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  • How can I check if PHP was compiled with the UNICODE version of the Win32 API?

    - by Wesley Murch
    This is related to this Stack Overflow post: glob() can't find file names with multibyte characters on Windows? I'm having issues with PHP and files that have multibyte characters on Windows. Here's my test case: print_r(scandir('./uploads/')); print_r(glob('./uploads/*')); Correct Output on remote UNIX server: Array ( [0] => . [1] => .. [2] => filename-äöü.jpg [3] => filename.jpg [4] => test?test.jpg [5] => ??? ?????.jpg [6] => ?????????.jpg [7] => ???.jpg ) Array ( [0] => ./uploads/filename-äöü.jpg [1] => ./uploads/filename.jpg [2] => ./uploads/test?test.jpg [3] => ./uploads/??? ?????.jpg [4] => ./uploads/?????????.jpg [5] => ./uploads/???.jpg ) Incorrect Output locally on Windows: Array ( [0] => . [1] => .. [2] => ??? ?????.jpg [3] => ???.jpg [4] => ?????????.jpg [5] => filename-äöü.jpg [6] => filename.jpg [7] => test?test.jpg ) Array ( [0] => ./uploads/filename-äöü.jpg [1] => ./uploads/filename.jpg ) Here's a relevant excerpt from the answer I chose to accept (which actually is a quote from an article that was posted online over 2 years ago): From the comments on this article: http://www.rooftopsolutions.nl/blog/filesystem-encoding-and-php The output from your PHP installation on Windows is easy to explain : you installed the wrong version of PHP, and used a version not compiled to use the Unicode version of the Win32 API. For this reason, the filesystem calls used by PHP will use the legacy "ANSI" API and so the C/C++ libraries linked with this version of PHP will first try to convert yout UTF-8-encoded PHP string into the local "ANSI" codepage selected in the running environment (see the CHCP command before starting PHP from a command line window) Your version of Windows is MOST PROBABLY NOT responsible of this weird thing. Actually, this is YOUR version of PHP which is not compiled correctly, and that uses the legacy ANSI version of the Win32 API (for compatibility with the legacy 16-bit versions of Windows 95/98 whose filesystem support in the kernel actually had no direct support for Unicode, but used an internal conversion layer to convert Unicode to the local ANSI codepage before using the actual ANSI version of the API). Recompile PHP using the compiler option to use the UNICODE version of the Win32 API (which should be the default today, and anyway always the default for PHP installed on a server that will NEVER be Windows 95 or Windows 98...) I can't confirm whether this is my problem or not. I used phpinfo() and did not find anything interesting, but I wasn't sure what to look for. I've been using XAMPP for easy installations, so I'm really not sure exactly how it was installed. I'm using Windows 7, 64 bit - so forgive my ignorance, but I'm not even sure if "Win32" is relevant here. How can I check if my current version of PHP was compiled with the configuration mentioned above? PHP Version: 5.3.8 System: Windows NT WES-PC 6.1 build 7601 (Windows 7 Home Premium Edition Service Pack 1) i586 Build Date: Aug 23 2011 11:47:20 Compiler: MSVC9 (Visual C++ 2008) Architecture: x86 Configure Command: cscript /nologo configure.js "--enable-snapshot-build" "--disable-isapi" "--enable-debug-pack" "--disable-isapi" "--without-mssql" "--without-pdo-mssql" "--without-pi3web" "--with-pdo-oci=D:\php-sdk\oracle\instantclient10\sdk,shared" "--with-oci8=D:\php-sdk\oracle\instantclient10\sdk,shared" "--with-oci8-11g=D:\php-sdk\oracle\instantclient11\sdk,shared" "--enable-object-out-dir=../obj/" "--enable-com-dotnet" "--with-mcrypt=static" "--disable-static-analyze"

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  • Alias wordpress folder from within another website

    - by Bretticus
    I have a little dilemma. I wrote a custom PHP MVC framework and built a CMS on top of it. I decided to give nginx+fpm a whirl too. Which is the root of my dilemma. I was asked to incorporate a wordpress blog into my website (yah.) It has much content and it's not feasible in the short amount of time I have to bring all the content into my CMS. Because of using Apache for years, I'm, admittedly, a little lost using nginx. My website has the file path: /opt/directories/mysite/public/ The wordpress files are located at: /opt/directories/mysite/news/ I know I just need to setup location(s) that targets /news[/*] and then forces all matching URI's to the index.php within. Can someone point me in the right direction perhaps? My configuration is below: server { listen 80; server_name staging.mysite.com index index.php; root /opt/directories/mysite/public; access_log /var/log/nginx/mysite/access.log; error_log /var/log/nginx/mysite/error.log; add_header X-NodeName directory01; location = /favicon.ico { log_not_found off; access_log off; } location = /robots.txt { allow all; log_not_found off; access_log off; } location / { try_files $uri $uri/ /index.php?route=$uri&$args; } location ~ /news { try_files $uri $uri/ @news; } location @news { fastcgi_pass unix:/tmp/php-fpm.sock; fastcgi_split_path_info ^(/news)(/.*)$; fastcgi_param SCRIPT_FILENAME /opt/directories/mysite/news/index.php; fastcgi_param PATH_INFO $fastcgi_path_info; } include fastcgi_params; include php.conf; location ~* ^.+.(jpg|jpeg|gif|css|png|js|ico|xml)$ { access_log off; expires 30d; } ## Disable viewing .htaccess & .htpassword location ~ /\.ht { deny all; } } My php.conf file: location ~ \.php { fastcgi_param QUERY_STRING $query_string; fastcgi_param REQUEST_METHOD $request_method; fastcgi_param CONTENT_TYPE $content_type; fastcgi_param CONTENT_LENGTH $content_length; fastcgi_param SCRIPT_NAME $fastcgi_script_name; fastcgi_param SCRIPT_FILENAME $document_root$fastcgi_script_name; fastcgi_param REQUEST_URI $request_uri; fastcgi_param DOCUMENT_URI $document_uri; fastcgi_param DOCUMENT_ROOT $document_root; fastcgi_param SERVER_PROTOCOL $server_protocol; fastcgi_param GATEWAY_INTERFACE CGI/1.1; fastcgi_param SERVER_SOFTWARE nginx; fastcgi_param REMOTE_ADDR $remote_addr; fastcgi_param REMOTE_PORT $remote_port; fastcgi_param SERVER_ADDR $server_addr; fastcgi_param SERVER_PORT $server_port; fastcgi_param SERVER_NAME $server_name; fastcgi_pass unix:/tmp/php-fpm.sock; # If you must use PATH_INFO and PATH_TRANSLATED then add # the following within your location block above # (make sure $ does not exist after \.php or /index.php/some/path/ will not match): #fastcgi_split_path_info ^(.+\.php)(/.+)$; #fastcgi_param PATH_INFO $fastcgi_path_info; #fastcgi_param PATH_TRANSLATED $document_root$fastcgi_path_info; } fastcgi_params file: fastcgi_connect_timeout 60; fastcgi_send_timeout 180; fastcgi_read_timeout 180; fastcgi_buffer_size 128k; fastcgi_buffers 4 256k; fastcgi_busy_buffers_size 256k; fastcgi_temp_file_write_size 256k; fastcgi_intercept_errors on; Thanks, in large part, to @Kromey, I have adjusted my location /news/ but I am still not getting the desired result. I was able to learn to tack a ~ my /news location as I discovered that my php location was being matched first. With this setup, I now get a 200 status, but the page is blank. Any ideas?

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  • Drupal & nginx : a sound "general purpose" configuration?

    - by sbrattla
    After a bit back and forth with configuring Drupal and nginx to work together, I've come up with the below configuration for a site. It works well, both with private and public file systems. However, as I am fairly new to nginx I'd like to hear if there is something with this configuration that I should change (for Please note! I'm aiming towards getting feedback on a general purpose Drupal configuration. That is, a configuration which others who are trying out Drupal + nginx can "copy paste" to get up and running. server { listen 80; server_name www.example.* example.*; access_log /home/example/www/logs/access.log; error_log /home/example/www/logs/error.log; root /home/example/www/public_html; # Site Icon location = /favicon.ico { log_not_found off; access_log off; } # Search Engines location = /robots.txt { allow all; log_not_found off; access_log off; } # Drush location = /backup { deny all; } # Very rarely should these ever be accessed from outside # the local network. location ~* \.(txt|log)$ { allow 10.0.0.0/8; allow 172.16.0.0/12; allow 192.168.0.0/16; deny all; } location ~ \..*/.*\.php$ { return 403; } # Default location location / { try_files $uri @rewrite; } # Files managed by Drupal will be served via PHP. location ~* /system/files/ { access_log off; try_files $uri @rewrite; } ## Images and static content is treated different location ~* \.(jpg|jpeg|gif|css|png|js|ico|xml)$ { access_log off; expires 30d; } # Some Drupal modules enforce no slash (/) at the end # of the URL. location @rewrite { rewrite_log on; rewrite ^/(.*)$ /index.php?q=$1; } # PHP5-FPM is used to handle PHP. location ~ \.php$ { fastcgi_split_path_info ^(.+\.php)(.*)$; fastcgi_pass unix:/var/run/example.sock; fastcgi_index index.php; fastcgi_param SCRIPT_FILENAME $document_root$fastcgi_script_name; include fastcgi_params; fastcgi_param QUERY_STRING $query_string; fastcgi_param REQUEST_METHOD $request_method; fastcgi_param CONTENT_TYPE $content_type; fastcgi_param CONTENT_LENGTH $content_length; fastcgi_intercept_errors on; fastcgi_ignore_client_abort off; fastcgi_connect_timeout 60; fastcgi_send_timeout 180; fastcgi_read_timeout 180; fastcgi_buffer_size 128k; fastcgi_buffers 4 256k; fastcgi_busy_buffers_size 256k; fastcgi_temp_file_write_size 256k; } ## Disable viewing .htaccess & .htpassword location ~ /\.ht { deny all; } }

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  • batch file to deploy files

    - by Martin Michalak
    hi I have created batch file which pulls info from *.txt file and deploy code from the source to destination: SET Source=%1 if exist %Source% ( ECHO Source for WEB exists ) else ( ECHO Wrong build%Source% doesn't exist GOTO Menu ) SET Server=%2 SET AppPool=%3 SET Destination=%4 SET Folder=%5 SET ENV=%6 SET AppName=%7 SET Envlog=%8 ECHO Deployment of WEB > %Envlog% %Date% %Time% echo. @ECHO Stopping App Pools @ECHO Stopping App Pools >> %Envlog% %Date% %Time% D:\ICTTools\PSEXEC.EXE -d \\%Server% cmd.exe /c c:\windows\system32\inetsrv\appcmd STOP apppool /apppool.name:%AppPool% echo. @ECHO App Pools will be stopped in the background @ECHO App Pools will be stopped in the background >> %Envlog% %Date% %Time% Pause echo. IF EXIST "%Destination%" ( ECHO Deleting %AppName% %Folder% RMDIR %Destination% /s /q ECHO Destination Folder %Folder% Deleted ECHO Destination Folder %Folder% Deleted >> %Envlog% %Date% %Time% ) else ( ECHO Destination Folder %Destination% does not exist, please check ECHO Destination Folder %Destination% does not exist, please check >> %Envlog% %Date% %Time% Pause ) echo. @ECHO Starting Robocopy for %AppName% @ECHO Starting Robocopy for %AppName% >> %Envlog% %Date% %Time% echo. START /WAIT /MIN ROBOCOPY.EXE %Source% %Destination% *.* /S /NP /R:3 /W:5 /LOG:"Logs\Robo%AppName%%ENV%.log" D:\Tools\Windiff\windiff.exe %Source% %Destination% echo. @ECHO Finished with Robocopy @ECHO Finished with Robocopy >> %Envlog% %Date% %Time% echo. @ECHO Checking if App pools stopped: @ECHO Checking if App pools stopped: >> %Envlog% %Date% %Time% D:\ICTTools\PSEXEC.EXE \\%Server% c:\windows\system32\inetsrv\appcmd LIST apppool /apppool.name:%AppPool% @echo off set /p ask=All app pools stopped? (y/n) if %ask%==y (echo Great, please continue with deployemnt) else echo Before continuing please check why app pools did not stop @echo App pools stopped?: %ask% >> %Envlog% %Date% %Time% DEL %Source%\web.config echo. @ECHO Production Config check if exist "%Destination%\%ENV%-Web.config" ( echo. ECHO The Application production configuration file does exist. ECHO The Application production configuration file does exist. >> %Envlog% %Date% %Time% COPY %Destination%\%ENV%-Web.config web.config echo. ECHO Production %ENV%-Web.config has been renamed to web.config ECHO Production %ENV%-Web.config has been renamed to web.config >> %Envlog% %Date% %Time% ) else ( ECHO The Application production configuration file is missing in Production %AppName% ECHO The Application production configuration file is missing in Production %AppName% >> %Envlog% %Date% %Time% explorer %Destination% Pause ) echo. @ECHO Confirm that configs were renamed correclty, if yes please hit any key to START APP Pools @ECHO Confirm that configs were renamed correclty, if yes please hit any key to START APP Pools >> %Envlog% %Date% %Time% Pause echo. @ECHO Start %AppName% Application Pool >> %Envlog% %Date% %Time% D:\ICTTools\PSEXEC.EXE \\%Server% c:\windows\system32\inetsrv\appcmd START apppool /apppool.name:%AppPool% @echo off set /p ask=All app pools started? (y/n) if %ask%==y (echo Great, please continue with deployemnt) else echo Before continuing please check why app pools did not start @echo App pools started?: %ask% >> %Envlog% %Date% %Time% Pause echo. @ECHO Build Version for %AppName% @ECHO Build Version for %AppName% >> %Envlog% %Date% %Time% type %Destination%\buildinfo.xml echo. ECHO ............................................... @ECHO ...........Deployment Compelted................ @ECHO ...........Deployment Compelted................>> %Envlog% %Date% %Time% ECHO ............................................... here are my issues: Lets say I am running code for 3 servers, then for each instance: For all three servers I am performing destination folder delete even so destination folder is always the same, the code should only delete it in the 1st instance (when code is deployed to first server) then I would prefer if script would check if the code from the source and destination is the same and if it is it should delete the folder or not. Then based on 1: a) deleting web.config and renaming should only happen if code in destination is new b) Robocopy should not override files if they are the same I think there is /Xo option to do that any idea how to achieve that? :)

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  • IPv6: Should I have private addresses?

