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• The best way to separate admin functionality from a public site?

  - by AndrewO

  I'm working on a site that's grown both in terms of user-base and functionality to the point where it's becoming evident that some of the admin tasks should be separate from the public website. I was wondering what the best way to do this would be. For example, the site has a large social component to it, and a public sales interface. But at the same time, there's back office tasks, bulk upload processing, dashboards (with long running queries), and customer relations tools in the admin section that I would like to not be effected by spikes in public traffic (or effect the public-facing response time). The site is running on a fairly standard Rails/MySQL/Linux stack, but I think this is more of an architecture problem than an implementation one: mainly, how does one keep the data and business logic in sync between these different applications? Some strategies that I'm evaluating: 1) Create a slave database of the public facing database on another machine. Extract out all of the model and library code so that it can be shared between the applications. Create new controllers and views for the admin interfaces. I have limited experience with replication and am not even sure that it's supposed to be used this way (most of the time I've seen it, it's been for scaling out the read capabilities of the same application, rather than having multiple different ones). I'm also worried about the potential for latency issues if the slave is not on the same network. 2) Create new more task/department-specific applications and use a message oriented middleware to integrate them. I read Enterprise Integration Patterns awhile back and they seemed to advocate this for distributed systems. (Alternatively, in some cases the basic Rails-style RESTful API functionality might suffice.) But, I have nightmares about data synchronization issues and the massive re-architecting that this would entail. 3) Some mixture of the two. For example, the only public information necessary for some of the back office tasks is a read-only completion time or status. Would it make sense to have that on a completely separate system and send the data to public? Meanwhile, the user/group admin functionality would be run on a separate system sharing the database? The downside is, this seems to keep many of the concerns I have with the first two, especially the re-architecting. I'm sure the answers are going to be highly dependent on a site's specific needs, but I'd love to hear success (or failure) stories.

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• Upgrading php from php 5.3 to 5.4 .7

  - by Takingsides

  So, quickly so to speak I have noticed this topic around, I have searched and there are plenty of solutions. However these solutions do not work for me, not only that but I'm intending to learn more about the Debian based OS. Questions I would like to know how to upgrade php5.3 to php 5.4.7 compiling it from source, myself without using a third-party ppa. Is the way (explained below) the correct way of configuring php5.4? I'm new to compiling from source. Set-up I run Ubuntu Server 12.04 64bit. I've currently got: PHP 5.3 MySQL-Server Apache2 Memcached The Problem So I initially installed php5.3 using apt-get. I now wish to upgrade the php 5.4 due to the advantage of traits in OOP and the struct with Arrays and all the other recent patches and such. Possible Solutions I've seen this ondrej/ppa repository, which I refuse to use, given the fact that it may work, but it's an unknown/untrusted source. ALso, i'm not learning how to administer from source, using configure, make and install accordingly. I've seen a solution compiling from source, which is essentially how I was hoping to go about it with some guidance. Conclusion So I didn't just expect to be spoon-fed, and I went out and did some manual reading and atleast started the ball rolling myself; this how far i've got. The first thing I did was su into root (to save the typing sudo all the darn time). $ sudo su The next thing I did was download the latest version of php (5.4.7) and extracted it's contents ready to configure before installing it. $ mkdir php5-new && cd !$ $ wget -O php-5.4.7.tar.bz2 http://php.net/get/php-5.4.7.tar.bz2/from/uk3.php.net/mirror $ bzip2 -d php-5.4.7.tar.bz2 $ tar xvf php-5.4.7.tar.gz $ cd php-5.4.7 $ ./configure --help Finally I decided to have a bash, I looked through the list of options and decided I needed to list ALL of the things I wanted to include in the configuration. $ ./configure --with-mysql --with-apache2 --with-libxml --with-openssl --with-zlib --with-bz2 --with-curl --with-dom --with-gd --with-imap --with-imap-ssl --with-mcrypt --with-mysqli --with-pdo-mysql --with-libxml --enable-ftp --enable-mbstring --enable-soap Finally, the results... When the configuration process had finished, it threw an error: configure: error: xml2-config not found. Please check your libxml2 installation.

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• Setting up Netbeans/Eclipse for Linux Kernel Development

  - by red.october

  Hi: I'm doing some Linux kernel development, and I'm trying to use Netbeans. Despite declared support for Make-based C projects, I cannot create a fully functional Netbeans project. This is despite compiling having Netbeans analyze a kernel binary that was compiled with full debugging information. Problems include: files are wrongly excluded: Some files are incorrectly greyed out in the project, which means Netbeans does not believe they should be included in the project, when in fact they are compiled into the kernel. The main problem is that Netbeans will miss any definitions that exist in these files, such as data structures and functions, but also miss macro definitions. cannot find definitions: Pretty self-explanatory - often times, Netbeans cannot find the definition of something. This is partly a result of the above problem. can't find header files: self-explanatory I'm wondering if anyone has had success with setting up Netbeans for Linux kernel development, and if so, what settings they used. Ultimately, I'm looking for Netbeans to be able to either parse the Makefile (preferred) or extract the debug information from the binary (less desirable, since this can significantly slow down compilation), and automatically determine which files are actually compiled and which macros are actually defined. Then, based on this, I would like to be able to find the definitions of any data structure, variable, function, etc. and have complete auto-completion. Let me preface this question with some points: I'm not interested in solutions involving Vim/Emacs. I know some people like them, but I'm not one of them. As the title suggest, I would be also happy to know how to set-up Eclipse to do what I need While I would prefer perfect coverage, something that only misses one in a million definitions is obviously fine SO's useful "Related Questions" feature has informed me that the following question is related: http://stackoverflow.com/questions/149321/what-ide-would-be-good-for-linux-kernel-driver-development. Upon reading it, the question is more of a comparison between IDE's, whereas I'm looking for how to set-up a particular IDE. Even so, the user Wade Mealing seems to have some expertise in working with Eclipse on this kind of development, so I would certainly appreciate his (and of course all of your) answers. Cheers

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• Init script & the green [ OK ]

  - by Lord Loh.

  I am trying to install fast-cgi for nginx on an EC2 instance. I followed the steps explained here, but that is meant for Debian and does not work out of the box for a red-hat based system. I modified the script a bit to look like - #!/bin/bash ### BEGIN INIT INFO # Provides: php-fcgi # Required-Start: $nginx # Required-Stop: $nginx # Default-Start: 2 3 4 5 # Default-Stop: 0 1 6 # Short-Description: starts php over fcgi # Description: starts php over fcgi ### END INIT INFO . /etc/rc.d/init.d/functions (( EUID )) && echo .You need to have root priviliges.. && exit 1 BIND=/tmp/php.socket USER=nginx PHP_FCGI_CHILDREN=15 PHP_FCGI_MAX_REQUESTS=1000 PHP_CGI=/usr/bin/php-cgi PHP_CGI_NAME=`basename $PHP_CGI` PHP_CGI_ARGS="- USER=$USER PATH=/usr/bin PHP_FCGI_CHILDREN=$PHP_FCGI_CHILDREN PHP_FCGI_MAX_REQUESTS=$PHP_FCGI_MAX_REQUESTS $PHP_CGI -b $BIND" RETVAL=0 start() { echo -n "Starting PHP FastCGI: " #ORIGINAL LINE #daemon $PHP_CGI --quiet --start --background --chuid "$USER" --exec /usr/bin/env -- $PHP_CGI_ARGS #MODIFIED LINE daemon --user=$USER $PHP_CGI -b $BIND& RETVAL=$? echo [ $RETVAL -eq 0 ] && touch /var/lock/subsys/php-fcgi #echo "$PHP_CGI_NAME." } stop() { echo -n "Stopping PHP FastCGI: " killall -q -w -u $USER $PHP_CGI RETVAL=$? echo "$PHP_CGI_NAME." rm /var/lock/subsys/php-fcgi } case "$1" in start) start ;; stop) stop ;; restart) stop start ;; *) echo "Usage: php-fastcgi {start|stop|restart}" exit 1 ;; esac exit $RETVAL The problem I have now is - service php-fcgi start keeps the shell blocked. If I run service php-fcgi start & and then ps aux, I see the php-cgi process running bound to the socket. I see the start command stop only when I execute service php-fcgi stop. How do I solve this blocking issue? I have tried adding an & at the end of the line spawning the daemon. But other scripts do not seem to be doing this. This is the most complicated script I am attempting to modify yet :-( How do I get the script to display the green [ OK ]? I checked scripts like httpd and saw that all they were doing was something as shown below. But I never see a green [ OK ] when I execute php-fcgi. I also discovered that putting echo_success with functions sourced displays the green [ OK ] but I do not see any other scripts in the /etc/rc.d/init.d/ executing echo_success or echo_failure. What have I got wrong? Also, How do i specify PHP_FCGI_CHILDREN with daemon? echo [ $RETVAL -eq 0 ] && touch /var/lock/subsys/

