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• How to generate a Program template by generating an abstract class

  - by Byron-Lim Timothy Steffan

  i have the following problem. The 1st step is to implement a program, which follows a specific protocol on startup. Therefore, functions as onInit, onConfigRequest, etc. will be necessary. (These are triggered e.g. by incoming message on a TCP Port) My goal is to generate a class for example abstract one, which has abstract functions as onInit(), etc. A programmer should just inherit from this base class and should merely override these abstract functions of the base class. The rest as of the protocol e.g. should be simply handled in the background (using the code of the base class) and should not need to appear in the programmers code. What is the correct design strategy for such tasks? and how do I deal with, that the static main method is not inheritable? What are the key-tags for this problem? (I have problem searching for a solution since I lack clear statements on this problem) Goal is to create some sort of library/class, which - included in ones code - results in executables following the protocol. EDIT (new explanation): Okay let me try to explain more detailled: In this case programs should be clients within a client server architecture. We have a client server connection via TCP/IP. Each program needs to follow a specific protocol upon program start: As soon as my program starts and gets connected to the server it will receive an Init Message (TcpClient), when this happens it should trigger the function onInit(). (Should this be implemented by an event system?) After onInit() a acknowledgement message should be sent to the server. Afterwards there are some other steps as e.g. a config message from the server which triggers an onConfig and so on. Let's concentrate on the onInit function. The idea is, that onInit (and onConfig and so on) should be the only functions the programmer should edit while the overall protocol messaging is hidden for him. Therefore, I thought using an abstract class with the abstract methods onInit(), onConfig() in it should be the right thing. The static Main class I would like to hide, since within it e.g. there will be some part which connects to the tcp port, which reacts on the Init Message and which will call the onInit function. 2 problems here: 1. the static main class cant be inherited, isn it? 2. I cannot call abstract functions from the main class in the abstract master class. Let me give an Pseudo-example for my ideas: public abstract class MasterClass { static void Main(string[] args){ 1. open TCP connection 2. waiting for Init Message from server 3. onInit(); 4. Send Acknowledgement, that Init Routine has ended successfully 5. waiting for Config message from server 6..... } public abstract void onInit(); public abstract void onConfig(); } I hope you get the idea now! The programmer should afterwards inherit from this masterclass and merely need to edit the functions onInit and so on. Is this way possible? How? What else do you recommend for solving this? EDIT: The strategy ideo provided below is a good one! Check out my comment on that.

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• FreeBSD high load loopback interface

  - by user1740915

  I have a problem with a FreeBSD server. There is a FreeBSD 9.0 amd64, two network cards em1 (internet), em0 (local network) configured firewall ipfw, natd, squid (not transparent), the server acts as a gateway for access to the Internet. Next problem: upload via squid is very low. At this moment I see next: natd, dhcpd load the cpu at that time when uploading through squid and there are a lot of traffic through the loopback interface. ipfw show output 0100 655389684 36707144666 allow ip from any to any via lo0 00200 0 0 deny ip from any to 127.0.0.0/8 00300 0 0 deny ip from 127.0.0.0/8 to any 00400 0 0 deny ip from any to ::1 00500 0 0 deny ip from ::1 to any 00600 4 292 allow ipv6-icmp from :: to ff02::/16 00700 0 0 allow ipv6-icmp from fe80::/10 to fe80::/10 00800 1 76 allow ipv6-icmp from fe80::/10 to ff02::/16 00900 0 0 allow ipv6-icmp from any to any ip6 icmp6types 1 01000 0 0 allow ipv6-icmp from any to any ip6 icmp6types 2,135,136 01100 1615 76160 deny ip from 192.168.1.1 to any in via em1 01200 0 0 deny ip from 199.69.99.11 to any in via em0 01300 46652 3705426 deny ip from any to 172.16.0.0/12 via em1 01400 3936404 345618870 deny ip from any to 192.168.0.0/16 via em1 01500 4 336 deny ip from any to 0.0.0.0/8 via em1 01600 4129 387621 deny ip from any to 169.254.0.0/16 via em1 01700 0 0 deny ip from any to 192.0.2.0/24 via em1 01800 917566 33777571 deny ip from any to 224.0.0.0/4 via em1 01900 147872 22029252 deny ip from any to 240.0.0.0/4 via em1 02000 1132194739 1190981955947 divert 8668 ip4 from any to any via em1 02100 3 248 deny ip from 172.16.0.0/12 to any via em1 02200 35925 2281289 deny ip from 192.168.0.0/16 to any via em1 02300 1808 122494 deny ip from 0.0.0.0/8 to any via em1 02400 3 174 deny ip from 169.254.0.0/16 to any via em1 02500 0 0 deny ip from 192.0.2.0/24 to any via em1 02600 0 0 deny ip from 224.0.0.0/4 to any via em1 02700 0 0 deny ip from 240.0.0.0/4 to any via em1 02800 960156249 1095316736582 allow tcp from any to any established 02900 64236062 8243196577 allow ip from any to any frag 03000 34 1756 allow tcp from any to me dst-port 25 setup 03100 193 11580 allow tcp from any to me dst-port 53 setup 03200 63 4222 allow udp from any to me dst-port 53 03300 64 8350 allow udp from me 53 to any 03400 417 24140 allow tcp from any to me dst-port 80 setup 03500 211 10472 allow ip from any to me dst-port 3389 setup 05300 77 4488 allow ip from any to me dst-port 1723 setup 05400 3 156 allow ip from any to me dst-port 8443 setup 05500 9882 590596 allow tcp from any to me dst-port 22 setup 05600 1 60 allow ip from any to me dst-port 2000 setup 05700 0 0 allow ip from any to me dst-port 2201 setup 07400 4241779 216690096 deny log logamount 1000 ip4 from any to any in via em1 setup proto tcp 07500 21135656 1048824936 allow tcp from any to any setup 07600 474447 35298081 allow udp from me to any dst-port 53 keep-state 07700 532 40612 allow udp from me to any dst-port 123 keep-state 65535 1990638432 1122305322718 allow ip from any to any systat -ifstat when uploading via squid Load Average ||| Interface Traffic Peak Total tun0 in 79.507 KB/s 232.479 KB/s 42.314 GB out 2.022 MB/s 2.424 MB/s 59.662 GB lo0 in 4.450 MB/s 4.450 MB/s 43.723 GB out 4.450 MB/s 4.450 MB/s 43.723 GB em1 in 2.629 MB/s 2.982 MB/s 464.533 GB out 2.493 MB/s 2.875 MB/s 484.673 GB em0 in 240.458 KB/s 296.941 KB/s 442.368 GB out 512.508 KB/s 850.857 KB/s 416.122 GB top output PID USERNAME THR PRI NICE SIZE RES STATE C TIME WCPU COMMAND 66885 root 1 92 0 26672K 2784K CPU3 3 528:43 65.48% natd 9160 dhcpd 1 45 0 31032K 9280K CPU1 1 7:40 32.96% dhcpd 66455 root 1 20 0 18344K 2856K select 1 119:27 1.37% openvpn 16043 squid 1 20 0 44404K 17884K kqread 2 0:22 0.29% squid squid.conf cat /usr/local/etc/squid/squid.conf # # Recommended minimum configuration: # acl manager proto cache_object acl localhost src 127.0.0.1/32 ::1 acl to_localhost dst 127.0.0.0/8 0.0.0.0/32 ::1 # Example rule allowing access from your local networks. # Adapt to list your (internal) IP networks from where browsing # should be allowed acl localnet src 10.0.0.0/8 # RFC1918 possible internal network acl localnet src 172.16.0.0/12 # RFC1918 possible internal network acl localnet src 192.168.0.0/16 # RFC1918 possible internal network acl localnet src fc00::/7 # RFC 4193 local private network range acl localnet src fe80::/10 # RFC 4291 link-local (directly plugged) machines acl SSL_ports port 443 acl Safe_ports port 80 # http acl Safe_ports port 21 # ftp acl Safe_ports port 443 # https acl Safe_ports port 70 # gopher acl Safe_ports port 210 # wais acl Safe_ports port 1025-65535 # unregistered ports acl Safe_ports port 280 # http-mgmt acl Safe_ports port 488 # gss-http acl Safe_ports port 591 # filemaker acl Safe_ports port 777 # multiling http acl CONNECT method CONNECT # # Recommended minimum Access Permission configuration: # # Only allow cachemgr access from localhost http_access allow manager localhost http_access deny manager # Deny requests to certain unsafe ports http_access deny !Safe_ports # Deny CONNECT to other than secure SSL ports http_access deny CONNECT !SSL_ports # We strongly recommend the following be uncommented to protect innocent # web applications running on the proxy server who think the only # one who can access services on "localhost" is a local user http_access deny to_localhost # # INSERT YOUR OWN RULE(S) HERE TO ALLOW ACCESS FROM YOUR CLIENTS # # Example rule allowing access from your local networks. # Adapt localnet in the ACL section to list your (internal) IP networks # from where browsing should be allowed http_access allow localnet http_access allow localhost # And finally deny all other access to this proxy http_access deny all # Squid normally listens to port 3128 http_port 192.168.1.1:3128 # Uncomment and adjust the following to add a disk cache directory. #cache_dir ufs /var/squid/cache 100 16 256 # Leave coredumps in the first cache dir coredump_dir /var/squid/cache I understand that the traffic passes through the SQUID several times. But can not find why.

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• Openfire and LDAP issues

  - by clsmith

  Thanks in advance for the help. Has anyone see this issue with openfire? Currently I use Openfire Fedora with Auth using windows 2003 and also use mysql for the database. When I bring up two clients and talk to each other the time is slow between messages. Sometimes it can take between 5-15 minutes for something sent to get to the person (this is with only two people on the openfire server). I ran a tcp dump using port 389 and see that the machine is running thousands of queries against ldap. When i plug it into wireshark I notice that it is transferring the entire contact list or checking on the status of the entire contact list ? When I run debug on openfire itself I am presented with only this small message in the log: 2010.06.08 07:01:17 LdapManager: Starting LDAP search... 2010.06.08 07:01:17 LdapManager: ... search finished 2010.06.08 07:01:17 LdapManager: Creating a DirContext in LdapManager.getContext()... 2010.06.08 07:01:17 LdapManager: Created hashtable with context values, attempting to create context... 2010.06.08 07:01:17 LdapManager: ... context created successfully, returning. 2010.06.08 07:01:17 LdapManager: Trying to find a groups's DN based on it's groupname. cn: Spark agents CLT, Base DN: OU="Hidden",DC="Hidden",DC="net"... 2010.06.08 07:01:17 LdapManager: Creating a DirContext in LdapManager.getContext()... 2010.06.08 07:01:17 LdapManager: Created hashtable with context values, attempting to create context... 2010.06.08 07:01:17 LdapManager: ... context created successfully, returning. 2010.06.08 07:01:17 LdapManager: Starting LDAP search... 2010.06.08 07:01:17 LdapManager: ... search finished 2010.06.08 07:01:17 LdapManager: Trying to find a groups's DN based on it's groupname. cn: Spark agents CLT, Base DN: OU="Hidden",DC="Hidden",DC="net"... 2010.06.08 07:01:17 LdapManager: Creating a DirContext in LdapManager.getContext()... 2010.06.08 07:01:17 LdapManager: Created hashtable with context values, attempting to create context... 2010.06.08 07:01:17 LdapManager: ... context created successfully, returning. 2010.06.08 07:01:17 LdapManager: Starting LDAP search... 2010.06.08 07:01:17 LdapManager: ... search finished I thought this was a configuration on my end and started to look into the cache settings on the openfire webpages. I tweaked the settings as recommend by the pages and still get the same issues. I doesnt seem to cache the contact list or this might be a feature never fixed or implemented. Has anyone gone through this before ? I have searched online and I see others have great experience with openfire with no issues like I have, or is it because noone checked the queries ? For the time being I created a new Domain Controller and moved openfire to that computer so it can run local queries. This seems to help reduce the speed alot, but when I run the server performance manager tool I see that with two people only using that openfire server I run 593.7 request per second. Thanks for your help, if I didnt provide enough data please let me know what you need and I can find it.

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• Help needed with drawRect:

  - by Andrew Coad

  Hi, I'm having a fundamental issue with use of drawRect: Any advice would be greatly appreciated. The application needs to draw a variety of .png images at different times, sometimes with animation, sometimes without. A design goal that I was hoping to adhere to is to have the code inside drawRect: very simple and "dumb" - i.e. just do drawing and no other application logic. To draw the image I am using the drawAtPoint: method of UIImage. Since this method does not take a CGContext as a parameter, it can only be called within the drawRect: method. So I have: - (void)drawRect:(CGRect)rect { [firstImage drawAtPoint:CGPointMake(firstOffsetX, firstOffsetY)]; } All fine and dandy for one image. To draw multiple images (over time) the approach I have taken is to maintain an array of dictionaries with each dictionary containing an image, the point location to draw at and a flag to enable/suppress drawing for that image. I add dictionaries to the array over time and trigger drawing via the setNeedsDisplay: method of UIView. Use of an array of dictionaries allows me to completely reconstruct the entire display at any time. drawRect: now becomes: - (void)drawRect:(CGRect)rect { for (NSMutableDictionary *imageDict in [self imageDisplayList]) { if ([[imageDict objectForKey:@"needsDisplay"] boolValue]) { [[imageDict objectForKey:@"image"] drawAtPoint:[[imageDict objectForKey:@"location"] CGPointValue]]; [imageDict setValue:[NSNumber numberWithBool:NO] forKey:@"needsDisplay"]; } } } Still OK. The code is simple and compact. Animating this is where I run into problems. The first problem is where do I put the animation code? Do I put it in UIView or UIViewController? If in UIView, do I put it in drawRect: or elsewhere? Because the actual animation depends on the overall state of the application, I would need nested switch statements which, if put in drawRect:, would look something like this: - (void)drawRect:(CGRect)rect { for (NSMutableDictionary *imageDict in [self imageDisplayList]) { if ([[imageDict objectForKey:@"needsDisplay"] boolValue]) { switch ([self currentState]) { case STATE_1: switch ([[imageDict objectForKey:@"animationID"] intValue]) { case ANIMATE_FADE_IN: [self setAlpha:0.0]; [UIView beginAnimations:[[imageDict objectForKey:@"animationID"] intValue] context:nil]; [UIView setAnimationDelegate:self]; [UIView setAnimationCurve:UIViewAnimationCurveEaseIn]; [UIView setAnimationDuration:2]; [self setAlpha:1.0]; break; case ANIMATE_FADE_OUT: [self setAlpha:1.0]; [UIView beginAnimations:[[imageDict objectForKey:@"animationID"] intValue] context:nil]; [UIView setAnimationDelegate:self]; [UIView setAnimationCurve:UIViewAnimationCurveEaseOut]; [UIView setAnimationDuration:2]; [self setAlpha:0.0]; break; case ANIMATE_OTHER: // similar code here break; default: break; } break; case STATE_2: // similar code here break; default: break; } [[imageDict objectForKey:@"image"] drawAtPoint:[[imageDict objectForKey:@"location"] CGPointValue]]; [imageDict setValue:[NSNumber numberWithBool:NO] forKey:@"needsDisplay"]; } } [UIView commitAnimations]; } In addition, to make multiple sequential animations work correctly, there would need to be an outer controlling mechanism involving the animation delegate animationDidStop: callback that would set the needsDisplay entries in the dictionaries to allow/suppress drawing (and animation). The point that we are at now is that it all starts to look very ugly. More specifically: drawRect: starts to bloat quickly and contain code that is not "just drawing" code the UIView needs implicit awareness of the application state the overall process of drawing is now spread across three methods at a minimum And on to the point of this post: how can I do this better? What would the experts out there recommend in terms of overall structure? How can I keep application state information out of the view? Am I looking at this problem from the wrong direction. Is there some completely different approach that I should consider?

