast deployments. Cost advantages include paying for just what you use. This is especially critical for organizations with variable or seasonal usage. Companies don't have to invest to support peak computing periods. Costs are also more predictable and controllable. Increased agility includes access to the latest technology and experts without making significant up front investments.   While Cloud Computing is certainly very alluring with a clear value proposition, it is not without its challenges. An IDC survey of 244 IT executives/CIOs and their line-of-business (LOB) colleagues identified a number of issues:   Security - 74% identified security as an issue involving data privacy and resource access control. Integration - 61% found that it is hard to integrate Cloud Apps with in-house applications. Operational Costs - 50% are worried that On Demand will actually cost more given the impact of poor data quality on the rest of the enterprise. Compliance - 49% felt that compliance with required regulatory, legal and general industry requirements (such as PCI, HIPAA and Sarbanes-Oxley) would be a major issue. When control is lost, the ability of a provider to directly manage how and where data is deployed, used and destroyed is negatively impacted.  There are others, but I singled out these four top issues because Master Data Management, properly incorporated into a Cloud Computing infrastructure, can significantly ameliorate all of these problems. Cloud Computing can literally rain raw data across the enterprise.   According to fellow blogger, Mike Ferguson, "the fracturing of data caused by the adoption of cloud computing raises the importance of MDM in keeping disparate data synchronized."   David Linthicum, CTO Blue Mountain Labs blogs that "the lack of MDM will become more of an issue as cloud computing rises. We're moving from complex federated on-premise systems, to complex federated on-premise and cloud-delivered systems."    Left unmanaged, non-standard, inconsistent, ungoverned data with questionable quality can pollute analytical systems, increase operational costs, and reduce the ROI in Cloud and On-Premise applications. As cloud computing becomes more relevant, and more data, applications, services, and processes are moved out to cloud computing platforms, the need for MDM becomes ever more important. Oracle's MDM suite is designed to deal with all four of the above Cloud issues listed in the IDC survey.   Security - MDM manages all master data attribute privacy and resource access control issues. Integration - MDM pre-integrates Cloud Apps with each other and with On Premise applications at the data level. Operational Costs - MDM significantly reduces operational costs by increasing data quality, thereby improving enterprise business processes efficiency. Compliance - MDM, with its built in Data Governance capabilities, insures that the data is governed according to organizational standards. This facilitates rapid and accurate reporting for compliance purposes. Oracle MDM creates governed high quality master data. A unified cleansed and standardized data view is produced. The Oracle Customer Hub creates a single view of the customer. The Oracle Product Hub creates high quality product data designed to support all go-to-market processes. Oracle Supplier Hub dramatically reduces the chances of 'supplier exceptions'. Oracle Site Hub masters locations. And Oracle Hyperion Data Relationship Management masters financial reference data and manages enterprise hierarchies across operational areas from ERP to EPM and CRM to SCM. Oracle Fusion Middleware connects Cloud and On Premise applications to MDM Hubs and brings high quality master data to your enterprise business processes.   An independent analyst once said "Poor data quality is like dirt on the windshield. You may be able to drive for a long time with slowly degrading vision, but at some point, you either have to stop and clear the windshield or risk everything."  Cloud Computing has the potential to significantly degrade data quality across the enterprise over time. Deploying a Master Data Management solution prior to or in conjunction with a move to the Cloud can insure that the data flowing into the enterprise from the Cloud is clean and governed. This will in turn insure that expected returns on the investment in Cloud Computing will be realized.       Oracle MDM has proven its metal in this area and has the customers to back that up. In fact, I will be hosting a webcast on Tuesday, April 10th at 10 am PT with one of our top Cloud customers, the Church Pension Group. They have moved all mainline applications to a hosted model and use Oracle MDM to insure the master data is managed and cleansed before it is propagated to other cloud and internal systems. I invite you join Martin Hossfeld, VP, IT Operations, and Danette Patterson, Enterprise Data Manager as they review business drivers for MDM and hosted applications, how they did it, the benefits achieved, and lessons learned. You can register for this free webcast here.  Hope to see you there.

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• DTracing a PHPUnit Test: Looking at Functional Programming

  - by cj

  Here's a quick example of using DTrace Dynamic Tracing to work out what a PHP code base does. I was reading the article Functional Programming in PHP by Patkos Csaba and wondering how efficient this stype of programming is. I thought this would be a good time to fire up DTrace and see what is going on. Since DTrace is "always available" even in production machines (once PHP is compiled with --enable-dtrace), this was easy to do. I have Oracle Linux with the UEK3 kernel and PHP 5.5 with DTrace static probes enabled, as described in DTrace PHP Using Oracle Linux 'playground' Pre-Built Packages I installed the Functional Programming sample code and Sebastian Bergmann's PHPUnit. Although PHPUnit is included in the Functional Programming example, I found it easier to separately download and use its phar file: cd ~/Desktop wget -O master.zip https://github.com/tutsplus/functional-programming-in-php/archive/master.zip wget https://phar.phpunit.de/phpunit.phar unzip master.zip I created a DTrace D script functree.d: #pragma D option quiet self int indent; BEGIN { topfunc = $1; } php$target:::function-entry /copyinstr(arg0) == topfunc/ { self->follow = 1; } php$target:::function-entry /self->follow/ { self->indent += 2; printf("%*s %s%s%s\n", self->indent, "->", arg3?copyinstr(arg3):"", arg4?copyinstr(arg4):"", copyinstr(arg0)); } php$target:::function-return /self->follow/ { printf("%*s %s%s%s\n", self->indent, "<-", arg3?copyinstr(arg3):"", arg4?copyinstr(arg4):"", copyinstr(arg0)); self->indent -= 2; } php$target:::function-return /copyinstr(arg0) == topfunc/ { self->follow = 0; } This prints a PHP script function call tree starting from a given PHP function name. This name is passed as a parameter to DTrace, and assigned to the variable topfunc when the DTrace script starts. With this D script, choose a PHP function that isn't recursive, or modify the script to set self->follow = 0 only when all calls to that function have unwound. From looking at the sample FunSets.php code and its PHPUnit test driver FunSetsTest.php, I settled on one test function to trace: function testUnionContainsAllElements() { ... } I invoked DTrace to trace function calls invoked by this test with # dtrace -s ./functree.d -c 'php phpunit.phar \ /home/cjones/Desktop/functional-programming-in-php-master/FunSets/Tests/FunSetsTest.php' \ '"testUnionContainsAllElements"' The core of this command is a call to PHP to run PHPUnit on the FunSetsTest.php script. Outside that, DTrace is called and the PID of PHP is passed to the D script $target variable so the probes fire just for this invocation of PHP. Note the quoting around the PHP function name passed to DTrace. The parameter must have double quotes included so DTrace knows it is a string. The output is: PHPUnit 3.7.28 by Sebastian Bergmann. ......-> FunSetsTest::testUnionContainsAllElements -> FunSets::singletonSet <- FunSets::singletonSet -> FunSets::singletonSet <- FunSets::singletonSet -> FunSets::union <- FunSets::union -> FunSets::contains -> FunSets::{closure} -> FunSets::contains -> FunSets::{closure} <- FunSets::{closure} <- FunSets::contains <- FunSets::{closure} <- FunSets::contains -> PHPUnit_Framework_Assert::assertTrue -> PHPUnit_Framework_Assert::isTrue <- PHPUnit_Framework_Assert::isTrue -> PHPUnit_Framework_Assert::assertThat -> PHPUnit_Framework_Constraint::count <- PHPUnit_Framework_Constraint::count -> PHPUnit_Framework_Constraint::evaluate -> PHPUnit_Framework_Constraint_IsTrue::matches <- PHPUnit_Framework_Constraint_IsTrue::matches <- PHPUnit_Framework_Constraint::evaluate <- PHPUnit_Framework_Assert::assertThat <- PHPUnit_Framework_Assert::assertTrue -> FunSets::contains -> FunSets::{closure} -> FunSets::contains -> FunSets::{closure} <- FunSets::{closure} <- FunSets::contains -> FunSets::contains -> FunSets::{closure} <- FunSets::{closure} <- FunSets::contains <- FunSets::{closure} <- FunSets::contains -> PHPUnit_Framework_Assert::assertTrue -> PHPUnit_Framework_Assert::isTrue <- PHPUnit_Framework_Assert::isTrue -> PHPUnit_Framework_Assert::assertThat -> PHPUnit_Framework_Constraint::count <- PHPUnit_Framework_Constraint::count -> PHPUnit_Framework_Constraint::evaluate -> PHPUnit_Framework_Constraint_IsTrue::matches <- PHPUnit_Framework_Constraint_IsTrue::matches <- PHPUnit_Framework_Constraint::evaluate <- PHPUnit_Framework_Assert::assertThat <- PHPUnit_Framework_Assert::assertTrue -> FunSets::contains -> FunSets::{closure} -> FunSets::contains -> FunSets::{closure} <- FunSets::{closure} <- FunSets::contains -> FunSets::contains -> FunSets::{closure} <- FunSets::{closure} <- FunSets::contains <- FunSets::{closure} <- FunSets::contains -> PHPUnit_Framework_Assert::assertFalse -> PHPUnit_Framework_Assert::isFalse -> {closure} -> main <- main <- {closure} <- PHPUnit_Framework_Assert::isFalse -> PHPUnit_Framework_Assert::assertThat -> PHPUnit_Framework_Constraint::count <- PHPUnit_Framework_Constraint::count -> PHPUnit_Framework_Constraint::evaluate -> PHPUnit_Framework_Constraint_IsFalse::matches <- PHPUnit_Framework_Constraint_IsFalse::matches <- PHPUnit_Framework_Constraint::evaluate <- PHPUnit_Framework_Assert::assertThat <- PHPUnit_Framework_Assert::assertFalse <- FunSetsTest::testUnionContainsAllElements ... Time: 1.85 seconds, Memory: 3.75Mb OK (9 tests, 23 assertions) The periods correspond to the successful tests before and after (and from) the test I was tracing. You can see the function entry ("->") and return ("<-") points. Cross checking with the testUnionContainsAllElements() source code confirms the two singletonSet() calls, one union() call, two assertTrue() calls and finally an assertFalse() call. These assertions have a contains() call as a parameter, so contains() is called before the PHPUnit assertion functions are run. You can see contains() being called recursively, and how the closures are invoked. If you want to focus on the application logic and suppress the PHPUnit function trace, you could turn off tracing when assertions are being checked by adding D clauses checking the entry and exit of assertFalse() and assertTrue(). But if you want to see all of PHPUnit's code flow, you can modify the functree.d code that sets and unsets self-follow, and instead change it to toggle the variable in request-startup and request-shutdown probes: php$target:::request-startup { self->follow = 1 } php$target:::request-shutdown { self->follow = 0 } Be prepared for a large amount of output!