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• ReadFile doesn't work asynchronously on Win7 and Win2k8

  - by f0b0s

  According to MSDN ReadFile can read data 2 different ways: synchronously and asynchronously. I need the second one. The folowing code demonstrates usage with OVERLAPPED struct: #include <windows.h> #include <stdio.h> #include <time.h> void Read() { HANDLE hFile = CreateFileA("c:\\1.avi", GENERIC_READ, 0, NULL, OPEN_EXISTING, FILE_FLAG_OVERLAPPED, NULL); if ( hFile == INVALID_HANDLE_VALUE ) { printf("Failed to open the file\n"); return; } int dataSize = 256 * 1024 * 1024; char* data = (char*)malloc(dataSize); memset(data, 0xFF, dataSize); OVERLAPPED overlapped; memset(&overlapped, 0, sizeof(overlapped)); printf("reading: %d\n", time(NULL)); BOOL result = ReadFile(hFile, data, dataSize, NULL, &overlapped); printf("sent: %d\n", time(NULL)); DWORD bytesRead; result = GetOverlappedResult(hFile, &overlapped, &bytesRead, TRUE); // wait until completion - returns immediately printf("done: %d\n", time(NULL)); CloseHandle(hFile); } int main() { Read(); } On Windows XP output is: reading: 1296651896 sent: 1296651896 done: 1296651899 It means that ReadFile didn't block and returned imediatly at the same second, whereas reading process continued for 3 seconds. It is normal async reading. But on windows 7 and windows 2008 I get following results: reading: 1296661205 sent: 1296661209 done: 1296661209. It is a behavior of sync reading. MSDN says that async ReadFile sometimes can behave as sync (when the file is compressed or encrypted for example). But the return value in this situation should be TRUE and GetLastError() == NO_ERROR. On Windows 7 I get FALSE and GetLastError() == ERROR_IO_PENDING. So WinApi tells me that it is an async call, but when I look at the test I see that it is not! I'm not the only one who found this "bug": read the comment on ReadFile MSDN page. So what's the solution? Does anybody know? It is been 14 months after Denis found this strange behavior.

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• How to salvage SQL server 2008 query from KILLED/ROLLBACK state without waiting half a day?

  - by littlegreen

  I have a stored procedure that inserts batches of millions of rows, emerging from a certain query, into an SQL database. It has one parameter selecting the batch; when this parameter is omitted, it will gather a list of batches and recursively call itself, in order to iterate over batches. In (pseudo-)code, it looks something like this: CREATE PROCEDURE spProcedure AS BEGIN IF @code = 0 BEGIN ... WHILE @@Fetch_Status=0 BEGIN EXEC spProcedure @code FETCH NEXT ... INTO @code END END ELSE BEGIN -- Disable indexes ... INSERT INTO table SELECT (...) -- Enable indexes ... Now it can happen that this procedure is slow, for whatever reason: it can't get a lock, one of the indexes it uses is misdefined or disabled. In that case, I want to be able kill the procedure, truncate and recreate the resulting table, and try again. However, when I try and kill the procedure, the process frequently oozes into a KILLED/ROLLBACK state from which there seems to be no return. From Google I have learned to do an sp_lock, find the spid, and then kill it with KILL <spid>. But when I try to kill it, it tells me SPID 75: transaction rollback in progress. Estimated rollback completion: 0%. Estimated time remaining: 554 seconds. I did find a forum message hinting that another spid should be killed before the other one can start a rollback. But that didn't work for me either, plus I do not understand, why that would be the case... could it be because I am recursively calling my own stored procedure? (But it should be having the same spid, right?) In any case, my process is just sitting there, being dead, not responding to kills, and locking the table. This is very frustrating, as I want to go on developing my queries, not waiting hours on my server sitting dead while pretending to be finishing a supposed rollback. Is there some way in which I can tell the server not to store any rollback information for my query? Or not to allow any other queries to interfere with the rollback, so that it will not take so long? Or how to rewrite my query in a better way, or how kill the process successfully without restarting the server?

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• How would you implement this "WorkerChain" functionality in .NET?

  - by Dan Tao

  Sorry for the vague question title -- not sure how to encapsulate what I'm asking below succinctly. (If someone with editing privileges can think of a more descriptive title, feel free to change it.) The behavior I need is this. I am envisioning a worker class that accepts a single delegate task in its constructor (for simplicity, I would make it immutable -- no more tasks can be added after instantiation). I'll call this task T. The class should have a simple method, something like GetToWork, that will exhibit this behavior: If the worker is not currently running T, then it will start doing so right now. If the worker is currently running T, then once it is finished, it will start T again immediately. GetToWork can be called any number of times while the worker is running T; the simple rule is that, during any execution of T, if GetToWork was called at least once, T will run again upon completion (and then if GetToWork is called while T is running that time, it will repeat itself again, etc.). Now, this is pretty straightforward with a boolean switch. But this class needs to be thread-safe, by which I mean, steps 1 and 2 above need to comprise atomic operations (at least I think they do). There is an added layer of complexity. I have need of a "worker chain" class that will consist of many of these workers linked together. As soon as the first worker completes, it essentially calls GetToWork on the worker after it; meanwhile, if its own GetToWork has been called, it restarts itself as well. Logically calling GetToWork on the chain is essentially the same as calling GetToWork on the first worker in the chain (I would fully intend that the chain's workers not be publicly accessible). One way to imagine how this hypothetical "worker chain" would behave is by comparing it to a team in a relay race. Suppose there are four runners, W1 through W4, and let the chain be called C. If I call C.StartWork(), what should happen is this: If W1 is at his starting point (i.e., doing nothing), he will start running towards W2. If W1 is already running towards W2 (i.e., executing his task), then once he reaches W2, he will signal to W2 to get started, immediately return to his starting point and, since StartWork has been called, start running towards W2 again. When W1 reaches W2's starting point, he'll immediately return to his own starting point. If W2 is just sitting around, he'll start running immediately towards W3. If W2 is already off running towards W3, then W2 will simply go again once he's reached W3 and returned to his starting point. The above is probably a little convoluted and written out poorly. But hopefully you get the basic idea. Obviously, these workers will be running on their own threads. Also, I guess it's possible this functionality already exists somewhere? If that's the case, definitely let me know!

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• cd Command Linux and Mystery Flags

  - by Jason R. Mick

  Platform: CentOS 6.2 Shell:tcsh I'm playing around with cd for a BASH script, and noticed the wondrous cd - option, but was left with many questions... Why the cd -? Isn't this redundant with cd ..? EDIT [As FatalError points out, these two commands don't do the same things... so the answer is "no"] Can you delve farther back into your history with - flag, a la in a browser? e.g. When I type cd -, it takes me to my previous directory, but then if I enter that command again, it takes me to the directory I just came from, creating a sort of loop. Is a shorthand for going back multiple levels supported?EDITI realize I can go back with cd .., but was hoping this could be a gateway to a less verbose deep back, e.g. cd -3 vs. cd ../../../ ... hopefully that clarifies what I'm asking....EDIT2As to the current feedback, while .. is a special directory, I don't see a reason why the built-in cd to the terminal couldn't use a shorthand for ../../ ... ../ e.g. cd ..5 or why the built-in also couldn't have a history (a la auto pushd/popd) that could be turned on and used like cd -3. I get that this could be somewhat of security/privacy risk, but I don't see how it's any worst than storing a command history, which most shells/terminals do. The manpage for cd, accessible via man cd and help cd (it's the same for either command), only lists -L and -P flags. However when I type in cd --help it outputs Usage: cd [-plvn][-|<dir>].. Am I right in assuming the other flags and the - (back) option are nonstandard? What are the -n and -v flags for? Both seem to take me back to my home directory, that's all I've been able to figure out via experimentation. A quick read on web resources [1][2] offered just the same sort of info that the man page did and didn't answer my questions. Note: The second Linux-centric resource above claimed cd only had two options (obviously not true in current CentOS) hence my assumption that this functionality could be non-standard.

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• git private server error: "Permission denied (publickey)."