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• Opinions on Dual-Salt authentication for low sensitivity user accounts?

  - by Heleon

  EDIT - Might be useful for someone in the future... Looking around the bcrypt class in php a little more, I think I understand what's going on, and why bcrypt is secure. In essence, I create a random blowfish salt, which contains the number of crypt rounds to perform during the encryption step, which is then hashed using the crypt() function in php. There is no need for me to store the salt I used in the database, because it's not directly needed to decrypt, and the only way to gain a password match to an email address (without knowing the salt values or number of rounds) would be to brute force plain text passwords against the hash stored in the database using the crypt() function to verify, which, if you've got a strong password, would just be more effort than it's worth for the user information i'm storing... I am currently working on a web project requiring user accounts. The application is CodeIgniter on the server side, so I am using Ion Auth as the authentication library. I have written an authentication system before, where I used 2 salts to secure the passwords. One was a server-wide salt which sat as an environment variable in the .htaccess file, and the other was a randomly generated salt which was created at user signup. This was the method I used in that authentication system for hashing the password: $chars = "abcdefghijklmnopqrstuvwxyzABCDEFGHIJKLMNOPQRSTUVWXYZ0123456789"; //create a random string to be used as the random salt for the password hash $size = strlen($chars); for($i = 0; $i < 22; $i++) { $str .= $chars[rand(0, $size - 1)]; } //create the random salt to be used for the crypt $r_blowfish_salt = "$2a$12$" . $str . "$"; //grab the website salt $salt = getenv('WEBSITE_SALT'); //combine the website salt, and the password $password_to_hash = $pwd . $salt; //crypt the password string using blowfish $password = crypt($password_to_hash, $r_blowfish_salt); I have no idea whether this has holes in it or not, but regardless, I moved over to Ion Auth for a more complete set of functions to use with CI. I noticed that Ion only uses a single salt as part of its hashing mechanism (although does recommend that encryption_key is set in order to secure the database session.) The information that will be stored in my database is things like name, email address, location by country, some notes (which will be recommended that they do not contain sensitive information), and a link to a Facebook, Twitter or Flickr account. Based on this, i'm not convinced it's necessary for me to have an SSL connection on the secure pages of my site. My question is, is there a particular reason why only 1 salt is being used as part as the Ion Auth library? Is it implied that I write my own additional salting in front of the functionality it provides, or am I missing something? Furthermore, is it even worth using 2 salts, or once an attacker has the random salt and the hashed password, are all bets off anyway? (I assume not, but worth checking if i'm worrying about nothing...)

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• evaluating cost/benefits of using extension methods in C# => 3.0

  - by BillW

  Hi, In what circumstances (usage scenarios) would you choose to write an extension rather than sub-classing an object ? < full disclosure : I am not an MS employee; I do not know Mitsu Furota personally; I do know the author of the open-source Componax library mentioned here, but I have no business dealings with him whatsoever; I am not creating, or planning to create any commercial product using extensions : in sum : this post is from pure intellectal curiousity related to my trying to (continually) become aware of "best practices" I find the idea of extension methods "cool," and obviously you can do "far-out" things with them as in the many examples you can in Mitsu Furota's (MS) blog postslink text. A personal friend wrote the open-source Componax librarylink text, and there's some remarkable facilities in there; but he is in complete command of his small company with total control over code guidelines, and every line of code "passes through his hands." While this is speculation on my part : I think/guess other issues might come into play in a medium-to-large software team situation re use of Extensions. Looking at MS's guidelines at link text, you find : In general, you will probably be calling extension methods far more often than implementing your own. ... In general, we recommend that you implement extension methods sparingly and only when you have to. Whenever possible, client code that must extend an existing type should do so by creating a new type derived from the existing type. For more information, see Inheritance (C# Programming Guide). ... When the compiler encounters a method invocation, it first looks for a match in the type's instance methods. If no match is found, it will search for any extension methods that are defined for the type, and bind to the first extension method that it finds. And at Ms's link text : Extension methods present no specific security vulnerabilities. They can never be used to impersonate existing methods on a type, because all name collisions are resolved in favor of the instance or static method defined by the type itself. Extension methods cannot access any private data in the extended class. Factors that seem obvious to me would include : I assume you would not write an extension unless you expected it be used very generally and very frequently. On the other hand : couldn't you say the same thing about sub-classing ? Knowing we can compile them into a seperate dll, and add the compiled dll, and reference it, and then use the extensions : is "cool," but does that "balance out" the cost inherent in the compiler first having to check to see if instance methods are defined as described above. Or the cost, in case of a "name clash," of using the Static invocation methods to make sure your extension is invoked rather than the instance definition ? How frequent use of Extensions would affect run-time performance or memory use : I have no idea. So, I'd appreciate your thoughts, or knowing about how/when you do, or don't do, use Extensions, compared to sub-classing. thanks, Bill

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• CentOS - Configuring Puppet to play nice with SELinux

  - by Mike Purcell

  I am running into an issue every time I attempt to start the puppetmasterd service, for which I receive the following error message: root@service1 ~ # -> /etc/init.d/puppetmaster start Starting puppetmaster: Could not prepare for execution: Got 1 failure(s) while initializing: change from absent to directory failed: Could not set 'directory on ensure: Permission denied - /etc/puppet/ssl [FAILED] Apparently there was a known issue with this scenario as outlined in this bug report, however in the bug report it states the issue has been resolved in selinux-policy-3.9.16-29.fc15, but the latest CentOS default upstream version is 3.7.19-155.el6_3.4. So I am trying to figure out the best solution. I can either create a local security policy to allow puppetmasterd the access it needs, or keep researching and install a newer version of selinux-policy outside of the default upstream channel. Anyone have any recommendations? Please don't recommend disabling SELinux... ----- Update ----- Here is the puppet.conf: [main] # The Puppet log directory. # The default value is '$vardir/log'. logdir = /var/log/puppet # Where Puppet PID files are kept. # The default value is '$vardir/run'. rundir = /var/run/puppet # Where SSL certificates are kept. # The default value is '$confdir/ssl'. ssldir = $vardir/ssl [master] certname=puppetmaster.ownij.lan dns_alt_names=puppetmaster.ownij.lan [agent] # The file in which puppetd stores a list of the classes # associated with the retrieved configuratiion. Can be loaded in # the separate ``puppet`` executable using the ``--loadclasses`` # option. # The default value is '$confdir/classes.txt'. classfile = $vardir/classes.txt # Where puppetd caches the local configuration. An # extension indicating the cache format is added automatically. # The default value is '$confdir/localconfig'. localconfig = $vardir/localconfig server=puppetmaster.ownij.lan And here are the denials per the audit log: type=AVC msg=audit(1349751364.985:666): avc: denied { search } for pid=15093 comm="puppetmasterd" name="/" dev=dm-2 ino=2 scontext=unconfined_u:system_r:puppetmaster_t:s0 tcontext=system_u:object_r:home_root_t:s0 tclass=dir type=SYSCALL msg=audit(1349751364.985:666): arch=c000003e syscall=4 success=no exit=-13 a0=1391420 a1=7fffef09ed10 a2=7fffef09ed10 a3=120c500 items=0 ppid=15092 pid=15093 auid=500 uid=0 gid=0 euid=0 suid=0 fsuid=0 egid=0 sgid=0 fsgid=0 tty=pts1 ses=13 comm="puppetmasterd" exe="/usr/bin/ruby" subj=unconfined_u:system_r:puppetmaster_t:s0 key=(null) type=AVC msg=audit(1349751365.302:667): avc: denied { search } for pid=15093 comm="puppetmasterd" name="/" dev=dm-2 ino=2 scontext=unconfined_u:system_r:puppetmaster_t:s0 tcontext=system_u:object_r:home_root_t:s0 tclass=dir type=SYSCALL msg=audit(1349751365.302:667): arch=c000003e syscall=4 success=no exit=-13 a0=1d18530 a1=7fffef0d04d0 a2=7fffef0d04d0 a3=8 items=0 ppid=15092 pid=15093 auid=500 uid=0 gid=0 euid=0 suid=0 fsuid=0 egid=0 sgid=0 fsgid=0 tty=pts1 ses=13 comm="puppetmasterd" exe="/usr/bin/ruby" subj=unconfined_u:system_r:puppetmaster_t:s0 key=(null) type=AVC msg=audit(1349751365.465:668): avc: denied { search } for pid=15093 comm="puppetmasterd" name="/" dev=dm-2 ino=2 scontext=unconfined_u:system_r:puppetmaster_t:s0 tcontext=system_u:object_r:home_root_t:s0 tclass=dir type=SYSCALL msg=audit(1349751365.465:668): arch=c000003e syscall=4 success=no exit=-13 a0=1af3930 a1=7fffef0c5c70 a2=7fffef0c5c70 a3=8 items=0 ppid=15092 pid=15093 auid=500 uid=0 gid=0 euid=0 suid=0 fsuid=0 egid=0 sgid=0 fsgid=0 tty=pts1 ses=13 comm="puppetmasterd" exe="/usr/bin/ruby" subj=unconfined_u:system_r:puppetmaster_t:s0 key=(null) type=AVC msg=audit(1349751365.467:669): avc: denied { search } for pid=15093 comm="puppetmasterd" name="/" dev=dm-2 ino=2 scontext=unconfined_u:system_r:puppetmaster_t:s0 tcontext=system_u:object_r:home_root_t:s0 tclass=dir type=SYSCALL msg=audit(1349751365.467:669): arch=c000003e syscall=4 success=no exit=-13 a0=1b17aa0 a1=7fffef0c5c70 a2=7fffef0c5c70 a3=8 items=0 ppid=15092 pid=15093 auid=500 uid=0 gid=0 euid=0 suid=0 fsuid=0 egid=0 sgid=0 fsgid=0 tty=pts1 ses=13 comm="puppetmasterd" exe="/usr/bin/ruby" subj=unconfined_u:system_r:puppetmaster_t:s0 key=(null) type=AVC msg=audit(1349751366.401:670): avc: denied { write } for pid=15093 comm="puppetmasterd" name="puppet" dev=dm-0 ino=132035 scontext=unconfined_u:system_r:puppetmaster_t:s0 tcontext=system_u:object_r:puppet_etc_t:s0 tclass=dir type=SYSCALL msg=audit(1349751366.401:670): arch=c000003e syscall=83 success=no exit=-13 a0=2d7a400 a1=1f9 a2=2d7a40f a3=7fffef0a6df0 items=0 ppid=15092 pid=15093 auid=500 uid=0 gid=0 euid=0 suid=0 fsuid=0 egid=0 sgid=0 fsgid=0 tty=pts1 ses=13 comm="puppetmasterd" exe="/usr/bin/ruby" subj=unconfined_u:system_r:puppetmaster_t:s0 key=(null) And the audit log if I pass through audit2allow: root@service1 ~ # -> fgrep puppetmasterd /var/log/audit/audit.log | audit2allow -m puppetmasterd module puppetmasterd 1.0; require { type home_root_t; type puppetmaster_t; type puppet_etc_t; type puppet_var_run_t; type httpd_sys_content_t; class lnk_file { relabelfrom relabelto }; class file { relabelfrom read getattr open }; class dir { write read search getattr setattr }; } #============= puppetmaster_t ============== allow puppetmaster_t home_root_t:dir { search getattr }; allow puppetmaster_t httpd_sys_content_t:dir read; allow puppetmaster_t httpd_sys_content_t:file { read getattr open }; #!!!! The source type 'puppetmaster_t' can write to a 'dir' of the following types: # puppet_log_t, puppet_var_lib_t, puppet_var_run_t, puppetmaster_tmp_t allow puppetmaster_t puppet_etc_t:dir { write setattr }; allow puppetmaster_t puppet_etc_t:lnk_file { relabelfrom relabelto }; allow puppetmaster_t puppet_var_run_t:file relabelfrom;

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• How can I disable 'output escaping' in minidom

  - by William

  I'm trying to build an xml document from scratch using xml.dom.minidom. Everything was going well until I tried to make a text node with a ® (Registered Trademark) symbol in. My objective is for when I finally hit print mydoc.toxml() this particular node will actually contain a ® symbol. First I tried: import xml.dom.minidom as mdom data = '®' which gives the rather obvious error of: File "C:\src\python\HTMLGen\test2.py", line 3 SyntaxError: Non-ASCII character '\xae' in file C:\src\python\HTMLGen\test2.py on line 3, but no encoding declared; see http://www.python.or g/peps/pep-0263.html for details I have of course also tried changing the encoding of my python script to 'utf-8' using the opening line comment method, but this didn't help. So I thought import xml.dom.minidom as mdom data = '&#174;' #Both accepted xml encodings for registered trademark data = '&reg;' text = mdom.Text() text.data = data print data print text.toxml() But because when I print text.toxml(), the ampersands are being escaped, I get this output: &reg; &amp;reg; My question is, does anybody know of a way that I can force the ampersands not to be escaped in the output, so that I can have my special character reference carry through to the XML document? Basically, for this node, I want print text.toxml() to produce output of &reg; or &#174; in a happy and cooperative way! EDIT 1: By the way, if minidom actually doesn't have this capacity, I am perfectly happy using another module that you can recommend which does. EDIT 2: As Hugh suggested, I tried using data = u'®' (while also using data # -*- coding: utf-8 -*- Python source tags). This almost helped in the sense that it actually caused the ® symbol itself to be outputted to my xml. This is actually not the result I am looking for. As you may have guessed by now (and perhaps I should have specified earlier) this xml document happens to be an HTML page, which needs to work in a browser. So having ® in the document ends up causing rubbish in the browser (® to be precise!). I also tried: data = unichr(174) text.data = data.encode('ascii','xmlcharrefreplace') print text.toxml() But of course this lead to the same origional problem where all that happens is the ampersand gets escaped by .toxml(). My ideal scenario would be some way of escaping the ampersand so that the XML printing function won't "escape" it on my behalf for the document (in other words, achieving my original goal of having &reg; or &#174; appear in the document). Seems like soon I'm going to have to resort to regular expressions! EDIT 2a: Or perhaps not. Seems like getting my html meta information correct <META http-equiv="Content-Type" Content="text/html; charset=UTF-8"> could help, but I'm not sure yet how this fits in with the xml structure...