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• Segfault when iterating over a map<string, string> and drawing its contents using SDL_TTF

  - by Michael Stahre

  I'm not entirely sure this question belongs on gamedev.stackexchange, but I'm technically working on a game and working with SDL, so it might not be entirely offtopic. I've written a class called DebugText. The point of the class is to have a nice way of printing values of variables to the game screen. The idea is to call SetDebugText() with the variables in question every time they change or, as is currently the case, every time the game's Update() is called. The issue is that when iterating over the map that contains my variables and their latest updated values, I get segfaults. See the comments in DrawDebugText() below, it specifies where the error happens. I've tried splitting the calls to it-first and it-second into separate lines and found that the problem doesn't always happen when calling it-first. It alters between it-first and it-second. I can't find a pattern. It doesn't fail on every call to DrawDebugText() either. It might fail on the third time DrawDebugText() is called, or it might fail on the fourth. Class header: #ifndef CLIENT_DEBUGTEXT_H #define CLIENT_DEBUGTEXT_H #include <Map> #include <Math.h> #include <sstream> #include <SDL.h> #include <SDL_ttf.h> #include "vector2.h" using std::string; using std::stringstream; using std::map; using std::pair; using game::Vector2; namespace game { class DebugText { private: TTF_Font* debug_text_font; map<string, string>* debug_text_list; public: void SetDebugText(string var, bool value); void SetDebugText(string var, float value); void SetDebugText(string var, int value); void SetDebugText(string var, Vector2 value); void SetDebugText(string var, string value); int DrawDebugText(SDL_Surface*, SDL_Rect*); void InitDebugText(); void Clear(); }; } #endif Class source file: #include "debugtext.h" namespace game { // Copypasta function for handling the toString conversion template <class T> inline string to_string (const T& t) { stringstream ss (stringstream::in | stringstream::out); ss << t; return ss.str(); } // Initializes SDL_TTF and sets its font void DebugText::InitDebugText() { if(TTF_WasInit()) TTF_Quit(); TTF_Init(); debug_text_font = TTF_OpenFont("LiberationSans-Regular.ttf", 16); TTF_SetFontStyle(debug_text_font, TTF_STYLE_NORMAL); } // Iterates over the current debug_text_list and draws every element on the screen. // After drawing with SDL you need to get a rect specifying the area on the screen that was changed and tell SDL that this part of the screen needs to be updated. this is done in the game's Draw() function // This function sets rects_to_update to the new list of rects provided by all of the surfaces and returns the number of rects in the list. These two parameters are used in Draw() when calling on SDL_UpdateRects(), which takes an SDL_Rect* and a list length int DebugText::DrawDebugText(SDL_Surface* screen, SDL_Rect* rects_to_update) { if(debug_text_list == NULL) return 0; if(!TTF_WasInit()) InitDebugText(); rects_to_update = NULL; // Specifying the font color SDL_Color font_color = {0xff, 0x00, 0x00, 0x00}; // r, g, b, unused int row_count = 0; string line; // The iterator variable map<string, string>::iterator it; // Gets the iterator and iterates over it for(it = debug_text_list->begin(); it != debug_text_list->end(); it++) { // Takes the first value (the name of the variable) and the second value (the value of the parameter in string form) //---------THIS LINE GIVES ME SEGFAULTS----- line = it->first + ": " + it->second; //------------------------------------------ // Creates a surface with the text on it that in turn can be rendered to the screen itself later SDL_Surface* debug_surface = TTF_RenderText_Solid(debug_text_font, line.c_str(), font_color); if(debug_surface == NULL) { // A standard check for errors fprintf(stderr, "Error: %s", TTF_GetError()); return NULL; } else { // If SDL_TTF did its job right, then we now set a destination rect row_count++; SDL_Rect dstrect = {5, 5, 0, 0}; // x, y, w, h dstrect.x = 20; dstrect.y = 20*row_count; // Draws the surface with the text on it to the screen int res = SDL_BlitSurface(debug_surface,NULL,screen,&dstrect); if(res != 0) { //Just an error check fprintf(stderr, "Error: %s", SDL_GetError()); return NULL; } // Creates a new rect to specify the area that needs to be updated with SDL_Rect* new_rect_to_update = (SDL_Rect*) malloc(sizeof(SDL_Rect)); new_rect_to_update->h = debug_surface->h; new_rect_to_update->w = debug_surface->w; new_rect_to_update->x = dstrect.x; new_rect_to_update->y = dstrect.y; // Just freeing the surface since it isn't necessary anymore SDL_FreeSurface(debug_surface); // Creates a new list of rects with room for the new rect SDL_Rect* newtemp = (SDL_Rect*) malloc(row_count*sizeof(SDL_Rect)); // Copies the data from the old list of rects to the new one memcpy(newtemp, rects_to_update, (row_count-1)*sizeof(SDL_Rect)); // Adds the new rect to the new list newtemp[row_count-1] = *new_rect_to_update; // Frees the memory used by the old list free(rects_to_update); // And finally redirects the pointer to the old list to the new list rects_to_update = newtemp; newtemp = NULL; } } // When the entire map has been iterated over, return the number of lines that were drawn, ie. the number of rects in the returned rect list return row_count; } // The SetDebugText used by all the SetDebugText overloads // Takes two strings, inserts them into the map as a pair void DebugText::SetDebugText(string var, string value) { if (debug_text_list == NULL) { debug_text_list = new map<string, string>(); } debug_text_list->erase(var); debug_text_list->insert(pair<string, string>(var, value)); } // Writes the bool to a string and calls SetDebugText(string, string) void DebugText::SetDebugText(string var, bool value) { string result; if (value) result = "True"; else result = "False"; SetDebugText(var, result); } // Does the same thing, but uses to_string() to convert the float void DebugText::SetDebugText(string var, float value) { SetDebugText(var, to_string(value)); } // Same as above, but int void DebugText::SetDebugText(string var, int value) { SetDebugText(var, to_string(value)); } // Vector2 is a struct of my own making. It contains the two float vars x and y void DebugText::SetDebugText(string var, Vector2 value) { SetDebugText(var + ".x", to_string(value.x)); SetDebugText(var + ".y", to_string(value.y)); } // Empties the list. I don't actually use this in my code. Shame on me for writing something I don't use. void DebugText::Clear() { if(debug_text_list != NULL) debug_text_list->clear(); } }

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• Granular Clipboard Control in Oracle IRM