  - by goddfree

  I followed the instructions here in order to set up a private git server on my Amazon EC2 instance. However, I am having problems when trying to SSH into the git account. Specifically, I get the error "Permission denied (publickey)." Here are the permissions of my files/folders on the EC2 server: drwx------ 4 git git 4096 Aug 13 19:52 /home/git/ drwx------ 2 git git 4096 Aug 13 19:52 /home/git/.ssh -rw------- 1 git git 400 Aug 13 19:51 /home/git/.ssh/authorized_keys Here are the permissions of my files/folders on my own computer: drwx------ 5 CYT staff 170 Aug 13 14:51 .ssh -rw------- 1 CYT staff 1679 Aug 13 13:53 .ssh/id_rsa -rw-r--r-- 1 CYT staff 400 Aug 13 13:53 .ssh/id_rsa.pub -rw-r--r-- 1 CYT staff 1585 Aug 13 13:53 .ssh/known_hosts When checking my logs in /var/log/secure, I used to get the following error message every time I tried to SSH: Authentication refused: bad ownership or modes for file /home/git/.ssh/authorized_keys However, after making a few permission changes, I no longer get this error message. Despite this, I am still getting the "Permission denied (publickey)." message every time I try to SSH. The command I am using to SSH is ssh -T git@my-ip. Here is the full log I get when I run ssh -vT [email protected]: OpenSSH_6.2p2, OSSLShim 0.9.8r 8 Dec 2011 debug1: Reading configuration data /etc/ssh_config debug1: /etc/ssh_config line 20: Applying options for * debug1: Connecting to my-ip [my-ip] port 22. debug1: Connection established. debug1: identity file /Users/CYT/.ssh/id_rsa type -1 debug1: identity file /Users/CYT/.ssh/id_rsa-cert type -1 debug1: identity file /Users/CYT/.ssh/id_dsa type -1 debug1: identity file /Users/CYT/.ssh/id_dsa-cert type -1 debug1: Enabling compatibility mode for protocol 2.0 debug1: Local version string SSH-2.0-OpenSSH_6.2 debug1: Remote protocol version 2.0, remote software version OpenSSH_6.2 debug1: match: OpenSSH_6.2 pat OpenSSH* debug1: SSH2_MSG_KEXINIT sent debug1: SSH2_MSG_KEXINIT received debug1: kex: server->client aes128-ctr [email protected] none debug1: kex: client->server aes128-ctr [email protected] none debug1: SSH2_MSG_KEX_DH_GEX_REQUEST(1024<1024<8192) sent debug1: expecting SSH2_MSG_KEX_DH_GEX_GROUP debug1: SSH2_MSG_KEX_DH_GEX_INIT sent debug1: expecting SSH2_MSG_KEX_DH_GEX_REPLY debug1: Server host key: RSA 08:ad:8a:bc:ab:4d:5f:73:24:b2:78:69:46:1a:a5:5a debug1: Host 'my-ip' is known and matches the RSA host key. debug1: Found key in /Users/CYT/.ssh/known_hosts:1 debug1: ssh_rsa_verify: signature correct debug1: SSH2_MSG_NEWKEYS sent debug1: expecting SSH2_MSG_NEWKEYS debug1: SSH2_MSG_NEWKEYS received debug1: Roaming not allowed by server debug1: SSH2_MSG_SERVICE_REQUEST sent debug1: SSH2_MSG_SERVICE_ACCEPT received debug1: Authentications that can continue: publickey debug1: Next authentication method: publickey debug1: Trying private key: /Users/CYT/.ssh/id_rsa debug1: Trying private key: /Users/CYT/.ssh/id_dsa debug1: No more authentication methods to try. Permission denied (publickey). I have spent a few hours going through threads on various sites, including SO and SF, looking for a solution. It seems that the permissions for my files are all okay, but I just can't figure out the problem. Any help would be greatly appreciated. Edit: EEAA: Here are the outputs you requested: $ getent passwd git git:x:503:504::/home/git:/bin/bash $ grep ssh ~git/.ssh/authorized_keys | wc -l grep: /home/git/.ssh/authorized_keys: Permission denied 0

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• UIVIewController not released when view is dismissed

  - by Nelson Ko

  I have a main view, mainWindow, which presents a couple of buttons. Both buttons create a new UIViewController (mapViewController), but one will start a game and the other will resume it. Both buttons are linked via StoryBoard to the same View. They are segued to modal views as I'm not using the NavigationController. So in a typical game, if a person starts a game, but then goes back to the main menu, he triggers: [self dismissViewControllerAnimated:YES completion:nil ]; to return to the main menu. I would assume the view controller is released at this point. The user resumes the game with the second button by opening another instance of mapViewController. What is happening, tho, is some touch events will trigger methods on the original instance (and write status updates to them - therefore invisible to the current view). When I put a breakpoint in the mapViewController code, I can see the instance will be one or the other (one of which should be released). I have tried putting a delegate to the mainWindow clearing the view: [self.delegate clearMapView]; where in the mainWindow - (void) clearMapView{ gameWindow = nil; } I have also tried self.view=nil; in the mapViewController. The mapViewController code contains MVC code, where the model is static. I wonder if this may prevent ARC from releasing the view. The model.m contains: static CanShieldModel *sharedInstance; + (CanShieldModel *) sharedModel { @synchronized(self) { if (!sharedInstance) sharedInstance = [[CanShieldModel alloc] init]; return sharedInstance; } return sharedInstance; } Another post which may have a lead, but so far not successful, is UIViewController not being released when popped I have in ViewDidLoad: // checks to see if app goes inactive - saves. [[NSNotificationCenter defaultCenter] addObserver:self selector:@selector(resignActive) name:UIApplicationWillResignActiveNotification object:nil]; with the corresponding in ViewDidUnload: [[NSNotificationCenter defaultCenter] removeObserver:self name:UIApplicationWillResignActiveNotification object:nil]; Does anyone have any suggestions? EDIT: - (void) prepareForSegue:(UIStoryboardSegue *)segue sender:(id)sender{ NSString *identifier = segue.identifier; if ([identifier isEqualToString: @"Start Game"]){ gameWindow = (ViewController *)[segue destinationViewController]; gameWindow.newgame=-1; gameWindow.delegate = self; } else if ([identifier isEqualToString: @"Resume Game"]){ gameWindow = (ViewController *)[segue destinationViewController]; gameWindow.newgame=0; gameWindow.delegate = self;

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• Akka framework support for finding duplicate messages

  - by scala_is_awesome

  I'm trying to build a high-performance distributed system with Akka and Scala. If a message requesting an expensive (and side-effect-free) computation arrives, and the exact same computation has already been requested before, I want to avoid computing the result again. If the computation requested previously has already completed and the result is available, I can cache it and re-use it. However, the time window in which duplicate computation can be requested may be arbitrarily small. e.g. I could get a thousand or a million messages requesting the same expensive computation at the same instant for all practical purposes. There is a commercial product called Gigaspaces that supposedly handles this situation. However there seems to be no framework support for dealing with duplicate work requests in Akka at the moment. Given that the Akka framework already has access to all the messages being routed through the framework, it seems that a framework solution could make a lot of sense here. Here is what I am proposing for the Akka framework to do: 1. Create a trait to indicate a type of messages (say, "ExpensiveComputation" or something similar) that are to be subject to the following caching approach. 2. Smartly (hashing etc.) identify identical messages received by (the same or different) actors within a user-configurable time window. Other options: select a maximum buffer size of memory to be used for this purpose, subject to (say LRU) replacement etc. Akka can also choose to cache only the results of messages that were expensive to process; the messages that took very little time to process can be re-processed again if needed; no need to waste precious buffer space caching them and their results. 3. When identical messages (received within that time window, possibly "at the same time instant") are identified, avoid unnecessary duplicate computations. The framework would do this automatically, and essentially, the duplicate messages would never get received by a new actor for processing; they would silently vanish and the result from processing it once (whether that computation was already done in the past, or ongoing right then) would get sent to all appropriate recipients (immediately if already available, and upon completion of the computation if not). Note that messages should be considered identical even if the "reply" fields are different, as long as the semantics/computations they represent are identical in every other respect. Also note that the computation should be purely functional, i.e. free from side-effects, for the caching optimization suggested to work and not change the program semantics at all. If what I am suggesting is not compatible with the Akka way of doing things, and/or if you see some strong reasons why this is a very bad idea, please let me know. Thanks, Is Awesome, Scala

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• Visual studio 2010 colourizers, intellisense and the rest. Where to start!!