    - by AlReece45
    Right now, we have a rack of servers. Every server right now has at least 2 IP addresses, one for the public interface, another for the private. The servers that have SSL websites on them have more IP addresses. We also have virtual servers, that are configured similarly. Private Network The private range is currently just used for backups and monitoring. Its a gigabit port, the interface usage does not usually get very high. There are other technologies we're considering using that would use this port: iSCSI (implementations usually recommends dedicating an interface to it, which would be yet another IP network), VPN to get access to the private range (something I'd rather avoid) dedicated database servers LDAP centralized configuration (like puppet) centralized logging We don't have any private addresses in our DNS records (only public addresses). For our servers to utilize the correct IP address for the right interface (and not hard code the IP address) probably requires setting up a private DNS server (So now we add 2 different dns entries to 2 different systems). Public Network Our public range has a variety of services include web, email, and ftp. There is a hardware firewall between our network and the "public" network. We have (relatively secure) method to instruct the firewall to open and close administrative access (web interfaces, ssh, etc) for our current IP address. With either solution discussed, the host-based firewalls will be configured as well. The public network currently runs at a dedicated 20Mbps link. There are a couple of legacy servers with fast-ethernet ports, but they are scheduled for decommissioning. All of the other production boxes have at least 2 Gigabit Ethernet ports. The more traffic-heavy servers have 4-6 available (none is using more than the 2 Gigabit ports right now). IPv6 I want to get an IPv6 prefix from our ISP. So at least every "server" has at least one IPv6 interface. We'll still need to keep the IPv4 addressees up and available for legacy clients (web servers and email at the very least). We have two IP networks right now. Adding the public IPv6 address would make it three. Just use IPv6? I'm thinking about just dumping the private IPv4 range and using the IPv6 range as the primary means of all communications. If an interface starts reaching its capacity, utilize the newly free interfaces to create a trunk. It has the advantage that if either the public or private traffic needs to exceed 1Gbps. The traffic for each interface is already analyzed on a regular basis to predict future bandwidth use. In the rare instances where bandwidth unexpected peaks: utilize QoS to ensure traffic (like our limited SSH access) is prioritized correctly so the problem can be corrected (if possible, our WAN is the bottleneck right now). It also has the advantage of not needing to make an entry for every private address. We may have private DNS (or just LDAP), but it'll be much more limited in scope with less entries to duplicate. Summary I'm trying to make this network as "simple" as possible. At the same time, I want to make sure its reliable, upgradeable, scalable, and (eventually) redundant. Having one IPv6 network, and a legacy IPv4 network seems to be the best solution to me. Regarding using assigned IPv6 addresses for both networks, sharing the available bandwidth on one (more trunked if needed): Are there any technical disadvantages (limitations, buffers, scalability)? Are there any other security considerations (asides from firewalls mentioned above) to consider? Are there regulations or other security requirements (like PCI-DSS) that this doesn't meet? Is there typical software for setting up a Linux network that doesn't have IPv6 support yet? (logging, ldap, puppet) Some other thing I didn't consider?

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  • Choice of an OS for a home ZFS NAS

    - by OlafM
    I am preparing a home NAS with an old Athlon 64 X2 3800+, 4 GB ECC RAM, Asus M2V MX motherboard, and a single 3 TB WDC Green (another one as mirror may be installed in the future). It's the cheapest solution I found that includes ECC memory and the higher energy consumption is offset by the lower (zero) cost of acquisition. The system will be used for: music storage and stream to other desktop computers; storage of the scanned dia slides (3-4k slides, 180 MB TIFF each one plus reduced quality JPEG version); stream of these photos to a local iPad 2 (maybe Plex App? not yet sure); (one additional) remote backup via rsync/ssh or ZFS send/receive. It will be controlled via remote ssh, maybe VNC, no monitor attached. Absolute requirement is a reliable ZFS solution, plus the ability to easily install packets/software/virtual machines and to update remotely (I will be the admin and I don't live near the NAS). I have mainly three options: NAS4free/FreeNAS OpenIndiana Solaris Express 11 (yeah yeah I know the license requirements, I will write a perl script on it to count it as development machine). Problems: NAS4free/FreeNAS (I tested only NAS4free) required embedded installation for remote upgrading, but full install for easy addition of software packets. Since I need at least AirVideo Server (linux/win) and Plex App (win/linux) to stream the photos and some videos to iPad (they both require virtualbox), but I cannot be there to install updates, NAS4free/FreeNAS are excluded. http://www.nas4free.org/general_information.html explains the issue: embedded can be remotely updated, full cannot. Solaris has also another advantage: Crashplan client supports Solaris and I'm already using it for other backups. I would like to leave the option open, even if I will be doing backups probably through zfs send/receive. NexentaStor was left out because zfs send/receive are not included in the free version. The question is now Solaris 11 Express over OpenIndiana. To ease the management, I will be using http://www.napp-it.org Which one would you suggest and why? I found lots of informations and it's difficult for me to decide. I think (from the napp-it manual) that Solaris has some additional options for SMB shares, but are they really needed at home? I think I won't even use ACLs, since normal unix-style permissions are enough. OpenIndiana has maybe more frequent updates (Solaris offers only security updates between releases), but again, do I need them? I don't think so. Moreover, this is a NAS that has to work and nothing else, I cannot risk having problems that require me to access the server. Isn't OpenIndiana a bit more... cutting edge (in the Solaris world)? I'm just asking, no need to focus on this for the answer :-) I would limit myself to these two options (SE11.1/OI) also because I will be making a NAS for me in the future (where high performances with Mac shares are also required) and Solaris has kernel support for AFP. I will use this server to gather experience as well. After this long question, thanks in advance! If you need additional info, let me know and I will update this post. UPDATES Given the first answers, I will strongly suggest the person paying the hardware to insert a second HD. Better 2x2TB than 1x3TB (3 TB is oversized anyway). I was trying to keep the initial costs down to spread them over a longer period, but better having something good from the beginning.

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  • Specifying a Postfix Instance to send outbound email

    - by Catherine Jefferson
    I have a CentOS 6.5 server running Postfix 2.6x (the default distribution) with five public IPv4 IPs bound to it. Each IP has DNS and rDNS set separately. Each uses a different hostname at a different domain. I have five Postfix instances, one bound to each IP, like this example: 192.168.34.104 red.example.com /etc/postfix 192.168.36.48 green.example.net /etc/postfix-green 192.168.36.49 pink.example.org /etc/postfix-pink 192.168.36.50 orange.example.info /etc/postfix-orange 192.168.36.51 blue.example.us /etc/postfix-blue I've tested each IP by telneting to port 25. Postfix answers and banners properly with the correct hostname. Email is received on all of these instances with no problems and is routed to the correct place. This setup, minus the final instance, has existed for a couple of years and works. I never bothered to set up outbound email to go through any but the main instance, however; there was no need. Now I need to send email from blue.example.us that actually leaves from that interface and IP, such that the Received headers show blue.example.us as the sending mailhost, so that SPF and DKIM validate, etc etc. The email that will be sent from blue.example.com is a feedback loop sent by a single shell account on the server (account5), an account that is dedicated to sending this email. The account receives the feedback loop emails from servers on other networks, saves the bodies of those emails, and then generates a new outbound email header, appends the saved body, and sends the email. It's sending by piping each email to sendmail -oi -t. We're doing it this way to mask the identities of the initial servers. The procmail script that processes these emails works correctly. However, I cannot configure this account to send email through the proper Postfix instance/IP/interface. The exact same account and script sends email through the main Postfix instance /etc/postfix without any issues. When I change MAIL_CONFIG to point to /etc/postfix-blue in either .bash_profile or the Procmail script that handles this email, though, I get this error: sendmail: fatal: User account5(###) is not allowed to submit mail I've read the manuals on Postfix.org, searched Google, and tried the suggestions in three previous answers here on ServerFault.com: Postfix - specify interface to deliver outbound mail on Postfix user is not allowed to submit mail Postfix rejects php mails I have been careful to stop and restart Postfix after each configuration change, and tested the results. Nothing has worked. The main postfix instance happily accepts outbound email from account5. The postfix-blue instance continues to reject email from account5 with the sendmail error above. As tempting as it is to blame machine hostility, I know that I must be missing something or doing something wrong. Does anybody have any suggestions as to what it might be? Please feel free to ask for further information about my setup if you need it. =-=-=-=-=-=-=-=-=-= At the request of the responder, here are main.cf and master.cf for a) the main postfix instance ("red.example.com") and b) the FBL instance ("blue.example.us") [NOTE: All parameters not specified below were left at the default Postfix 2.6 settings] MAIN: master.cf smtp inet n - n - - smtpd main.cf myhostname = red.example.com mydomain = example.com inet_interfaces = $myhostname, localhost inet_protocols = all lmtp_host_lookup = native smtp_host_lookup = native ignore_mx_lookup_error = yes mydestination = $myhostname, localhost.$mydomain, localhost local_recipient_maps = mynetworks = 192.168.34.104/32 relay_domains = example.com, example.info, example.net, example.org, example.us relayhost = [192.168.34.102] # Separate physical server, main mailserver. relay_recipient_maps = hash:/etc/postfix/relay_recipients alias_maps = hash:/etc/aliases alias_database = hash:/etc/aliases smtpd_banner = $myhostname ESMTP $mail_name multi_instance_wrapper = ${command_directory}/postmulti -p -- multi_instance_enable = yes multi_instance_directories = /etc/postfix-green /etc/postfix-pink /etc/postfix-orange /etc/postfix-blue FBL: master.cf 184.173.119.103:25 inet n - n - - smtpd main.cf myhostname = blue.example.us mydomain = blue.example.us <= Deliberately set to subdomain only. myorigin = $mydomain inet_interfaces = $myhostname lmtp_host_lookup = native smtp_host_lookup = native ignore_mx_lookup_error = yes mydestination = $myhostname local_recipient_maps = unix:passwd.byname $alias_maps $virtual_alias_maps mynetworks = 192.168.36.51/32, 192.168.35.20/31 <= Second IP is backup MX servers relay_domains = $mydestination recipient_canonical_maps = hash:/etc/postfix-blue/canonical virtual_alias_maps = hash:/etc/postfix-fbl/virtual alias_maps = hash:/etc/aliases, hash:/etc/postfix-blue/canonical alias_maps = hash:/etc/aliases, hash:/etc/postfix-blue/canonical mailbox_command = /usr/bin/procmail -a "$EXTENSION" DEFAULT=$HOME/Mail/ MAILDIR=$HOME/Mail smtpd_banner = $myhostname ESMTP $mail_name authorized_submit_users = multi_instance_name = postfix-blue multi_instance_enable = yes

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  • website connection reset on first load

    - by Tar
    i'm using nginx with php-cgi. lately a problem has arose where if you don't view my site for a while, like 3-4 minutes, and then open it again, the first request you send will return connection reset by peer in the browser. if you refresh, operation is normal for all subsequent requests. this happens every time and it isn't just an isolated incident, it happens to everyone using my site. i've tried to restart nginx and php-cgi but to no avail. does anyone know what the problem could be? i can provide whatever information necessary. it's worth noting that there's nothing in error log besides that message about client closing the connection early. nginx.conf user nobody; worker_processes 4; error_log /var/log/nginx/error.log; pid /var/run/nginx.pid; events { worker_connections 2048; } http { include /etc/nginx/mime.types; error_page 404 /404.html; error_page 403 /403.html; error_page 444 /444.html; error_page 502 /502.html; default_type application/octet-stream; log_format main '$remote_addr - $remote_user [$time_local] "$request" ' '$status $body_bytes_sent "$http_referer" ' '"$http_user_agent" "$http_x_forwarded_for"'; access_log /var/log/nginx/access.log main; large_client_header_buffers 8 8k; sendfile on; tcp_nopush on; tcp_nodelay on; keepalive_timeout 30; server_tokens off; gzip on; gzip_proxied any; gzip_comp_level 6; gzip_buffers 64 8k; gzip_min_length 1024; gzip_http_version 1.1; gzip_types text/plain text/css application/json application/x-javascript text/xml application/xml application/xml+rss text/javascript; include /etc/nginx/conf.d/*.conf; } default.conf server { listen 80; server_name domain.com; error_log /var/log/nginx/error.log debug; access_log /var/log/nginx/access.log; location / { if ($request_method !~ ^(GET|HEAD|POST)$ ) { return 444; } if ($http_user_agent ~* Havij|hvj|acunetix|wget|HTtrack) { return 403; } root /home/admin06/public_html; autoindex off; index index.php; # Images and static content is treated different location ~* ^.+.(jpg|jpeg|gif|css|png|js|ico|xml)$ { access_log off; expires 30d; root /home/admin06/public_html; } location /nginx_status { stub_status on; access_log off;] deny all; } location ~ \.php$ { fastcgi_split_path_info ^(.+\.php)(/.*)$; #try_files $uri =404; fastcgi_pass backend; fastcgi_index index.php; fastcgi_param SCRIPT_FILENAME /home/site/public_html$fastcgi_script_name; include fastcgi_params; fastcgi_param QUERY_STRING $query_string; fastcgi_param REQUEST_METHOD $request_method; fastcgi_param CONTENT_TYPE $content_type; fastcgi_param CONTENT_LENGTH $content_length; fastcgi_intercept_errors on; fastcgi_ignore_client_abort off; fastcgi_connect_timeout 60; fastcgi_send_timeout 60; fastcgi_read_timeout 60; fastcgi_buffer_size 128k; fastcgi_buffers 4 256k; fastcgi_busy_buffers_size 256k; fastcgi_temp_file_write_size 256k; } ## Disable viewing .htaccess & .htpassword location ~ /\.ht { deny all; } location ~ error_log { deny all; } location ~ access_log { deny all; } location ~ \.cgi { deny all; } location ~ \.db { deny all; } }

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  • In Linux, what's the best way to delegate administration responsibilities, like for Apache, a database, or some other application?