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• Recommended integration mechanism for bi-directional, authenticated, encrypted connection in C clien

  - by rcampbell

  Let me first give an example. Imagine you have a single server running a JVM application. This server keeps a collection of N equations, once for each client: Client #1: 2x Client #2: 1 + y Client #3: z/4 This server includes an HTTP interface so that random visitors can type https://www.acme.com/client/3 int their browsers and see the latest evaluated result of z/4. The tricky part is that either the client or the server may change the variable value at any time, informing the other party immediately. More specifically, Client #3 - a C app - can initially tell the server that z = 20. An hour later that same client informs the server that z = 23. Likewise the server can later inform the client that z = 28. As caf pointed out in the comments, there can be a race condition when values are changed by the client and server simultaneously. The solution would be for both client and server to send the operation performed in their message, which would need to be executed by the other party. To keep things simple, let's limit the operations to (commutative) addition, allowing us to disregard message ordering. For example, the client seeds the server with z = 20: server:z=20, client:z=20 server sends {+3} message (so z=23 locally) & client sends {-2} message (so z=18 locally) at the exact same time server receives {-2} message at some point, adds to his local copy so z=21 client receives {+3} message at some point, adds to his local copy so z=21 As long as all messages are eventually evaluated by both parties, the correct answer will eventually be given to the users of the client and server since we limited ourselves to commutative operations (addition of 3 and -2). This does mean that both client and server can be returning incorrect answers in the time it takes for messages to be exchanged and processed. While undesirable, I believe this is unavoidable. Some possible implementations of this idea include: Open an encrypted, always on TCP socket connection for communication Pros: no additional infrastructure needed, client and server know immediately if there is a problem (disconnect) with the other party, fairly straightforward (except the the encryption), native support from both JVM and C platforms Cons: pretty low-level so you end up writing a lot yourself (protocol, delivery verification, retry-on-failure logic), probably have a lot of firewall headaches during client app installation Asynchronous messaging (ex: ActiveMQ) Pros: transactional, both C & Java integration, free up the client and server apps from needing retry logic or delivery verification, pretty straightforward encryption, easy extensibility via message filters/routers/etc Cons: need additional infrastructure (message server) which must never fail, Database or file system as asynchronous integration point Same pros/cons as above but messier RESTful Web Service Pros: simple, possible reuse of the server's existing REST API, SSL figures out the encryption problem for you (maybe use RSA key a la GitHub for authentication?) Cons: Client now needs to run a C HTTP REST server w/SSL, client and server need retry logic. Axis2 has both a Java and C version, but you may be limited to SOAP. What other techniques should I be evaluating? What real world experiences have you had with these mechanisms? Which do you recommend for this problem and why?

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• MySQL access classes in PHP

  - by Mike

  I have a connection class for MySQL that looks like this: class MySQLConnect { private $connection; private static $instances = 0; function __construct() { if(MySQLConnect::$instances == 0) { //Connect to MySQL server $this->connection = mysql_connect(MySQLConfig::HOST, MySQLConfig::USER, MySQLConfig::PASS) or die("Error: Unable to connect to the MySQL Server."); MySQLConnect::$instances = 1; } else { $msg = "Close the existing instance of the MySQLConnector class."; die($msg); } } public function singleQuery($query, $databasename) { mysql_select_db(MySQLConfig::DB, $this->connection) or die("Error: Could not select database " . MySQLConfig::DB . " from the server."); $result = mysql_query($query) or die('Query failed.'); return $result; } public function createResultSet($query, $databasename) { $rs = new MySQLResultSet($query, MySQLConfig::DB, $this->connection ) ; return $rs; } public function close() { MySQLConnect::$instances = 0; if(isset($this->connection) ) { mysql_close($this->connection) ; unset($this->connection) ; } } public function __destruct() { $this->close(); } } The MySQLResultSet class looks like this: class MySQLResultSet implements Iterator { private $query; private $databasename; private $connection; private $result; private $currentRow; private $key = 0; private $valid; public function __construct($query, $databasename, $connection) { $this->query = $query; //Select the database $selectedDatabase = mysql_select_db($databasename, $connection) or die("Error: Could not select database " . $this->dbname . " from the server."); $this->result = mysql_query($this->query) or die('Query failed.'); $this->rewind(); } public function getResult() { return $this->result; } // public function getRow() // { // return mysql_fetch_row($this->result); // } public function getNumberRows() { return mysql_num_rows($this->result); } //current() returns the current row public function current() { return $this->currentRow; } //key() returns the current index public function key() { return $this->key; } //next() moves forward one index public function next() { if($this->currentRow = mysql_fetch_array($this->result) ) { $this->valid = true; $this->key++; }else{ $this->valid = false; } } //rewind() moves to the starting index public function rewind() { $this->key = 0; if(mysql_num_rows($this->result) > 0) { if(mysql_data_seek($this->result, 0) ) { $this->valid = true; $this->key = 0; $this->currentRow = mysql_fetch_array($this->result); } } else { $this->valid = false; } } //valid returns 1 if the current position is a valid array index //and 0 if it is not valid public function valid() { return $this->valid; } } The following class is an example of how I am accessing the database: class ImageCount { public function getCount() { $mysqlConnector = new MySQLConnect(); $query = "SELECT * FROM images;"; $resultSet = $mysqlConnector->createResultSet($query, MySQLConfig::DB); $mysqlConnector->close(); return $resultSet->getNumberRows(); } } I use the ImageCount class like this: if(!ImageCount::getCount()) { //Do something } Question: Is this an okay way to access the database? Could anybody recommend an alternative method if it is bad? Thank-you.

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• Data munging and data import scripting

  - by morpheous

  I need to write some scripts to carry out some tasks on my server (running Ubuntu server 8.04 TLS). The tasks are to be run periodically, so I will be running the scripts as cron jobs. I have divided the tasks into "group A" and "group B" - because (in my mind at least), they are a bit different. Task Group A import data from a file and possibly reformat it - by reformatting, I mean doing things like santizing the data, possibly normalizing it and or running calculations on 'columns' of the data Import the munged data into a database. For now, I am mostly using mySQL for the vast majority of imports - although some files will be imported into a sqlLite database. Note: The files will be mostly text files, although some of the files are in a binary format (my own proprietary format, written by a C++ application I developed). Task Group B Extract data from the database Perform calculations on the data and either insert or update tables in the database. My coding experience is is primarily as a C/C++ developer, although I have been using PHP as well for the last 2 years or so. I am from a windows background so I am still finding my feet in the linux environment. My question is this - I need to write scripts to perform the tasks I described above. Although I suppose I could write a few C++ applications to be used in the shell scripts, I think it may be better to write them in a scripting language (maybe this is a flawed assumption?). My thinking is that it would be easier to modify thins in a script - no need to rebuild etc for changes to functionality. Additionally, C++ data munging in C++ tends to involve more lines of code than "natural" scripting languages such as Perl, Python etc. Assuming that the majority of people on here agree that scripting is the way to go, herein lies my dilema. Which scripting language to use to perform the tasks above (giving my background). My gut instinct tells me that Perl (shudder) would be the most obvious choice for performing all of the above tasks. BUT (and that is a big BUT). The mere mention of Perl makes my toes curl, as I had a very, very bag experience with it a while back. The syntax seems quite unnatural to me - despite how many times I have tried to learn it - so if possible, I would really like to give it a miss. PHP (which I already know), also am not sure is a good candidate for scripting on the CLI (I have not seen many examples on how to do this etc - so I may be wrong). The last thing I must mention is that IF I have to learn a new language in order to do this, I cannot afford (time constraint) to spend more than a day, in learning the key commands/features required in order to do this (I can always learn the details of the language later, once I have actually deployed the scripts). So, which scripting language would you recommend (PHP, Python, Perl, [insert your favorite here]) - and most importantly WHY?. Or, should I just stick to writing little C++ applications that I call in a shell script?. Lastly, if you have suggested a scripting language, can you please show with a FEW lines (Perl mongers - I'm looking in your direction [nothing to cryptic!] ;) ) how I can use the language you suggested to do what I want to do. Hopefully, the lines you present will convince me that it can be done easily and elegantly in the language you suggested.

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• Magento Onepage Success Conversion Tracking Design Pattern

  - by user1734954

  My intent is to track conversions through multiple channels by inserting third party javascript (for example google analytics, optimizely, pricegrabber etc.) into the footer of onepage success . I've accomplished this by adding a block to the footer reference inside of the checkout success node within local.xml and everything works appropriately. My questions are more about efficiency and extensibility. It occurred to me that it would be better to combine all of the blocks into a single block reference and then use a various methods acting on a single call to the various related models to provide the data needed for insertion into the javascript for each of the conversion tracking scripts. Some examples of the common data that conversion tracking may rely on(pseudo): Order ID , Order Total, Order.LineItem.Name(foreach) and so on Currently for each of the scripts I've made a call to the appropriate model passing the customers last order id as the load value and the calling a get() assigning the return value to a variable and then iterating through the data to match the values with the expectations of the given third party service. All of the data should be pulled once when checkout is complete each third party services may expect different data in different formats Here is an example of one of the conversion tracking template files which loads at the footer of checkout success. $order = Mage::getModel('sales/order')->loadByIncrementId(Mage::getSingleton('checkout/session')->getLastRealOrderId()); $amount = number_format($order->getGrandTotal(),2); $customer = Mage::helper('customer')->getCustomer()->getData(); ?> <script type="text/javascript"> popup_email = '<?php echo($customer['email']);?>'; popup_order_number = '<?php echo $this->getOrderId() ?>'; </script> <!-- PriceGrabber Merchant Evaluation Code --> <script type="text/javascript" charset="UTF-8" src="https://www.pricegrabber.com/rating_merchrevpopjs.php?retid=<something>"></script> <noscript><a href="http://www.pricegrabber.com/rating_merchrev.php?retid=<something>" target=_blank> <img src="https://images.pricegrabber.com/images/mr_noprize.jpg" border="0" width="272" height="238" alt="Merchant Evaluation"></a></noscript> <!-- End PriceGrabber Code --> Having just a single piece of code like this is not that big of a deal, but we are doing similar things with a number of different third party services. Pricegrabber is one of the simpler examples. A more sophisticated tracking service expects a comma separated list of all of the product names, ids, prices, categories , order id etc. I would like to make it all more manageable so my idea to do the following: combine all of the template files into a single file Develop a helper class or library to deliver the data to the conversion template Goals Include Extensibility Minimal Model Calls Minimal Method Calls The Questions 1. Is a Mage helper the best route to take? 2. Is there any design pattern you may recommend for the "helper" class? 3. Why would this the design pattern you've chosen be best for this instance?

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• How to figure out who owns a worker thread that is still running when my app exits?

  - by Dave

  Not long after upgrading to VS2010, my application won't shut down cleanly. If I close the app and then hit pause in the IDE, I see this: The problem is, there's no context. The call stack just says [External code], which isn't too helpful. Here's what I've done so far to try to narrow down the problem: deleted all extraneous plugins to minimize the number of worker threads launched set breakpoints in my code anywhere I create worker threads (and delegates + BeginInvoke, since I think they are labeled "Worker Thread" in the debugger anyway). None were hit. set IsBackground = true for all threads While I could do the next brute force step, which is to roll my code back to a point where this didn't happen and then look over all of the change logs, this isn't terribly efficient. Can anyone recommend a better way to figure this out, given the notable lack of information presented by the debugger? The only other things I can think of include: read up on WinDbg and try to use it to stop anytime a thread is started. At least, I thought that was possible... :) comment out huge blocks of code until the app closes properly, then start uncommenting until it doesn't. UPDATE Perhaps this information will be of use. I decided to use WinDbg and attach to my application. I then closed it, and switched to thread 0 and dumped the stack contents. Here's what I have: ThreadCount: 6 UnstartedThread: 0 BackgroundThread: 1 PendingThread: 0 DeadThread: 4 Hosted Runtime: no PreEmptive GC Alloc Lock ID OSID ThreadOBJ State GC Context Domain Count APT Exception 0 1 1c70 005a65c8 6020 Enabled 02dac6e0:02dad7f8 005a03c0 0 STA 2 2 1b20 005b1980 b220 Enabled 00000000:00000000 005a03c0 0 MTA (Finalizer) XXXX 3 08504048 19820 Enabled 00000000:00000000 005a03c0 0 Ukn XXXX 4 08504540 19820 Enabled 00000000:00000000 005a03c0 0 Ukn XXXX 5 08516a90 19820 Enabled 00000000:00000000 005a03c0 0 Ukn XXXX 6 08517260 19820 Enabled 00000000:00000000 005a03c0 0 Ukn 0:008> ~0s eax=c0674960 ebx=00000000 ecx=00000000 edx=00000000 esi=0040f320 edi=005a65c8 eip=76c37e47 esp=0040f23c ebp=0040f258 iopl=0 nv up ei pl nz na po nc cs=0023 ss=002b ds=002b es=002b fs=0053 gs=002b efl=00000202 USER32!NtUserGetMessage+0x15: 76c37e47 83c404 add esp,4 0:000> !clrstack OS Thread Id: 0x1c70 (0) Child SP IP Call Site 0040f274 76c37e47 [InlinedCallFrame: 0040f274] 0040f270 6baa8976 DomainBoundILStubClass.IL_STUB_PInvoke(System.Windows.Interop.MSG ByRef, System.Runtime.InteropServices.HandleRef, Int32, Int32)*** WARNING: Unable to verify checksum for C:\Windows\assembly\NativeImages_v4.0.30319_32\WindowsBase\d17606e813f01376bd0def23726ecc62\WindowsBase.ni.dll 0040f274 6ba924c5 [InlinedCallFrame: 0040f274] MS.Win32.UnsafeNativeMethods.IntGetMessageW(System.Windows.Interop.MSG ByRef, System.Runtime.InteropServices.HandleRef, Int32, Int32) 0040f2c4 6ba924c5 MS.Win32.UnsafeNativeMethods.GetMessageW(System.Windows.Interop.MSG ByRef, System.Runtime.InteropServices.HandleRef, Int32, Int32) 0040f2dc 6ba8e5f8 System.Windows.Threading.Dispatcher.GetMessage(System.Windows.Interop.MSG ByRef, IntPtr, Int32, Int32) 0040f318 6ba8d579 System.Windows.Threading.Dispatcher.PushFrameImpl(System.Windows.Threading.DispatcherFrame) 0040f368 6ba8d2a1 System.Windows.Threading.Dispatcher.PushFrame(System.Windows.Threading.DispatcherFrame) 0040f374 6ba7fba0 System.Windows.Threading.Dispatcher.Run() 0040f380 62e6ccbb System.Windows.Application.RunDispatcher(System.Object)*** WARNING: Unable to verify checksum for C:\Windows\assembly\NativeImages_v4.0.30319_32\PresentationFramewo#\7f91eecda3ff7ce478146b6458580c98\PresentationFramework.ni.dll 0040f38c 62e6c8ff System.Windows.Application.RunInternal(System.Windows.Window) 0040f3b0 62e6c682 System.Windows.Application.Run(System.Windows.Window) 0040f3c0 62e6c30b System.Windows.Application.Run() 0040f3cc 001f00bc MyApplication.App.Main() [C:\code\trunk\MyApplication\obj\Debug\GeneratedInternalTypeHelper.g.cs @ 24] 0040f608 66c421db [GCFrame: 0040f608] EDIT -- not sure if this helps, but the main thread's call stack looks like this: [Managed to Native Transition] > WindowsBase.dll!MS.Win32.UnsafeNativeMethods.GetMessageW(ref System.Windows.Interop.MSG msg, System.Runtime.InteropServices.HandleRef hWnd, int uMsgFilterMin, int uMsgFilterMax) + 0x15 bytes WindowsBase.dll!System.Windows.Threading.Dispatcher.GetMessage(ref System.Windows.Interop.MSG msg, System.IntPtr hwnd, int minMessage, int maxMessage) + 0x48 bytes WindowsBase.dll!System.Windows.Threading.Dispatcher.PushFrameImpl(System.Windows.Threading.DispatcherFrame frame = {System.Windows.Threading.DispatcherFrame}) + 0x85 bytes WindowsBase.dll!System.Windows.Threading.Dispatcher.PushFrame(System.Windows.Threading.DispatcherFrame frame) + 0x49 bytes WindowsBase.dll!System.Windows.Threading.Dispatcher.Run() + 0x4c bytes PresentationFramework.dll!System.Windows.Application.RunDispatcher(object ignore) + 0x17 bytes PresentationFramework.dll!System.Windows.Application.RunInternal(System.Windows.Window window) + 0x6f bytes PresentationFramework.dll!System.Windows.Application.Run(System.Windows.Window window) + 0x26 bytes PresentationFramework.dll!System.Windows.Application.Run() + 0x1b bytes I did a search on it and found some posts related to WPF GUIs hanging, and maybe that'll give me some more clues.