  - by Owen

  Ok, before I begin I realize that there is a lot of documentation on this subject but I have thus far failed to get even basic colourization working for VS2010. My goal is to simply get to a point where I can open a document and everything is coloured red, from here I can implement the relevant parsing logic. Here's what I have tried/found: 1) Downloaded all the relevent SDK's and such- Found the ook sample (http://code.msdn.microsoft.com/ookLanguage) - didn't build, didn't work. 2) Knowing almost nothing about MEF read through "Implementing a Language Service By Using the Managed Package Framework" - http://msdn.microsoft.com/en-us/library/bb166533(v=VS.100).aspx This was pretty much a copy and paste of all the basic stuff here, and also updating some references which were out of date with the sample see: http://social.msdn.microsoft.com/Forums/en-US/vsx/thread/a310fe67-afd2-4592-b295-3fc86fec7996 Now, I have got to a point where when running the package MEF appears to have hooked up correctly (I know this because with the debugger open I can see that the packages initialize and FDoIdle methods are being hit). When I open a file of the extension I have registered with the ProvideLanguageExtensionAttribute everything dies as if in an endless loop, yet no debug symbols hit (though they are loaded). Looking at the ook sample and the MEF examples they seem to be totally different approaches to the same problem. In the ook sample there are notions of Clasifications and Completion controllers which aren't mentioned in the MEF example. Also, they don't seem to create a Package or Language service, so I have no idea how it should work? With the MEF example, my assumption is that I need to hook into the "IScanner.ScanTokenAndProvideInfoAboutIt" to provide syntax highlighting? Which would be fine if I could ever hit this method. So my first question I guess is which approach should I be taking here? Or do they both somehow tie together? My second questions is, where can I find a basic fully working project that implements bog standard basic syntax highlighting and intellisense or VS2010? Thirdly, in the MEF example when I created a Package there were a bunch of test projects created for me. I appears that the integration tests launch the VS2010 test rig somehow, but the test fails. It would be good to write my service with tests but I have no idea what/how I can test each interaction so any references to testing Language services would be helpful. Finally, please throw any resource/book links my way that I may find useful. Cheers, Chris. N.B. Sorry I realize this is part question part rant, but I have never been so confused.

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• Strange behavior: save video recorded within app?

  - by Josue Espinosa

  I allow the user to record a video within my app, then later play it again. When a user records a video, I save the URL of the video, then play the video later from the saved URL. I save the video both in the Photos app and in my app. If I delete the video within the photos app, it still plays. After about 7 days, the video gets deleted. I think I am saving in my tmp directory, but i'm not sure. Here is what I am doing: -(void)imagePickerController:(UIImagePickerController *)picker didFinishPickingMediaWithInfo:(NSDictionary *)info { NSString *mediaType = [info objectForKey: UIImagePickerControllerMediaType]; [self dismissViewControllerAnimated:YES completion:nil]; // Handle a movie capture if (CFStringCompare ((__bridge_retained CFStringRef) mediaType, kUTTypeMovie, 0) == kCFCompareEqualTo) { NSString *moviePath = [NSString stringWithFormat:@"%@",[[info objectForKey:UIImagePickerControllerMediaURL] path]]; NSURL *videoURL = [info objectForKey:UIImagePickerControllerMediaURL]; NSData *videoData = [NSData dataWithContentsOfURL:videoURL]; _justRecordedVideoURL = [NSString stringWithFormat:@"%@",videoURL]; AppDelegate *appDelegate = [[UIApplication sharedApplication] delegate]; _managedObjectContext = [appDelegate managedObjectContext]; Video *video = [NSEntityDescription insertNewObjectForEntityForName:@"Video" inManagedObjectContext:_managedObjectContext]; [video setVideoData:videoData]; [video setVideoURL:[NSString stringWithFormat:@"%@",videoURL]]; NSDateFormatter *dateFormatter = [[NSDateFormatter alloc] init]; dateFormatter.dateStyle = NSDateFormatterLongStyle; [dateFormatter setDateStyle:NSDateFormatterLongStyle]; NSDate *date = [dateFormatter dateFromString:[dateFormatter stringFromDate:[NSDate date]]]; NSString *dateAdded = [dateFormatter stringFromDate:date]; [video setDate_recorded:dateAdded]; if(_currentAthlete != nil){ [video setWhosVideo:_currentAthlete]; } NSError *error = nil; if(![_managedObjectContext save:&error]){ //handle dat error } NSArray *paths = NSSearchPathForDirectoriesInDomains(NSDocumentDirectory, NSUserDomainMask, YES); NSString *documentsDirectory = [paths objectAtIndex:0]; NSString *tempPath = [documentsDirectory stringByAppendingFormat:@"/vid1.mp4"]; BOOL success = [videoData writeToFile:tempPath atomically:NO]; if(success == FALSE){ NSLog(@"Video was not successfully saved."); } if (UIVideoAtPathIsCompatibleWithSavedPhotosAlbum(moviePath)) { UISaveVideoAtPathToSavedPhotosAlbum(moviePath, self, @selector(video:didFinishSavingWithError:contextInfo:), nil); } } } Am I saving it incorrectly? When I go to play the video, it works fine, after a couple days the video will play without audio, then eventually it will be gone. Any ideas why?

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• bluetooth connection using pybluez

  - by srj0408

  I am working on bluetooth not exactly on bluetooth stack-development but to use bluetooth in one of my project. I had done all that before using some of the py-bluez commands like hciconfig, hcitool scan , then simple-agents and using serial module inside python. But that was quite random. We were able to connect only one specific device based on its bluetooth address and there was no facility of reconnection once the devices are disconnected. Now i want to try out this stuff in a sequential manner like this (i am doing that all on a RPI and for at present on ubuntu 12.04.) i) Store some names in a file along with some other information with respect to that device. ii) Run a script to find out the device in locality with those names and if any one if found, report that. For this step, i had taken a reference from BTBook , made available from MIT. Below is the script for the same, but that script only search for the single name. from bluetooth import * target_name = "XT1033" target_address = None nearby_devices = discover_devices() for address in nearby_devices: if target_name == lookup_name( address ): target_address = address break if target_address is not None: print "found target bluetooth device with address ", target_address connect_socket(target_address); else: print "could not find target bluetooth device nearby" iii) Connect the device using client sock. But i dont have any device on which i can write a simple python script. My client can be any device that will be publishing data. Now i came through a script in the same book, that actually connect to a client requesting permission to connect to server. from bluetooth import * port = 1 server_sock=BluetoothSocket( RFCOMM ) server_sock.bind(("",port)) server_sock.listen(1) client_sock, client_info = server_sock.accept() print "Accepted connection from ", client_info data = client_sock.recv(1024) print "received [%s]" % data client_sock.close() server_sock.close() here client_sock, client_info = server_sock.accept() provide the client address and port requested to be connected. Can i pass address obtained from the earlier script to this, so that it connect server to the client? iv) Then if client get disconnected, re-connect(a simple polling can be used.) All this stuff can be done using bash and py-bluez functions but i want to do that in a sequential manner.I am not a master in python but i can do some small stuff. Can any one guide me for the same or can direct me to more usefull resource through which i can continue my coding part after finding the "X", "Y" named devices.

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• 'Timeout Expired' error against local SQL Express on only 2 LINQ Methods...

  - by Refracted Paladin

  I am going to sum up my problem first and then offer massive details and what I have already tried. Summary: I have an internal winform app that uses Linq 2 Sql to connect to a local SQL Express database. Each user has there own DB and the DB stay in sync through Merge Replication with a Central DB. All DB's are SQL 2005(sp2or3). We have been using this app for over 5 months now but recently our users are getting a Timeout expired. The timeout period elapsed prior to completion of the operation or the server is not responding. Detailed: The strange part is they get that in two differnt locations(2 differnt LINQ Methods) and only the first time they fire in a given time period(~5mins). One LINQ method is pulling all records that match a FK ID and then Manipulating them to form a Heirarchy View for a TreeView. The second is pulling all records that match a FK ID and dumping them into a DataGridView. The only things I can find in common with the 2 are that the first IS an IEnumerable and the second converts itself from IQueryable - IEnumerable - DataTable... I looked at the query's in Profiler and they 'seemed' normal. They are not very complicated querys. They are only pulling back 10 - 90 records, from one table. Any thoughts, suggestions, hints whatever would be greatly appreciated. I am at my wit's end on this.... public IList<CaseNoteTreeItem> GetTreeViewDataAsList(int personID) { var myContext = MatrixDataContext.Create(); var caseNotesTree = from cn in myContext.tblCaseNotes where cn.PersonID == personID orderby cn.ContactDate descending, cn.InsertDate descending select new CaseNoteTreeItem { CaseNoteID = cn.CaseNoteID, NoteContactDate = Convert.ToDateTime(cn.ContactDate). ToShortDateString(), ParentNoteID = cn.ParentNote, InsertUser = cn.InsertUser, ContactDetailsPreview = cn.ContactDetails.Substring(0, 75) }; return caseNotesTree.ToList<CaseNoteTreeItem>(); } AND THIS ONE public static DataTable GetAllCNotes(int personID) { using (var context = MatrixDataContext.Create()) { var caseNotes = from cn in context.tblCaseNotes where cn.PersonID == personID orderby cn.ContactDate select new { cn.ContactDate, cn.ContactDetails, cn.TimeSpentUnits, cn.IsCaseLog, cn.IsPreEnrollment, cn.PresentAtContact, cn.InsertDate, cn.InsertUser, cn.CaseNoteID, cn.ParentNote }; return caseNotes.ToList().CopyLinqToDataTable(); } }