    - by Andrew Banks
    In Linux, what's the best way to delegate administration responsibilities for Apache and other "applications"? File permissions? Sudo? A mix of both? Something else? At work we have two tiers of "administrators" Operating system administrators. These are your run-of-the-mill "server administrators." They are responsible for just the operating system. Application administrators. The people who build the web site. This includes not only writing the SQL, PHP, and HTML, but also setting up and running Apache and PostgreSQL or MySQL. The aforementioned OS admins will install this stuff, but it's mainly up to the app admins to edit all the config files, start and stop processes when needed, and so on. I am one of the app admins. This is different than what I am used to. I used to just write code. The sysadmin took care not only of the OS but also installing, setting up, and keeping up the server software. But he left. Now I'm in charge of setting up Apache and the database. The new sysadmins say they just handle the operating system. It's no problem. I welcome learning new stuff. But there is a learning curve, even for the OS admins. Apache, by default, seems to be set up for administration by root directly. All the config files and scripts are 644 and owned by root:root. I'm not given the root password, naturally, so the OS admins must somehow give my ordinary OS user account all the rights necessary to edit Apache's config files, start and stop it, read its log files, and so on. Right now they're using a mix of: (1) giving me certain sudo rights, (2) adding me to certain groups, and (3) changing the file permissions of various directories, to make them writable by one of the groups I'm in. This never goes smoothly. There's always a back-and-forth between me and the sysadmins. They say it's ready. Then I try certain things, and half of them I still can't do. So they make some more changes. Then finally I seem to be independent and can administer Apache and the database without pestering them anymore. It's the sheer complication and amount of changes that make me uncomfortable. Even though it finally works, more or less, it seems hackneyed. I feel like we're doing it wrong. It seems like the makers of the software would have anticipated this scenario (someone other than root administering it) and have a clean two- or three-step program to delegate responsibility to me. But it feels like we are really chewing up the filesystem and making it far and away from the default set-up. Any suggestions? Are we doing it the recommended way? P.S. For PostgreSQL it seems a little better. Its files are owned by a system user named postgres. So giving me the right to run sudo su - postgres gives me just about everything. I'm just now getting into MySQL, but it seems to be set up similarly. But it seems a little weird doing all my work as another user.

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  • nginx + php-fpm - where are my $_GET params?

    - by egis
    I have a strange problem here. I just moved from apache + mod_php to nginx + php-fpm. Everything went fine except this one problem. I have a site, let's say example.com. When I access it like example.com?test=get_param $_SERVER['REQUEST_URI'] is /?test=get_param and there is a $_GET['test'] also. But when I access example.com/ajax/search/?search=get_param $_SERVER['REQUEST_URI'] is /ajax/search/?search=get_param yet there is no $_GET['search'] (there is no $_GET array at all). I'm using Kohana framework. which routes /ajax/search to controller, but I've put phpinfo() at index.php so I'm checking for $_GET variables before framework does anything (this means that disapearing get params aren't frameworks fault). My nginx.conf is like this worker_processes 4; pid logs/nginx.pid; events { worker_connections 1024; } http { index index.html index.php; autoindex on; autoindex_exact_size off; include mime.types; default_type application/octet-stream; server_names_hash_bucket_size 128; log_format main '$remote_addr - $remote_user [$time_local] "$request" ' '$status $body_bytes_sent "$http_referer" ' '"$http_user_agent" "$http_x_forwarded_for"'; access_log logs/access.log main; error_log logs/error.log debug; sendfile on; tcp_nopush on; tcp_nodelay off; keepalive_timeout 2; gzip on; gzip_comp_level 2; gzip_proxied any; gzip_types text/plain text/css application/x-javascript text/xml application/xml application/xml+rss text/javascript; include sites-enabled/*; } and example.conf is like this server { listen 80; server_name www.example.com; rewrite ^ $scheme://example.com$request_uri? permanent; } server { listen 80; server_name example.com; root /var/www/example/; location ~ /\. { return 404; } location / { try_files $uri $uri/ /index.php; } location ~ \.php$ { fastcgi_pass 127.0.0.1:9000; fastcgi_index index.php; fastcgi_param SCRIPT_FILENAME $document_root$fastcgi_script_name; include /usr/local/nginx/conf/fastcgi_params; } location ~* ^/(modules|application|system) { return 403; } # serve static files directly location ~* ^.+.(jpg|jpeg|gif|css|png|js|ico|html|xml|txt)$ { access_log off; expires 30d; } } fastcgi_params is like this fastcgi_param QUERY_STRING $query_string; fastcgi_param REQUEST_METHOD $request_method; fastcgi_param CONTENT_TYPE $content_type; fastcgi_param CONTENT_LENGTH $content_length; fastcgi_param SCRIPT_NAME $fastcgi_script_name; fastcgi_param REQUEST_URI $request_uri; fastcgi_param DOCUMENT_URI $document_uri; fastcgi_param DOCUMENT_ROOT $document_root; fastcgi_param SERVER_PROTOCOL $server_protocol; fastcgi_param GATEWAY_INTERFACE CGI/1.1; fastcgi_param SERVER_SOFTWARE nginx/$nginx_version; fastcgi_param REMOTE_ADDR $remote_addr; fastcgi_param REMOTE_PORT $remote_port; fastcgi_param SERVER_ADDR $server_addr; fastcgi_param SERVER_PORT $server_port; fastcgi_param SERVER_NAME $server_name; # PHP only, required if PHP was built with --enable-force-cgi-redirect fastcgi_param REDIRECT_STATUS 200; fastcgi_connect_timeout 60; fastcgi_send_timeout 180; fastcgi_read_timeout 180; fastcgi_buffer_size 128k; fastcgi_buffers 4 256k; fastcgi_busy_buffers_size 256k; fastcgi_temp_file_write_size 256k; fastcgi_intercept_errors on; fastcgi_param QUERY_STRING $query_string; fastcgi_param PATH_INFO $fastcgi_path_info; What is the problem here? By the way there are few more sites on the same server, both Kohana based and plain php, that are working perfectly.

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  • nginx + php-fpm - where are my $_GET params?

    - by egis
    Hello everyone, I have a strange problem here. I just moved from apache + mod_php to nginx + php-fpm. Everything went fine except this one problem. I have a site, let's say example.com. When I access it like example.com?test=get_param $_SERVER['REQUEST_URI'] is /?test=get_param and there is a $_GET['test'] also. But when I access example.com/ajax/search/?search=get_param $_SERVER['REQUEST_URI'] is /ajax/search/?search=get_param yet there is no $_GET['search'] (there is no $_GET array at all). I'm using Kohana framework. which routes /ajax/search to controller, but I've put phpinfo() at index.php so I'm checking for $_GET variables before framework does anything (this means that disapearing get params aren't frameworks fault). My nginx.conf is like this worker_processes 4; pid logs/nginx.pid; events { worker_connections 1024; } http { index index.html index.php; autoindex on; autoindex_exact_size off; include mime.types; default_type application/octet-stream; server_names_hash_bucket_size 128; log_format main '$remote_addr - $remote_user [$time_local] "$request" ' '$status $body_bytes_sent "$http_referer" ' '"$http_user_agent" "$http_x_forwarded_for"'; access_log logs/access.log main; error_log logs/error.log debug; sendfile on; tcp_nopush on; tcp_nodelay off; keepalive_timeout 2; gzip on; gzip_comp_level 2; gzip_proxied any; gzip_types text/plain text/css application/x-javascript text/xml application/xml application/xml+rss text/javascript; include sites-enabled/*; } and example.conf is like this server { listen 80; server_name www.example.com; rewrite ^ $scheme://example.com$request_uri? permanent; } server { listen 80; server_name example.com; root /var/www/example/; location ~ /\. { return 404; } location / { try_files $uri $uri/ /index.php; } location ~ \.php$ { fastcgi_pass 127.0.0.1:9000; fastcgi_index index.php; fastcgi_param SCRIPT_FILENAME $document_root$fastcgi_script_name; include /usr/local/nginx/conf/fastcgi_params; } location ~* ^/(modules|application|system) { return 403; } # serve static files directly location ~* ^.+.(jpg|jpeg|gif|css|png|js|ico|html|xml|txt)$ { access_log off; expires 30d; } } fastcgi_params is like this fastcgi_param QUERY_STRING $query_string; fastcgi_param REQUEST_METHOD $request_method; fastcgi_param CONTENT_TYPE $content_type; fastcgi_param CONTENT_LENGTH $content_length; fastcgi_param SCRIPT_NAME $fastcgi_script_name; fastcgi_param REQUEST_URI $request_uri; fastcgi_param DOCUMENT_URI $document_uri; fastcgi_param DOCUMENT_ROOT $document_root; fastcgi_param SERVER_PROTOCOL $server_protocol; fastcgi_param GATEWAY_INTERFACE CGI/1.1; fastcgi_param SERVER_SOFTWARE nginx/$nginx_version; fastcgi_param REMOTE_ADDR $remote_addr; fastcgi_param REMOTE_PORT $remote_port; fastcgi_param SERVER_ADDR $server_addr; fastcgi_param SERVER_PORT $server_port; fastcgi_param SERVER_NAME $server_name; # PHP only, required if PHP was built with --enable-force-cgi-redirect fastcgi_param REDIRECT_STATUS 200; fastcgi_connect_timeout 60; fastcgi_send_timeout 180; fastcgi_read_timeout 180; fastcgi_buffer_size 128k; fastcgi_buffers 4 256k; fastcgi_busy_buffers_size 256k; fastcgi_temp_file_write_size 256k; fastcgi_intercept_errors on; fastcgi_param QUERY_STRING $query_string; fastcgi_param PATH_INFO $fastcgi_path_info; What is the problem here? By the way there are few more sites on the same server, both Kohana based and plain php, that are working perfectly.

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  • Hard drive write speed - finding a lighter antivirus?

    - by Shingetsu
    I recently have been getting a lot of system lag here (for example, the mouse and the display in general take about 15 seconds to react in the worst cases). After a lot of monitoring the resources, I found that the problem mainly happens when too much Disk I/O is being done. Three culprits have been identified: My browser had the highest write I/O with 35,000,000 I/O Write Bytes. Steam had the highest read I/O (when IDLE!!!) with 106,000,000 I/O Read Bytes. My antivirus (in both cases I will soon mention) was the runner up in both cases with: 30,000,000ish write and 80,000,000ish read. The first AV I had was Avast! which I had liked on my previous system. After noticing it taking so much I/O I switched to Panda (supposing it wouldn't use TOO much during idle phase). However it only used a bit less I/O. Just a lot less memory and cpu and somewhat more network. My browser at the moment is Maxthon 3 (which I like a lot). Before this I was running chrome which had similar data and much higher cpu when running in the background was enabled. I'm not going to be running steam all the time and there aren't many alternatives to it. I like my browser very much, but I AM willing to switch if there's an obvious problem (I'm in programming, however I'm not a very good sysadmin, especially not when it comes to windows). Finally, my system almost stops lagging when I turn off the antivirus (and preferably steam) (some remains but once in every 5-6 hours for a few seconds so it isn't a big problem). My question (has a few parts): Is it possible to configure steam to lower it's I/O usage? (and maybe network while we're at it?) Which antivirus (very preferably free) uses lowest I/O while idle (I leave PC alone during active scans so that isn't a problem). Is there an obvious problem with my current browser and, if so, is there a way to fix it or should I switch and, if so, to what? (P.S. I've been on FFox for some time too). Info on system: Windows 7 (32 bit T_T, I am getting a new one in a few months but I want to keep using system during that time though). Hard Drive (main) is a Raid0. (Also have an external 1TB one which contains steam (and steam alone). As such it doesn't get used by much anything other than steam and isn't a very large problem. However steam still uses some I/O of registry) CPU: Intel(R) Core(TM)2 CPU [email protected] RAM: 6GB (3.25GB usable) (this and CPU have little effect as shown in next section) Additional info: Memory usage during problematic times: 44% CPU usage during problematic times: 35% Page File: main drive: system managed. 1TB drive: none. The current system I'm using is about 6 years old and is mainly a place holder while I await the new one in a few months. Final words: this is my 1st post on Super User (this question wouldn't feel right on Stack Overflow where I usually stay). If it doesn't have it's place here please tell me. If anything is wrong with it, same. Edit Technically I'm looking for a live thread detection program with minimal IO usage. I already have good active scan capability: Kaspersky (the free scanner uses the paid database) and MalwareBytes. Edit 2 Noticed another one, it seems that windows media player has been using stuff even when off! Turning it off and restarting now. If the problem is fixed I'll tell you guys. The reason I didn't notice it before was because I didn't have resource manager in front of me at the MOMENT of the problem. Now I did and it was at the very top of the list!