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• How best to modernize the 2002-era J2EE app?

  - by user331465

  I have this friend.... I have this friend who works on a java ee application (j2ee) application started in the early 2000's. Currently they add a feature here and there, but have a large codebase. Over the years the team has shrunk by 70%. [Yes, the "i have this friend is". It's me, attempting to humorously inject teenage high-school counselor shame into the mix] Java, Vintage 2002 The application uses EJB 2.1, struts 1.x, DAO's etc with straight jdbc calls (mixture of stored procedures and prepared statements). No ORM. For caching they use a mixture of OpenSymphony OSCache and a home-grown cache layer. Over the last few years, they have spent effort to modernize the UI using ajax techniques and libraries. This largely involves javascript libaries (jquery, yui, etc). Client Side On the client side, the lack of upgrade path from struts1 to struts2 discouraged them from migrating to struts2. Other web frameworks became popular (wicket, spring , jsf). Struts2 was not the "clear winner". Migrating all the existing UI from Struts1 to Struts2/wicket/etc did not seem to present much marginal benefit at a very high cost. They did not want to have a patchwork of technologies-du-jour (subsystem X in Struts2, subsystem Y in Wicket, etc.) so developer write new features using Struts 1. Server Side On the server side, they looked into moving to ejb 3, but never had a big impetus. The developers are all comfortable with ejb-jar.xml, EJBHome, EJBRemote, that "ejb 2.1 as is" represented the path of least resistance. One big complaint about the ejb environment: programmers still pretend "ejb server runs in separate jvm than servlet engine". No app server (jboss/weblogic) has ever enforced this separation. The team has never deployed the ejb server on a separate box then the app server. The ear file contains multiple copies of the same jar file; one for the 'web layer' (foo.war/WEB-INF/lib) and one for the server side (foo.ear/). The app server only loads one jar. The duplications makes for ambiguity. Caching As for caching, they use several cache implementations: OpenSymphony cache and a homegrown cache. Jgroups provides clustering support Now What? The question: The team currently has spare cycles to to invest in modernizing the application? Where would the smart investor spend them? The main criteria: 1) productivity gains. Specifically reducing the time to develope new subsystems features and reduced maintenance. 2) performance/scalability. They do not care about fashion or techno-du-jour street cred. What do you all recommend? On the persistence side Switch everything (or new development only) to JPA/JPA2? Straight hibernate? Wait for Java EE 6? On the client/web-framework side: Migrate (some or all) to struts2? wicket? jsf/jsf2? As for caching: terracotta? ehcache? coherence? stick with what they have? how best to take advantage of the huge heap sizes that the 64-bit jvms offer? Thanks in advance.

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• hosting simple python scripts in a container to handle concurrency, configuration, caching, etc.

  - by Justin Grant

  My first real-world Python project is to write a simple framework (or re-use/adapt an existing one) which can wrap small python scripts (which are used to gather custom data for a monitoring tool) with a "container" to handle boilerplate tasks like: fetching a script's configuration from a file (and keeping that info up to date if the file changes and handle decryption of sensitive config data) running multiple instances of the same script in different threads instead of spinning up a new process for each one expose an API for caching expensive data and storing persistent state from one script invocation to the next Today, script authors must handle the issues above, which usually means that most script authors don't handle them correctly, causing bugs and performance problems. In addition to avoiding bugs, we want a solution which lowers the bar to create and maintain scripts, especially given that many script authors may not be trained programmers. Below are examples of the API I've been thinking of, and which I'm looking to get your feedback about. A scripter would need to build a single method which takes (as input) the configuration that the script needs to do its job, and either returns a python object or calls a method to stream back data in chunks. Optionally, a scripter could supply methods to handle startup and/or shutdown tasks. HTTP-fetching script example (in pseudocode, omitting the actual data-fetching details to focus on the container's API): def run (config, context, cache) : results = http_library_call (config.url, config.http_method, config.username, config.password, ...) return { html : results.html, status_code : results.status, headers : results.response_headers } def init(config, context, cache) : config.max_threads = 20 # up to 20 URLs at one time (per process) config.max_processes = 3 # launch up to 3 concurrent processes config.keepalive = 1200 # keep process alive for 10 mins without another call config.process_recycle.requests = 1000 # restart the process every 1000 requests (to avoid leaks) config.kill_timeout = 600 # kill the process if any call lasts longer than 10 minutes Database-data fetching script example might look like this (in pseudocode): def run (config, context, cache) : expensive = context.cache["something_expensive"] for record in db_library_call (expensive, context.checkpoint, config.connection_string) : context.log (record, "logDate") # log all properties, optionally specify name of timestamp property last_date = record["logDate"] context.checkpoint = last_date # persistent checkpoint, used next time through def init(config, context, cache) : cache["something_expensive"] = get_expensive_thing() def shutdown(config, context, cache) : expensive = cache["something_expensive"] expensive.release_me() Is this API appropriately "pythonic", or are there things I should do to make this more natural to the Python scripter? (I'm more familiar with building C++/C#/Java APIs so I suspect I'm missing useful Python idioms.) Specific questions: is it natural to pass a "config" object into a method and ask the callee to set various configuration options? Or is there another preferred way to do this? when a callee needs to stream data back to its caller, is a method like context.log() (see above) appropriate, or should I be using yield instead? (yeild seems natural, but I worry it'd be over the head of most scripters) My approach requires scripts to define functions with predefined names (e.g. "run", "init", "shutdown"). Is this a good way to do it? If not, what other mechanism would be more natural? I'm passing the same config, context, cache parameters into every method. Would it be better to use a single "context" parameter instead? Would it be better to use global variables instead? Finally, are there existing libraries you'd recommend to make this kind of simple "script-running container" easier to write?

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• CSS optimization - extra classes in dom or preprocessor-repetitive styling in css file?

  - by anna.mi

  I'm starting on a fairly large project and I'm considering the option of using LESS for pre-processing my css. the useful thing about LESS is that you can define a mixin that contains for example: .border-radius(@radius) { -webkit-border-radius: @radius; -moz-border-radius: @radius; -o-border-radius: @radius; -ms-border-radius: @radius; border-radius: @radius; } and then use it in a class declaration as .rounded-div { .border-radius(10px); } to get the outputted css as: .rounded-div { -webkit-border-radius: 10px; -moz-border-radius: 10px; -o-border-radius: 10px; -ms-border-radius: 10px; border-radius: 10px; } this is extremely useful in the case of browser prefixes. However this same concept could be used to encapsulate commonly-used css, for example: .column-container { overflow: hidden; display: block; width: 100%; } .column(@width) { float: left; width: @width; } and then use this mixin whenever i need columns in my design: .my-column-outer { .column-container(); background: red; } .my-column-inner { .column(50%); font-color: yellow; } (of course, using the preprocessor we could easily expand this to be much more useful, eg. pass the number of columns and the container width as variables and have LESS determine the width of each column depending on the number of columns and container width!) the problem with this is that when compliled, my final css file would have 100 such declarations, copy&pasted, making the file huge and bloated and repetitive. The alternative to this would be to use a grid system which has predefined classes for each column-layout option, eg .c-50 ( with a "float: left; width:50%;" definition ), .c-33, .c-25 to accomodate for a 2-column, 3-column and 4-column layout and then use these classes to my dom. i really mislike the idea of the extra classes, from experience it results to bloated dom (creating extra divs just to attach the grid classes to). Also the most basic tutorial for html/css would tell you that the dom should be separated from the styling - grid classes are styling related! to me, its the same as attaching a "border-radius-10" class to the .rounded-div example above! on the other hand, the large css file that would result from the repetitive code is also a disadvantage so i guess my question is, which one would you recommend? and which do you use? and, which solution is best for optimization? apart from the larger file size, has there even been any research on whether browser renders multiple classes faster than a large css file, or the other way round? tnx! i'd love to hear your opinion!

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• User welcome message in php

  - by user225269

  How do I create a user welcome message in php. So that the user who has been logged on will be able to see his username. I have this code, but it doesn't seem to work. <?php $con = mysql_connect("localhost","root","nitoryolai123$%^"); if (!$con) { die('Could not connect: ' . mysql_error()); } mysql_select_db("school", $con); $result = mysql_query("SELECT * FROM users WHERE Username='$username'"); while($row = mysql_fetch_array($result)) { echo $row['Username']; echo "<br />"; } ?> I'm trying to make use of the data that is inputted in this login form: <form name="form1" method="post" action="verifylogin.php"> <td> <table border="0" cellpadding="3" cellspacing="1" bgcolor=""> <tr> <td colspan="16" height="25" style="background:#5C915C; color:white; border:white 1px solid; text-align: left"><strong><font size="2">Login User</strong></td> </tr> <tr> <td width="30" height="35"><font size="2">Username:</td> <td width="30"><input name="myusername" type="text" id="idnum" maxlength="5"></td> </tr> <tr> <td width="30" height="35" ><font size="2">Password:</td> <td width="30"><input name="mypassword" type="password" id="lname" maxlength="15"></td> </tr> <td align="right" width="30"><td align="right" width="30"><input type="submit" name="Submit" value="Submit" /></td> <td align="right" width="30"><input type="reset" name="Reset" value="Reset"></td></td> </tr> </form> But this, verifylogin.php, seems to be in the way. <?php $host="localhost"; $username="root"; $password="nitoryolai123$%^"; $db_name="school"; $tbl_name="users"; mysql_connect("$host", "$username", "$password")or die("cannot connect"); mysql_select_db("$db_name")or die("cannot select DB"); $myusername=$_POST['myusername']; $mypassword=$_POST['mypassword']; $myusername = stripslashes($myusername); $mypassword = stripslashes($mypassword); $myusername = mysql_real_escape_string($myusername); $mypassword = mysql_real_escape_string($mypassword); $sql="SELECT * FROM $tbl_name WHERE username='$myusername' and password='$mypassword'"; $result=mysql_query($sql); $count=mysql_num_rows($result); if($count==1){ session_register("myusername"); session_register("mypassword"); header("location:userpage.php"); } else { echo "Wrong Username or Password"; } ?> How do I do it? I always get this error when I run it: Notice: Undefined variable: username in C:\wamp\www\exp\userpage.php on line 53 Can you recommend of an easier on how I can achieve the same thing?

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• C/PHP: How do I convert the following PHP JSON API script into a C plugin for apache?

  - by TeddyB

  I have a JSON API that I need to provide super fast access to my data through. The JSON API makes a simply query against the database based on the GET parameters provided. I've already optimized my database, so please don't recommend that as an answer. I'm using PHP-APC, which helps PHP by saving the bytecode, BUT - for a JSON API that is being called literally dozens of times per second (as indicated by my logs), I need to reduce the massive RAM consumption PHP is consuming ... as well as rewrite my JSON API in a language that execute much faster than PHP. My code is below. As you can see, is fairly straight forward. <?php define(ALLOWED_HTTP_REFERER, 'example.com'); if ( stristr($_SERVER['HTTP_REFERER'], ALLOWED_HTTP_REFERER) ) { try { $conn_str = DB . ':host=' . DB_HOST . ';dbname=' . DB_NAME; $dbh = new PDO($conn_str, DB_USERNAME, DB_PASSWORD); $params = array(); $sql = 'SELECT homes.home_id, address, city, state, zip FROM homes WHERE homes.display_status = true AND homes.geolat BETWEEN :geolatLowBound AND :geolatHighBound AND homes.geolng BETWEEN :geolngLowBound AND :geolngHighBound'; $params[':geolatLowBound'] = $_GET['geolatLowBound']; $params[':geolatHighBound'] = $_GET['geolatHighBound']; $params[':geolngLowBound'] =$_GET['geolngLowBound']; $params[':geolngHighBound'] = $_GET['geolngHighBound']; if ( isset($_GET['min_price']) && isset($_GET['max_price']) ) { $sql = $sql . ' AND homes.price BETWEEN :min_price AND :max_price '; $params[':min_price'] = $_GET['min_price']; $params[':max_price'] = $_GET['max_price']; } if ( isset($_GET['min_beds']) && isset($_GET['max_beds']) ) { $sql = $sql . ' AND homes.num_of_beds BETWEEN :min_beds AND :max_beds '; $params['min_beds'] = $_GET['min_beds']; $params['max_beds'] = $_GET['max_beds']; } if ( isset($_GET['min_sqft']) && isset($_GET['max_sqft']) ) { $sql = $sql . ' AND homes.sqft BETWEEN :min_sqft AND :max_sqft '; $params['min_sqft'] = $_GET['min_sqft']; $params['max_sqft'] = $_GET['max_sqft']; } $stmt = $dbh->prepare($sql); $stmt->execute($params); $result_set = $stmt->fetchAll(PDO::FETCH_ASSOC); /* output a JSON representation of the home listing data retrieved */ ob_start("ob_gzhandler"); // compress the output header('Content-type: text/javascript'); print "{'homes' : "; array_walk_recursive($result_set, "cleanOutputFromXSS"); print json_encode( $result_set ); print '}'; $dbh = null; } catch (PDOException $e) { die('Unable to retreive home listing information'); } } function cleanOutputFromXSS(&$value) { $value = htmlspecialchars($value, ENT_QUOTES, 'UTF-8'); } ?> How would I begin converting this PHP code over to C, since C is both better on memory management (since you do it yourself) and much, much faster to execute?