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• jQuery Accordion /Tabs - Call different function upon load of content

  - by Scott

  I have the following markup: <body> <div id="tabs" style="float: left"> <ul> <li><a href="ajax/question_0.html" title="Question 1"><span>Question 1</span></a></li> <li><a href="ajax/question_1.html" title="Question 2"><span>Question 2</span></a></li> <li><a href="ajax/question_2.html" title="Question 3"><span>Question 3</span></a></li> <li><a href="ajax/question_3.html" title="Question 4"><span>Question 4</span></a></li> <li><a href="ajax/question_4.html" title="Question 5"><span>Question 5</span></a></li> <li><a href="ajax/question_5.html" title="Question 6"><span>Question 6</span></a></li> </ul> <div id="dynamicContent"> </div> </div> <div class="demo-description"> <p>Click tabs to see answers to Questions</p> </div> </body> I would like to utilize an accordion or the tabs plugin from the UI. Upon completion of the load event, I'd like to call a JavaScript function, but a different function for each tab -- almost like calling onDocumentReady for a page. I have the following JavaScript: $(function() { $('#tabs').tabs().bind('tabsselect', function(event, ui) { console.log("Loading: ajax/question_" + ui.index + ".html"); return true; //Ensure that the tab gets selected }); ); That's properly loading the Ajax file and all, but whenever I attempt to do something such as evaluating a statement in the JS, it seems to be ignored. Is there any way I can do this, so that once the file is loaded my function is called -- needs to be a different function for each tab.

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• C++: incorrect swapping of nodes in linked list

  - by Dragon

  I have 2 simple structures: struct Address { char city[255]; }; typedef Address* AddressPtr; struct Person { char fullName[255]; Address* address; Person* next; }; typedef Person* PersonPtr; The Person structure forms the Linked list where new elements are added to the beginning of the list. What I want to do is to sort them by fullName. At first I tried to swap links, but I lost the beginning of the list and as a result my list was sorted partially. Then I decided to sort list by swapping the values of nodes. But I get strange results. For a list with names: Test3, Test2, Test1, I get Test3, Test3, Test3. Here is my sorting code: void sortByName(PersonPtr& head) { TaskPtr currentNode, nextNode; for(currentNode = head; currentNode->next != NULL; currentNode = currentNode->next) { for(nextNode = currentNode->next; nextNode != NULL; nextNode = nextNode->next) { if(strcmp(currentNode->fullName, nextNode->fullName) > 0) { swapNodes(currentNode, nextNode); } } } } void swapNodes(PersonPtr& node1, PersonPtr& node2) { PersonPtr temp_node = node2; strcpy(node2->fullName, node1->fullName); strcpy(node1->fullName, temp_node->fullName); strcpy(node2->address->city, node1->address->city); strcpy(node1->address->city, temp_node->address->city); } After the sorting completion, nodes values are a little bit strange. UPDATED This is how I swapped links void swapNodes(PersonPtr& node1, PersonPtr& node2) { PersonPtr temp_person; AddressPtr temp_address; temp_person = node2; node2 = node1; node1 = temp_person; temp_address = node2->address; node2->address = node1->address; node1->address = temp_address; }

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• Asterisk server firewall script allows 2-way audio from incoming calls, but not on outgoing?

  - by cappie

  I'm running an Asterisk PBX on a virtual machine directly connected to the Internet and I really want to prevent script kiddies, l33t h4x0rz and actual hackers access to my server. The basic way I protect my calling-bill now is by using 32 character passwords, but I would much rather have a way to protect The firewall script I'm currently using is stated below, however, without the established connection firewall rule (mentioned rule #1), I cannot receive incoming audio from the target during outgoing calls: #!/bin/bash # first, clean up! iptables -F iptables -X iptables -t nat -F iptables -t nat -X iptables -t mangle -F iptables -t mangle -X iptables -P INPUT ACCEPT iptables -P FORWARD DROP # we're not a router iptables -P OUTPUT ACCEPT # don't allow invalid connections iptables -A INPUT -m state --state INVALID -j DROP # always allow connections that are already set up (MENTIONED RULE #1) iptables -A INPUT -m state --state RELATED,ESTABLISHED -j ACCEPT # always accept ICMP iptables -A INPUT -p icmp -j ACCEPT # always accept traffic on these ports #iptables -A INPUT -p tcp --dport 80 -j ACCEPT iptables -A INPUT -p tcp --dport 22 -j ACCEPT # always allow DNS traffic iptables -A INPUT -p udp --sport 53 -j ACCEPT iptables -A OUTPUT -p udp --dport 53 -j ACCEPT # allow return traffic to the PBX iptables -A INPUT -p udp -m udp --dport 50000:65536 -j ACCEPT iptables -A INPUT -p udp -m udp --dport 10000:20000 -j ACCEPT iptables -A INPUT -p udp --destination-port 5060:5061 -j ACCEPT iptables -A INPUT -p tcp --destination-port 5060:5061 -j ACCEPT iptables -A INPUT -m multiport -p udp --dports 10000:20000 iptables -A INPUT -m multiport -p tcp --dports 10000:20000 # IP addresses of the office iptables -A INPUT -s 95.XXX.XXX.XXX/32 -j ACCEPT # accept everything from the trunk IP's iptables -A INPUT -s 195.XXX.XXX.XXX/32 -j ACCEPT iptables -A INPUT -s 195.XXX.XXX.XXX/32 -j ACCEPT # accept everything on localhost iptables -A INPUT -i lo -j ACCEPT # accept all outgoing traffic iptables -A OUTPUT -j ACCEPT # DROP everything else #iptables -A INPUT -j DROP I would like to know what firewall rule I'm missing for this all to work.. There is so little documentation on which ports (incoming and outgoing) asterisk actually needs.. (return ports included). Are there any firewall/iptables specialists here that see major problems with this firewall script? It's so frustrating not being able to find a simple firewall solution that enabled me to have a PBX running somewhere on the Internet which is firewalled in such a way that it can ONLY allows connections from and to the office, the DNS servers and the trunk(s) (and only support SSH (port 22) and ICMP traffic for the outside world). Hopefully, using this question, we can solve this problem once and for all.

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• Yet another Memory Leak Issue (memory is still gone when program terminates)- C program on SLES

  - by user1426181

  I run my C program on Suse Linux Enterprise that compresses several thousand large files (between 10MB and 100MB in size), and the program gets slower and slower as the program runs (it's running multi-threaded with 32 threads on a Intel Sandy Bridge board). When the program completes, and it's run again, it's still very slow. When I watch the program running, I see that the memory is being depleted while the program runs, which you would think is just a classic memory leak problem. But, with a normal malloc()/free() mismatch, I would expect all the memory to return when the program terminates. But, most of the memory doesn't get reclaimed when the program completes. The free or top command shows Mem: 63996M total, 63724M used, 272M free when the program is slowed down to a halt, but, after the termination, the free memory only grows back to about 3660M. When the program is rerun, the free memory is quickly used up. The top program only shows that the program, while running, is using at most 4% or so of the memory. I thought that it might be a memory fragmentation problem, but, I built a small test program that simulates all the memory allocation activity in the program (many randomized aspects were built in - size/quantity), and it always returns all the memory upon completion. So, I don't think that's it. Questions: Can there be a malloc()/free() mismatch that will lose memory permanently, i.e. even after the process completes? What other things in a C program (not C++) can cause permanent memory loss, i.e. after the program completes, and even the terminal window closes? Only a reboot brings the memory back. I've read other posts about files not being closed causing problems, but, I don't think I have that problem. Is it valid to be looking at top and free for the memory statistics, i.e. do they accurately describe the memory situation? They do seem to correspond to the slowness of the program. If the program only shows a 4% memory usage, will something like valgrind find this problem?