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  • Exchange Server 2007 Send and Receive Connectors

    - by Mistiry
    I have gotten awesome advice from users on here for getting Exchange on Windows SBS 2008 set up. I think this is the final piece and I'm ready for roll-out! I need to set up Exchange so that it RECEIVES mail from our existing mail server as a Forward [aliases on the existing mail server to forward mail from [email protected] to [email protected]] (not using the POP3 Connector), and SENDS mail through that server as well (sends from [email protected] to [email protected] and then out to the world, showing in the headers as from [email protected] or at absolute least have the reply-to set as this). Alternatively, as long as the .net email address doesn't show in the From and replies are directed to the .com account, email can go from Exchange to the outside world without directing through the existing mail server. External Domain: domain.com Internal Domain: domain.local Internet Domain Name Set in SBS Console: domain.net When I go to http://remote.domain.net I get the Remote Web Workspace, and can login to both Sharepoint and OWA. I can send an email from OWA to a GMail account. I receive it from [email protected], which is an alias of [email protected]. I cannot, however, send an email from OWA to ANY domain.com email addresses. I am also not receiving any email to this Exchange account (except for NDRs). When I try sending an email to a domain.com account, here is the error (I had to replace all < and with { and }): Delivery has failed to these recipients or distribution lists: [email protected] The recipient's e-mail address was not found in the recipient's e-mail system. Microsoft Exchange will not try to redeliver this message for you. Please check the e-mail address and try resending this message, or provide the following diagnostic text to your system administrator. Generating server: IFEXCHANGE.domain.local [email protected] #550 5.1.1 RESOLVER.ADR.RecipNotFound; not found ## Original message headers: Received: from IFEXCHANGE.domain.local ([fe80::4d34:abc5:f7fd:e51a]) by IFEXCHANGE.domain.local ([fe80::4d34:abc5:f7fd:e51a%10]) with mapi; Tue, 17 Aug 2010 14:14:14 -0400 Content-Type: application/ms-tnef; name="winmail.dat" Content-Transfer-Encoding: binary From: John Doe {[email protected]} To: "[email protected]" {[email protected]} Date: Tue, 17 Aug 2010 14:14:12 -0400 Subject: asdf Thread-Topic: asdf Thread-Index: AQHLPjf+h6hA5MJ1JUu1WS4I4CiWeA== Message-ID: {E4E10393768D784D8760A51938BA456A029934BA30@IFEXCHANGE.domain.local} Accept-Language: en-US Content-Language: en-US X-MS-Has-Attach: X-MS-TNEF-Correlator: {E4E10393768D784D8760A51938BA456A029934BA30@IFEXCHANGE.domain.local} MIME-Version: 1.0 I hope I explained the situation well enough for someone to be able to explain to me what I'm missing. If I could, I'd be putting a 10K bounty, but unfortunately I've got only 74 reputation (hey, I'm a newbie here!). I'm pretty sure the obvious "RecipNotFound" error is why its not working, my question is how to resolve this. The email account exists, it receives mail just fine, yet when I send it from the Exchange server it fails. EDIT In OC-Hub Transport, the Email Address Policies has 2 entries. "Windows SBS Email Address Policy" is set up to: Include All Recipient Types, no conditions, and SMTP %[email protected]. "Default Policy" set to: Include All Recipient Types, no conditions, and SMTP @domain.net. Three Authoritative Accepted domains domain.com domain.local (Default) domain.net Remote Domains tab has two entries. Default with domain * Windows SBS Company Web Domain with domain companyweb.

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  • Monitor randomly shutting down, computer accepting no input, need to restart to get working

    - by Sebastian Lamerichs
    First off, spec list: OS: Windows 7 Ultimate 64-bit SP1 CPU: i7-4820k @ 3.7GHz (stock) GPU: Two 3GB Radeon HD 7970s @ 1.05GHz Mobo: AsRock X79 Extreme6 HDD: 2TB Seagate Barracuda 7200rpm RAM: 16GB quad-channel Kingston 1600MHz PSU: Antec HCG 900W Monitors: Acer S220HQL 1920x1080 + ViewSonic VA2251 1920x1080. Plugged into different GPUs. My problem is that, on a daily-ish basis, my monitors will turn off and not turn back on. My computer will still be running, GPU/CPU/case fans all still going, but the monitors will not turn back on. Additionally, it seems to cease all network activity. It doesn't seem to log any errors at all. I've verified that this is not a monitor issue, as when I press the num/caps/scroll lock buttons on my keyboard, the lights don't change, so the computer is clearly not accepting input. I have noticed a few other people on the internet with this problem, and some have claimed that it was solved by disabling PCI-Express Link State Power Management, but the issue still occurs for me after this. Whilst my CPU and GPUs both run at 100% 24/7, the temperatures are certainly not at dangerous levels, with the CPU averaging 65°C and the GPUs at 70°C and 78°C average. All components are brand new. I have tried forcing MSI Afterburner to start when Windows starts and to force a constant voltage, as this fixed the issue for a few days for another user, but he reported back saying that it had stopped working properly again, so I'm not putting too much faith in this working. Many people have said to adjust display sleep mode settings, but this will clearly not work, as the keyboard lights would still work if the monitors were the issue. The closest I can get to a log file for this issue is the following Folding@Home logs: 14:45:21:WU01:FS00:0x17:Completed 1120000 out of 2000000 steps (56%) 14:46:43:WU00:FS01:0x17:Completed 480000 out of 2000000 steps (24%) 14:46:49:WU01:FS00:0x17:Completed 1140000 out of 2000000 steps (57%) 14:48:30:WU01:FS00:0x17:Completed 1160000 out of 2000000 steps (58%) 14:49:55:WU01:FS00:0x17:Completed 1180000 out of 2000000 steps (59%) As you can see, the second GPU (FS01) stops computation approximately three and a half minutes before the issue occurs (it should be completing 1% every 80-120 seconds), and the first GPU (FS00) continues for a few minutes more before the logs just end. As far as I can tell, the computer has a network failure at the time the first GPU stops working, the latest IRC message I received from this time was at 14:47:58. That being said, there could have just not been any messages between then and 14:50:00, so I'm going to be connecting a laptop to the same bouncer to double-check if it happens again. The GPUs functioned perfectly well in another computer for a significant period of time, so I'm fairly confident that they aren't the issue, which means that this is being caused by either software or the motherboard, or possibly RAM. I really hope it's software. I heard from a forum board that there was a patch from Microsoft that fixed this problem, but "I've forgot which KB it was or the google search terms I used to find the patch, LOL.", so that's not much help. Haven't seen it mentioned by anyone else on about a dozen threads about this issue either. The computer is plugged in via a surge-protected power board, and I've run several other computers and pieces of hardware through it with no issues, so that is not the cause. I have just set the hard disk to never turn off, although I don't believe that that will solve the issue. Strangely, this has only happened when I'm not at the computer (which is actually a minority of the time). Until today it had only happened when I had not been actively using the computer for 6 hours, but today it happened within 10-30 minutes of me last using the computer actively. I have enabled file logging from MSI Afterburner, so hopefully this will shed some light on the issue, but I'm not too optimistic. I've heard that it could be a motherboard problem, but I figured I should ask around before RMAing it. Any help?

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  • BIND no longer responds to AXFR Requests

    - by djsumdog
    Recently we moved our primary external DNS server. It has three caching DNS slaves in front of it provided by our ISP. They've told us they've started getting access denied requests when doing zone transfers (AXFR). If I add in my own IPs to the allow-transfer list, I also get a transfer failed when using dig with the AXFR argument. Here is what my bind configuration looks like: options { directory "/var/lib/named"; dump-file "/var/log/named_dump.db"; zone-statistics yes; statistics-file "/var/log/named.stats"; listen-on-v6 { any; }; notify-source 10.19.0.68 port 53; querylog yes; notify yes; allow-transfer { 127.0.0.1; //localhost 1.1.1.1; //public dns slave 1 2.2.2.2; //public dns slave 2 3.3.3.3; //public dns slave 3 }; also-notify { 1.1.1.1; //public dns slave 1 2.2.2.2; //public dns slave 2 3.3.3.3; //public dns slave 3 }; include "/etc/named.d/forwarders.conf"; }; logging { channel simple_log { file "/var/log/bind.log" versions 10 size 3m; severity info; print-time yes; print-severity yes; print-category yes; }; category default{ simple_log; }; channel log_zone_transfers { file "/var/log/axfr.log" versions 10 size 3m; print-time yes; print-category yes; print-severity yes; }; category xfer-out { log_zone_transfers; }; channel log_notify { file "/var/log/notify.log" versions 10 size 3m; print-time yes; print-category yes; print-severity yes; }; category notify { log_notify; }; channel queries { file "/var/log/queries.log" versions 10 size 30m; print-time yes; severity info; print-category yes; print-severity yes; }; category queries { queries; }; }; zone "." in { type hint; file "root.hint"; }; zone "localhost" in { type master; file "localhost.zone"; }; zone "0.0.127.in-addr.arpa" in { type master; file "127.0.0.zone"; }; include "/etc/named.conf.include"; zone "example.net " { type master; file "/var/lib/named/master/example.net.hosts"; }; zone "example.com " { type master; file "/var/lib/named/master/example.com.hosts"; }; ## -- other master files -- And the errors in the xfer log look like the following: 29-Oct-2012 14:20:02.806 xfer-out: info: client 1.1.1.1#59069: bad zone transfer request: 'example.com./IN': non-authoritative zone (NOTAUTH) I've tried adding allow-transfer parameters directly on the zone files and still get failed transfers. Any idea what I'm doing wrong?

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  • Using TypeScript in ASP.NET MVC Projects

    - by shiju
    In the previous blog post Microsoft TypeScript : A Typed Superset of JavaScript, I have given a brief introduction on TypeScript. In this post, I will demonstrate how to use TypeScript with ASP.NET MVC projects and how we can compile TypeScript within the ASP.NET MVC projects. Using TypeScript with ASP.NET MVC 3 Projects The Visual Studio plug-in for TypeScript provides an ASP.NET MVC 3 project template for TypeScript that lets you to compile TypeScript from the Visual Studio. The following screen shot shows the TypeScript template for ASP.NET MVC 3 project The “TypeScript Internet Application” template is just a ASP.NET MVC 3 internet application project template which will allows to compile TypeScript programs to JavaScript when you are building your ASP.NET MVC projects. This project template will have the following section in the .csproject file <None Include="Scripts\jquery.d.ts" /> <TypeScriptCompile Include="Scripts\site.ts" /> <Content Include="Scripts\site.js"> <DependentUpon>site.ts</DependentUpon> </Content> .csharpcode, .csharpcode pre { font-size: small; color: black; font-family: consolas, "Courier New", courier, monospace; background-color: #ffffff; /*white-space: pre;*/ } .csharpcode pre { margin: 0em; } .csharpcode .rem { color: #008000; } .csharpcode .kwrd { color: #0000ff; } .csharpcode .str { color: #006080; } .csharpcode .op { color: #0000c0; } .csharpcode .preproc { color: #cc6633; } .csharpcode .asp { background-color: #ffff00; } .csharpcode .html { color: #800000; } .csharpcode .attr { color: #ff0000; } .csharpcode .alt { background-color: #f4f4f4; width: 100%; margin: 0em; } .csharpcode .lnum { color: #606060; } <Target Name="BeforeBuild"> <Exec Command="&amp;quot;$(PROGRAMFILES)\ Microsoft SDKs\TypeScript\0.8.0.0\tsc&amp;quot; @(TypeScriptCompile ->'&quot;%(fullpath)&quot;', ' ')" /> </Target> .csharpcode, .csharpcode pre { font-size: small; color: black; font-family: consolas, "Courier New", courier, monospace; background-color: #ffffff; /*white-space: pre;*/ } .csharpcode pre { margin: 0em; } .csharpcode .rem { color: #008000; } .csharpcode .kwrd { color: #0000ff; } .csharpcode .str { color: #006080; } .csharpcode .op { color: #0000c0; } .csharpcode .preproc { color: #cc6633; } .csharpcode .asp { background-color: #ffff00; } .csharpcode .html { color: #800000; } .csharpcode .attr { color: #ff0000; } .csharpcode .alt { background-color: #f4f4f4; width: 100%; margin: 0em; } .csharpcode .lnum { color: #606060; } The “BeforeBuild” target will allows you to compile TypeScript programs when you are building your ASP.NET MVC projects. The TypeScript project template will provide a typing reference file for the jQuery library named “jquery.d.ts”. The following default app.ts file referenced to jquery.d.ts 1: ///<reference path='jquery.d.ts' /> 2:   3: $(document).ready(function () { 4:   5: $(".btn-slide").click(function () { 6: $("#main").slideToggle("slow"); 7: $(this).toggleClass("active"); 8: }); 9:   10: }); .csharpcode, .csharpcode pre { font-size: small; color: black; font-family: consolas, "Courier New", courier, monospace; background-color: #ffffff; /*white-space: pre;*/ } .csharpcode pre { margin: 0em; } .csharpcode .rem { color: #008000; } .csharpcode .kwrd { color: #0000ff; } .csharpcode .str { color: #006080; } .csharpcode .op { color: #0000c0; } .csharpcode .preproc { color: #cc6633; } .csharpcode .asp { background-color: #ffff00; } .csharpcode .html { color: #800000; } .csharpcode .attr { color: #ff0000; } .csharpcode .alt { background-color: #f4f4f4; width: 100%; margin: 0em; } .csharpcode .lnum { color: #606060; } Using TypeScript with ASP.NET MVC 4 Projects The current preview version of TypeScript is not providing a project template for ASP.NET MVC 4 projects. But you can use TypeScript with ASP.NET MVC 4 projects by editing the project’s .csproject file. You can take the necessary settings from ASP.NET MVC 3 project file. I have just added the following section in the end of the .csproj file of a ASP.NET MVC 4 project, which will allows to compile all TypeScript when building ASP.NET MVC 4 project. <ItemGroup> <TypeScriptCompile Include="$(ProjectDir)\**\*.ts" /> </ItemGroup> <Target Name="BeforeBuild"> <Exec Command="&amp;quot;$(PROGRAMFILES)\ Microsoft SDKs\TypeScript\0.8.0.0\tsc&amp;quot; @(TypeScriptCompile ->'&quot;%(fullpath)&quot;', ' ')" /> </Target> .csharpcode, .csharpcode pre { font-size: small; color: black; font-family: consolas, "Courier New", courier, monospace; background-color: #ffffff; /*white-space: pre;*/ } .csharpcode pre { margin: 0em; } .csharpcode .rem { color: #008000; } .csharpcode .kwrd { color: #0000ff; } .csharpcode .str { color: #006080; } .csharpcode .op { color: #0000c0; } .csharpcode .preproc { color: #cc6633; } .csharpcode .asp { background-color: #ffff00; } .csharpcode .html { color: #800000; } .csharpcode .attr { color: #ff0000; } .csharpcode .alt { background-color: #f4f4f4; width: 100%; margin: 0em; } .csharpcode .lnum { color: #606060; }