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• Problem with table in php

  - by user225269

  I'm trying to create a table in php that would show the data on the mysql database based on the check box that is checked by the user. As you can see in this screen shot, it will have problems when you did not check on a checkbox before the one that will be the last: http://www.mypicx.com/04282010/1/ Here is my code: if($_POST['general'] == 'ADDRESS'){ $result2 = mysql_query("SELECT * FROM student WHERE ADDRESS='$saddress'"); ?> <table border='1'> <tr> <th>IDNO</th> <th>YEAR</th> <th>SECTION</th> <?php if ( $ShowLastName ) echo "<th>LASTNAME</th>" ?> <?php if ( $ShowFirstName ) echo "<th>FIRSTNAME</th>" ?> <?php if ( $ShowMidName ) echo "<th>MIDNAME</th>" ?> <?php if ( $ShowAddress ) echo "<th>ADDRESS</th>" ?> <?php if ( $ShowGender ) echo "<th>GENDER</th>" ?> <?php if ( $ShowReligion ) echo "<th>RELIGION</th>" ?> <?php if ( $ShowBday ) echo "<th>BIRTHDAY</th>" ?> <?php if ( $ShowContact ) echo "<th>CONTACT</th>" ?> </tr> <?php while($row = mysql_fetch_array($result2)) {?> <tr> <td><?php echo $row['IDNO']?> </td> <td><?php echo $row['YEAR'] ?> </td> <td><?php echo $row['SECTION'] ?></td> <td><?php if ( $ShowLastName ) echo $row['LASTNAME'] ?></td> <td><?php if ( $ShowFirstName ) echo $row['FIRSTNAME'] ?></td> <td><?php if ( $ShowMidName ) echo $row['MI'] ?></td> <td><?php if ( $ShowAddress ) echo $row['ADDRESS'] ?></td> <td><?php if ( $ShowGender ) echo $row['GENDER'] ?></td> <td><?php if ( $ShowReligion ) echo $row['RELIGION'] ?></td> <td><?php if ( $ShowBday ) echo $row['BIRTHDAY'] ?></td> <td><?php if ( $ShowContact ) echo $row['S_CONTACTNUM'] ?></td> </tr> <?PHP } ?> </table> <?PHP } mysql_close($con); ?> What can you recommend so that the output will not look like this when you one of the checkbox before a checkbox is not clicked: http://www.mypicx.com/04282010/2/

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• Version Assemblies with TFS 2010 Continuous Integration

  - by Steve Michelotti

  When I first heard that TFS 2010 had moved to Workflow Foundation for Team Build, I was *extremely* skeptical. I’ve loved MSBuild and didn’t quite understand the reasons for this change. In fact, given that I’ve been exclusively using Cruise Control for Continuous Integration (CI) for the last 5+ years of my career, I was skeptical of TFS for CI in general. However, after going through the learning process for TFS 2010 recently, I’m starting to become a believer. I’m also starting to see some of the benefits with Workflow Foundation for the overall processing because it gives you constructs not available in MSBuild such as parallel tasks, better control flow constructs, and a slightly better customization story. The first customization I had to make to the build process was to version the assemblies of my solution. This is not new. In fact, I’d recommend reading Mike Fourie’s well known post on Versioning Code in TFS before you get started. This post describes several foundational aspects of versioning assemblies regardless of your version of TFS. The main points are: 1) don’t use source control operations for your version file, 2) use a schema like <Major>.<Minor>.<IncrementalNumber>.0, and 3) do not keep AssemblyVersion and AssemblyFileVersion in sync. To do this in TFS 2010, the best post I’ve found has been Jim Lamb’s post of building a custom TFS 2010 workflow activity. Overall, this post is excellent but the primary issue I have with it is that the assembly version numbers produced are based in a date and look like this: “2010.5.15.1”. This is definitely not what I want. I want to be able to communicate to the developers and stakeholders that we are producing the “1.1 release” or “1.2 release” – which would have an assembly version number of “1.1.317.0” for example. In this post, I’ll walk through the process of customizing the assembly version number based on this method – customizing the concepts in Lamb’s post to suit my needs. I’ll also be combining this with the concepts of Fourie’s post – particularly with regards to the standards around how to version the assemblies. The first thing I’ll do is add a file called SolutionAssemblyVersionInfo.cs to the root of my solution that looks like this: 1: using System; 2: using System.Reflection; 3: [assembly: AssemblyVersion("1.1.0.0")] 4: [assembly: AssemblyFileVersion("1.1.0.0")] I’ll then add that file as a Visual Studio link file to each project in my solution by right-clicking the project, “Add – Existing Item…” then when I click the SolutionAssemblyVersionInfo.cs file, making sure I “Add As Link”: Now the Solution Explorer will show our file. We can see that it’s a “link” file because of the black arrow in the icon within all our projects. Of course you’ll need to remove the AssemblyVersion and AssemblyFileVersion attributes from the AssemblyInfo.cs files to avoid the duplicate attributes since they now leave in the SolutionAssemblyVersionInfo.cs file. This is an extremely common technique so that all the projects in our solution can be versioned as a unit. At this point, we’re ready to write our custom activity. The primary consideration is that I want the developer and/or tech lead to be able to easily be in control of the Major.Minor and then I want the CI process to add the third number with a unique incremental number. We’ll leave the fourth position always “0” for now – it’s held in reserve in case the day ever comes where we need to do an emergency patch to Production based on a branched version.   Writing the Custom Workflow Activity Similar to Lamb’s post, I’m going to write two custom workflow activities. The “outer” activity (a xaml activity) will be pretty straight forward. It will check if the solution version file exists in the solution root and, if so, delegate the replacement of version to the AssemblyVersionInfo activity which is a CodeActivity highlighted in red below:   Notice that the arguments of this activity are the “solutionVersionFile” and “tfsBuildNumber” which will be passed in. The tfsBuildNumber passed in will look something like this: “CI_MyApplication.4” and we’ll need to grab the “4” (i.e., the incremental revision number) and put that in the third position. Then we’ll need to honor whatever was specified for Major.Minor in the SolutionAssemblyVersionInfo.cs file. For example, if the SolutionAssemblyVersionInfo.cs file had “1.1.0.0” for the AssemblyVersion (as shown in the first code block near the beginning of this post), then we want to resulting file to have “1.1.4.0”. Before we do anything, let’s put together a unit test for all this so we can know if we get it right: 1: [TestMethod] 2: public void Assembly_version_should_be_parsed_correctly_from_build_name() 3: { 4: // arrange 5: const string versionFile = "SolutionAssemblyVersionInfo.cs"; 6: WriteTestVersionFile(versionFile); 7: var activity = new VersionAssemblies(); 8: var arguments = new Dictionary<string, object> { 9: { "tfsBuildNumber", "CI_MyApplication.4"}, 10: { "solutionVersionFile", versionFile} 11: }; 12:   13: // act 14: var result = WorkflowInvoker.Invoke(activity, arguments); 15:   16: // assert 17: Assert.AreEqual("1.2.4.0", (string)result["newAssemblyFileVersion"]); 18: var lines = File.ReadAllLines(versionFile); 19: Assert.IsTrue(lines.Contains("[assembly: AssemblyVersion(\"1.2.0.0\")]")); 20: Assert.IsTrue(lines.Contains("[assembly: AssemblyFileVersion(\"1.2.4.0\")]")); 21: } 22: 23: private void WriteTestVersionFile(string versionFile) 24: { 25: var fileContents = "using System.Reflection;\n" + 26: "[assembly: AssemblyVersion(\"1.2.0.0\")]\n" + 27: "[assembly: AssemblyFileVersion(\"1.2.0.0\")]"; 28: File.WriteAllText(versionFile, fileContents); 29: }   At this point, the code for our AssemblyVersion activity is pretty straight forward: 1: [BuildActivity(HostEnvironmentOption.Agent)] 2: public class AssemblyVersionInfo : CodeActivity 3: { 4: [RequiredArgument] 5: public InArgument<string> FileName { get; set; } 6:   7: [RequiredArgument] 8: public InArgument<string> TfsBuildNumber { get; set; } 9:   10: public OutArgument<string> NewAssemblyFileVersion { get; set; } 11:   12: protected override void Execute(CodeActivityContext context) 13: { 14: var solutionVersionFile = this.FileName.Get(context); 15: 16: // Ensure that the file is writeable 17: var fileAttributes = File.GetAttributes(solutionVersionFile); 18: File.SetAttributes(solutionVersionFile, fileAttributes & ~FileAttributes.ReadOnly); 19:   20: // Prepare assembly versions 21: var majorMinor = GetAssemblyMajorMinorVersionBasedOnExisting(solutionVersionFile); 22: var newBuildNumber = GetNewBuildNumber(this.TfsBuildNumber.Get(context)); 23: var newAssemblyVersion = string.Format("{0}.{1}.0.0", majorMinor.Item1, majorMinor.Item2); 24: var newAssemblyFileVersion = string.Format("{0}.{1}.{2}.0", majorMinor.Item1, majorMinor.Item2, newBuildNumber); 25: this.NewAssemblyFileVersion.Set(context, newAssemblyFileVersion); 26:   27: // Perform the actual replacement 28: var contents = this.GetFileContents(newAssemblyVersion, newAssemblyFileVersion); 29: File.WriteAllText(solutionVersionFile, contents); 30:   31: // Restore the file's original attributes 32: File.SetAttributes(solutionVersionFile, fileAttributes); 33: } 34:   35: #region Private Methods 36:   37: private string GetFileContents(string newAssemblyVersion, string newAssemblyFileVersion) 38: { 39: var cs = new StringBuilder(); 40: cs.AppendLine("using System.Reflection;"); 41: cs.AppendFormat("[assembly: AssemblyVersion(\"{0}\")]", newAssemblyVersion); 42: cs.AppendLine(); 43: cs.AppendFormat("[assembly: AssemblyFileVersion(\"{0}\")]", newAssemblyFileVersion); 44: return cs.ToString(); 45: } 46:   47: private Tuple<string, string> GetAssemblyMajorMinorVersionBasedOnExisting(string filePath) 48: { 49: var lines = File.ReadAllLines(filePath); 50: var versionLine = lines.Where(x => x.Contains("AssemblyVersion")).FirstOrDefault(); 51:   52: if (versionLine == null) 53: { 54: throw new InvalidOperationException("File does not contain [assembly: AssemblyVersion] attribute"); 55: } 56:   57: return ExtractMajorMinor(versionLine); 58: } 59:   60: private static Tuple<string, string> ExtractMajorMinor(string versionLine) 61: { 62: var firstQuote = versionLine.IndexOf('"') + 1; 63: var secondQuote = versionLine.IndexOf('"', firstQuote); 64: var version = versionLine.Substring(firstQuote, secondQuote - firstQuote); 65: var versionParts = version.Split('.'); 66: return new Tuple<string, string>(versionParts[0], versionParts[1]); 67: } 68:   69: private string GetNewBuildNumber(string buildName) 70: { 71: return buildName.Substring(buildName.LastIndexOf(".") + 1); 72: } 73:   74: #endregion 75: }   At this point the final step is to incorporate this activity into the overall build template. Make a copy of the DefaultTempate.xaml – we’ll call it DefaultTemplateWithVersioning.xaml. Before the build and labeling happens, drag the VersionAssemblies activity in. Then set the LabelName variable to “BuildDetail.BuildDefinition.Name + "-" + newAssemblyFileVersion since the newAssemblyFileVersion was produced by our activity.   Configuring CI Once you add your solution to source control, you can configure CI with the build definition window as shown here. The main difference is that we’ll change the Process tab to reflect a different build number format and choose our custom build process file:   When the build completes, we’ll see the name of our project with the unique revision number:   If we look at the detailed build log for the latest build, we’ll see the label being created with our custom task:     We can now look at the history labels in TFS and see the project name with the labels (the Assignment activity I added to the workflow):   Finally, if we look at the physical assemblies that are produced, we can right-click on any assembly in Windows Explorer and see the assembly version in its properties:   Full Traceability We now have full traceability for our code. There will never be a question of what code was deployed to Production. You can always see the assembly version in the properties of the physical assembly. That can be traced back to a label in TFS where the unique revision number matches. The label in TFS gives you the complete snapshot of the code in your source control repository at the time the code was built. This type of process for full traceability has been used for many years for CI – in fact, I’ve done similar things with CCNet and SVN for quite some time. This is simply the TFS implementation of that pattern. The new features that TFS 2010 give you to make these types of customizations in your build process are quite easy once you get over the initial curve.