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• How do you return a string from a function correctly in Dynamic C?

  - by aquanar

  I have a program I am trying to debug, but Dynamic C apparently treats strings differently than normal C does (well, character arrays, anyway). I have a function that I made to make an 8 character long (well, 10 to include the \0 ) string of 0s and 1s to show me the contents of an 8-bit char variable. (IE, I give it the number 13, it returns the string "0001101\0" ) When I use the code below, it prints out !{happy face] 6 times (well, the second one is the happy face alone for some reason), each return comes back as 0xDEAE or "!\x02. I thought it would dereference it and return the appropriate string, but it appears to just be sending the pointer and attempting to parse it. This may seem silly, but my experience was actually in C++ and Java, so going back to C brings up a few issues that were dealt with in later programming languages that I'm not entirely sure how to deal with (like the lack of string variables). How could I fix this code, or how would be a better way to do what I am trying to do (I thought maybe sending in a pointer to a character array and working on it from the function might work, but I thought I should ask to see if maybe I'm just trying to reinvent the wheel). Currently I have it set up like this: this is an excerpt from the main() display[0] = '\0'; for(i=0;i<6;i++) { sprintf(s, "%s ", *char_to_bits(buffer[i])); strcat(display, s); } DispStr(8,5, display); and this is the offending function: char *char_to_bits(char x) { char bits[16]; strcpy(bits,"00000000\0"); if (x & 0x01) bits[7]='1'; if (x & 0x02) bits[6]='1'; if (x & 0x04) bits[5]='1'; if (x & 0x08) bits[4]='1'; if (x & 0x10) bits[3]='1'; if (x & 0x20) bits[2]='1'; if (x & 0x40) bits[1]='1'; if (x & 0x80) bits[0]='1'; return bits; } and just for the sake of completion, the other function is used to output to the stdio window at a specific location: void DispStr(int x, int y, char *s) { x += 0x20; y += 0x20; printf ("\x1B=%c%c%s", x, y, s); }

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• Parallelism in .NET – Part 14, The Different Forms of Task

  - by Reed

  Before discussing Task creation and actual usage in concurrent environments, I will briefly expand upon my introduction of the Task class and provide a short explanation of the distinct forms of Task.  The Task Parallel Library includes four distinct, though related, variations on the Task class. In my introduction to the Task class, I focused on the most basic version of Task.  This version of Task, the standard Task class, is most often used with an Action delegate.  This allows you to implement for each task within the task decomposition as a single delegate. Typically, when using the new threading constructs in .NET 4 and the Task Parallel Library, we use lambda expressions to define anonymous methods.  The advantage of using a lambda expression is that it allows the Action delegate to directly use variables in the calling scope.  This eliminates the need to make separate Task classes for Action<T>, Action<T1,T2>, and all of the other Action<…> delegate types.  As an example, suppose we wanted to make a Task to handle the ”Show Splash” task from our earlier decomposition.  Even if this task required parameters, such as a message to display, we could still use an Action delegate specified via a lambda: // Store this as a local variable string messageForSplashScreen = GetSplashScreenMessage(); // Create our task Task showSplashTask = new Task( () => { // We can use variables in our outer scope, // as well as methods scoped to our class! this.DisplaySplashScreen(messageForSplashScreen); }); .csharpcode, .csharpcode pre { font-size: small; color: black; font-family: consolas, "Courier New", courier, monospace; background-color: #ffffff; /*white-space: pre;*/ } .csharpcode pre { margin: 0em; } .csharpcode .rem { color: #008000; } .csharpcode .kwrd { color: #0000ff; } .csharpcode .str { color: #006080; } .csharpcode .op { color: #0000c0; } .csharpcode .preproc { color: #cc6633; } .csharpcode .asp { background-color: #ffff00; } .csharpcode .html { color: #800000; } .csharpcode .attr { color: #ff0000; } .csharpcode .alt { background-color: #f4f4f4; width: 100%; margin: 0em; } .csharpcode .lnum { color: #606060; } This provides a huge amount of flexibility.  We can use this single form of task for any task which performs an operation, provided the only information we need to track is whether the task has completed successfully or not.  This leads to my first observation: Use a Task with a System.Action delegate for any task for which no result is generated. This observation leads to an obvious corollary: we also need a way to define a task which generates a result.  The Task Parallel Library provides this via the Task<TResult> class. Task<TResult> subclasses the standard Task class, providing one additional feature – the ability to return a value back to the user of the task.  This is done by switching from providing an Action delegate to providing a Func<TResult> delegate.  If we decompose our problem, and we realize we have one task where its result is required by a future operation, this can be handled via Task<TResult>.  For example, suppose we want to make a task for our “Check for Update” task, we could do: Task<bool> checkForUpdateTask = new Task<bool>( () => { return this.CheckWebsiteForUpdate(); }); Later, we would start this task, and perform some other work.  At any point in the future, we could get the value from the Task<TResult>.Result property, which will cause our thread to block until the task has finished processing: // This uses Task<bool> checkForUpdateTask generated above... // Start the task, typically on a background thread checkForUpdateTask.Start(); // Do some other work on our current thread this.DoSomeWork(); // Discover, from our background task, whether an update is available // This will block until our task completes bool updateAvailable = checkForUpdateTask.Result; This leads me to my second observation: Use a Task<TResult> with a System.Func<TResult> delegate for any task which generates a result. Task and Task<TResult> provide a much cleaner alternative to the previous Asynchronous Programming design patterns in the .NET framework.  Instead of trying to implement IAsyncResult, and providing BeginXXX() and EndXXX() methods, implementing an asynchronous programming API can be as simple as creating a method that returns a Task or Task<TResult>.  The client side of the pattern also is dramatically simplified – the client can call a method, then either choose to call task.Wait() or use task.Result when it needs to wait for the operation’s completion. While this provides a much cleaner model for future APIs, there is quite a bit of infrastructure built around the current Asynchronous Programming design patterns.  In order to provide a model to work with existing APIs, two other forms of Task exist.  There is a constructor for Task which takes an Action<Object> and a state parameter.  In addition, there is a constructor for creating a Task<TResult> which takes a Func<Object, TResult> as well as a state parameter.  When using these constructors, the state parameter is stored in the Task.AsyncState property. While these two overloads exist, and are usable directly, I strongly recommend avoiding this for new development.  The two forms of Task which take an object state parameter exist primarily for interoperability with traditional .NET Asynchronous Programming methodologies.  Using lambda expressions to capture variables from the scope of the creator is a much cleaner approach than using the untyped state parameters, since lambda expressions provide full type safety without introducing new variables.

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• Top 10 Reasons SQL Developer is Perfect for Oracle Beginners