  - by martin.abrahams

  One of the main leak prevention controls that customers are looking for is clipboard control. After all, there is little point in controlling access to a document if authorised users can simply make unprotected copies by use of the cut and paste mechanism. Oddly, for such a fundamental requirement, many solutions only offer very simplistic clipboard control - and require the customer to make an awkward choice between usability and security. In many cases, clipboard control is simply an ON-OFF option. By turning the clipboard OFF, you disable one of the most valuable edit functions known to man. Try working for any length of time without copying and pasting, and you'll soon appreciate how valuable that function is. Worse, some solutions disable the clipboard completely - not just for the protected document but for all of the various applications you have open at the time. Normal service is only resumed when you close the protected document. In this way, policy enforcement bleeds out of the particular assets you need to protect and interferes with the entire user experience. On the other hand, turning the clipboard ON satisfies a fundamental usability requirement - but also makes it really easy for users to create unprotected copies of sensitive information, maliciously or otherwise. All they need to do is paste into another document. If creating unprotected copies is this simple, you have to question how much you are really gaining by applying protection at all. You may not be allowed to edit, forward, or print the protected asset, but all you need to do is create a copy and work with that instead. And that activity would not be tracked in any way. So, a simple ON-OFF control creates a real tension between usability and security. If you are only using IRM on a small scale, perhaps security can outweigh usability - the business can put up with the restriction if it only applies to a handful of important documents. But try extending protection to large numbers of documents and large user communities, and the restriction rapidly becomes really unwelcome. I am aware of one solution that takes a different tack. Rather than disable the clipboard, pasting is always permitted, but protection is automatically applied to any document that you paste into. At first glance, this sounds great - protection travels with the content. However, at any scale this model may not be so appealing once you've had to deal with support calls from users who have accidentally applied protection to documents that really don't need it - which would be all too easily done. This may help control leakage, but it also pollutes the system with documents that have policies applied with no obvious rhyme or reason, and it can seriously inconvenience the business by making non-sensitive documents difficult to access. And what policy applies if you paste some protected content into an already protected document? Which policy applies? There are no prizes for guessing that Oracle IRM takes a rather different approach. The Oracle IRM Approach Oracle IRM offers a spectrum of clipboard controls between the extremes of ON and OFF, and it leverages the classification-based rights model to give granular control that satisfies both security and usability needs. Firstly, we take it for granted that if you have EDIT rights, of course you can use the clipboard within a given document. Why would we force you to retype a piece of content that you want to move from HERE... to HERE...? If the pasted content remains in the same document, it is equally well protected whether it be at the beginning, middle, or end - or all three. So, the first point is that Oracle IRM always enables the clipboard if you have the right to edit the file. Secondly, whether we enable or disable the clipboard, we only affect the protected document. That is, you can continue to use the clipboard in the usual way for unprotected documents and applications regardless of whether the clipboard is enabled or disabled for the protected document(s). And if you have multiple protected documents open, each may have the clipboard enabled or disabled independently, according to whether you have Edit rights for each. So, even for the simplest cases - the ON-OFF cases - Oracle IRM adds value by containing the effect to the protected documents rather than to the whole desktop environment. Now to the granular options between ON and OFF. Thanks to our classification model, we can define rights that enable pasting between documents in the same classification - ie. between documents that are protected by the same policy. So, if you are working on this month's financial report and you want to pull some data from last month's report, you can simply cut and paste between the two documents. The two documents are classified the same way, subject to the same policy, so the content is equally safe in both documents. However, if you try to paste the same data into an unprotected document or a document in a different classification, you can be prevented. Thus, the control balances legitimate user requirements to allow pasting with legitimate information security concerns to keep data protected. We can take this further. You may have the right to paste between related classifications of document. So, the CFO might want to copy some financial data into a board document, where the two documents are sealed to different classifications. The CFO's rights may well allow this, as it is a reasonable thing for a CFO to want to do. But policy might prevent the CFO from copying the same data into a classification that is accessible to external parties. The above option, to copy between classifications, may be for specific classifications or open-ended. That is, your rights might enable you to go from A to B but not to C, or you might be allowed to paste to any classification subject to your EDIT rights. As for so many features of Oracle IRM, our classification-based rights model makes this type of granular control really easy to manage - you simply define that pasting is permitted between classifications A and B, but omit C. Or you might define that pasting is permitted between all classifications, but not to unprotected locations. The classification model enables millions of documents to be controlled by a few such rules. Finally, you MIGHT have the option to paste anywhere - such that unprotected copies may be created. This is rare, but a legitimate configuration for some users, some use cases, and some classifications - but not something that you have to permit simply because the alternative is too restrictive. As always, these rights are defined in user roles - so different users are subject to different clipboard controls as required in different classifications. So, where most solutions offer just two clipboard options - ON-OFF or ON-but-encrypt-everything-you-touch - Oracle IRM offers real granularity that leverages our classification model. Indeed, I believe it is the lack of a classification model that makes such granularity impractical for other IRM solutions, because the matrix of rules for controlling pasting would be impossible to manage - there are so many documents to consider, and more are being created all the time.

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• C#/.NET Little Wonders: Using &lsquo;default&rsquo; to Get Default Values

  - by James Michael Hare

  Once again, in this series of posts I look at the parts of the .NET Framework that may seem trivial, but can help improve your code by making it easier to write and maintain. The index of all my past little wonders posts can be found here. Today’s little wonder is another of those small items that can help a lot in certain situations, especially when writing generics.  In particular, it is useful in determining what the default value of a given type would be. The Problem: what’s the default value for a generic type? There comes a time when you’re writing generic code where you may want to set an item of a given generic type.  Seems simple enough, right?  We’ll let’s see! Let’s say we want to query a Dictionary<TKey, TValue> for a given key and get back the value, but if the key doesn’t exist, we’d like a default value instead of throwing an exception. So, for example, we might have a the following dictionary defined: 1: var lookup = new Dictionary<int, string> 2: { 3: { 1, "Apple" }, 4: { 2, "Orange" }, 5: { 3, "Banana" }, 6: { 4, "Pear" }, 7: { 9, "Peach" } 8: }; And using those definitions, perhaps we want to do something like this: 1: // assume a default 2: string value = "Unknown"; 3:  4: // if the item exists in dictionary, get its value 5: if (lookup.ContainsKey(5)) 6: { 7: value = lookup[5]; 8: } But that’s inefficient, because then we’re double-hashing (once for ContainsKey() and once for the indexer).  Well, to avoid the double-hashing, we could use TryGetValue() instead: 1: string value; 2:  3: // if key exists, value will be put in value, if not default it 4: if (!lookup.TryGetValue(5, out value)) 5: { 6: value = "Unknown"; 7: } But the “flow” of using of TryGetValue() can get clunky at times when you just want to assign either the value or a default to a variable.  Essentially it’s 3-ish lines (depending on formatting) for 1 assignment.  So perhaps instead we’d like to write an extension method to support a cleaner interface that will return a default if the item isn’t found: 1: public static class DictionaryExtensions 2: { 3: public static TValue GetValueOrDefault<TKey, TValue>(this Dictionary<TKey, TValue> dict, 4: TKey key, TValue defaultIfNotFound) 5: { 6: TValue value; 7:  8: // value will be the result or the default for TValue 9: if (!dict.TryGetValue(key, out value)) 10: { 11: value = defaultIfNotFound; 12: } 13:  14: return value; 15: } 16: } 17:  So this creates an extension method on Dictionary<TKey, TValue> that will attempt to get a value using the given key, and will return the defaultIfNotFound as a stand-in if the key does not exist. This code compiles, fine, but what if we would like to go one step further and allow them to specify a default if not found, or accept the default for the type?  Obviously, we could overload the method to take the default or not, but that would be duplicated code and a bit heavy for just specifying a default.  It seems reasonable that we could set the not found value to be either the default for the type, or the specified value. So what if we defaulted the type to null? 1: public static TValue GetValueOrDefault<TKey, TValue>(this Dictionary<TKey, TValue> dict, 2: TKey key, TValue defaultIfNotFound = null) // ... No, this won’t work, because only reference types (and Nullable<T> wrapped types due to syntactical sugar) can be assigned to null.  So what about a calling parameterless constructor? 1: public static TValue GetValueOrDefault<TKey, TValue>(this Dictionary<TKey, TValue> dict, 2: TKey key, TValue defaultIfNotFound = new TValue()) // ... No, this won’t work either for several reasons.  First, we’d expect a reference type to return null, not an “empty” instance.  Secondly, not all reference types have a parameter-less constructor (string for example does not).  And finally, a constructor cannot be determined at compile-time, while default values can. The Solution: default(T) – returns the default value for type T Many of us know the default keyword for its uses in switch statements as the default case.  But it has another use as well: it can return us the default value for a given type.  And since it generates the same defaults that default field initialization uses, it can be determined at compile-time as well. For example: 1: var x = default(int); // x is 0 2:  3: var y = default(bool); // y is false 4:  5: var z = default(string); // z is null 6:  7: var t = default(TimeSpan); // t is a TimeSpan with Ticks == 0 8:  9: var n = default(int?); // n is a Nullable<int> with HasValue == false Notice that for numeric types the default is 0, and for reference types the default is null.  In addition, for struct types, the value is a default-constructed struct – which simply means a struct where every field has their default value (hence 0 Ticks for TimeSpan, etc.). So using this, we could modify our code to this: 1: public static class DictionaryExtensions 2: { 3: public static TValue GetValueOrDefault<TKey, TValue>(this Dictionary<TKey, TValue> dict, 4: TKey key, TValue defaultIfNotFound = default(TValue)) 5: { 6: TValue value; 7:  8: // value will be the result or the default for TValue 9: if (!dict.TryGetValue(key, out value)) 10: { 11: value = defaultIfNotFound; 12: } 13:  14: return value; 15: } 16: } Now, if defaultIfNotFound is unspecified, it will use default(TValue) which will be the default value for whatever value type the dictionary holds.  So let’s consider how we could use this: 1: lookup.GetValueOrDefault(1); // returns “Apple” 2:  3: lookup.GetValueOrDefault(5); // returns null 4:  5: lookup.GetValueOrDefault(5, “Unknown”); // returns “Unknown” 6:  Again, do not confuse a parameter-less constructor with the default value for a type.  Remember that the default value for any type is the compile-time default for any instance of that type (0 for numeric, false for bool, null for reference types, and struct will all default fields for struct).  Consider the difference: 1: // both zero 2: int i1 = default(int); 3: int i2 = new int(); 4:  5: // both “zeroed” structs 6: var dt1 = default(DateTime); 7: var dt2 = new DateTime(); 8:  9: // sb1 is null, sb2 is an “empty” string builder 10: var sb1 = default(StringBuilder()); 11: var sb2 = new StringBuilder(); So in the above code, notice that the value types all resolve the same whether using default or parameter-less construction.  This is because a value type is never null (even Nullable<T> wrapped types are never “null” in a reference sense), they will just by default contain fields with all default values. However, for reference types, the default is null and not a constructed instance.  Also it should be noted that not all classes have parameter-less constructors (string, for instance, doesn’t have one – and doesn’t need one). Summary Whenever you need to get the default value for a type, especially a generic type, consider using the default keyword.  This handy word will give you the default value for the given type at compile-time, which can then be used for initialization, optional parameters, etc. Technorati Tags: C#,CSharp,.NET,Little Wonders,default