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• Upgrading from TFS 2010 RC to TFS 2010 RTM done

  - by Martin Hinshelwood

  Today is the big day, with the Launch of Visual Studio 2010 already done in Asia, and rolling around the world towards us, we are getting ready for the RTM (Released). We have had TFS 2010 in Production for nearly 6 months and have had only minimal problems. Update 12th April 2010  – Added Scott Hanselman’s tweet about the MSDN download release time. SSW was the first company in the world outside of Microsoft to deploy Visual Studio 2010 Team Foundation Server to production, not once, but twice. I am hoping to make it 3 in a row, but with all the hype around the new version, and with it being a production release and not just a go-live, I think there will be a lot of competition. Developers: MSDN will be updated with #vs2010 downloads and details at 10am PST *today*! @shanselman - Scott Hanselman Same as before, we need to Uninstall 2010 RC and install 2010 RTM. The installer will take care of all the complexity of actually upgrading any schema changes. If you are upgrading from TFS 2008 to TFS2010 you can follow our Rules To Better TFS 2010 Migration and read my post on our successes.   We run TFS 2010 in a Hyper-V virtual environment, so we have the advantage of running a snapshot as well as taking a DB backup. Done - Snapshot the hyper-v server Microsoft does not support taking a snapshot of a running server, for very good reason, and Brian Harry wrote a post after my last upgrade with the reason why you should never snapshot a running server. Done - Uninstall Visual Studio Team Explorer 2010 RC You will need to uninstall all of the Visual Studio 2010 RC client bits that you have on the server. Done - Uninstall TFS 2010 RC Done - Install TFS 2010 RTM Done - Configure TFS 2010 RTM Pick the Upgrade option and point it at your existing “tfs_Configuration” database to load all of the existing settings Done - Upgrade the SharePoint Extensions Upgrade Build Servers (Pending) Test the server The back out plan, and you should always have one, is to restore the snapshot. Upgrading to Team Foundation Server 2010 – Done The first thing you need to do is off the TFS server and then log into the Hyper-v server and create a snapshot. Figure: Make sure you turn the server off and delete all old snapshots before you take a new one I noticed that the snapshot that was taken before the Beta 2 to RC upgrade was still there. You should really delete old snapshots before you create a new one, but in this case the SysAdmin (who is currently tucked up in bed) asked me not to. I guess he is worried about a developer messing up his server Turn your server on and wait for it to boot in anticipation of all the nice shiny RTM’ness that is coming next. The upgrade procedure for TFS2010 is to uninstal the old version and install the new one. Figure: Remove Visual Studio 2010 Team Foundation Server RC from the system.   Figure: Most of the heavy lifting is done by the Uninstaller, but make sure you have removed any of the client bits first. Specifically Visual Studio 2010 or Team Explorer 2010.  Once the uninstall is complete, this took around 5 minutes for me, you can begin the install of the RTM. Running the 64 bit OS will allow the application to use more than 2GB RAM, which while not common may be of use in heavy load situations. Figure: It is always recommended to install the 64bit version of a server application where possible. I do not think it is likely, with SharePoint 2010 and Exchange 2010  and even Windows Server 2008 R2 being 64 bit only, I do not think there will be another release of a server app that is 32bit. You then need to choose what it is you want to install. This depends on how you are running TFS and on how many servers. In our case we run TFS and the Team Foundation Build Service (controller only) on out TFS server along with Analysis services and Reporting Services. But our SharePoint server lives elsewhere. Figure: This always confuses people, but in reality it makes sense. Don’t install what you do not need. Every extra you install has an impact of performance. If you are integrating with SharePoint you will need to run this install on every Front end server in your farm and don’t forget to upgrade your Build servers and proxy servers later. Figure: Selecting only Team Foundation Server (TFS) and Team Foundation Build Services (TFBS)   It is worth noting that if you have a lot of builds kicking off, and hence a lot of get operations against your TFS server, you can use a proxy server to cache the source control on another server in between your TFS server and your build servers. Figure: Installing Microsoft .NET Framework 4 takes the most time. Figure: Now run Windows Update, and SSW Diagnostic to make sure all your bits and bobs are up to date. Note: SSW Diagnostic will check your Power Tools, Add-on’s, Check in Policies and other bits as well. Configure Team Foundation Server 2010 – Done Now you can configure the server. If you have no key you will need to pick “Install a Trial Licence”, but it is only £500, or free with a MSDN subscription. Anyway, if you pick Trial you get 90 days to get your key. Figure: You can pick trial and add your key later using the TFS Server Admin. Here is where the real choices happen. We are doing an Upgrade from a previous version, so I will pick Upgrade the same as all you folks that are using the RC or TFS 2008. Figure: The upgrade wizard takes your existing 2010 or 2008 databases and upgraded them to the release.   Once you have entered your database server name you can click “List available databases” and it will show what it can upgrade. Figure: Select your database from the list and at this point, make sure you have a valid backup. At this point you have not made ANY changes to the databases. At this point the configuration wizard will load configuration from your existing database if you have one. If you are upgrading TFS 2008 refer to Rules To Better TFS 2010 Migration. Mostly during the wizard the default values will suffice, but depending on the configuration you want you can pick different options. Figure: Set the application tier account and Authentication method to use. We use NTLM to keep things simple as we host our TFS server externally for our remote developers.  Figure: Setting your TFS server URL’s to be the remote URL’s allows the reports to be accessed without using VPN. Very handy for those remote developers. Figure: Detected the existing Warehouse no problem. Figure: Again we love green ticks. It gives us a warm fuzzy feeling. Figure: The username for connecting to Reporting services should be a domain account (if you are on a domain that is). Figure: Setup the SharePoint integration to connect to your external SharePoint server. You can take the option to connect later.   You then need to run all of your readiness checks. These check can save your life! it will check all of the settings that you have entered as well as checking all the external services are configures and running properly. There are two reasons that TFS 2010 is so easy and painless to install where previous version were not. Microsoft changes the install to two steps, Install and configuration. The second reason is that they have pulled out all of the stops in making the install run all the checks necessary to make sure that once you start the install that it will complete. if you find any errors I recommend that you report them on http://connect.microsoft.com so everyone can benefit from your misery.   Figure: Now we have everything setup the configuration wizard can do its work.  Figure: Took a while on the “Web site” stage for some point, but zipped though after that.  Figure: last wee bit. TFS Needs to do a little tinkering with the data to complete the upgrade. Figure: All upgraded. I am not worried about the yellow triangle as SharePoint was being a little silly Exception Message: TF254021: The account name or password that you specified is not valid. (type TfsAdminException) Exception Stack Trace:    at Microsoft.TeamFoundation.Management.Controls.WizardCommon.AccountSelectionControl.TestLogon(String connectionString)    at System.ComponentModel.BackgroundWorker.WorkerThreadStart(Object argument) [Info   @16:10:16.307] Benign exception caught as part of verify: Exception Message: TF255329: The following site could not be accessed: http://projects.ssw.com.au/. The server that you specified did not return the expected response. Either you have not installed the Team Foundation Server Extensions for SharePoint Products on this server, or a firewall is blocking access to the specified site or the SharePoint Central Administration site. For more information, see the Microsoft Web site (http://go.microsoft.com/fwlink/?LinkId=161206). (type TeamFoundationServerException) Exception Stack Trace:    at Microsoft.TeamFoundation.Client.SharePoint.WssUtilities.VerifyTeamFoundationSharePointExtensions(ICredentials credentials, Uri url)    at Microsoft.TeamFoundation.Admin.VerifySharePointSitesUrl.Verify() Inner Exception Details: Exception Message: TF249064: The following Web service returned an response that is not valid: http://projects.ssw.com.au/_vti_bin/TeamFoundationIntegrationService.asmx. This Web service is used for the Team Foundation Server Extensions for SharePoint Products. Either the extensions are not installed, the request resulted in HTML being returned, or there is a problem with the URL. Verify that the following URL points to a valid SharePoint Web application and that the application is available: http://projects.ssw.com.au. If the URL is correct and the Web application is operating normally, verify that a firewall is not blocking access to the Web application. (type TeamFoundationServerInvalidResponseException) Exception Data Dictionary: ResponseStatusCode = InternalServerError I’ll look at SharePoint after, probably the SharePoint box just needs a restart or a kick If there is a problem with SharePoint it will come out in testing, But I will definatly be passing this on to Microsoft.   Upgrading the SharePoint connector to TFS 2010 You will need to upgrade the Extensions for SharePoint Products and Technologies on all of your SharePoint farm front end servers. To do this uninstall  the TFS 2010 RC from it in the same way as the server, and then install just the RTM Extensions. Figure: Only install the SharePoint Extensions on your SharePoint front end servers. TFS 2010 supports both SharePoint 2007 and SharePoint 2010.   Figure: When you configure SharePoint it uploads all of the solutions and templates. Figure: Everything is uploaded Successfully. Figure: TFS even remembered the settings from the previous installation, fantastic.   Upgrading the Team Foundation Build Servers to TFS 2010 Just like on the SharePoint servers you will need to upgrade the Build Server to the RTM. Just uninstall TFS 2010 RC and then install only the Team Foundation Build Services component. Unlike on the SharePoint server you will probably have some version of Visual Studio installed. You will need to remove this as well. (Coming Soon) Connecting Visual Studio 2010 / 2008 / 2005 and Eclipse to TFS2010 If you have developers still on Visual Studio 2005 or 2008 you will need do download the respective compatibility pack: Visual Studio Team System 2005 Service Pack 1 Forward Compatibility Update for Team Foundation Server 2010 Visual Studio Team System 2008 Service Pack 1 Forward Compatibility Update for Team Foundation Server 2010 If you are using Eclipse you can download the new Team Explorer Everywhere install for connecting to TFS. Get your developers to check that you have the latest version of your applications with SSW Diagnostic which will check for Service Packs and hot fixes to Visual Studio as well.   Technorati Tags: TFS,TFS2010,TFS 2010,Upgrade

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• Access violation in DirectX OMSetRenderTargets

  - by IDWMaster

  I receive the following error (Unhandled exception at 0x527DAE81 (d3d11_1sdklayers.dll) in Lesson2.Triangles.exe: 0xC0000005: Access violation reading location 0x00000000) when running the Triangle sample application for DirectX 11 in D3D_FEATURE_LEVEL_9_1. This error occurs at the OMSetRenderTargets function, as shown below, and does not happen if I remove that function from the program (but then, the screen is blue, and does not render the triangle) //// THIS CODE AND INFORMATION IS PROVIDED "AS IS" WITHOUT WARRANTY OF //// ANY KIND, EITHER EXPRESSED OR IMPLIED, INCLUDING BUT NOT LIMITED TO //// THE IMPLIED WARRANTIES OF MERCHANTABILITY AND/OR FITNESS FOR A //// PARTICULAR PURPOSE. //// //// Copyright (c) Microsoft Corporation. All rights reserved #include #include #include "DirectXSample.h" #include "BasicMath.h" #include "BasicReaderWriter.h" using namespace Microsoft::WRL; using namespace Windows::UI::Core; using namespace Windows::Foundation; using namespace Windows::ApplicationModel::Core; using namespace Windows::ApplicationModel::Infrastructure; // This class defines the application as a whole. ref class Direct3DTutorialViewProvider : public IViewProvider { private: CoreWindow^ m_window; ComPtr m_swapChain; ComPtr m_d3dDevice; ComPtr m_d3dDeviceContext; ComPtr m_renderTargetView; public: // This method is called on application launch. void Initialize( _In_ CoreWindow^ window, _In_ CoreApplicationView^ applicationView ) { m_window = window; } // This method is called after Initialize. void Load(_In_ Platform::String^ entryPoint) { } // This method is called after Load. void Run() { // First, create the Direct3D device. // This flag is required in order to enable compatibility with Direct2D. UINT creationFlags = D3D11_CREATE_DEVICE_BGRA_SUPPORT; #if defined(_DEBUG) // If the project is in a debug build, enable debugging via SDK Layers with this flag. creationFlags |= D3D11_CREATE_DEVICE_DEBUG; #endif // This array defines the ordering of feature levels that D3D should attempt to create. D3D_FEATURE_LEVEL featureLevels[] = { D3D_FEATURE_LEVEL_11_1, D3D_FEATURE_LEVEL_11_0, D3D_FEATURE_LEVEL_10_1, D3D_FEATURE_LEVEL_10_0, D3D_FEATURE_LEVEL_9_3, D3D_FEATURE_LEVEL_9_1 }; ComPtr d3dDevice; ComPtr d3dDeviceContext; DX::ThrowIfFailed( D3D11CreateDevice( nullptr, // specify nullptr to use the default adapter D3D_DRIVER_TYPE_HARDWARE, nullptr, // leave as nullptr if hardware is used creationFlags, // optionally set debug and Direct2D compatibility flags featureLevels, ARRAYSIZE(featureLevels), D3D11_SDK_VERSION, // always set this to D3D11_SDK_VERSION &d3dDevice, nullptr, &d3dDeviceContext ) ); // Retrieve the Direct3D 11.1 interfaces. DX::ThrowIfFailed( d3dDevice.As(&m_d3dDevice) ); DX::ThrowIfFailed( d3dDeviceContext.As(&m_d3dDeviceContext) ); // After the D3D device is created, create additional application resources. CreateWindowSizeDependentResources(); // Create a Basic Reader-Writer class to load data from disk. This class is examined // in the Resource Loading sample. BasicReaderWriter^ reader = ref new BasicReaderWriter(); // Load the raw vertex shader bytecode from disk and create a vertex shader with it. auto vertexShaderBytecode = reader-ReadData("SimpleVertexShader.cso"); ComPtr vertexShader; DX::ThrowIfFailed( m_d3dDevice-CreateVertexShader( vertexShaderBytecode-Data, vertexShaderBytecode-Length, nullptr, &vertexShader ) ); // Create an input layout that matches the layout defined in the vertex shader code. // For this lesson, this is simply a float2 vector defining the vertex position. const D3D11_INPUT_ELEMENT_DESC basicVertexLayoutDesc[] = { { "POSITION", 0, DXGI_FORMAT_R32G32_FLOAT, 0, 0, D3D11_INPUT_PER_VERTEX_DATA, 0 }, }; ComPtr inputLayout; DX::ThrowIfFailed( m_d3dDevice-CreateInputLayout( basicVertexLayoutDesc, ARRAYSIZE(basicVertexLayoutDesc), vertexShaderBytecode-Data, vertexShaderBytecode-Length, &inputLayout ) ); // Load the raw pixel shader bytecode from disk and create a pixel shader with it. auto pixelShaderBytecode = reader-ReadData("SimplePixelShader.cso"); ComPtr pixelShader; DX::ThrowIfFailed( m_d3dDevice-CreatePixelShader( pixelShaderBytecode-Data, pixelShaderBytecode-Length, nullptr, &pixelShader ) ); // Create vertex and index buffers that define a simple triangle. float3 triangleVertices[] = { float3(-0.5f, -0.5f,13.5f), float3( 0.0f, 0.5f,0), float3( 0.5f, -0.5f,0), }; D3D11_BUFFER_DESC vertexBufferDesc = {0}; vertexBufferDesc.ByteWidth = sizeof(float3) * ARRAYSIZE(triangleVertices); vertexBufferDesc.Usage = D3D11_USAGE_DEFAULT; vertexBufferDesc.BindFlags = D3D11_BIND_VERTEX_BUFFER; vertexBufferDesc.CPUAccessFlags = 0; vertexBufferDesc.MiscFlags = 0; vertexBufferDesc.StructureByteStride = 0; D3D11_SUBRESOURCE_DATA vertexBufferData; vertexBufferData.pSysMem = triangleVertices; vertexBufferData.SysMemPitch = 0; vertexBufferData.SysMemSlicePitch = 0; ComPtr vertexBuffer; DX::ThrowIfFailed( m_d3dDevice-CreateBuffer( &vertexBufferDesc, &vertexBufferData, &vertexBuffer ) ); // Once all D3D resources are created, configure the application window. // Allow the application to respond when the window size changes. m_window-SizeChanged += ref new TypedEventHandler( this, &Direct3DTutorialViewProvider::OnWindowSizeChanged ); // Specify the cursor type as the standard arrow cursor. m_window-PointerCursor = ref new CoreCursor(CoreCursorType::Arrow, 0); // Activate the application window, making it visible and enabling it to receive events. m_window-Activate(); // Enter the render loop. Note that tailored applications should never exit. while (true) { // Process events incoming to the window. m_window-Dispatcher-ProcessEvents(CoreProcessEventsOption::ProcessAllIfPresent); // Specify the render target we created as the output target. ID3D11RenderTargetView* targets[1] = {m_renderTargetView.Get()}; m_d3dDeviceContext-OMSetRenderTargets( 1, targets, NULL // use no depth stencil ); // Clear the render target to a solid color. const float clearColor[4] = { 0.071f, 0.04f, 0.561f, 1.0f }; //Code fails here m_d3dDeviceContext-ClearRenderTargetView( m_renderTargetView.Get(), clearColor ); m_d3dDeviceContext-IASetInputLayout(inputLayout.Get()); // Set the vertex and index buffers, and specify the way they define geometry. UINT stride = sizeof(float3); UINT offset = 0; m_d3dDeviceContext-IASetVertexBuffers( 0, 1, vertexBuffer.GetAddressOf(), &stride, &offset ); m_d3dDeviceContext-IASetPrimitiveTopology(D3D11_PRIMITIVE_TOPOLOGY_TRIANGLELIST); // Set the vertex and pixel shader stage state. m_d3dDeviceContext-VSSetShader( vertexShader.Get(), nullptr, 0 ); m_d3dDeviceContext-PSSetShader( pixelShader.Get(), nullptr, 0 ); // Draw the cube. m_d3dDeviceContext-Draw(3,0); // Present the rendered image to the window. Because the maximum frame latency is set to 1, // the render loop will generally be throttled to the screen refresh rate, typically around // 60Hz, by sleeping the application on Present until the screen is refreshed. DX::ThrowIfFailed( m_swapChain-Present(1, 0) ); } } // This method is called before the application exits. void Uninitialize() { } private: // This method is called whenever the application window size changes. void OnWindowSizeChanged( _In_ CoreWindow^ sender, _In_ WindowSizeChangedEventArgs^ args ) { m_renderTargetView = nullptr; CreateWindowSizeDependentResources(); } // This method creates all application resources that depend on // the application window size. It is called at app initialization, // and whenever the application window size changes. void CreateWindowSizeDependentResources() { if (m_swapChain != nullptr) { // If the swap chain already exists, resize it. DX::ThrowIfFailed( m_swapChain-ResizeBuffers( 2, 0, 0, DXGI_FORMAT_R8G8B8A8_UNORM, 0 ) ); } else { // If the swap chain does not exist, create it. DXGI_SWAP_CHAIN_DESC1 swapChainDesc = {0}; swapChainDesc.Stereo = false; swapChainDesc.BufferUsage = DXGI_USAGE_RENDER_TARGET_OUTPUT; swapChainDesc.Scaling = DXGI_SCALING_NONE; swapChainDesc.Flags = 0; // Use automatic sizing. swapChainDesc.Width = 0; swapChainDesc.Height = 0; // This is the most common swap chain format. swapChainDesc.Format = DXGI_FORMAT_R8G8B8A8_UNORM; // Don't use multi-sampling. swapChainDesc.SampleDesc.Count = 1; swapChainDesc.SampleDesc.Quality = 0; // Use two buffers to enable flip effect. swapChainDesc.BufferCount = 2; // We recommend using this swap effect for all applications. swapChainDesc.SwapEffect = DXGI_SWAP_EFFECT_FLIP_SEQUENTIAL; // Once the swap chain description is configured, it must be // created on the same adapter as the existing D3D Device. // First, retrieve the underlying DXGI Device from the D3D Device. ComPtr dxgiDevice; DX::ThrowIfFailed( m_d3dDevice.As(&dxgiDevice) ); // Ensure that DXGI does not queue more than one frame at a time. This both reduces // latency and ensures that the application will only render after each VSync, minimizing // power consumption. DX::ThrowIfFailed( dxgiDevice-SetMaximumFrameLatency(1) ); // Next, get the parent factory from the DXGI Device. ComPtr dxgiAdapter; DX::ThrowIfFailed( dxgiDevice-GetAdapter(&dxgiAdapter) ); ComPtr dxgiFactory; DX::ThrowIfFailed( dxgiAdapter-GetParent( __uuidof(IDXGIFactory2), &dxgiFactory ) ); // Finally, create the swap chain. DX::ThrowIfFailed( dxgiFactory-CreateSwapChainForImmersiveWindow( m_d3dDevice.Get(), DX::GetIUnknown(m_window), &swapChainDesc, nullptr, // allow on all displays &m_swapChain ) ); } // Once the swap chain is created, create a render target view. This will // allow Direct3D to render graphics to the window. ComPtr backBuffer; DX::ThrowIfFailed( m_swapChain-GetBuffer( 0, __uuidof(ID3D11Texture2D), &backBuffer ) ); DX::ThrowIfFailed( m_d3dDevice-CreateRenderTargetView( backBuffer.Get(), nullptr, &m_renderTargetView ) ); // After the render target view is created, specify that the viewport, // which describes what portion of the window to draw to, should cover // the entire window. D3D11_TEXTURE2D_DESC backBufferDesc = {0}; backBuffer-GetDesc(&backBufferDesc); D3D11_VIEWPORT viewport; viewport.TopLeftX = 0.0f; viewport.TopLeftY = 0.0f; viewport.Width = static_cast(backBufferDesc.Width); viewport.Height = static_cast(backBufferDesc.Height); viewport.MinDepth = D3D11_MIN_DEPTH; viewport.MaxDepth = D3D11_MAX_DEPTH; m_d3dDeviceContext-RSSetViewports(1, &viewport); } }; // This class defines how to create the custom View Provider defined above. ref class Direct3DTutorialViewProviderFactory : IViewProviderFactory { public: IViewProvider^ CreateViewProvider() { return ref new Direct3DTutorialViewProvider(); } }; [Platform::MTAThread] int main(array^) { auto viewProviderFactory = ref new Direct3DTutorialViewProviderFactory(); Windows::ApplicationModel::Core::CoreApplication::Run(viewProviderFactory); return 0; }