  - by thatjeffsmith

  Learning new technologies can be daunting. If you’ve never used a Mac before, you’ll probably be a bit baffled at first. But, you’re probably at least coming from a desktop computing background (Windows), so you common frame of reference. But what if you’re just now learning to use a relational database? Yes, you’ve played with Access a bit, but now your employer or college instructor has charged you with becoming proficient with Oracle database. Here’s 10 reasons why I think Oracle SQL Developer is the perfect vehicle to help get you started. 1. It’s free No need to break into one of these… No start-up costs, no need to wrangle budget dollars from your company. Students don’t have any money after books and lab fees anyway. And most employees don’t like having to ask for ‘special’ software anyway. So avoid all of that and make sure the free stuff doesn’t suit your needs first. Upgrades are available on a regular base, also at no cost, and support is freely available via our public forums. 2. It will run pretty much anywhere Windows – check. OSX (Apple) – check. Unix – check. Linux – check. No need to start up a windows VM to run your Windows-only software in your lab machine. 3. Anyone can install it There’s no installer, no registry to be updated, no admin privs to be obtained. If you can download and extract files to your machine or USB storage device, you can run it. You can be up and running with SQL Developer in under 5 minutes. Here’s a video tutorial to see how to get started. 4. It’s ubiquitous I admit it, I learned a new word yesterday and I wanted an excuse to use it. SQL Developer’s everywhere. It’s had over 2,500,000 downloads in the past year, and is the one of the most downloaded items from OTN. This means if you need help, there’s someone sitting nearby you that can assist, and since they’re in the same tool as you, they’ll be speaking the same language. 5. Simple User Interface Up-up-down-down-Left-right-left-right-A-B-A-B-START will get you 30 lives, but you already knew that, right? You connect, you see your objects, you click on your objects. Or, you can use the worksheet to write your queries and programs in. There’s only one toolbar, and just a few buttons. If you’re like me, video games became less fun when each button had 6 action items mapped to it. I just want the good ole ‘A’, ‘B’, ‘SELECT’, and ‘START’ controls. If you’re new to Oracle, you shouldn’t have the double-workload of learning a new complicated tool as well. 6. It’s not a ‘black box’ Click through your objects, but also get the SQL that drives the GUI As you use the wizards to accomplish tasks for you, you can view the SQL statement being generated on your behalf. Just because you have a GUI, doesn’t mean you’re ceding your responsibility to learn the underlying code that makes the database work. 7. It’s four tools in one It’s not just a query tool. Maybe you need to design a data model first? Or maybe you need to migrate your Sybase ASE database to Oracle for a new project? Or maybe you need to create some reports? SQL Developer does all of that. So once you get comfortable with one part of the tool, the others will be much easier to pick up as your needs change. 8. Great learning resources available Videos, blogs, hands-on learning labs – you name it, we got it. Why wait for someone to train you, when you can train yourself at your own pace? 9. You can use it to teach yourself SQL Instead of being faced with the white-screen-of-panic, you can visually build your queries by dragging and dropping tables and views into the Query Builder. Yes, ‘just like Access’ – only better. And as you build your query, toggle to the Worksheet panel and see the SQL statement. Again, SQL Developer is not a black box. If you prefer to learn by trial and error, the worksheet will attempt to suggest the next bit of your SQL statement with it’s completion insight feature. And if you have syntax errors, those will be highlighted – just like your misspelled words in your favorite word processor. 10. It scales to match your experience level You won’t be a n00b forever. In 6-8 months, when you’re ready to tackle something a bit more complicated, like XML DB or Oracle Spatial, the tool is already there waiting on you. No need to go out and find the ‘advanced’ tool. 11. Wait, you said this was a ‘Top 10′ list? Yes. Yes, I did. I’m using this ‘trick’ to get you to continue reading because I’m going to say something you might not want to hear. Are you ready? Tools won’t replace experience, failure, hard work, and training. Just because you have the keys to the car, doesn’t mean you’re ready to head out on the race track. While SQL Developer reduces the barriers to entry, it does not completely remove them. Many experienced folks simply do not like tools. Rather, they don’t like the people that pick up tools without the know-how to properly use them. If you don’t understand what ‘TRUNCATE’ means, don’t try it out. Try picking up a book first. Of course, it’s very nice to have your own sandbox to play in, so you don’t upset the other children. That’s why I really like our Dev Days Database Virtual Box image. It’s your own database to learn and experiment with.

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• 8 Mac System Features You Can Access in Recovery Mode

  - by Chris Hoffman

  A Mac’s Recovery Mode is for more than just reinstalling Mac OS X. You’ll find many other useful troubleshooting utilities here — you can use these even if your Mac can’t boot normally. To access Recovery Mode, restart your Mac and press and hold the Command + R keys during the boot-up process. This is one of several hidden startup options on a Mac. Reinstall Mac OS X Most people know Recovery Mode as the place you go to reinstall OS X on your Mac. Recovery Mode will download the OS X installer files from teh Intenret if you don’t have them locally, so they don’t take up space on your disk and you’ll never have to hunt for an opearign system disc. Better yet, it will download up-to-date installation files so you don’t have to spend hours installing operating system updates later. Microsoft could learn a lot from Apple here. Restore From a Time Machine Backup Instead of reinstalling OS X, you can choose to restore your Mac from a time machine backup. This is like restoring a system image on another operating system. You’ll need an external disk containing a backup image created on the current computer to do this. Browse the Web The Get Help Online link opens the Safari web browser to Apple’s documentation site. It’s not limited to Apple’s website, though — you can navigate to any website you like. This feature allows you to access and use a browser on your Mac even if it isn’t booting properly. It’s ideal for looking up troubleshooting information. Manage Your Disks The Disk Utility option opens the same Disk Utility you can access from within Mac OS X. It allows you to partition disks, format them, scan disks for problems, wipe drives, and set up drives in a RAID configuration. If you need to edit partitions from outside your operating system, you can just boot into the recovery environment — you don’t have to download a special partitioning tool and boot into it. Choose the Default Startup Disk Click the Apple menu on the bar at the top of your screen and select Startup Disk to access the Choose Startup Disk tool. Use this tool to choose your computer’s default startup disk and reboot into another operating system. For example, it’s useful if you have Windows installed alongside Mac OS X with Boot Camp. Add or Remove an EFI Firmware Password You can also add a firmware password to your Mac. This works like a BIOS password or UEFI password on a Windows or Linux PC. Click the Utilities menu on the bar at the top of your screen and select Firmware Password Utility to open this tool. Use the tool to turn on a firmware password, which will prevent your computer from starting up from a different hard disk, CD, DVD, or USB drive without the password you provide. This prevents people form booting up your Mac with an unauthorized operating system. If you’ve already enabled a firmware password, you can remove it from here. Use Network Tools to Troubleshoot Your Connection Select Utilities > Network Utility to open a network diagnostic tool. This utility provides a graphical way to view your network connection information. You can also use the netstat, ping, lookup, traceroute, whois, finger, and port scan utilities from here. These can be helpful to troubleshoot Internet connection problems. For example, the ping command can demonstrate whether you can communicate with a remote host and show you if you’re experiencing packet loss, while the traceroute command can show you where a connection is failing if you can’t connect to a remote server. Open a Terminal If you’d like to get your hands dirty, you can select Utilities > Terminal to open a terminal from here. This terminal allows you to do more advanced troubleshooting. Mac OS X uses the bash shell, just as typical Linux distributions do. Most people will just need to use the Reinstall Mac OS X option here, but there are many other tools you can benefit from. If the Recovery Mode files on your Mac are damaged or unavailable, your Mac will automatically download them from Apple so you can use the full recovery environment.

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• Teamviewer: cannot control monitor 1, but can control monitor 2