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• What's up with LDoms: Part 2 - Creating a first, simple guest

  - by Stefan Hinker

  Welcome back! In the first part, we discussed the basic concepts of LDoms and how to configure a simple control domain.  We saw how resources were put aside for guest systems and what infrastructure we need for them.  With that, we are now ready to create a first, very simple guest domain.  In this first example, we'll keep things very simple.  Later on, we'll have a detailed look at things like sizing, IO redundancy, other types of IO as well as security. For now,let's start with this very simple guest.  It'll have one core's worth of CPU, one crypto unit, 8GB of RAM, a single boot disk and one network port.  CPU and RAM are easy.  The network port we'll create by attaching a virtual network port to the vswitch we created in the primary domain.  This is very much like plugging a cable into a computer system on one end and a network switch on the other.  For the boot disk, we'll need two things: A physical piece of storage to hold the data - this is called the backend device in LDoms speak.  And then a mapping between that storage and the guest domain, giving it access to that virtual disk.  For this example, we'll use a ZFS volume for the backend.  We'll discuss what other options there are for this and how to chose the right one in a later article.  Here we go: root@sun # ldm create mars root@sun # ldm set-vcpu 8 mars root@sun # ldm set-mau 1 mars root@sun # ldm set-memory 8g mars root@sun # zfs create rpool/guests root@sun # zfs create -V 32g rpool/guests/mars.bootdisk root@sun # ldm add-vdsdev /dev/zvol/dsk/rpool/guests/mars.bootdisk \ mars.root@primary-vds root@sun # ldm add-vdisk root mars.root@primary-vds mars root@sun # ldm add-vnet net0 switch-primary mars That's all, mars is now ready to power on.  There are just three commands between us and the OK prompt of mars:  We have to "bind" the domain, start it and connect to its console.  Binding is the process where the hypervisor actually puts all the pieces that we've configured together.  If we made a mistake, binding is where we'll be told (starting in version 2.1, a lot of sanity checking has been put into the config commands themselves, but binding will catch everything else).  Once bound, we can start (and of course later stop) the domain, which will trigger the boot process of OBP.  By default, the domain will then try to boot right away.  If we don't want that, we can set "auto-boot?" to false.  Finally, we'll use telnet to connect to the console of our newly created guest.  The output of "ldm list" shows us what port has been assigned to mars.  By default, the console service only listens on the loopback interface, so using telnet is not a large security concern here. root@sun # ldm set-variable auto-boot\?=false mars root@sun # ldm bind mars root@sun # ldm start mars root@sun # ldm list NAME STATE FLAGS CONS VCPU MEMORY UTIL UPTIME primary active -n-cv- UART 8 7680M 0.5% 1d 4h 30m mars active -t---- 5000 8 8G 12% 1s root@sun # telnet localhost 5000 Trying 127.0.0.1... Connected to localhost. Escape character is '^]'. ~Connecting to console "mars" in group "mars" .... Press ~? for control options .. {0} ok banner SPARC T3-4, No Keyboard Copyright (c) 1998, 2011, Oracle and/or its affiliates. All rights reserved. OpenBoot 4.33.1, 8192 MB memory available, Serial # 87203131. Ethernet address 0:21:28:24:1b:50, Host ID: 85241b50. {0} ok We're done, mars is ready to install Solaris, preferably using AI, of course ;-)  But before we do that, let's have a little look at the OBP environment to see how our virtual devices show up here: {0} ok printenv auto-boot? auto-boot? = false {0} ok printenv boot-device boot-device = disk net {0} ok devalias root /virtual-devices@100/channel-devices@200/disk@0 net0 /virtual-devices@100/channel-devices@200/network@0 net /virtual-devices@100/channel-devices@200/network@0 disk /virtual-devices@100/channel-devices@200/disk@0 virtual-console /virtual-devices/console@1 name aliases We can see that setting the OBP variable "auto-boot?" to false with the ldm command worked.  Of course, we'd normally set this to "true" to allow Solaris to boot right away once the LDom guest is started.  The setting for "boot-device" is the default "disk net", which means OBP would try to boot off the devices pointed to by the aliases "disk" and "net" in that order, which usually means "disk" once Solaris is installed on the disk image.  The actual devices these aliases point to are shown with the command "devalias".  Here, we have one line for both "disk" and "net".  The device paths speak for themselves.  Note that each of these devices has a second alias: "net0" for the network device and "root" for the disk device.  These are the very same names we've given these devices in the control domain with the commands "ldm add-vnet" and "ldm add-vdisk".  Remember this, as it is very useful once you have several dozen disk devices... To wrap this up, in this part we've created a simple guest domain, complete with CPU, memory, boot disk and network connectivity.  This should be enough to get you going.  I will cover all the more advanced features and a little more theoretical background in several follow-on articles.  For some background reading, I'd recommend the following links: LDoms 2.2 Admin Guide: Setting up Guest Domains Virtual Console Server: vntsd manpage - This includes the control sequences and commands available to control the console session. OpenBoot 4.x command reference - All the things you can do at the ok prompt

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• How to Easily Put a Windows PC into Kiosk Mode With Assigned Access

  - by Chris Hoffman

  Windows 8.1′s Assigned Access feature allows you to easily lock a Windows PC to a single application, such as a web browser. This feature makes it easy for anyone to configure Windows 8.1 devices as point-of-sale or other kiosk systems. In the past, setting up a Windows PC in kiosk mode involved much more work, requiring the use of third-party software, group policy, or Linux distributions designed around kiosk mode. Assigned Access is available on Windows 8.1 RT, Windows 8.1 Professional, and Windows 8.1 Enterprise. The standard edition of Windows 8.1 doesn’t support Assigned Access. Create a User Account for Assigned Access Rather than turn your entire computer into a locked-down kiosk system, Assigned Access allows you to create a separate user account that can only launch a single app — such as a web browser. To set this up, you must be logged into Windows as a user with administrator permissions. First, open the PC settings app — swipe in from the right or press Windows Key + C to open the charms bar, tap Settings, and tap Change PC settings. In the PC settings app, select Accounts and select Other accounts. Use the Add an account button to create a new Windows account. Select  the “Sign in without a Microsoft account” option and select Local account to create a local user account. You could also create a Microsoft account, but you may not want to do this if you just want a locked-down account with only browser access. If you need to install apps from the Windows Store to use in Assigned Access mode, you’ll have to set up a Microsoft account instead of a local account. A local account will still allow you access to the preinstalled apps, such as Internet Explorer. You may want to create a user account with a blank password. This would make it simple for anyone to access kiosk mode, even if the system becomes locked or needs to be rebooted. The account will be created as a standard user account with limited permissions. Leave it as a standard user account — don’t make it an administrator account. Set Up Assigned Access Once you’ve created an account, you’ll first need to sign into it. If you don’t, you’ll see a “This account has no apps” message when trying to enable Assigned Access. Go back to the welcome screen, log in to the new account you created, and allow Windows to go through the first-time account setup process. If you want to use a non-default app in kiosk mode, install it while logged in as that user account. Once you’re done, log out of the other account, log back in as your administrator account, and go back to the Other accounts screen. Click the Set up an account for assigned access option to continue. Select the user account you created and select the app you want to limit the account to. For a web-based kiosk, this can be a web browser such as the Modern version of Internet Explorer. Businesses can also create their own Modern apps and set them to run in kiosk mode in this way. Note that Microsoft’s documentation says “web browsers are not good choices for assigned access” because they require more permissions than average Modern (or “Windows Store”) apps. However, if you want to provide a kiosk for web-browsing, using Assigned Access is a much better option than using Guest Mode and offering up a full Windows desktop. When you’re done, restart your PC and log in as the Assigned Access account. Windows will automatically open the app you chose and won’t allow a user to leave that app. Standard Windows 8 features like the charms bar, app switcher, and Start screen won’t appear. Pressing the Windows key once will do nothing. To sign out of Assigned Access mode, press the Windows key five times — quickly — while signed in. You’ll be sent back to the standard login screen. The account will actually still be logged in and the app will remain running — this method just “locks” the screen and allows another user to log in. Automatically Log Into Assigned Access Whenever your Windows device boots, you can log into the Assigned Access account and turn it into a kiosk system. While this isn’t ideal for all kiosk systems, you may want the device to automatically launch the specific app when it boots without requiring any login process. To do so, you’ll just need to have Windows automatically log into the Assigned Access account when it boots. This option is hidden and not available in the standard Control Panel. You’ll need to use the hidden netplwiz Control Panel tool to set up automatic login on boot. If you didn’t create a password for the user account, leave the Password field empty while configuring this. Security Considerations If you’re using this feature to turn a Windows 8.1 system into a kiosk and leaving it open to the public, remember to consider security. Anyone could come up to the system, press the Windows key five times, and try to log into your standard administrator user account. Ensure the administrator user account has a strong password so people won’t be able to get past the kiosk system’s limitations and tamper with the system. Even Windows 8′s detractors have to admit that it’s an ideal system for a touch-screen kiosk device, running either a browser or another specific application. Assigned Access finally makes this easy to set up on Windows systems in the real world — no IT experience, third-party software, or Linux distributions necessary.     