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• When should I use Areas in TFS instead of Team Projects

  - by Martin Hinshelwood

  Well, it depends…. If you are a small company that creates a finite number of internal projects then you will find it easier to create a single project for each of your products and have TFS do the heavy lifting with reporting, SharePoint sites and Version Control. But what if you are not… Update 9th March 2010 Michael Fourie gave me some feedback which I have integrated. Ed Blankenship via @edblankenship offered encouragement and a nice quote. Ewald Hofman gave me a couple of Cons, and maybe a few more soon. Ewald’s company, Avanade, currently uses Areas, but it looks like the manual management is getting too much and the project is getting cluttered. What if you are likely to have hundreds of projects, possibly with a multitude of internal and external projects? You might have 1 project for a customer or 10. This is the situation that most consultancies find themselves in and thus they need a more sustainable and maintainable option. What I am advocating is that we should have 1 “Team Project” per customer, and use areas to create “sub projects” within that single “Team Project”. "What you describe is what we generally do internally and what we recommend. We make very heavy use of area path to categorize the work within a larger project." - Brian Harry, Microsoft Technical Fellow & Product Unit Manager for Team Foundation Server   "We tend to use areas to segregate multiple projects in the same team project and it works well." - Tiago Pascoal, Visual Studio ALM MVP   "In general, I believe this approach provides consistency [to multi-product engagements] and lowers the administration and maintenance costs. All good." - Michael Fourie, Visual Studio ALM MVP   “@MrHinsh BTW, I'm very much a fan of very large, if not huge, team projects in TFS. Just FYI :) Use Areas & Iterations.” Ed Blankenship, Visual Studio ALM MVP   This would mean that SSW would have a single Team Project called “SSW” that contains all of our internal projects and consequently all of the Areas and Iteration move down one hierarchy to accommodate this. Where we would have had “\SSW\Sprint 1” we now have “\SSW\SqlDeploy\Sprint1” with “SqlDeploy” being our internal project. At the moment SSW has over 70 internal projects and more than 170 total projects in TFS. This method has long term benefits that help to simplify the support model for companies that often have limited internal support time and many projects. But, there are implications as TFS does not provide this model “out-of-the-box”. These implications stretch across Areas, Iterations, Queries, Project Portal and Version Control. Michael made a good comment, he said: I agree with your approach, assuming that in a multi-product engagement with a client, they are happy to adopt the same process template across all products. If they are not, then it’ll either be easy to convince them or there is a valid reason for having a different template - Michael Fourie, Visual Studio ALM MVP   At SSW we have a standard template that we use and this is applied across the board, to all of our projects. We even apply any changes to the core process template to all of our existing projects as well. If you have multiple projects for the same clients on multiple templates and you want to keep it that way, then this approach will not work for you. However, if you want to standardise as we have at SSW then this approach may benefit you as well. Implications around Areas Areas should be used for topological classification/isolation of work items. You can think of this as architecture areas, organisational areas or even the main features of your application. In our scenario there is an additional top level item that represents the Project / Product that we want to chop our Team Project into. Figure: Creating a sub area to represent a product/project is easy. <teamproject> <teamproject>\<Functional Area/module whatever> Becomes: <teamproject> <teamproject>\<ProjectName>\ <teamproject>\<ProjectName>\<Functional Area/module whatever> Implications around Iterations Iterations should be used for chronological classification/isolation of work items. This could include isolated time boxes, milestones or release timelines and really depends on the logical flow of your project or projects. Due to the new level in Area we need to add the same level to Iteration. This is primarily because it is unlikely that the sprints in each of your projects/products will start and end at the same time. This is just a reality of managing multiple projects. Figure: Adding the same Area value to Iteration as the top level item adds flexibility to Iteration. <teamproject>\Sprint 1 Or <teamproject>\Release 1\Sprint 1 Becomes: <teamproject>\<ProjectName>\Sprint 1 Or <teamproject>\<ProjectName>\Release 1\Sprint 1 Implications around Queries Queries are used to filter your work items based on a specified level of granularity. There are a number of queries that are built into a project created using the MSF Agile 5.0 template, but we now have multiple projects and it would be a pain to have to edit all of the work items every time we changed project, and that would only allow one team to work on one project at a time.   Figure: The Queries that are created in a normal MSF Agile 5.0 project do not quite suit our new needs. In order for project contributors to be able to query based on their project we need a couple of things. The first thing I did was to create an “_Area Template” folder that has a copy of the project layout with all the queries setup to filter based on the “_Area Template” Area and the “_Sprint template” you can see in the Area and Iteration views. Figure: The template is currently easily drag and drop, but you then need to edit the queries to point at the right Area and Iteration. This needs a tool. I then created an “Areas” folder to hold all of the area specific queries. So, when you go to create a new TFS Sub-Project you just drag “_Area Template” while holding “Ctrl” and drop it onto “Areas”. There is a little setup here. That said I managed it in around 10 minutes which is not so bad, and I can imagine it being quite easy to build a tool to create these queries Figure: These new queries can be configured in around 10 minutes, which includes setting up the Area and Iteration as well. Version Control What about your source code? Well, that is the easiest of the lot. Just create a sub folder for each of your projects/products.   Figure: Creating sub folders in source control is easy as “Right click | Create new folder”. <teamproject>\DEV\Main\ Becomes: <teamproject>\<ProjectName>\DEV\Main\ Conclusion I think it is up to each company to make a call on how you want to configure your Team Projects and it depends completely on how many projects/products you are going to have for each customer including yourself. If we decide to utilise this route it will require some configuration to get our 170+ projects into this format, and I will probably be writing some tools to help. Pros You only have one project to upgrade when a process template changes – After going through an upgrade of over 170 project prior to the changes in the RC I can tell you that that many projects is no fun. Standardises your Process Template – You will always have the same Process implementation across projects/products without exception You get tighter control over the permissions – Yes, you can do this on a standard Team Project, but it gets a lot easier with practice. You can “move” work items from one “product” to another – Have we not always wanted to do that. You can rename your projects – Wahoo: everyone wants to do this, now you can. One set of Reporting Services reports to manage – You set an area and iteration to run reports anyway, so you may as well set both. Simplified Check-In Policies– There is only one set of check-in policies per client. This simplifies administration of policies. Simplified Alerts – As alerts are applied across multiple projects this simplifies your alert rules as per client. Cons All of these cons could be mitigated by a custom tool that helps automate creation of “Sub-projects” within Team Projects. This custom tool could create areas, Iteration, permissions, SharePoint and queries. It just does not exist yet :) You need to configure the Areas and Iterations You need to configure the permissions You may need to configure sub sites for SharePoint (depends on your requirement) – If you have two projects/products in the same Team Project then you will not see the burn down for each one out-of-the-box, but rather a cumulative for the Team Project. This is not really that much of a problem as you would have to configure your burndown graphs for your current iteration anyway. note: When you create a sub site to a TFS linked portal it will inherit the settings of its parent site :) This is fantastic as it means that you can easily create sub sites and then set the Area and Iteration path in each of the reports to be the correct one. Every team wants their own customization (via Ewald Hofman) - small teams of 2 persons against teams of 30 – or even outsourcing – need their own process, you cannot allow that because everybody gets the same work item types. note: Luckily at SSW this is not a problem as our template is standardised across all projects and customers. Large list of builds (via Ewald Hofman) – As the build list in Team Explorer is just a flat list it can get very cluttered. note: I would mitigate this by removing any build that has not been run in over 30 days. The build template and workflow will still be available in version control, but it will clean the list. Feedback Now that I have explained this method, what do you think? What other pros and cons can you see? What do you think of this approach? Will you be using it? What tools would you like to support you?   Technorati Tags: Visual Studio ALM,TFS Administration,TFS,Team Foundation Server,Project Planning,TFS Customisation