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  • Integrate SharePoint 2010 with Team Foundation Server 2010

    - by Martin Hinshelwood
    Our client is using a brand new shiny installation of SharePoint 2010, so we need to integrate our upgraded Team Foundation Server 2010 instance into it. In order to do that you need to run the Team Foundation Server 2010 install on the SharePoint 2010 server and choose to install only the “Extensions for SharePoint Products and Technologies”. We want out upgraded Team Project Collection to create any new portal in this SharePoint 2010 server farm. There a number of goodies above and beyond a solution file that requires the install, with the main one being the TFS2010 client API. These goodies allow proper integration with the creation and viewing of Work Items from SharePoint a new feature with TFS 2010. This works in both SharePoint 2007 and SharePoint 2010 with the level of integration dependant on the version of SharePoint that you are running. There are three levels of integration with “SharePoint Services 3.0” or “SharePoint Foundation 2010” being the lowest. This level only offers reporting services framed integration for reporting along with Work Item Integration and document management. The highest is Microsoft Office SharePoint Services (MOSS) Enterprise with Excel Services integration providing some lovely dashboards. Figure: Dashboards take the guessing out of Project Planning and estimation. Plus writing these reports would be boring!   The Extensions that you need are on the same installation media as the main TFS install and the only difference is the options you pick during the install. Figure: Installing the TFS 2010 Extensions for SharePoint Products and Technologies onto SharePoint 2010   Annoyingly you may need to reboot a couple of times, but on this server the process was MUCH smother than on our internal server. I think this was mostly to do with this being a clean install. Once it is installed you need to run the configuration. This will add all of the Solution and Templates that are needed for SharePoint to work properly with TFS. Figure: This is where all the TFS 2010 goodies are added to your SharePoint 2010 server and the TFS 2010 object model is installed.   Figure: All done, you have everything installed, but you still need to configure it Now that we have the TFS 2010 SharePoint Extensions installed on our SharePoint 2010 server we need to configure them both so that they will talk happily to each other. Configuring the SharePoint 2010 Managed path for Team Foundation Server 2010 In order for TFS to automatically create your project portals you need a wildcard managed path setup. This is where TFS will create the portal during the creation of a new Team project. To find the managed paths page for any application you need to first select the “Managed web applications”  link from the SharePoint 2010 Central Administration screen. Figure: Find the “Manage web applications” link under the “Application Management” section. On you are there you will see that the “Managed Paths” are there, they are just greyed out and selecting one of the applications will enable it to be clicked. Figure: You need to select an application for the SharePoint 2010 ribbon to activate.   Figure: You need to select an application before you can get to the Managed Paths for that application. Now we need to add a managed path for TFS 2010 to create its portals under. I have gone for the obvious option of just calling the managed path “TFS02” as the TFS 2010 server is the second TFS server that the client has installed, TFS 2008 being the first. This links the location to the server name, and as you can’t have two projects of the same name in two separate project collections there is unlikely to be any conflicts. Figure: Add a “tfs02” wildcard inclusion path to your SharePoint site. Configure the Team Foundation Server 2010 connection to SharePoint 2010 In order to have you new TFS 2010 Server talk to and create sites in SharePoint 2010 you need to tell the TFS server where to put them. As this TFS 2010 server was installed in out-of-the-box mode it has a SharePoint Services 3.0 (the free one) server running on the same box. But we want to change that so we can use the external SharePoint 2010 instance. Just open the “Team Foundation Server Administration Console” and navigate to the “SharePoint Web Applications” section. Here you click “Add” and enter the details for the Managed path we just created. Figure: If you have special permissions on your SharePoint you may need to add accounts to the “Service Accounts” section.    Before we can se this new SharePoint 2010 instance to be the default for our upgraded Team Project Collection we need to configure SharePoint to take instructions from our TFS server. Configure SharePoint 2010 to connect to Team Foundation Server 2010 On your SharePoint 2010 server open the Team Foundation Server Administration Console and select the “Extensions for SharePoint Products and Technologies” node. Here we need to “grant access” for our TFS 2010 server to create sites. Click the “Grant access” link and  fill out the full URL to the  TFS server, for example http://servername.domain.com:8080/tfs, and if need be restrict the path that TFS sites can be created on. Remember that when the users create a new team project they can change the default and point it anywhere they like as long as it is an authorised SharePoint location. Figure: Grant access for your TFS 2010 server to create sites in SharePoint 2010 Now that we have an authorised location for our team project portals to be created we need to tell our Team Project Collection that this is where it should stick sites by default for any new Team Projects created. Configure the Team Foundation Server 2010 Team Project Collection to create new sites in SharePoint 2010 Back on out TFS 2010 server we need to setup the defaults for our upgraded Team Project Collection to the new SharePoint 2010 integration we have just set up. On the TFS 2010 server open up the “Team Foundation Server Administration Console” again and navigate to the “Team Project Collections” node. Once you are there you will see a list of all of your TPC’s and in our case we have a DefaultCollection as well as out named and Upgraded collection for TFS 2008. If you select the “SharePoint Site” tab we can see that it is not currently configured. Figure: Our new Upgrade TFS2008 Team Project Collection does not have SharePoint configured Select to “Edit Default Site Location” and select the new integration point that we just set up for SharePoint 2010. Once you have selected the “SharePoint Web Application” (the thing we just configured) then it will give you an example based on that configuration point and the name of the Team Project Collection that we are configuring. Figure: Set the default location for new Team Project Portals to be created for this Team Project Collection This is where the reason for configuring the Extensions on the SharePoint 2010 server before doing this last bit becomes apparent. TFS 2010 is going to create a site at our http://sharepointserver/tfs02/ location called http://sharepointserver/tfs02/[TeamProjectCollection], or whatever we had specified, and it would have had difficulty doing this if we had not given it permission first. Figure: If there is no Team Project Collection site at this location the TFS 2010 server is going to create one This will create a nice Team Project Collection parent site to contain the Portals for any new Team Projects that are created. It is with noting that it will not create portals for existing Team Projects as this process is run during the Team Project Creation wizard. Figure: Just a basic parent site to host all of your new Team Project Portals as sub sites   You will need to add all of the users that will be creating Team Projects to be Administrators of this site so that they will not get an error during the Project Creation Wizard. You may also want to customise this as a proper portal to your projects if you are going to be having lots of them, but it is really just a default placeholder so you have a top level site that you can backup and point at. You have now integrated SharePoint 2010 and team Foundation Server 2010! You can now go forth and multiple your Team Projects for this Team Project Collection or you can continue to add portals to your other Collections.   Technorati Tags: TFS 2010,Sharepoint 2010,VS ALM

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  • Android - creating a custom preferences activity screen

    - by Bill Osuch
    Android applications can maintain their own internal preferences (and allow them to be modified by users) with very little coding. In fact, you don't even need to write an code to explicitly save these preferences, it's all handled automatically! Create a new Android project, with an intial activity title Main. Create two more activities: ShowPrefs, which extends Activity Set Prefs, which extends PreferenceActivity Add these two to your AndroidManifest.xml file: <activity android:name=".SetPrefs"></activity> <activity android:name=".ShowPrefs"></activity> Now we'll work on fleshing out each activity. First, open up the main.xml layout file and add a couple of buttons to it: <?xml version="1.0" encoding="utf-8"?> <LinearLayout xmlns:android="http://schemas.android.com/apk/res/android"    android:orientation="vertical"    android:layout_width="fill_parent"    android:layout_height="fill_parent"> <Button android:text="Edit Preferences"    android:id="@+id/prefButton"    android:layout_width="wrap_content"    android:layout_height="wrap_content"    android:layout_gravity="center_horizontal"/> <Button android:text="Show Preferences"    android:id="@+id/showButton"    android:layout_width="wrap_content"    android:layout_height="wrap_content"    android:layout_gravity="center_horizontal"/> </LinearLayout> Next, create a couple button listeners in Main.java to handle the clicks and start the other activities: Button editPrefs = (Button) findViewById(R.id.prefButton);       editPrefs.setOnClickListener(new View.OnClickListener() {              public void onClick(View view) {                  Intent myIntent = new Intent(view.getContext(), SetPrefs.class);                  startActivityForResult(myIntent, 0);              }      });           Button showPrefs = (Button) findViewById(R.id.showButton);      showPrefs.setOnClickListener(new View.OnClickListener() {              public void onClick(View view) {                  Intent myIntent = new Intent(view.getContext(), ShowPrefs.class);                  startActivityForResult(myIntent, 0);              }      }); Now, we'll create the actual preferences layout. You'll need to create a file called preferences.xml inside res/xml, and you'll likely have to create the xml directory as well. Add the following xml: <?xml version="1.0" encoding="utf-8"?> <PreferenceScreen xmlns:android="http://schemas.android.com/apk/res/android"> </PreferenceScreen> First we'll add a category, which is just a way to group similar preferences... sort of a horizontal bar. Add this inside the PreferenceScreen tags: <PreferenceCategory android:title="First Category"> </PreferenceCategory> Now add a Checkbox and an Edittext box (inside the PreferenceCategory tags): <CheckBoxPreference    android:key="checkboxPref"    android:title="Checkbox Preference"    android:summary="This preference can be true or false"    android:defaultValue="false"/> <EditTextPreference    android:key="editTextPref"    android:title="EditText Preference"    android:summary="This allows you to enter a string"    android:defaultValue="Nothing"/> The key is how you will refer to the preference in code, the title is the large text that will be displayed, and the summary is the smaller text (this will make sense when you see it). Let's say we've got a second group of preferences that apply to a different part of the app. Add a new category just below the first one: <PreferenceCategory android:title="Second Category"> </PreferenceCategory> In there we'll a list with radio buttons, so add: <ListPreference    android:key="listPref"    android:title="List Preference"    android:summary="This preference lets you select an item in a array"    android:entries="@array/listArray"    android:entryValues="@array/listValues" /> When complete, your full xml file should look like this: <?xml version="1.0" encoding="utf-8"?> <PreferenceScreen xmlns:android="http://schemas.android.com/apk/res/android">  <PreferenceCategory android:title="First Category"> <CheckBoxPreference    android:key="checkboxPref"    android:title="Checkbox Preference"    android:summary="This preference can be true or false"    android:defaultValue="false"/> <EditTextPreference    android:key="editTextPref"    android:title="EditText Preference"    android:summary="This allows you to enter a string"    android:defaultValue="Nothing"/>  </PreferenceCategory>  <PreferenceCategory android:title="Second Category">   <ListPreference    android:key="listPref"    android:title="List Preference"    android:summary="This preference lets you select an item in a array"    android:entries="@array/listArray"    android:entryValues="@array/listValues" />  </PreferenceCategory> </PreferenceScreen> However, when you try to save it, you'll get an error because you're missing your array definition. To fix this, add a file called arrays.xml in res/values, and paste in the following: <?xml version="1.0" encoding="utf-8"?> <resources>  <string-array name="listArray">      <item>Value 1</item>      <item>Value 2</item>      <item>Value 3</item>  </string-array>  <string-array name="listValues">      <item>1</item>      <item>2</item>      <item>3</item>  </string-array> </resources> Finally (for the preferences screen at least...) add the code that will display the preferences layout to the SetPrefs.java file:  @Override     public void onCreate(Bundle savedInstanceState) {      super.onCreate(savedInstanceState);      addPreferencesFromResource(R.xml.preferences);      } OK, so now we've got an activity that will set preferences, and save them without the need to write custom save code. Let's throw together an activity to work with the saved preferences. Create a new layout called showpreferences.xml and give it three Textviews: <?xml version="1.0" encoding="utf-8"?> <LinearLayout xmlns:android="http://schemas.android.com/apk/res/android"     android:orientation="vertical"     android:layout_width="fill_parent"     android:layout_height="fill_parent"> <TextView   android:id="@+id/textview1"     android:layout_width="fill_parent"     android:layout_height="wrap_content"     android:text="textview1"/> <TextView   android:id="@+id/textview2"     android:layout_width="fill_parent"     android:layout_height="wrap_content"     android:text="textview2"/> <TextView   android:id="@+id/textview3"     android:layout_width="fill_parent"     android:layout_height="wrap_content"     android:text="textview3"/> </LinearLayout> Open up the ShowPrefs.java file and have it use that layout: setContentView(R.layout.showpreferences); Then add the following code to load the DefaultSharedPreferences and display them: SharedPreferences prefs = PreferenceManager.getDefaultSharedPreferences(this);    TextView text1 = (TextView)findViewById(R.id.textview1); TextView text2 = (TextView)findViewById(R.id.textview2); TextView text3 = (TextView)findViewById(R.id.textview3);    text1.setText(new Boolean(prefs.getBoolean("checkboxPref", false)).toString()); text2.setText(prefs.getString("editTextPref", "<unset>"));; text3.setText(prefs.getString("listPref", "<unset>")); Fire up the application in the emulator and click the Edit Preferences button. Set various things, click the back button, then the Edit Preferences button again. Notice that your choices have been saved.   Now click the Show Preferences button, and you should see the results of what you set:   There are two more preference types that I did not include here: RingtonePreference - shows a radioGroup that lists your ringtones PreferenceScreen - allows you to embed a second preference screen inside the first - it opens up a new set of preferences when clicked