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• 6 Ways to Free Up Hard Drive Space Used by Windows System Files

  - by Chris Hoffman

  We’ve previously covered the standard ways to free up space on Windows. But if you have a small solid-state drive and really want more hard space, there are geekier ways to reclaim hard drive space. Not all of these tips are recommended — in fact, if you have more than enough hard drive space, following these tips may actually be a bad idea. There’s a tradeoff to changing all of these settings. Erase Windows Update Uninstall Files Windows allows you to uninstall patches you install from Windows Update. This is helpful if an update ever causes a problem — but how often do you need to uninstall an update, anyway? And will you really ever need to uninstall updates you’ve installed several years ago? These uninstall files are probably just wasting space on your hard drive. A recent update released for Windows 7 allows you to erase Windows Update files from the Windows Disk Cleanup tool. Open Disk Cleanup, click Clean up system files, check the Windows Update Cleanup option, and click OK. If you don’t see this option, run Windows Update and install the available updates. Remove the Recovery Partition Windows computers generally come with recovery partitions that allow you to reset your computer back to its factory default state without juggling discs. The recovery partition allows you to reinstall Windows or use the Refresh and Reset your PC features. These partitions take up a lot of space as they need to contain a complete system image. On Microsoft’s Surface Pro, the recovery partition takes up about 8-10 GB. On other computers, it may be even larger as it needs to contain all the bloatware the manufacturer included. Windows 8 makes it easy to copy the recovery partition to removable media and remove it from your hard drive. If you do this, you’ll need to insert the removable media whenever you want to refresh or reset your PC. On older Windows 7 computers, you could delete the recovery partition using a partition manager — but ensure you have recovery media ready if you ever need to install Windows. If you prefer to install Windows from scratch instead of using your manufacturer’s recovery partition, you can just insert a standard Window disc if you ever want to reinstall Windows. Disable the Hibernation File Windows creates a hidden hibernation file at C:\hiberfil.sys. Whenever you hibernate the computer, Windows saves the contents of your RAM to the hibernation file and shuts down the computer. When it boots up again, it reads the contents of the file into memory and restores your computer to the state it was in. As this file needs to contain much of the contents of your RAM, it’s 75% of the size of your installed RAM. If you have 12 GB of memory, that means this file takes about 9 GB of space. On a laptop, you probably don’t want to disable hibernation. However, if you have a desktop with a small solid-state drive, you may want to disable hibernation to recover the space. When you disable hibernation, Windows will delete the hibernation file. You can’t move this file off the system drive, as it needs to be on C:\ so Windows can read it at boot. Note that this file and the paging file are marked as “protected operating system files” and aren’t visible by default. Shrink the Paging File The Windows paging file, also known as the page file, is a file Windows uses if your computer’s available RAM ever fills up. Windows will then “page out” data to disk, ensuring there’s always available memory for applications — even if there isn’t enough physical RAM. The paging file is located at C:\pagefile.sys by default. You can shrink it or disable it if you’re really crunched for space, but we don’t recommend disabling it as that can cause problems if your computer ever needs some paging space. On our computer with 12 GB of RAM, the paging file takes up 12 GB of hard drive space by default. If you have a lot of RAM, you can certainly decrease the size — we’d probably be fine with 2 GB or even less. However, this depends on the programs you use and how much memory they require. The paging file can also be moved to another drive — for example, you could move it from a small SSD to a slower, larger hard drive. It will be slower if Windows ever needs to use the paging file, but it won’t use important SSD space. Configure System Restore Windows seems to use about 10 GB of hard drive space for “System Protection” by default. This space is used for System Restore snapshots, allowing you to restore previous versions of system files if you ever run into a system problem. If you need to free up space, you could reduce the amount of space allocated to system restore or even disable it entirely. Of course, if you disable it entirely, you’ll be unable to use system restore if you ever need it. You’d have to reinstall Windows, perform a Refresh or Reset, or fix any problems manually. Tweak Your Windows Installer Disc Want to really start stripping down Windows, ripping out components that are installed by default? You can do this with a tool designed for modifying Windows installer discs, such as WinReducer for Windows 8 or RT Se7en Lite for Windows 7. These tools allow you to create a customized installation disc, slipstreaming in updates and configuring default options. You can also use them to remove components from the Windows disc, shrinking the size of the resulting Windows installation. This isn’t recommended as you could cause problems with your Windows installation by removing important features. But it’s certainly an option if you want to make Windows as tiny as possible. Most Windows users can benefit from removing Windows Update uninstallation files, so it’s good to see that Microsoft finally gave Windows 7 users the ability to quickly and easily erase these files. However, if you have more than enough hard drive space, you should probably leave well enough alone and let Windows manage the rest of these settings on its own. Image Credit: Yutaka Tsutano on Flickr     

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• Identity Propagation across Web and Web Service - 11g

  - by Prakash Yamuna

  I was on a customer call recently and this topic came up. In fact since this topic seems to come up fairly frequently - I thought I would describe the recommended model for doing SSO for Web Apps and then doing Identity Propagation across the Back end web services. The Image below shows a typical flow: Here is a more detailed drill down of what happens at each step of the flow (the number in red in the diagram maps to the description below of the behind the scenes processing that happens in the stack). [1] The Web App is protected with OAM and so the typical SSO scenario is applicable. The Web App URL is protected in OAM. The Web Gate intercepts the request from the Browser to the Web App - if there is an OAM (SSO) token - then the Web Gate validates the OAM token. If there is no SSO token - then the user is directed to the login page - user enters credentials, user is authenticated and OAM token is created for that browser session. [2] Once the Web Gate validates the OAM token - the token is propagated to the WLS Server where the Web App is running. You need to ensure that you have configured the OAM Identity Asserter in the Weblogic domain. If the OAM Identity Asserter is configured, this will end up creating a JAAS Subject. Details can be found at: http://docs.oracle.com/cd/E23943_01/doc.1111/e15478/webgate.htm#CACIAEDJ [3] The Web Service client (in the Web App) is secured with one of the OWSM SAML Client Policies. If secured in this fashion, the OWSM Agent creates a SAML Token from the JAAS Subject (created in [2] by the OAM Identity Asserter) and injects it into the SOAP message. Steps for securing a JEE JAX-WS Proxy Client using OWSM Policies are documented at: http://docs.oracle.com/cd/E23943_01/web.1111/b32511/attaching.htm#BABBHHHC Note: As shown in the diagram - instead of building a JEE Web App - you can also use WebCenter and build portlets. If you are using WebCenter then you can follow the same architecture. Only the steps for securing WebCenter Portlets with OWSM is different. 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:0in 5.4pt 0in 5.4pt; mso-para-margin:0in; 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-hansi-font-family:Calibri; mso-hansi-theme-font:minor-latin;} http://docs.oracle.com/cd/E23943_01/webcenter.1111/e12405/wcadm_security_wss.htm#CIHEBAHB [4] The SOA Composite App is secured with OWSM SAML Service policy. OWSM Agent intercepts the incoming SOAP request and validates the SAML token and creates a JAAS Subject. [5] When the SOA Composite App tries to invoke the OSB Proxy Service, the SOA Composite App "Reference" is secured with OWSM SAML Client Policy. Here again OWSM Agent will create a new SAML Token from the JAAS Subject created in [4] by the OWSM Agent and inject it into the SOAP message. Steps for securing SOA Composite Apps (Service, Reference, Component) are documented at: 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:0in 5.4pt 0in 5.4pt; mso-para-margin:0in; 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-hansi-font-family:Calibri; mso-hansi-theme-font:minor-latin;} http://docs.oracle.com/cd/E23943_01/web.1111/b32511/attaching.htm#CEGDGIHD [6] When the request reaches the OSB Proxy Service, the Proxy Service is again secured with the OWSM SAML Token Service Policy. So the same steps are performed as in [4]. The end result is a JAAS Subject. 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:0in 5.4pt 0in 5.4pt; mso-para-margin:0in; 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-hansi-font-family:Calibri; mso-hansi-theme-font:minor-latin;} [7] When OSB needs to invoke the Business App Web Service, it goes through the OSB Business Service. The OSB Business Service is secured with OWSM SAML Client Policy and step [5] is repeated. Steps for securing OSB Proxy Service and OSB Business Services are document at: http://docs.oracle.com/cd/E23943_01/admin.1111/e15867/proxy_services.htm#OSBAG1097[8] Finally when the message reaches the Business App Web Service, this service is protected by OWSM SAML Service policy and step [4] is repeated by the OWSM Agent. Steps for securing Weblogic Web Services, ADF Web Services, etc are documented at: http://docs.oracle.com/cd/E23943_01/web.1111/b32511/attaching.htm#CEGCJDIF In the above description for purposes of brevity - I have not described which OWSM SAML policies one should use; OWSM ships with a number of SAML policies, I briefly described some of the trade-offs involved with the various SAML policies here. The diagram above and the accompanying description of what is happening in each step of the flow - assumes you are using "SAML SV" or SAML Bearer" based policies without an STS.