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• 500 Metro Style WP7 Icons

  - by Bil Simser

  I was inspired by The Noun Project, a project that offers up “Metro-style” icons in SVG format. The project is licensed under a public domain license and while it’s a great project, all of the content is in SVG format. Jon Galloway has a great post (from 2007) talking about the differences between SVG and XAML so I highly recommend that for some background. I thought it would be helpful to the WPF/Windows Phone 7/Silverlight community to provide the content in alternative formats for use in your applications. The Goods I’ve put together a package of the 500 icons (502 actually) in PNG, XAML and the original SVG format along with a couple of sample projects so you can see them in action. There’s a WPF desktop app: And a Windows Phone 7 app: Building It To get all the content first I wrote up a quick program to suck the original SVG files. Luckily they’re all in a common path just named 1.SVG, 2.SVG, and so on. Easy sleazy to grab the contents. Once I had 500 SVG files I used the latest copy of XamlTune, an open source CodePlex project that has a command line conversion tool to convert the directory of SVG files into XAML (the tool also created a PNG file of each SVG so that’s just icing on the cake). Conversions The conversion from SVG to XAML isn’t 100%. While you can just drop the content into a WPF app, it doesn’t work that way for WP7. There are just some small adjustments I made to each format so you’ll have to do the same. Follow the information below or refer to the sample applications. As a sample, here’s an icon we want to use: Here’s the original SVG file: <svg version="1.0" id="Layer_1" xmlns="http://www.w3.org/2000/svg" xmlns:xlink="http://www.w3.org/1999/xlink" x="0px" y="0px" width="100px" height="94.616px" viewBox="0 0 100 94.616" enable-background="new 0 0 100 94.616" xml:space="preserve"> <path d="M25.076,15.639c4.324,0.009,7.824-3.488,7.82-7.82C32.9,3.512,29.4,0.012,25.076,0c-4.313,0.012-7.814,3.512-7.821,7.819 C17.262,12.15,20.763,15.648,25.076,15.639L25.076,15.639z"/> <path d="M4.593,43.388h6.861l4.137-15.135h1.716L13.22,43.388h24.318l-4.389-15.135h1.817l2.32,7.415 c1.08,3.131,3.852,3.851,6.003,1.162l8.375-10.142c2.651-3.42-2.104-7.021-4.844-4.035l-4.993,5.952 c0.007,0.095-0.96-3.278-0.96-3.278c-1.135-3.978-4.918-7.903-10.595-7.922H19.576c-5.071,0.019-9.043,4.434-9.888,7.214 L4.593,43.388L4.593,43.388z"/> <polygon points="56.206,22.753 56.206,7.163 49.192,7.163 49.192,22.753 56.206,22.753 "/> <path d="M79.87,15.738c4.332-0.014,7.831-3.516,7.82-7.82c0.011-4.332-3.488-7.833-7.82-7.82c-4.306-0.013-7.806,3.488-7.821,7.82 C72.064,12.222,75.564,15.725,79.87,15.738L79.87,15.738z"/> <path d="M89.759,89.556v-43.19h5.751V22.804c0.007-3.079-2.757-5.448-6.71-5.449H70.436c-3.65,0.001-4.539,1.186-5.551,2.168 L49.597,37.889c-3.098,3.848,2.428,8.333,5.55,4.743L69.88,25.226v64.43c-0.019,6.475,9.06,6.686,9.081,0.201v-36.58h1.765v36.379 C80.748,96.109,89.772,96.13,89.759,89.556L89.759,89.556z"/> <polygon points="100,54.035 100,45.155 0,45.155 0,54.035 100,54.035 "/> </svg> Here’s the XAML that XamlTune created. It can be used in any WPF app without any changes: <Canvas Name="Layer_1" Width="100" Height="94.616" ClipToBounds="True" xmlns="http://schemas.microsoft.com/winfx/2006/xaml/presentation"> <Path Fill="#FF000000"> <Path.Data> <PathGeometry FillRule="Nonzero" Figures="M25.076,15.639C29.4,15.648 32.9,12.151 32.896,7.819 32.9,3.512 29.4,0.012 25.076,0 20.763,0.012 17.262,3.512 17.255,7.819 17.262,12.15 20.763,15.648 25.076,15.639L25.076,15.639z" /> </Path.Data> </Path> <Path Fill="#FF000000"> <Path.Data> <PathGeometry FillRule="Nonzero" Figures="M4.593,43.388L11.454,43.388 15.591,28.253 17.307,28.253 13.22,43.388 37.538,43.388 33.149,28.253 34.966,28.253 37.286,35.668C38.366,38.799,41.138,39.519,43.289,36.83L51.664,26.688C54.315,23.268,49.56,19.667,46.82,22.653L41.827,28.605C41.834,28.7 40.867,25.327 40.867,25.327 39.732,21.349 35.949,17.424 30.272,17.405L19.576,17.405C14.505,17.424,10.533,21.839,9.688,24.619L4.593,43.388 4.593,43.388z" /> </Path.Data> </Path> <Path Fill="#FF000000"> <Path.Data> <PathGeometry FillRule="Nonzero" Figures="M56.206,22.753L56.206,7.163 49.192,7.163 49.192,22.753 56.206,22.753z" /> </Path.Data> </Path> <Path Fill="#FF000000"> <Path.Data> <PathGeometry FillRule="Nonzero" Figures="M79.87,15.738C84.202,15.724 87.701,12.222 87.69,7.918 87.701,3.586 84.202,0.0849999999999991 79.87,0.097999999999999 75.564,0.084999999999999 72.064,3.586 72.049,7.918 72.064,12.222 75.564,15.725 79.87,15.738L79.87,15.738z" /> </Path.Data> </Path> <Path Fill="#FF000000"> <Path.Data> <PathGeometry FillRule="Nonzero" Figures="M89.759,89.556L89.759,46.366 95.51,46.366 95.51,22.804C95.517,19.725,92.753,17.356,88.8,17.355L70.436,17.355C66.786,17.356,65.897,18.541,64.885,19.523L49.597,37.889C46.499,41.737,52.025,46.222,55.147,42.632L69.88,25.226 69.88,89.656C69.861,96.131,78.94,96.342,78.961,89.857L78.961,53.277 80.726,53.277 80.726,89.656C80.748,96.109,89.772,96.13,89.759,89.556L89.759,89.556z" /> </Path.Data> </Path> <Path Fill="#FF000000"> <Path.Data> <PathGeometry FillRule="Nonzero" Figures="M100,54.035L100,45.155 0,45.155 0,54.035 100,54.035z" /> </Path.Data> </Path> </Canvas> The XAML works AS-IS in a WPF application but there are some changes I did to get it to work in a WP7 app. Here’s the modified XAML in a WP7 application: <Canvas Grid.Row="0" Grid.Column="0" Name="Icon_1" Width="100" Height="94.616"> <Path Fill="#FF000000" Data="M25.076,15.639C29.4,15.648 32.9,12.151 32.896,7.819 32.9,3.512 29.4,0.012 25.076,0 20.763,0.012 17.262,3.512 17.255,7.819 17.262,12.15 20.763,15.648 25.076,15.639L25.076,15.639z"> </Path> <Path Fill="#FF000000" Data="M4.593,43.388L11.454,43.388 15.591,28.253 17.307,28.253 13.22,43.388 37.538,43.388 33.149,28.253 34.966,28.253 37.286,35.668C38.366,38.799,41.138,39.519,43.289,36.83L51.664,26.688C54.315,23.268,49.56,19.667,46.82,22.653L41.827,28.605C41.834,28.7 40.867,25.327 40.867,25.327 39.732,21.349 35.949,17.424 30.272,17.405L19.576,17.405C14.505,17.424,10.533,21.839,9.688,24.619L4.593,43.388 4.593,43.388z"> </Path> <Path Fill="#FF000000" Data="M56.206,22.753L56.206,7.163 49.192,7.163 49.192,22.753 56.206,22.753z"> </Path> <Path Fill="#FF000000" Data="M79.87,15.738C84.202,15.724 87.701,12.222 87.69,7.918 87.701,3.586 84.202,0.0849999999999991 79.87,0.097999999999999 75.564,0.084999999999999 72.064,3.586 72.049,7.918 72.064,12.222 75.564,15.725 79.87,15.738L79.87,15.738z"> </Path> <Path Fill="#FF000000" Data="M89.759,89.556L89.759,46.366 95.51,46.366 95.51,22.804C95.517,19.725,92.753,17.356,88.8,17.355L70.436,17.355C66.786,17.356,65.897,18.541,64.885,19.523L49.597,37.889C46.499,41.737,52.025,46.222,55.147,42.632L69.88,25.226 69.88,89.656C69.861,96.131,78.94,96.342,78.961,89.857L78.961,53.277 80.726,53.277 80.726,89.656C80.748,96.109,89.772,96.13,89.759,89.556L89.759,89.556z"> </Path> <Path Fill="#FF000000" Data="M100,54.035L100,45.155 0,45.155 0,54.035 100,54.035z"> </Path> </Canvas> All I did was take the data portion and put it directly into a Data attribute on the Path. Note that while it does show up in the app (on the emulator or device) it wouldn’t show up in Visual Studio for me. Maybe some XAML guru out there can tell me why. You can just as easily use the PNG files in WP7 but if you want the crispness of vector graphics, go for the XAML version. Of course with XamlTune being open source you could always modify the output of that program to cater it to your app. If you do make a change that’s worthy please consider submitting a patch to the project so everyone can benefit. Hope this helps and happy programming! Resources and Links Sample Project and Icons XamlTune an open source project to convert SVG to XAML The Noun Project source of the original files Jon Galloways post on SVG and XAML StackOverflow question on converting SVG to XAML

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• Pluralsight Meet the Author Podcast on Structuring JavaScript Code

  - by dwahlin

  I had the opportunity to talk with Fritz Onion from Pluralsight about one of my recent courses titled Structuring JavaScript Code for one of their Meet the Author podcasts. We talked about why JavaScript patterns are important for building more re-useable and maintainable apps, pros and cons of different patterns, and how to go about picking a pattern as a project is started. The course provides a solid walk-through of converting what I call “Function Spaghetti Code” into more modular code that’s easier to maintain, more re-useable, and less susceptible to naming conflicts. Patterns covered in the course include the Prototype Pattern, Revealing Module Pattern, and Revealing Prototype Pattern along with several other tips and techniques that can be used. Meet the Author:  Dan Wahlin on Structuring JavaScript Code   The transcript from the podcast is shown below: [Fritz]  Hello, this is Fritz Onion with another Pluralsight author interview. Today we’re talking with Dan Wahlin about his new course, Structuring JavaScript Code. Hi, Dan, it’s good to have you with us today. [Dan]  Thanks for having me, Fritz. [Fritz]  So, Dan, your new course, which came out in December of 2011 called Structuring JavaScript Code, goes into several patterns of usage in JavaScript as well as ways of organizing your code and what struck me about it was all the different techniques you described for encapsulating your code. I was wondering if you could give us just a little insight into what your motivation was for creating this course and sort of why you decided to write it and record it. [Dan]  Sure. So, I got started with JavaScript back in the mid 90s. In fact, back in the days when browsers that most people haven’t heard of were out and we had JavaScript but it wasn’t great. I was on a project in the late 90s that was heavy, heavy JavaScript and we pretty much did what I call in the course function spaghetti code where you just have function after function, there’s no rhyme or reason to how those functions are structured, they just kind of flow and it’s a little bit hard to do maintenance on it, you really don’t get a lot of reuse as far as from an object perspective. And so coming from an object-oriented background in JAVA and C#, I wanted to put something together that highlighted kind of the new way if you will of writing JavaScript because most people start out just writing functions and there’s nothing with that, it works, but it’s definitely not a real reusable solution. So the course is really all about how to move from just kind of function after function after function to the world of more encapsulated code and more reusable and hopefully better maintenance in the process. [Fritz]  So I am sure a lot of people have had similar experiences with their JavaScript code and will be looking forward to seeing what types of patterns you’ve put forth. Now, a couple I noticed in your course one is you start off with the prototype pattern. Do you want to describe sort of what problem that solves and how you go about using it within JavaScript? [Dan]  Sure. So, the patterns that are covered such as the prototype pattern and the revealing module pattern just as two examples, you know, show these kind of three things that I harp on throughout the course of encapsulation, better maintenance, reuse, those types of things. The prototype pattern specifically though has a couple kind of pros over some of the other patterns and that is the ability to extend your code without touching source code and what I mean by that is let’s say you’re writing a library that you know either other teammates or other people just out there on the Internet in general are going to be using. With the prototype pattern, you can actually write your code in such a way that we’re leveraging the JavaScript property and by doing that now you can extend my code that I wrote without touching my source code script or you can even override my code and perform some new functionality. Again, without touching my code.  And so you get kind of the benefit of the almost like inheritance or overriding in object oriented languages with this prototype pattern and it makes it kind of attractive that way definitely from a maintenance standpoint because, you know, you don’t want to modify a script I wrote because I might roll out version 2 and now you’d have to track where you change things and it gets a little tricky. So with this you just override those pieces or extend them and get that functionality and that’s kind of some of the benefits that that pattern offers out of the box. [Fritz]  And then the revealing module pattern, how does that differ from the prototype pattern and what problem does that solve differently? [Dan]  Yeah, so the prototype pattern and there’s another one that’s kind of really closely lined with revealing module pattern called the revealing prototype pattern and it also uses the prototype key word but it’s very similar to the one you just asked about the revealing module pattern. [Fritz]  Okay. [Dan]  This is a really popular one out there. In fact, we did a project for Microsoft that was very, very heavy JavaScript. It was an HMTL5 jQuery type app and we use this pattern for most of the structure if you will for the JavaScript code and what it does in a nutshell is allows you to get that encapsulation so you have really a single function wrapper that wraps all your other child functions but it gives you the ability to do public versus private members and this is kind of a sort of debate out there on the web. Some people feel that all JavaScript code should just be directly accessible and others kind of like to be able to hide their, truly their private stuff and a lot of people do that. You just put an underscore in front of your field or your variable name or your function name and that kind of is the defacto way to say hey, this is private. With the revealing module pattern you can do the equivalent of what objective oriented languages do and actually have private members that you literally can’t get to as an external consumer of the JavaScript code and then you can expose only those members that you want to be public. Now, you don’t get the benefit though of the prototype feature, which is I can’t easily extend the revealing module pattern type code if you don’t like something I’m doing, chances are you’re probably going to have to tweak my code to fix that because we’re not leveraging prototyping but in situations where you’re writing apps that are very specific to a given target app, you know, it’s not a library, it’s not going to be used in other apps all over the place, it’s a pattern I actually like a lot, it’s very simple to get going and then if you do like that public/private feature, it’s available to you. [Fritz]  Yeah, that’s interesting. So it’s almost, you can either go private by convention just by using a standard naming convention or you can actually enforce it by using the prototype pattern. [Dan]  Yeah, that’s exactly right. [Fritz]  So one of the things that I know I run across in JavaScript and I’m curious to get your take on is we do have all these different techniques of encapsulation and each one is really quite different when you’re using closures versus simply, you know, referencing member variables and adding them to your objects that the syntax changes with each pattern and the usage changes. So what would you recommend for people starting out in a brand new JavaScript project? Should they all sort of decide beforehand on what patterns they’re going to stick to or do you change it based on what part of the library you’re working on? I know that’s one of the points of confusion in this space. [Dan]  Yeah, it’s a great question. In fact, I just had a company ask me about that. So which one do I pick and, of course, there’s not one answer fits all. [Fritz]  Right. [Dan]  So it really depends what you just said is absolutely in my opinion correct, which is I think as a, especially if you’re on a team or even if you’re just an individual a team of one, you should go through and pick out which pattern for this particular project you think is best. Now if it were me, here’s kind of the way I think of it. If I were writing a let’s say base library that several web apps are going to use or even one, but I know that there’s going to be some pieces that I’m not really sure on right now as I’m writing I and I know people might want to hook in that and have some better extension points, then I would look at either the prototype pattern or the revealing prototype. Now, really just a real quick summation between the two the revealing prototype also gives you that public/private stuff like the revealing module pattern does whereas the prototype pattern does not but both of the prototype patterns do give you the benefit of that extension or that hook capability. So, if I were writing a library that I need people to override things or I’m not even sure what I need them to override, I want them to have that option, I’d probably pick a prototype, one of the prototype patterns. If I’m writing some code that is very unique to the app and it’s kind of a one off for this app which is what I think a lot of people are kind of in that mode as writing custom apps for customers, then my personal preference is the revealing module pattern you could always go with the module pattern as well which is very close but I think the revealing module patterns a little bit cleaner and we go through that in the course and explain kind of the syntax there and the differences. [Fritz]  Great, that makes a lot of sense. [Fritz]  I appreciate you taking the time, Dan, and I hope everyone takes a chance to look at your course and sort of make these decisions for themselves in their next JavaScript project. Dan’s course is, Structuring JavaScript Code and it’s available now in the Pluralsight Library. So, thank you very much, Dan. [Dan]  Thanks for having me again.

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