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  • Ajax Control Toolkit Now Supports jQuery

    - by Stephen.Walther
    I’m excited to announce the September 2013 release of the Ajax Control Toolkit, which now supports building new Ajax Control Toolkit controls with jQuery. You can download the latest release of the Ajax Control Toolkit from http://AjaxControlToolkit.CodePlex.com or you can install the Ajax Control Toolkit directly within Visual Studio by executing the following NuGet command: The New jQuery Extender Base Class This release of the Ajax Control Toolkit introduces a new jQueryExtender base class. This new base class enables you to create Ajax Control Toolkit controls with jQuery instead of the Microsoft Ajax Library. Currently, only one control in the Ajax Control Toolkit has been rewritten to use the new jQueryExtender base class (only one control has been jQueryized). The ToggleButton control is the first of the Ajax Control Toolkit controls to undergo this dramatic transformation. All of the other controls in the Ajax Control Toolkit are written using the Microsoft Ajax Library. We hope to gradually rewrite these controls as jQuery controls over time. You can view the new jQuery ToggleButton live at the Ajax Control Toolkit sample site: http://www.asp.net/ajaxLibrary/AjaxControlToolkitSampleSite/ToggleButton/ToggleButton.aspx Why are we rewriting Ajax Control Toolkits with jQuery? There are very few developers actively working with the Microsoft Ajax Library while there are thousands of developers actively working with jQuery. Because we want talented developers in the community to continue to contribute to the Ajax Control Toolkit, and because almost all JavaScript developers are familiar with jQuery, it makes sense to support jQuery with the Ajax Control Toolkit. Also, we believe that the Ajax Control Toolkit is a great framework for Web Forms developers who want to build new ASP.NET controls that use JavaScript. The Ajax Control Toolkit has great features such as automatic bundling, minification, caching, and compression. We want to make it easy for ASP.NET developers to build new controls that take advantage of these features. Instantiating Controls with data-* Attributes We took advantage of the new JQueryExtender base class to change the way that Ajax Control Toolkit controls are instantiated. In the past, adding an Ajax Control Toolkit to a page resulted in inline JavaScript being injected into the page. For example, adding the ToggleButton control to a page injected the following HTML and script: <input id="ctl00_SampleContent_CheckBox1" name="ctl00$SampleContent$CheckBox1" type="checkbox" checked="checked" /> <script type="text/javascript"> //<![CDATA[ Sys.Application.add_init(function() { $create(Sys.Extended.UI.ToggleButtonBehavior, {"CheckedImageAlternateText":"Check", "CheckedImageUrl":"ToggleButton_Checked.gif", "ImageHeight":19, "ImageWidth":19, "UncheckedImageAlternateText":"UnCheck", "UncheckedImageUrl":"ToggleButton_Unchecked.gif", "id":"ctl00_SampleContent_ToggleButtonExtender1"}, null, null, $get("ctl00_SampleContent_CheckBox1")); }); //]]> </script> Notice the call to the JavaScript $create() method at the bottom of the page. When using the Microsoft Ajax Library, this call to the $create() method is necessary to create the Ajax Control Toolkit control. This inline script looks pretty ugly to a modern JavaScript developer. Inline script! Horrible! The jQuery version of the ToggleButton injects the following HTML and script into the page: <input id="ctl00_SampleContent_CheckBox1" name="ctl00$SampleContent$CheckBox1" type="checkbox" checked="checked" data-act-togglebuttonextender="imageWidth:19, imageHeight:19, uncheckedImageUrl:'ToggleButton_Unchecked.gif', checkedImageUrl:'ToggleButton_Checked.gif', uncheckedImageAlternateText:'I don&#39;t understand why you don&#39;t like ASP.NET', checkedImageAlternateText:'It&#39;s really nice to hear from you that you like ASP.NET'" /> Notice that there is no script! There is no call to the $create() method. In fact, there is no inline JavaScript at all. The jQuery version of the ToggleButton uses an HTML5 data-* attribute instead of an inline script. The ToggleButton control is instantiated with a data-act-togglebuttonextender attribute. Using data-* attributes results in much cleaner markup (You don’t need to feel embarrassed when selecting View Source in your browser). Ajax Control Toolkit versus jQuery So in a jQuery world why is the Ajax Control Toolkit needed at all? Why not just use jQuery plugins instead of the Ajax Control Toolkit? For example, there are lots of jQuery ToggleButton plugins floating around the Internet. Why not just use one of these jQuery plugins instead of using the Ajax Control Toolkit ToggleButton control? There are three main reasons why the Ajax Control Toolkit continues to be valuable in a jQuery world: Ajax Control Toolkit controls run on both the server and client jQuery plugins are client only. A jQuery plugin does not include any server-side code. If you need to perform any work on the server – think of the AjaxFileUpload control – then you can’t use a pure jQuery solution. Ajax Control Toolkit controls provide a better Visual Studio experience You don’t get any design time experience when you use jQuery plugins within Visual Studio. Ajax Control Toolkit controls, on the other hand, are designed to work with Visual Studio. For example, you can use the Visual Studio Properties window to set Ajax Control Toolkit control properties. Ajax Control Toolkit controls shield you from working with JavaScript I like writing code in JavaScript. However, not all developers like JavaScript and some developers want to completely avoid writing any JavaScript code at all. The Ajax Control Toolkit enables you to take advantage of JavaScript (and the latest features of HTML5) in your ASP.NET Web Forms websites without writing a single line of JavaScript. Better ToolkitScriptManager Documentation With this release, we have added more detailed documentation for using the ToolkitScriptManager. In particular, we added documentation that describes how to take advantage of the new bundling, minification, compression, and caching features of the Ajax Control Toolkit. The ToolkitScriptManager documentation is part of the Ajax Control Toolkit sample site and it can be read here: http://www.asp.net/ajaxLibrary/AjaxControlToolkitSampleSite/ToolkitScriptManager/ToolkitScriptManager.aspx Other Fixes This release of the Ajax Control Toolkit includes several important bug fixes. For example, the Ajax Control Toolkit Twitter control was completely rewritten with this release. Twitter is in the process of retiring the first version of their API. You can read about their plans here: https://dev.twitter.com/blog/planning-for-api-v1-retirement We completely rewrote the Ajax Control Toolkit Twitter control to use the new Twitter API. To take advantage of the new Twitter API, you must get a key and access token from Twitter and add the key and token to your web.config file. Detailed instructions for using the new version of the Ajax Control Toolkit Twitter control can be found here: http://www.asp.net/ajaxLibrary/AjaxControlToolkitSampleSite/Twitter/Twitter.aspx   Summary We’ve made some really great changes to the Ajax Control Toolkit over the last two releases to modernize the toolkit. In the previous release, we updated the Ajax Control Toolkit to use a better bundling, minification, compression, and caching system. With this release, we updated the Ajax Control Toolkit to support jQuery. We also continue to update the Ajax Control Toolkit with important bug fixes. I hope you like these changes and I look forward to hearing your feedback.

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  • LINQ to SQL and missing Many to Many EntityRefs

    - by Rick Strahl
    Ran into an odd behavior today with a many to many mapping of one of my tables in LINQ to SQL. Many to many mappings aren’t transparent in LINQ to SQL and it maps the link table the same way the SQL schema has it when creating one. In other words LINQ to SQL isn’t smart about many to many mappings and just treats it like the 3 underlying tables that make up the many to many relationship. Iain Galloway has a nice blog entry about Many to Many relationships in LINQ to SQL. I can live with that – it’s not really difficult to deal with this arrangement once mapped, especially when reading data back. Writing is a little more difficult as you do have to insert into two entities for new records, but nothing that can’t be handled in a small business object method with a few lines of code. When I created a database I’ve been using to experiment around with various different OR/Ms recently I found that for some reason LINQ to SQL was completely failing to map even to the linking table. As it turns out there’s a good reason why it fails, can you spot it below? (read on :-}) Here is the original database layout: There’s an items table, a category table and a link table that holds only the foreign keys to the Items and Category tables for a typical M->M relationship. When these three tables are imported into the model the *look* correct – I do get the relationships added (after modifying the entity names to strip the prefix): The relationship looks perfectly fine, both in the designer as well as in the XML document: <Table Name="dbo.wws_Item_Categories" Member="ItemCategories"> <Type Name="ItemCategory"> <Column Name="ItemId" Type="System.Guid" DbType="uniqueidentifier NOT NULL" CanBeNull="false" /> <Column Name="CategoryId" Type="System.Guid" DbType="uniqueidentifier NOT NULL" CanBeNull="false" /> <Association Name="ItemCategory_Category" Member="Categories" ThisKey="CategoryId" OtherKey="Id" Type="Category" /> <Association Name="Item_ItemCategory" Member="Item" ThisKey="ItemId" OtherKey="Id" Type="Item" IsForeignKey="true" /> </Type> </Table> <Table Name="dbo.wws_Categories" Member="Categories"> <Type Name="Category"> <Column Name="Id" Type="System.Guid" DbType="UniqueIdentifier NOT NULL" IsPrimaryKey="true" IsDbGenerated="true" CanBeNull="false" /> <Column Name="ParentId" Type="System.Guid" DbType="UniqueIdentifier" CanBeNull="true" /> <Column Name="CategoryName" Type="System.String" DbType="NVarChar(150)" CanBeNull="true" /> <Column Name="CategoryDescription" Type="System.String" DbType="NVarChar(MAX)" CanBeNull="true" /> <Column Name="tstamp" AccessModifier="Internal" Type="System.Data.Linq.Binary" DbType="rowversion" CanBeNull="true" IsVersion="true" /> <Association Name="ItemCategory_Category" Member="ItemCategory" ThisKey="Id" OtherKey="CategoryId" Type="ItemCategory" IsForeignKey="true" /> </Type> </Table> However when looking at the code generated these navigation properties (also on Item) are completely missing: [global::System.Data.Linq.Mapping.TableAttribute(Name="dbo.wws_Item_Categories")] [global::System.Runtime.Serialization.DataContractAttribute()] public partial class ItemCategory : Westwind.BusinessFramework.EntityBase { private System.Guid _ItemId; private System.Guid _CategoryId; public ItemCategory() { } [global::System.Data.Linq.Mapping.ColumnAttribute(Storage="_ItemId", DbType="uniqueidentifier NOT NULL")] [global::System.Runtime.Serialization.DataMemberAttribute(Order=1)] public System.Guid ItemId { get { return this._ItemId; } set { if ((this._ItemId != value)) { this._ItemId = value; } } } [global::System.Data.Linq.Mapping.ColumnAttribute(Storage="_CategoryId", DbType="uniqueidentifier NOT NULL")] [global::System.Runtime.Serialization.DataMemberAttribute(Order=2)] public System.Guid CategoryId { get { return this._CategoryId; } set { if ((this._CategoryId != value)) { this._CategoryId = value; } } } } Notice that the Item and Category association properties which should be EntityRef properties are completely missing. They’re there in the model, but the generated code – not so much. So what’s the problem here? The problem – it appears – is that LINQ to SQL requires primary keys on all entities it tracks. In order to support tracking – even of the link table entity – the link table requires a primary key. Real obvious ain’t it, especially since the designer happily lets you import the table and even shows the relationship and implicitly the related properties. Adding an Id field as a Pk to the database and then importing results in this model layout: which properly generates the Item and Category properties into the link entity. It’s ironic that LINQ to SQL *requires* the PK in the middle – the Entity Framework requires that a link table have *only* the two foreign key fields in a table in order to recognize a many to many relation. EF actually handles the M->M relation directly without the intermediate link entity unlike LINQ to SQL. [updated from comments – 12/24/2009] Another approach is to set up both ItemId and CategoryId in the database which shows up in LINQ to SQL like this: This also work in creating the Category and Item fields in the ItemCategory entity. Ultimately this is probably the best approach as it also guarantees uniqueness of the keys and so helps in database integrity. It took me a while to figure out WTF was going on here – lulled by the designer to think that the properties should be when they were not. It’s actually a well documented feature of L2S that each entity in the model requires a Pk but of course that’s easy to miss when the model viewer shows it to you and even the underlying XML model shows the Associations properly. This is one of the issue with L2S of course – you have to play by its rules and once you hit one of those rules there’s no way around them – you’re stuck with what it requires which in this case meant changing the database.© Rick Strahl, West Wind Technologies, 2005-2010Posted in ADO.NET  LINQ  

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  • Ajax Control Toolkit and Superexpert