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• Thread placement policies on NUMA systems - update

  - by Dave

  In a prior blog entry I noted that Solaris used a "maximum dispersal" placement policy to assign nascent threads to their initial processors. The general idea is that threads should be placed as far away from each other as possible in the resource topology in order to reduce resource contention between concurrently running threads. This policy assumes that resource contention -- pipelines, memory channel contention, destructive interference in the shared caches, etc -- will likely outweigh (a) any potential communication benefits we might achieve by packing our threads more densely onto a subset of the NUMA nodes, and (b) benefits of NUMA affinity between memory allocated by one thread and accessed by other threads. We want our threads spread widely over the system and not packed together. Conceptually, when placing a new thread, the kernel picks the least loaded node NUMA node (the node with lowest aggregate load average), and then the least loaded core on that node, etc. Furthermore, the kernel places threads onto resources -- sockets, cores, pipelines, etc -- without regard to the thread's process membership. That is, initial placement is process-agnostic. Keep reading, though. This description is incorrect. On Solaris 10 on a SPARC T5440 with 4 x T2+ NUMA nodes, if the system is otherwise unloaded and we launch a process that creates 20 compute-bound concurrent threads, then typically we'll see a perfect balance with 5 threads on each node. We see similar behavior on an 8-node x86 x4800 system, where each node has 8 cores and each core is 2-way hyperthreaded. So far so good; this behavior seems in agreement with the policy I described in the 1st paragraph. I recently tried the same experiment on a 4-node T4-4 running Solaris 11. Both the T5440 and T4-4 are 4-node systems that expose 256 logical thread contexts. To my surprise, all 20 threads were placed onto just one NUMA node while the other 3 nodes remained completely idle. I checked the usual suspects such as processor sets inadvertently left around by colleagues, processors left offline, and power management policies, but the system was configured normally. I then launched multiple concurrent instances of the process, and, interestingly, all the threads from the 1st process landed on one node, all the threads from the 2nd process landed on another node, and so on. This happened even if I interleaved thread creating between the processes, so I was relatively sure the effect didn't related to thread creation time, but rather that placement was a function of process membership. I this point I consulted the Solaris sources and talked with folks in the Solaris group. The new Solaris 11 behavior is intentional. The kernel is no longer using a simple maximum dispersal policy, and thread placement is process membership-aware. Now, even if other nodes are completely unloaded, the kernel will still try to pack new threads onto the home lgroup (socket) of the primordial thread until the load average of that node reaches 50%, after which it will pick the next least loaded node as the process's new favorite node for placement. On the T4-4 we have 64 logical thread contexts (strands) per socket (lgroup), so if we launch 48 concurrent threads we will find 32 placed on one node and 16 on some other node. If we launch 64 threads we'll find 32 and 32. That means we can end up with our threads clustered on a small subset of the nodes in a way that's quite different that what we've seen on Solaris 10. So we have a policy that allows process-aware packing but reverts to spreading threads onto other nodes if a node becomes too saturated. It turns out this policy was enabled in Solaris 10, but certain bugs suppressed the mixed packing/spreading behavior. There are configuration variables in /etc/system that allow us to dial the affinity between nascent threads and their primordial thread up and down: see lgrp_expand_proc_thresh, specifically. In the OpenSolaris source code the key routine is mpo_update_tunables(). This method reads the /etc/system variables and sets up some global variables that will subsequently be used by the dispatcher, which calls lgrp_choose() in lgrp.c to place nascent threads. Lgrp_expand_proc_thresh controls how loaded an lgroup must be before we'll consider homing a process's threads to another lgroup. Tune this value lower to have it spread your process's threads out more. To recap, the 'new' policy is as follows. Threads from the same process are packed onto a subset of the strands of a socket (50% for T-series). Once that socket reaches the 50% threshold the kernel then picks another preferred socket for that process. Threads from unrelated processes are spread across sockets. More precisely, different processes may have different preferred sockets (lgroups). Beware that I've simplified and elided details for the purposes of explication. The truth is in the code. Remarks: It's worth noting that initial thread placement is just that. If there's a gross imbalance between the load on different nodes then the kernel will migrate threads to achieve a better and more even distribution over the set of available nodes. Once a thread runs and gains some affinity for a node, however, it becomes "stickier" under the assumption that the thread has residual cache residency on that node, and that memory allocated by that thread resides on that node given the default "first-touch" page-level NUMA allocation policy. Exactly how the various policies interact and which have precedence under what circumstances could the topic of a future blog entry. The scheduler is work-conserving. The x4800 mentioned above is an interesting system. Each of the 8 sockets houses an Intel 7500-series processor. Each processor has 3 coherent QPI links and the system is arranged as a glueless 8-socket twisted ladder "mobius" topology. Nodes are either 1 or 2 hops distant over the QPI links. As an aside the mapping of logical CPUIDs to physical resources is rather interesting on Solaris/x4800. On SPARC/Solaris the CPUID layout is strictly geographic, with the highest order bits identifying the socket, the next lower bits identifying the core within that socket, following by the pipeline (if present) and finally the logical thread context ("strand") on the core. But on Solaris on the x4800 the CPUID layout is as follows. [6:6] identifies the hyperthread on a core; bits [5:3] identify the socket, or package in Intel terminology; bits [2:0] identify the core within a socket. Such low-level details should be of interest only if you're binding threads -- a bad idea, the kernel typically handles placement best -- or if you're writing NUMA-aware code that's aware of the ambient placement and makes decisions accordingly. Solaris introduced the so-called critical-threads mechanism, which is expressed by putting a thread into the FX scheduling class at priority 60. The critical-threads mechanism applies to placement on cores, not on sockets, however. That is, it's an intra-socket policy, not an inter-socket policy. Solaris 11 introduces the Power Aware Dispatcher (PAD) which packs threads instead of spreading them out in an attempt to be able to keep sockets or cores at lower power levels. Maximum dispersal may be good for performance but is anathema to power management. PAD is off by default, but power management polices constitute yet another confounding factor with respect to scheduling and dispatching. If your threads communicate heavily -- one thread reads cache lines last written by some other thread -- then the new dense packing policy may improve performance by reducing traffic on the coherent interconnect. On the other hand if your threads in your process communicate rarely, then it's possible the new packing policy might result on contention on shared computing resources. Unfortunately there's no simple litmus test that says whether packing or spreading is optimal in a given situation. The answer varies by system load, application, number of threads, and platform hardware characteristics. Currently we don't have the necessary tools and sensoria to decide at runtime, so we're reduced to an empirical approach where we run trials and try to decide on a placement policy. The situation is quite frustrating. Relatedly, it's often hard to determine just the right level of concurrency to optimize throughput. (Understanding constructive vs destructive interference in the shared caches would be a good start. We could augment the lines with a small tag field indicating which strand last installed or accessed a line. Given that, we could augment the CPU with performance counters for misses where a thread evicts a line it installed vs misses where a thread displaces a line installed by some other thread.)