  - by DaveT

  I'm using the web client of Teamviewer from my work computer trying to control my home computer. I have 2 monitors on the remote desktop, but for some reason only have control on the second monitor. When I switch to the main monitor (monitor 1), I cannot do anything and cannot even move the cursor. But I have no issues when I switch over to the second monitor (monitor 2). I used to have no issues with either, but in the past couple of months this has been causing me issues. Anyone have a suggestion? Thanks!! Also... Here is the log from the Teamviewer session. Showing me switching back and forth between the monitors. (just in case this will help). I had to remove the links in order to post the log since I don't have enough reputation points, but they were just teamviewer login weblinks. =============================================================================== 21.08 16:00:41,176: Version: 9.0.15099 21.08 16:00:41,177: Sandbox: remote 21.08 16:00:41,177: SysLanguage: en 21.08 16:00:41,177: VarLanguage: en 21.08 16:00:41,177: Flash Player: PlugIn (WIN 14,0,0,179) 21.08 16:00:41,178: UseLanguage: en 21.08 16:00:41,178: UseLanguage: en 21.08 16:00:41,182: TeamViewer hasPassword: true 21.08 16:00:41,418: ExternalConnect id=910035824 21.08 16:00:41,419: CT connect 910035824 masterURL: , sandbox = remote 21.08 16:00:41,425: MC.requestRoute(910035824) 21.08 16:00:41,426: MC.sendMasterCommand text=F=RequestRoute2&ID1=777&Client=TV& ID2=910035824&SA_AccountID=26641022&SA_PasswordMD5HashBase64Encoded=& SA_SessionSecret=f7H6Z7SYfX5ahQ7SJq/r/K20PBYg9fOZhp+DKLhf5ts=&SA_SessionID=1558929948& V=9.0.15099&OS=Flash 21.08 16:00:41,426: MC wait for ping completion 21.08 16:00:42,064: PS.socket event: [Event type="connect" bubbles=false cancelable=false eventPhase=2] 21.08 16:00:42,182: PingThread: TCP-Ping ok 21.08 16:00:42,183: MC.socket mode = TCP, MasterURL: 21.08 16:00:42,183: MC.connect: 21.08 16:00:43,058: PS.socket event: [Event type="connect" bubbles=false cancelable=false eventPhase=2] 21.08 16:00:43,058: MC.connectHandler: [Event type="connect" bubbles=false cancelable=false eventPhase=2] 21.08 16:00:43,236: MC.requestRouteResponse: [email protected]_10800_128000_762319420_910035824_10000__1_0_16778176_128000_16778176: 128000;2147483647:1280000;4:640000_786297_786297 21.08 16:00:43,239: CT init socket: TCP 21.08 16:00:43,513: PS.socket event: [Event type="connect" bubbles=false cancelable=false eventPhase=2] 21.08 16:00:43,514: CT.connectHandler: [Event type="connect" bubbles=false cancelable=false eventPhase=2] 21.08 16:00:43,519: Browser name: Netscape 21.08 16:00:43,936: CMD_IDENTIFY id=910035824 ver=2.41 21.08 16:00:44,666: CMD_CONFIRMENCRYPTION: encryption confirmed 21.08 16:00:44,667: Started resendrequest timer 21.08 16:00:45,063: Remote Version: TV 009.000 21.08 16:00:45,501: start classic authentication 21.08 16:00:45,502: Login::SendRequestToConsole(): url= 21.08 16:00:45,828: start srp authentication 21.08 16:00:46,983: checkFirstPacket ok, m_LastReceivedPacketID =4 21.08 16:00:47,148: Login::SendRequestToConsole(): url= 21.08 16:00:47,478: start srp authentication 21.08 16:00:48,210: Login::SendRequestToConsole(): url= 21.08 16:00:48,485: checkFirstPacket ok, m_LastReceivedPacketID =7 21.08 16:00:48,780: TVCmdAuthenticate_Authenticated: 1 21.08 16:00:49,321: Connected to 910035824, name=NEWMAN, os=14, version=9.0.31064 21.08 16:00:49,329: ConnectionAccessSettings: RemoteControl: AllowedFileTransfer: AllowedControlRemoteTV: AllowedSwitchSides: DeniedAllowDisableRemoteInput: AllowedAllowVPN: AllowedAllowPartnerViewDesktop: Allowed 21.08 16:00:52,195: unexpected TVCommand.CommandType == 56 21.08 16:00:52,231: CW received display params: 1680x1050x8 monitors: 2 (active:0) 21.08 16:00:52,301: Caching active, version=2 21.08 16:03:47,158: CW received display params: 1680x1050x8 monitors: 2 (active:1) 21.08 16:04:24,447: CW received display params: 1680x1050x8 monitors: 2 (active:0) 21.08 16:04:40,609: CW received display params: 3360x1050x8 monitors: 2 (active:-1) 21.08 16:04:59,802: CW received display params: 1680x1050x8 monitors: 2 (active:1) 21.08 16:04:59,933: CW received display params: 1680x1050x8 monitors: 2 (active:1) 21.08 16:05:58,419: CW received display params: 1680x1050x8 monitors: 2 (active:0) 21.08 16:06:36,824: CW received display params: 1680x1050x8 monitors: 2 (active:1) 21.08 16:07:07,232: CW received display params: 1680x1050x8 monitors: 2 (active:0)

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• Cloud Based Load Testing Using TF Service &amp; VS 2013

  - by Tarun Arora [Microsoft MVP]

  Originally posted on: http://geekswithblogs.net/TarunArora/archive/2013/06/30/cloud-based-load-testing-using-tf-service-amp-vs-2013.aspx One of the new features announced as part of the Visual Studio 2013 Ultimate Preview is ‘Cloud Based Load Testing’. In this blog post I’ll walk you through, What is Cloud Based Load Testing? How have I been using this feature? – Success story! Where can you find more resources on this feature? What is Cloud Based Load Testing? It goes without saying that performance testing your application not only gives you the confidence that the application will work under heavy levels of stress but also gives you the ability to test how scalable the architecture of your application is. It is important to know how much is too much for your application! Working with various clients in the industry I have realized that the biggest barriers in Load Testing & Performance Testing adoption are, High infrastructure and administration cost that comes with this phase of testing Time taken to procure & set up the test infrastructure Finding use for this infrastructure investment after completion of testing Is cloud the answer? 100% Visual Studio Compatible Scalable and Realistic Start testing in < 2 minutes Intuitive Pay only for what you need Use existing on premise tests on cloud There are a lot of vendors out there offering Cloud Based Load Testing, to name a few, Load Storm Soasta Blaze Meter Blitz And others… The question you may want to ask is, why should you go with Microsoft’s Cloud based Load Test offering. If you are a Microsoft shop or already have investments in Microsoft technologies, you’ll see great benefit in the natural integration this offers with existing Microsoft products such as Visual Studio and Windows Azure. For example, your existing Web tests authored in Visual Studio 2010 or Visual Studio 2012 will run on the cloud without requiring any modifications what so ever. Microsoft’s cloud test rig also supports API based testing, for example, if you are building a WPF application which consumes WCF services, you can write unit tests to invoke the WCF service, these tests can be run on the cloud test rig and loaded with ‘N’ concurrent users for performance testing. If you have your assets already hosted in the Azure and possibly in the same data centre as the Cloud test rig, your Azure app will not incur a usage cost because of the generated traffic since the traffic is coming from the same data centre. The licensing or pricing information on Microsoft’s cloud based Load test service is yet to be announced, but I would expect this to be priced attractively to match the market competition.   The only additional configuration required for running load tests on Microsoft Cloud based Load Tests service is to select the Test run location as Run tests using Visual Studio Team Foundation Service, How have I been using Microsoft’s Cloud based Load Test Service? I have been part of the Microsoft Cloud Based Load Test Service advisory council for the last 7 months. This gave the opportunity to see the product shape up from concept to working solution. I was also the first person outside of Microsoft to try this offering out. This gave me the opportunity to test real world application at various clients using the Microsoft Load Test Service and provide real world feedback to the Microsoft product team. One of the most recent systems I tested using the Load Test Service has been an insurance quote generation engine. This insurance quote generation engine is,   hosted in Windows Azure expected to get quote requests from across the globe expected to handle 5 Million quote requests in a day (not clear how this load will be distributed across the day) There was no way, I could simulate such kind of load from on premise without standing up additional hardware. But Microsoft’s Cloud based Load Test service allowed me to test my key performance testing scenarios, i.e. Simulate expected Load, Endurance Testing, Threshold Testing and Testing for Latency. Simulating expected load: approach to devising a load pattern My approach to devising a load test pattern has been to run the test scenario with 1 user to figure out the response time. Then work out how many users are required to reach the target load. So, for example, to invoke 1 quote from the quote engine software takes 0.5 seconds. Now if you do the math,   1 quote request by 1 user = 0.5 seconds   quotes generated by 1 user in 24 hour = 1 * (((2 * 60) * 60) * 24) = 172,800   quotes generated by 30 users in 24 hours = 172,800 * 30 =  5,184,000 This was a very simple example, if your application requires more concurrent users to test scenario’s such as caching, etc then you can devise your own load pattern, some examples of load test patterns can be found here.  Endurance Testing To test for endurance, I loaded the quote generation engine with an expected fixed user load and ran the test for very long duration such as over 48 hours and observed the affect of the long running test on the Azure infrastructure. Currently Microsoft Load Test service does not support metrics from the machine under test. I used Azure diagnostics to begin with, but later started using Cerebrata Azure Diagnostics Manager to capture the metrics of the machine under test. Threshold Testing To figure out how much user load the application could cope with before falling on its belly, I opted to step load the quote generation engine by incrementing user load with different variations of incremental user load per minute till the application crashed out and forced an IIS reset. Testing for Latency Currently the Microsoft Load Test service does not support generating geographically distributed load, I however, deployed the insurance quote generation engine in different Azure data centres and ran the same set of performance tests to measure for latency. Because I could compare load test results from different runs by exporting the results to excel (this feature is provided out of the box right from Visual Studio 2010) I could see the different in response times. More resources on Microsoft Cloud based Load Test Service A few important links to get you started, Download Visual Studio Ultimate 2013 Preview Getting started guide for load testing using Team Foundation Service Troubleshooting guide for FAQs and known issues Team Foundation Service forum for questions and support Detailed demo and presentation (link to Tech-Ed session recording) Detailed demo and presentation (link to Build session recording) There a few limits on the usage of Microsoft Cloud based Load Test service that you can read about here. If you have any feedback on Microsoft Cloud based Load Test service, feel free to share it with the product team via the Visual Studio User Voice forum. I hope you found this useful. Thank you for taking the time out and reading this blog post. If you enjoyed the post, remember to subscribe to http://feeds.feedburner.com/TarunArora. Stay tuned!

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