    - by Stephen Walther
    Microsoft has asked my company, Superexpert Consulting, to take ownership of the development and maintenance of the Ajax Control Toolkit moving forward. In this blog entry, I discuss our strategy for improving the Ajax Control Toolkit. Why the Ajax Control Toolkit? The Ajax Control Toolkit is one of the most popular projects on CodePlex. In fact, some have argued that it is among the most successful open-source projects of all time. It consistently receives over 3,500 downloads a day (not weekends -- workdays). A mind-boggling number of developers use the Ajax Control Toolkit in their ASP.NET Web Forms applications. Why does the Ajax Control Toolkit continue to be such a popular project? The Ajax Control Toolkit fills a strong need in the ASP.NET Web Forms world. The Toolkit enables Web Forms developers to build richly interactive JavaScript applications without writing any JavaScript. For example, by taking advantage of the Ajax Control Toolkit, a Web Forms developer can add modal dialogs, popup calendars, and client tabs to a web application simply by dragging web controls onto a page. The Ajax Control Toolkit is not for everyone. If you are comfortable writing JavaScript then I recommend that you investigate using jQuery plugins instead of the Ajax Control Toolkit. However, if you are a Web Forms developer and you don’t want to get your hands dirty writing JavaScript, then the Ajax Control Toolkit is a great solution. The Ajax Control Toolkit is Vast The Ajax Control Toolkit consists of 40 controls. That’s a lot of controls (For the sake of comparison, jQuery UI consists of only 8 controls – those slackers J). Furthermore, developers expect the Ajax Control Toolkit to work on browsers both old and new. For example, people expect the Ajax Control Toolkit to work with Internet Explorer 6 and Internet Explorer 9 and every version of Internet Explorer in between. People also expect the Ajax Control Toolkit to work on the latest versions of Mozilla Firefox, Apple Safari, and Google Chrome. And, people expect the Ajax Control Toolkit to work with different operating systems. Yikes, that is a lot of combinations. The biggest challenge which my company faces in supporting the Ajax Control Toolkit is ensuring that the Ajax Control Toolkit works across all of these different browsers and operating systems. Testing, Testing, Testing Because we wanted to ensure that we could easily test the Ajax Control Toolkit with different browsers, the very first thing that we did was to set up a dedicated testing server. The dedicated server -- named Schizo -- hosts 4 virtual machines so that we can run Internet Explorer 6, Internet Explorer 7, Internet Explorer 8, and Internet Explorer 9 at the same time (We also use the virtual machines to host the latest versions of Firefox, Chrome, Opera, and Safari). The five developers on our team (plus me) can each publish to a separate FTP website on the testing server. That way, we can quickly test how changes to the Ajax Control Toolkit affect different browsers. QUnit Tests for the Ajax Control Toolkit Introducing regressions – introducing new bugs when trying to fix existing bugs – is the concern which prevents me from sleeping well at night. There are so many people using the Ajax Control Toolkit in so many unique scenarios, that it is difficult to make improvements to the Ajax Control Toolkit without introducing regressions. In order to avoid regressions, we decided early on that it was extremely important to build good test coverage for the 40 controls in the Ajax Control Toolkit. We’ve been focusing a lot of energy on building automated JavaScript unit tests which we can use to help us discover regressions. We decided to write the unit tests with the QUnit test framework. We picked QUnit because it is quickly becoming the standard unit testing framework in the JavaScript world. For example, it is the unit testing framework used by the jQuery team, the jQuery UI team, and many jQuery UI plugin developers. We had to make several enhancements to the QUnit framework in order to test the Ajax Control Toolkit. For example, QUnit does not support tests which include postbacks. We modified the QUnit framework so that it works with IFrames so we could perform postbacks in our automated tests. At this point, we have written hundreds of QUnit tests. For example, we have written 135 QUnit tests for the Accordion control. The QUnit tests are included with the Ajax Control Toolkit source code in a project named AjaxControlToolkit.Tests. You can run all of the QUnit tests contained in the project by opening the Default.aspx page. Automating the QUnit Tests across Multiple Browsers Automated tests are useless if no one ever runs them. In order for the QUnit tests to be useful, we needed an easy way to run the tests automatically against a matrix of browsers. We wanted to run the unit tests against Internet Explorer 6, Internet Explorer 7, Internet Explorer 8, Internet Explorer 9, Firefox, Chrome, and Safari automatically. Expecting a developer to run QUnit tests against every browser after every check-in is just too much to expect. It takes 20 seconds to run the Accordion QUnit tests. We are testing against 8 browsers. That would require the developer to open 8 browsers and wait for the results after each change in code. Too much work. Therefore, we built a JavaScript Test Server. Our JavaScript Test Server project was inspired by John Resig’s TestSwarm project. The JavaScript Test Server runs our QUnit tests in a swarm of browsers (running on different operating systems) automatically. Here’s how the JavaScript Test Server works: 1. We created an ASP.NET page named RunTest.aspx that constantly polls the JavaScript Test Server for a new set of QUnit tests to run. After the RunTest.aspx page runs the QUnit tests, the RunTest.aspx records the test results back to the JavaScript Test Server. 2. We opened the RunTest.aspx page on instances of Internet Explorer 6, Internet Explorer 7, Internet Explorer 8, Internet Explorer 9, FireFox, Chrome, Opera, Google, and Safari. Now that we have the JavaScript Test Server setup, we can run all of our QUnit tests against all of the browsers which we need to support with a single click of a button. A New Release of the Ajax Control Toolkit Each Month The Ajax Control Toolkit Issue Tracker contains over one thousand five hundred open issues and feature requests. So we have plenty of work on our plates J At CodePlex, anyone can vote for an issue to be fixed. Originally, we planned to fix issues in order of their votes. However, we quickly discovered that this approach was inefficient. Constantly switching back and forth between different controls was too time-consuming. It takes time to re-familiarize yourself with a control. Instead, we decided to focus on two or three controls each month and really focus on fixing the issues with those controls. This way, we can fix sets of related issues and avoid the randomization caused by context switching. Our team works in monthly sprints. We plan to do another release of the Ajax Control Toolkit each and every month. So far, we have competed one release of the Ajax Control Toolkit which was released on April 1, 2011. We plan to release a new version in early May. Conclusion Fortunately, I work with a team of smart developers. We currently have 5 developers working on the Ajax Control Toolkit (not full-time, they are also building two very cool ASP.NET MVC applications). All the developers who work on our team are required to have strong JavaScript, jQuery, and ASP.NET MVC skills. In the interest of being as transparent as possible about our work on the Ajax Control Toolkit, I plan to blog frequently about our team’s ongoing work. In my next blog entry, I plan to write about the two Ajax Control Toolkit controls which are the focus of our work for next release.

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  • C# 4: The Curious ConcurrentDictionary

    - by James Michael Hare
    In my previous post (here) I did a comparison of the new ConcurrentQueue versus the old standard of a System.Collections.Generic Queue with simple locking.  The results were exactly what I would have hoped, that the ConcurrentQueue was faster with multi-threading for most all situations.  In addition, concurrent collections have the added benefit that you can enumerate them even if they're being modified. So I set out to see what the improvements would be for the ConcurrentDictionary, would it have the same performance benefits as the ConcurrentQueue did?  Well, after running some tests and multiple tweaks and tunes, I have good and bad news. But first, let's look at the tests.  Obviously there's many things we can do with a dictionary.  One of the most notable uses, of course, in a multi-threaded environment is for a small, local in-memory cache.  So I set about to do a very simple simulation of a cache where I would create a test class that I'll just call an Accessor.  This accessor will attempt to look up a key in the dictionary, and if the key exists, it stops (i.e. a cache "hit").  However, if the lookup fails, it will then try to add the key and value to the dictionary (i.e. a cache "miss").  So here's the Accessor that will run the tests: 1: internal class Accessor 2: { 3: public int Hits { get; set; } 4: public int Misses { get; set; } 5: public Func<int, string> GetDelegate { get; set; } 6: public Action<int, string> AddDelegate { get; set; } 7: public int Iterations { get; set; } 8: public int MaxRange { get; set; } 9: public int Seed { get; set; } 10:  11: public void Access() 12: { 13: var randomGenerator = new Random(Seed); 14:  15: for (int i=0; i<Iterations; i++) 16: { 17: // give a wide spread so will have some duplicates and some unique 18: var target = randomGenerator.Next(1, MaxRange); 19:  20: // attempt to grab the item from the cache 21: var result = GetDelegate(target); 22:  23: // if the item doesn't exist, add it 24: if(result == null) 25: { 26: AddDelegate(target, target.ToString()); 27: Misses++; 28: } 29: else 30: { 31: Hits++; 32: } 33: } 34: } 35: } Note that so I could test different implementations, I defined a GetDelegate and AddDelegate that will call the appropriate dictionary methods to add or retrieve items in the cache using various techniques. So let's examine the three techniques I decided to test: Dictionary with mutex - Just your standard generic Dictionary with a simple lock construct on an internal object. Dictionary with ReaderWriterLockSlim - Same Dictionary, but now using a lock designed to let multiple readers access simultaneously and then locked when a writer needs access. ConcurrentDictionary - The new ConcurrentDictionary from System.Collections.Concurrent that is supposed to be optimized to allow multiple threads to access safely. So the approach to each of these is also fairly straight-forward.  Let's look at the GetDelegate and AddDelegate implementations for the Dictionary with mutex lock: 1: var addDelegate = (key,val) => 2: { 3: lock (_mutex) 4: { 5: _dictionary[key] = val; 6: } 7: }; 8: var getDelegate = (key) => 9: { 10: lock (_mutex) 11: { 12: string val; 13: return _dictionary.TryGetValue(key, out val) ? val : null; 14: } 15: }; Nothing new or fancy here, just your basic lock on a private object and then query/insert into the Dictionary. Now, for the Dictionary with ReadWriteLockSlim it's a little more complex: 1: var addDelegate = (key,val) => 2: { 3: _readerWriterLock.EnterWriteLock(); 4: _dictionary[key] = val; 5: _readerWriterLock.ExitWriteLock(); 6: }; 7: var getDelegate = (key) => 8: { 9: string val; 10: _readerWriterLock.EnterReadLock(); 11: if(!_dictionary.TryGetValue(key, out val)) 12: { 13: val = null; 14: } 15: _readerWriterLock.ExitReadLock(); 16: return val; 17: }; And finally, the ConcurrentDictionary, which since it does all it's own concurrency control, is remarkably elegant and simple: 1: var addDelegate = (key,val) => 2: { 3: _concurrentDictionary[key] = val; 4: }; 5: var getDelegate = (key) => 6: { 7: string s; 8: return _concurrentDictionary.TryGetValue(key, out s) ? s : null; 9: };                    Then, I set up a test harness that would simply ask the user for the number of concurrent Accessors to attempt to Access the cache (as specified in Accessor.Access() above) and then let them fly and see how long it took them all to complete.  Each of these tests was run with 10,000,000 cache accesses divided among the available Accessor instances.  All times are in milliseconds. 1: Dictionary with Mutex Locking 2: --------------------------------------------------- 3: Accessors Mostly Misses Mostly Hits 4: 1 7916 3285 5: 10 8293 3481 6: 100 8799 3532 7: 1000 8815 3584 8:  9:  10: Dictionary with ReaderWriterLockSlim Locking 11: --------------------------------------------------- 12: Accessors Mostly Misses Mostly Hits 13: 1 8445 3624 14: 10 11002 4119 15: 100 11076 3992 16: 1000 14794 4861 17:  18:  19: Concurrent Dictionary 20: --------------------------------------------------- 21: Accessors Mostly Misses Mostly Hits 22: 1 17443 3726 23: 10 14181 1897 24: 100 15141 1994 25: 1000 17209 2128 The first test I did across the board is the Mostly Misses category.  The mostly misses (more adds because data requested was not in the dictionary) shows an interesting trend.  In both cases the Dictionary with the simple mutex lock is much faster, and the ConcurrentDictionary is the slowest solution.  But this got me thinking, and a little research seemed to confirm it, maybe the ConcurrentDictionary is more optimized to concurrent "gets" than "adds".  So since the ratio of misses to hits were 2 to 1, I decided to reverse that and see the results. So I tweaked the data so that the number of keys were much smaller than the number of iterations to give me about a 2 to 1 ration of hits to misses (twice as likely to already find the item in the cache than to need to add it).  And yes, indeed here we see that the ConcurrentDictionary is indeed faster than the standard Dictionary here.  I have a strong feeling that as the ration of hits-to-misses gets higher and higher these number gets even better as well.  This makes sense since the ConcurrentDictionary is read-optimized. Also note that I tried the tests with capacity and concurrency hints on the ConcurrentDictionary but saw very little improvement, I think this is largely because on the 10,000,000 hit test it quickly ramped up to the correct capacity and concurrency and thus the impact was limited to the first few milliseconds of the run. So what does this tell us?  Well, as in all things, ConcurrentDictionary is not a panacea.  It won't solve all your woes and it shouldn't be the only Dictionary you ever use.  So when should we use each? Use System.Collections.Generic.Dictionary when: You need a single-threaded Dictionary (no locking needed). You need a multi-threaded Dictionary that is loaded only once at creation and never modified (no locking needed). You need a multi-threaded Dictionary to store items where writes are far more prevalent than reads (locking needed). And use System.Collections.Concurrent.ConcurrentDictionary when: You need a multi-threaded Dictionary where the writes are far more prevalent than reads. You need to be able to iterate over the collection without locking it even if its being modified. Both Dictionaries have their strong suits, I have a feeling this is just one where you need to know from design what you hope to use it for and make your decision based on that criteria.

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  • Why can't I reinstall MySQL?