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• Underwriting in a New Frontier: Spurring Innovation

  - by [email protected]

  Normal 0 false false false EN-US X-NONE X-NONE MicrosoftInternetExplorer4 st1\:*{behavior:url(#ieooui) } /* 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:0in 5.4pt 0in 5.4pt; mso-para-margin:0in; mso-para-margin-bottom:.0001pt; mso-pagination:widow-orphan; font-size:10.0pt; font-family:"Calibri","sans-serif";} Susan Keuer, product strategy manager for Oracle Insurance, shares her experiences and insight from the 2010 Association of Home Office Underwriters (AHOU) Annual Conference, April 11-14, in San Antonio, Texas    How can I be more innovative in underwriting?  It's a common question I hear from insurance carriers, producers and others, so it was no surprise that it was the key theme at the recent 2010 AHOU Annual Conference.  This year's event drew more than 900 insurance professionals involved in the underwriting process across life and annuities, property and casualty and reinsurance from around the globe, including the U.S., Canada, Australia, Bahamas, and more, to San Antonio - a Texas city where innovation transformed a series of downtown drainage canals into its premiere River Walk tourist destination.   CNN's Medical Correspondent Dr. Sanjay Gupta kicked off the conference with a phenomenal opening session that drove home the theme of the conference, "Underwriting in a New Frontier:  Spurring Innovation."   Drawing from his own experience as a neurosurgeon treating critically injured medical patients in the field in Iraq, Gupta inspired audience members to think outside the box during the underwriting process. He shared a compelling story of operating on a soldier who had suffered a head-related trauma in a field hospital.  With minimal supplies available Gupta used a Black and Decker saw to operate on the soldier's head and reduce pressure on his swelling brain. Drawing from this example, Gupta encouraged underwriters to think creatively, be innovative, and consider new tools and sources of information, such as social networking sites, during the underwriting process. So as you are looking at risk take into consideration all resources you have available.    Gupta also stressed the concept of IKIGAI - noting that individuals who believe that their life is worth living are less likely to die than are their counterparts without this belief.  How does one quantify this approach to life or thought process when evaluating risk?  Could this be something to consider as a "category" in the near future? How can this same belief in your own work spur innovation?   The role of technology was a hot topic of discussion throughout the conference.  Sessions delved into the latest in underwriting software to the rise of social media and how it is being increasingly integrated into underwriting process and solutions.  In one session a trio of panelists representing the carrier, producer and vendor communities stressed the importance to underwriters of leveraging new technology and the plethora of online information sources, which all could be used to accurately, honestly and consistently evaluate the risk throughout the underwriting process.   Another focused on the explosion of social media noting:  1.    Social media is growing exponentially - About eight percent of Americans used social media five years ago. Today about 46 percent of Americans do so, with 85 percent of financial services professionals using social media in their work.  2.    It will impact your business - Underwriters reconfirmed over and over that they are increasingly using "free" tools that are available in cyberspace in lieu of more costly solutions, such as inspection reports conducted by individuals in the field.  3.    Information is instantly available on the Web, anytime, anywhere - LinkedIn was mentioned as a way to connect to peers in the underwriting community and producers alike.  Many carriers and agents also are using Facebook to promote their company to customers - and as a point-of-entry to allow them to perform some functionality - such as accessing product marketing information versus directing users to go to the carrier's own proprietary website.  Other carriers have released their tight brand marketing to allow their producers to drive more business to their personal Facebook site where they offer innovative tools such as Application Capture or asking medical information in a more relaxed fashion.     Other key topics at the conference included the economy, ongoing industry consolidation, real-estate valuations as an asset and input into the underwriting process, and producer trends.  All stressed a "back to basics" approach for low cost, term products.   Finally, Connie Merritt, RN, PHN, entertained the large group of atttendees with audience-engaging insight on how to "Tame the Lions in Your Life - Dealing with Complainers, Bullies, Grump and Curmudgeon." Merritt noted "we are too busy for our own good." She shared how her overachieving personality had impacted her life.  Audience members then were asked to pick red, yellow, blue, or green shapes, without knowing that each one represented a specific personality trait.  For example, those who picked blue were the peacemakers. Those who choose yellow were social - the hint was to "Be Quiet Longer."  She then offered these "lion taming" steps:   1.    Admit It 2.    Accept It 3.    Let Go 4.    Be Present (which paralleled Gupta's IKIGAI concept)   When thinking about underwriting I encourage you to be present in the moment and think creatively, but don't be afraid to look ahead to the future and be an innovator.  I hope to see you at next year's AHOU Annual Conference, May 1-4, 2011 at The Mirage in Las Vegas, Nev.     Susan Keuer is the product strategy manager for new business underwriting.  She brings more than 20 years of insurance industry experience working with leading insurance carriers and technology companies to her role on the product strategy team for life/annuities solutions within the Oracle Insurance Global Business Unit  

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• Integrating Coherence & Java EE 6 Applications using ActiveCache

  - by Ricardo Ferreira

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

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• Charms and the App Bar

  - by Dennis Vroegop

  Ok. I admit. I made a mistake in the last post about our planespotter app. I have dedicated a full part of the hub to Social. I also had a section called Friends but that made sense since I said that “Friends” is a special group of people that connect to each other through our app and only our app. Social however is sharing our spots with Twitter, Facebook and so on. Now, we could write that functionality in our app in a different section but there is one small problem with that: users don’t expect that. Ok, I admit. The mistake was quite deliberate to give me an excuse to write this part. But still: the mistake is one I see a lot. People are trying to do stuff in their application that they shouldn’t be doing. This always strike me as slightly odd: why do some work when others have already done it for you and you can just use it? After all: good developers are lazy (lazy people will always try to find the easiest way to do something and in development land this usually means the cleanest and best to support way…) So. What is that part that Microsoft has done for us and we don’t have to do ourselves? The answer lies on the right hand of your Win8 screen: This is a screenshot of my tablet (as you can see I am writing this right now….) When I swipe my finger from out of the screen on the right inside the screen (or move the mouse to the upper right corner) this menu will appear. Next to settings and the start menu button we’ll find the Search and the Share charms. These are two ways that your app can share the information it contains with the rest of the world, or at least: the rest of your system. So don’t write a Search feature in your app. Don’t write a Share feature in your app. It’s here already. Users, once they are used to Windows 8, will use that feature and expect it to work. If it doesn’t, they won’t like your app and you can kiss you dreams of everlasting fame goodbye. So use these two. What are they? Well, simply they are parts of a contract. In your app you say somewhere in code that you are supporting Search and Share. So when the user selects Share the system will interrogate the current app in the foreground if it supports this feature. Your app will say “But why, yes, I do!” Then the system will ask the app “Ok then, wisecrack, then share!” and you will have to provide the system with some information about the format. Other applications have subscribed to be at the receiving end of the Share contract. They have told the system that they support Sharing (receiving) and which formats they understand. If one or more of them support the formats you specify, the user will see them. The user clicks / taps on the app of their choice and data is moved from your app to the new one. So if you say you support Facebook and Twitter users can post data from your app to these networks by selecting Share. The same applies to Search. Don’t make a “search” button in your app but use the contract to tell the system that you support search and use that instead. Users will be grateful (remember that bar with men/women/creatures that are waiting for you?) The more and more people get to know Windows 8, the more they will use this. And if you are one of the people who wrote an app that helped them learn the system, well, that’s even better. So. We don’t have a Share or a Search button. We do have other buttons. Most important: we probably need a “New Spot” button. And a “Filter” might be useful. Or someway to open the camera so you can add a picture to the spot. Where will be put those? The answer is the “Appbar” . This is a application / context aware menu that slides up from the bottom of the screen when you move your finger / mouse from below the screen into it. From above downwards works just as well. Here you see an example of the appbar from the People app. (click on it for a larger version). This appears whenever you slide your finger up from below of down from above. This is where you put your commands. Remember, this is context aware so this menu will change when you are in different parts of your app or when you have selected different items. There are a few conventions when you create this appbar. First, the items on the right are “General” items, meaning they have little to do with what is on the screen right now. I think this would be a great place to add our “New Spot” icon. On the far left are items associated with the current selected item or screen. So if you have a spot selected, the button for Add Photo should be visible here and on the left hand side. Not everything is as clear as this, but this is what you should strive for. Group items together. And please note: this is the only place in Metro design where we are allowed to use lines as separators. So when you want to separate a group of icons from another group, add a line. Also note the simplicity of the buttons. No colors, no lights or shadows, no 3D. After a couple of years of fancy almost realistic looking icons people have finally decided that hey, this is a virtual world: it’s ok to look virtual as well. So make things as readable and clear as possible and don’t try to duplicate nature. It’s all about the information, remember? (If you don’t remember I’d like to point you to a older blog post of mine about the what and why of Metro). So.. think about the buttons a bit and think about Share and Search. What will you put there? Remember: this is the way the users interact with your apps and while you shouldn’t judge a book by its covers when it comes to people, this isn’t entirely so when it comes to apps. People DO judge an app by its looks and the way it feels. Take advantage of that. History has learned that a crappy app with a GREAT user interface gets better reviews than a GREAT app with a lousy UI… I know: developers will find this extremely unfair but that’s the world we live in (No, I am not saying you should deliver rubbish apps). Next time: we’ll start by building the darn thing!

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• Windows Azure Evolution &ndash; Deploy Web Sites (WAWS Part 3)

  - by Shaun

  This is the sixth post of my Windows Azure Evolution series. After talked a bit about the new caching preview feature in the previous one, let’s back to the Windows Azure Web Sites (WAWS).   Git and GitHub Integration In the third post I introduced the overview functionality of WAWS and demonstrated how to create a WordPress blog through the build-in application gallery. And in the fourth post I covered how to use the TFS service preview to deploy an ASP.NET MVC application to the web site through the TFS integration. WAWS also have the Git integration. I’m not going to talk very detailed about the Git and GitHub integration since there are a bunch of information on the internet you can refer to. To enable the Git just go to the web site item in the developer portal and click the “Set up Git publishing”. After specified the username and password the windows azure platform will establish the Git integration and provide some basic guide. As you can see, you can download the Git binaries, commit the files and then push to the remote repository. Regarding the GitHub, since it’s built on top of Git it should work. Maarten Balliauw have a wonderful post about how to integrate GitHub to Windows Azure Web Site you can find here.   WebMatrix 2 RC WebMatrix is a lightweight web application development tool provided by Microsoft. It utilizes WebDeploy or FTP to deploy the web application to the server. And in WebMatrix 2.0 RC it added the feature to work with Windows Azure. First of all we need to download the latest WebMatrix 2 through the Web Platform Installer 4.0. Just open the WebPI and search “WebMatrix”, or go to its home page download its web installer. Once we have WebMatrix 2, we need to download the publish file of our WAWS. Let’s go to the developer portal and open the web site we want to deploy and download the publish file from the link on the right hand side. This file contains the necessary information of publishing the web site through WebDeploy and FTP, which can be used in WebMatrix, Visual Studio, etc.. Once we have the publish file we can open the WebMatrix, click the Open Site, Remote Site. Then it will bring up a dialog where we can input the information of the remote site. Since we have our publish file already, we can click the “Import publish settings” and select the publish file, then we can see the site information will be populated automatically. Click OK, the WebMatrix will connect to the remote site, which is the WAWS we had deployed already, retrieve the folders and files information. We can open files in WebMatrix and modify. But since WebMatrix is a lightweight web application tool, we cannot update the backend C# code. So in this case, we will modify the frontend home page only. After saved our modification, WebMatrix will compare the files between in local and remote and then it will only upload the modified files to Windows Azure through the connection information in the publish file. Since it only update the files which were changed, this minimized the bandwidth and deployment duration. After few seconds we back to the website and the modification had been applied.   Visual Studio and WebDeploy The publish file we had downloaded can be used not only in WebMatrix but also Visual Studio. As we know in Visual Studio we can publish a web application by clicking the “Publish” item from the project context menu in the solution explorer, and we can specify the WebDeploy, FTP or File System for the publish target. Now we can use the WAWS publish file to let Visual Studio publish the web application to WAWS. Let’s create a new ASP.NET MVC Web Application in Visual Studio 2010 and then click the “Publish” in solution explorer. Once we have the Windows Azure SDK 1.7 installed, it will update the web application publish dialog. So now we can import the publish information from the publish file. Select WebDeploy as the publish method. We can select FTP as well, which is supported by Windows Azure and the FTP information was in the same publish file. In the last step the publish wizard can check the files which will be uploaded to the remote site before the actually publishing. This gives us a chance to review and amend the files. Same as the WebMatrix, Visual Studio will compare the files between local and WAWS and determined which had been changed and need to be published. Finally Visual Studio will publish the web application to windows azure through WebDeploy protocol. Once it finished we can browse our website.   FTP Deployment The publish file we downloaded contains the connection information to our web site via both WebDeploy and FTP. When using WebMatrix and Visual Studio we can select WebDeploy or FTP. WebDeploy method can be used very easily from WebMatrix and Visual Studio, with the file compare feature. But the FTP gives more flexibility. We can use any FTP client to upload files to windows azure regardless which client and OS we are using. Open the publish file in any text editor, we can find the connection information very easily. As you can see the publish file is actually a XML file with WebDeploy and FTP information in plain text attributes. And once we have the FTP URL, username and password, when can connect to the site and upload and download files. For example I opened FileZilla and connected to my WAWS through FTP. Then I can download files I am interested in and modify them on my local disk. Then upload back to windows azure through FileZilla. Then I can see the new page.   Summary In this simple and quick post I introduced vary approaches to deploy our web application to Windows Azure Web Site. It supports TFS integration which I mentioned previously. It also supports Git and GitHub, WebDeploy and FTP as well.   Hope this helps, Shaun All documents and related graphics, codes are provided "AS IS" without warranty of any kind. Copyright © Shaun Ziyan Xu. This work is licensed under the Creative Commons License.

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• Custom Model Binding of IEnumerable Properties in ASP.Net MVC 2

  - by Doug Lampe

  MVC 2 provides a GREAT feature for dealing with enumerable types.  Let's say you have an object with a parent/child relationship and you want to allow users to modify multiple children at the same time.  You can simply use the following syntax for any indexed enumerables (arrays, generic lists, etc.) and then your values will bind to your enumerable model properties. 1: <% using (Html.BeginForm("TestModelParameter", "Home")) 2: { %> 3: < table > 4: < tr >< th >ID</th><th>Name</th><th>Description</th></tr> 5: <% for (int i = 0; i < Model.Items.Count; i++) 6: { %> 7: < tr > 8: < td > 9: <%= i %> 10: </ td > 11: < td > 12: <%= Html.TextBoxFor(m => m.Items[i].Name) %> 13: </ td > 14: < td > 15: <%= Model.Items[i].Description %> 16: </ td > 17: </ tr > 18: <% } %> 19: </ table > 20: < input type ="submit" /> 21: <% } %> Then just update your model either by passing it into your action method as a parameter or explicitly with UpdateModel/TryUpdateModel. 1: public ActionResult TestTryUpdate() 2: { 3: ContainerModel model = new ContainerModel(); 4: TryUpdateModel(model); 5:   6: return View("Test", model); 7: } 8:   9: public ActionResult TestModelParameter(ContainerModel model) 10: { 11: return View("Test", model); 12: } Simple right?  Well, not quite.  The problem is the DefaultModelBinder and how it sets properties.  In this case our model has a property that is a generic list (Items).  The first bad thing the model binder does is create a new instance of the list.  This can be fixed by making the property truly read-only by removing the set accessor.  However this won't help because this behaviour continues.  As the model binder iterates through the items to "set" their values, it creates new instances of them as well.  This means you lose any information not passed via the UI to your controller so in the examplel above the "Description" property would be blank for each item after the form posts. One solution for this is custom model binding.  I have put together a solution which allows you to retain the structure of your model.  Model binding is a somewhat advanced concept so you may need to do some additional research to really understand what is going on here, but the code is fairly simple.  First we will create a binder for the parent object which will retain the state of the parent as well as some information on which children have already been bound. 1: public class ContainerModelBinder : DefaultModelBinder 2: { 3: /// <summary> 4: /// Gets an instance of the model to be used to bind child objects. 5: /// </summary> 6: public ContainerModel Model { get; private set; } 7:   8: /// <summary> 9: /// Gets a list which will be used to track which items have been bound. 10: /// </summary> 11: public List<ItemModel> BoundItems { get; private set; } 12:   13: public ContainerModelBinder() 14: { 15: BoundItems = new List<ItemModel>(); 16: } 17:   18: protected override object CreateModel(ControllerContext controllerContext, ModelBindingContext bindingContext, Type modelType) 19: { 20: // Set the Model property so child binders can find children. 21: Model = base.CreateModel(controllerContext, bindingContext, modelType) as ContainerModel; 22:   23: return Model; 24: } 25: } Next we will create the child binder and have it point to the parent binder to get instances of the child objects.  Note that this only works if there is only one property of type ItemModel in the parent class since the property to find the item in the parent is hard coded. 1: public class ItemModelBinder : DefaultModelBinder 2: { 3: /// <summary> 4: /// Gets the parent binder so we can find objects in the parent's collection 5: /// </summary> 6: public ContainerModelBinder ParentBinder { get; private set; } 7: 8: public ItemModelBinder(ContainerModelBinder containerModelBinder) 9: { 10: ParentBinder = containerModelBinder; 11: } 12:   13: protected override object CreateModel(ControllerContext controllerContext, ModelBindingContext bindingContext, Type modelType) 14: { 15: // Find the item in the parent collection and add it to the bound items list. 16: ItemModel item = ParentBinder.Model.Items.FirstOrDefault(i => !ParentBinder.BoundItems.Contains(i)); 17: ParentBinder.BoundItems.Add(item); 18: 19: return item; 20: } 21: } Finally, we will register these binders in Global.asax.cs so they will be used to bind the classes. 1: protected void Application_Start() 2: { 3: AreaRegistration.RegisterAllAreas(); 4:   5: ContainerModelBinder containerModelBinder = new ContainerModelBinder(); 6: ModelBinders.Binders.Add(typeof(ContainerModel), containerModelBinder); 7: ModelBinders.Binders.Add(typeof(ItemModel), new ItemModelBinder(containerModelBinder)); 8:   9: RegisterRoutes(RouteTable.Routes); 10: } I'm sure some of my fellow geeks will comment that this could be done more efficiently by simply rewriting some of the methods of the default model binder to get the same desired behavior.  I like my method shown here because it extends the binder class instead of modifying it so it minimizes the potential for unforseen problems. In a future post (if I ever get around to it) I will explore creating a generic version of these binders.

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