    - by Johannes Nielsen
    I've been looking all around the Internet for an answer but didn't find anything. I hope you can help me now. I have a server with MySQL. From one day to another, MySQL didn't let me enter with my root password anymore (accsess denied for user 'root'@'localhost' using password: 'YES'). So I tried two ways to reset the password: No.1: I typed: shell> /etc/init.d/mysqld stop To stop MySQL. Then I restarted it skipping the grant-tables: shell> mysqld_safe --skip-grant-tables So I was able to log in as root and change the password using: mysql> UPDATE mysql.user SET Password = PASSWORD('MyNewPassword') WHERE User = 'root'; FLUSH PRIVILEGES; I restarted MySQL and tried to log in as root with my new password - didn't work. So I tried the solution that's described here: http://dev.mysql.com/doc/refman/5.0/en/resetting-permissions.html (I don't want to post it here because this post is already pretty long). Didn't work either. Actually it made it worse, because since that day, every time I try to start MySQL, it doesn't even ask me for my password, but I get: shell> ERROR 2002 (HY000): Can't connect to local MySQL server through socket '/var/run/mysqld/mysqld.sock' (111) Well, I've looked up what it means and found that my mysqld.sock is missing. I tried to create it using touch but MySQL can't start with that socket. Now I'm trying to reinstall MySQL but everytime I type in shell> apt-get --purge remove mysql-server mysql-common mysql-client In that or any other order or every one of those three alone, I get: shell> Reading package lists... Done shell> Building dependency tree shell> Reading state information... Done shell> Package mysql-client is not installed, so not removed shell> Package mysql-server is not installed, so not removed shell> You might want to run 'apt-get -f install' to correct these: shell> The following packages have unmet dependencies: shell> libmysqlclient18 : Depends: mysql-common (>= 5.5.28-0ubuntu0.12.04.2) but it is not going to be installed shell> libmysqlclient18:i386 : Depends: mysql-common:i386 (>= 5.5.28-0ubuntu0.12.04.2) shell> mysql-client-5.5 : Depends: mysql-common (>= 5.5.28-0ubuntu0.12.04.2) but it is not going to be installed shell> mysql-server-5.5 : PreDepends: mysql-common (>= 5.5.28-0ubuntu0.12.04.2) but it is not going to be installed shell> psa-firewall : Depends: plesk-core (>= 11.0.9) but it is not installable shell> Depends: mysql-server but it is not going to be installed shell> psa-spamassassin : Depends: plesk-core (>= 11.0.9) but it is not installable shell> psa-vpn : Depends: plesk-core (>= 11.0.9) but it is not installable shell> Depends: plesk-base (>= 11.0.9) but it is not installable shell> Depends: mysql-server but it is not going to be installed shell> E: Unmet dependencies. Try 'apt-get -f install' with no packages (or specify a solution). So I said to my self "let's just remove those files with depenencies, too" (that psa-stuff since plesk is virtual and can't be uninstalled)... Guess what happened: shell> Reading package lists... Done shell> Building dependency tree shell> Reading state information... Done shell> Package mysql-client is not installed, so not removed shell> Package mysql-server is not installed, so not removed shell> You might want to run 'apt-get -f install' to correct these: shell> The following packages have unmet dependencies: shell> libmysqlclient18 : Depends: mysql-common (>= 5.5.28-0ubuntu0.12.04.2) but it is not going to be installed shell> libmysqlclient18:i386 : Depends: mysql-common:i386 (>= 5.5.28-0ubuntu0.12.04.2) shell> mysql-client-5.5 : Depends: mysql-common (>= 5.5.28-0ubuntu0.12.04.2) but it is not going to be installed shell> mysql-server-5.5 : PreDepends: mysql-common (>= 5.5.28-0ubuntu0.12.04.2) but it is not going to be installed shell> E: Unmet dependencies. Try 'apt-get -f install' with no packages (or specify a solution). Of course I tried apt-get -f install, too many times even. What am I doing wrong? No matter, which other packages I include into apt-get --purge remove, I always get new dependencies. Do I have to delete every MySQL-related directory and file manually? Hope there's someone out there who can help me! Cheers! EDIT: After trying apt-get purge mysql-server mysql-common mysql-client libmysqlclient18 libmysqlclient18:i386 mysql-client-5.5 mysql-server-5.5 psa-firewall psa-spamassassin psa-vpn Reading package lists... Done Building dependency tree Reading state information... Done Package mysql-client is not installed, so not removed Package mysql-server is not installed, so not removed You might want to run 'apt-get -f install' to correct these: The following packages have unmet dependencies: libdbd-mysql-perl : Depends: libmysqlclient18 (>= 5.5.13-1) but it is not going to be installed libmyodbc : Depends: libmysqlclient18 (>= 5.5.13-1) but it is not going to be installed libqt4-sql-mysql:i386 : Depends: libmysqlclient18:i386 (>= 5.5.13-1) but it is not going to be installed php5-mysql : Depends: libmysqlclient18 (>= 5.5.13-1) but it is not going to be installed ruby-mysql : Depends: libmysqlclient18 (>= 5.5.13-1) but it is not going to be installed E: Unmet dependencies. Try 'apt-get -f install' with no packages (or specify a solution). So I tried to remove all these and got: Building dependency tree Reading state information... Done Package mysql-client is not installed, so not removed Package mysql-server is not installed, so not removed You might want to run 'apt-get -f install' to correct these:qlclient18:i386 mysql The following packages have unmet dependencies: libmysql-ruby1.8 : Depends: ruby-mysql but it is not going to be installed E: Unmet dependencies. Try 'apt-get -f install' with no packages (or specify a solution). And actually I think removing that file, too solved my problem :-S Next time I'll try everything before asking :D Thank you Eric for keeping me couraged to just go on removing :D

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  • Basics of Join Predicate Pushdown in Oracle

    - by Maria Colgan
    Happy New Year to all of our readers! We hope you all had a great holiday season. We start the new year by continuing our series on Optimizer transformations. This time it is the turn of Predicate Pushdown. I would like to thank Rafi Ahmed for the content of this blog.Normally, a view cannot be joined with an index-based nested loop (i.e., index access) join, since a view, in contrast with a base table, does not have an index defined on it. A view can only be joined with other tables using three methods: hash, nested loop, and sort-merge joins. Introduction The join predicate pushdown (JPPD) transformation allows a view to be joined with index-based nested-loop join method, which may provide a more optimal alternative. In the join predicate pushdown transformation, the view remains a separate query block, but it contains the join predicate, which is pushed down from its containing query block into the view. The view thus becomes correlated and must be evaluated for each row of the outer query block. These pushed-down join predicates, once inside the view, open up new index access paths on the base tables inside the view; this allows the view to be joined with index-based nested-loop join method, thereby enabling the optimizer to select an efficient execution plan. The join predicate pushdown transformation is not always optimal. The join predicate pushed-down view becomes correlated and it must be evaluated for each outer row; if there is a large number of outer rows, the cost of evaluating the view multiple times may make the nested-loop join suboptimal, and therefore joining the view with hash or sort-merge join method may be more efficient. The decision whether to push down join predicates into a view is determined by evaluating the costs of the outer query with and without the join predicate pushdown transformation under Oracle's cost-based query transformation framework. The join predicate pushdown transformation applies to both non-mergeable views and mergeable views and to pre-defined and inline views as well as to views generated internally by the optimizer during various transformations. The following shows the types of views on which join predicate pushdown is currently supported. UNION ALL/UNION view Outer-joined view Anti-joined view Semi-joined view DISTINCT view GROUP-BY view Examples Consider query A, which has an outer-joined view V. The view cannot be merged, as it contains two tables, and the join between these two tables must be performed before the join between the view and the outer table T4. A: SELECT T4.unique1, V.unique3 FROM T_4K T4,            (SELECT T10.unique3, T10.hundred, T10.ten             FROM T_5K T5, T_10K T10             WHERE T5.unique3 = T10.unique3) VWHERE T4.unique3 = V.hundred(+) AND       T4.ten = V.ten(+) AND       T4.thousand = 5; The following shows the non-default plan for query A generated by disabling join predicate pushdown. When query A undergoes join predicate pushdown, it yields query B. Note that query B is expressed in a non-standard SQL and shows an internal representation of the query. B: SELECT T4.unique1, V.unique3 FROM T_4K T4,           (SELECT T10.unique3, T10.hundred, T10.ten             FROM T_5K T5, T_10K T10             WHERE T5.unique3 = T10.unique3             AND T4.unique3 = V.hundred(+)             AND T4.ten = V.ten(+)) V WHERE T4.thousand = 5; The execution plan for query B is shown below. In the execution plan BX, note the keyword 'VIEW PUSHED PREDICATE' indicates that the view has undergone the join predicate pushdown transformation. The join predicates (shown here in red) have been moved into the view V; these join predicates open up index access paths thereby enabling index-based nested-loop join of the view. With join predicate pushdown, the cost of query A has come down from 62 to 32.  As mentioned earlier, the join predicate pushdown transformation is cost-based, and a join predicate pushed-down plan is selected only when it reduces the overall cost. Consider another example of a query C, which contains a view with the UNION ALL set operator.C: SELECT R.unique1, V.unique3 FROM T_5K R,            (SELECT T1.unique3, T2.unique1+T1.unique1             FROM T_5K T1, T_10K T2             WHERE T1.unique1 = T2.unique1             UNION ALL             SELECT T1.unique3, T2.unique2             FROM G_4K T1, T_10K T2             WHERE T1.unique1 = T2.unique1) V WHERE R.unique3 = V.unique3 and R.thousand < 1; The execution plan of query C is shown below. In the above, 'VIEW UNION ALL PUSHED PREDICATE' indicates that the UNION ALL view has undergone the join predicate pushdown transformation. As can be seen, here the join predicate has been replicated and pushed inside every branch of the UNION ALL view. The join predicates (shown here in red) open up index access paths thereby enabling index-based nested loop join of the view. Consider query D as an example of join predicate pushdown into a distinct view. We have the following cardinalities of the tables involved in query D: Sales (1,016,271), Customers (50,000), and Costs (787,766).  D: SELECT C.cust_last_name, C.cust_city FROM customers C,            (SELECT DISTINCT S.cust_id             FROM sales S, costs CT             WHERE S.prod_id = CT.prod_id and CT.unit_price > 70) V WHERE C.cust_state_province = 'CA' and C.cust_id = V.cust_id; The execution plan of query D is shown below. As shown in XD, when query D undergoes join predicate pushdown transformation, the expensive DISTINCT operator is removed and the join is converted into a semi-join; this is possible, since all the SELECT list items of the view participate in an equi-join with the outer tables. Under similar conditions, when a group-by view undergoes join predicate pushdown transformation, the expensive group-by operator can also be removed. With the join predicate pushdown transformation, the elapsed time of query D came down from 63 seconds to 5 seconds. Since distinct and group-by views are mergeable views, the cost-based transformation framework also compares the cost of merging the view with that of join predicate pushdown in selecting the most optimal execution plan. Summary We have tried to illustrate the basic ideas behind join predicate pushdown on different types of views by showing example queries that are quite simple. Oracle can handle far more complex queries and other types of views not shown here in the examples. Again many thanks to Rafi Ahmed for the content of this blog post.

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  • AutoMapper MappingFunction from Source Type of NameValueCollection

    - by REA_ANDREW
    I have had a situation arise today where I need to construct a complex type from a source of a NameValueCollection.  A little while back I submitted a patch for the Agatha Project to include REST (JSON and XML) support for the service contract.  I realized today that as useful as it is, it did not actually support true REST conformance, as REST should support GET so that you can use JSONP from JavaScript directly meaning you can query cross domain services.  My original implementation for POX and JSON used the POST method and this immediately rules out JSONP as from reading, JSONP only works with GET Requests. This then raised another issue.  The current operation contract of Agatha and one of its main benefits is that you can supply an array of Request objects in a single request, limiting the about of server requests you need to make.  Now, at the present time I am thinking that this will not be the case for the REST imlementation but will yield the benefits of the fact that : The same Request objects can be used for SOAP and RST (POX, JSON) The construct of the JavaScript functions will be simpler and more readable It will enable the use of JSONP for cross domain REST Services The current contract for the Agatha WcfRequestProcessor is at time of writing the following: [ServiceContract] public interface IWcfRequestProcessor { [OperationContract(Name = "ProcessRequests")] [ServiceKnownType("GetKnownTypes", typeof(KnownTypeProvider))] [TransactionFlow(TransactionFlowOption.Allowed)] Response[] Process(params Request[] requests); [OperationContract(Name = "ProcessOneWayRequests", IsOneWay = true)] [ServiceKnownType("GetKnownTypes", typeof(KnownTypeProvider))] void ProcessOneWayRequests(params OneWayRequest[] requests); }   My current proposed solution, and at the very early stages of my concept is as follows: [ServiceContract] public interface IWcfRestJsonRequestProcessor { [OperationContract(Name="process")] [ServiceKnownType("GetKnownTypes", typeof(KnownTypeProvider))] [TransactionFlow(TransactionFlowOption.Allowed)] [WebGet(UriTemplate = "process/{name}/{*parameters}", BodyStyle = WebMessageBodyStyle.WrappedResponse, ResponseFormat = WebMessageFormat.Json)] Response[] Process(string name, NameValueCollection parameters); [OperationContract(Name="processoneway",IsOneWay = true)] [ServiceKnownType("GetKnownTypes", typeof(KnownTypeProvider))] [WebGet(UriTemplate = "process-one-way/{name}/{*parameters}", BodyStyle = WebMessageBodyStyle.WrappedResponse, ResponseFormat = WebMessageFormat.Json)] void ProcessOneWayRequests(string name, NameValueCollection parameters); }   Now this part I have not yet implemented, it is the preliminart step which I have developed which will allow me to take the name of the Request Type and the NameValueCollection and construct the complex type which is that of the Request which I can then supply to a nested instance of the original IWcfRequestProcessor  and work as it should normally.  To give an example of some of the urls which you I envisage with this method are: http://www.url.com/service.svc/json/process/getweather/?location=london http://www.url.com/service.svc/json/process/getproductsbycategory/?categoryid=1 http://www.url.om/service.svc/json/process/sayhello/?name=andy Another reason why my direction has gone to a single request for the REST implementation is because of restrictions which are imposed by browsers on the length of the url.  From what I have read this is on average 2000 characters.  I think that this is a very acceptable usage limit in the context of using 1 request, but I do not think this is acceptable for accommodating multiple requests chained together.  I would love to be corrected on that one, I really would but unfortunately from what I have read I have come to the conclusion that this is not the case. The mapping function So, as I say this is just the first pass I have made at this, and I am not overly happy with the try catch for detecting types without default constructors.  I know there is a better way but for the minute, it escapes me.  I would also like to know the correct way for adding mapping functions and not using the anonymous way that I have used.  To achieve this I have used recursion which I am sure is what other mapping function use. As you do have to go as deep as the complex type is. public static object RecurseType(NameValueCollection collection, Type type, string prefix) { try { var returnObject = Activator.CreateInstance(type); foreach (var property in type.GetProperties()) { foreach (var key in collection.AllKeys) { if (String.IsNullOrEmpty(prefix) || key.Length > prefix.Length) { var propertyNameToMatch = String.IsNullOrEmpty(prefix) ? key : key.Substring(property.Name.IndexOf(prefix) + prefix.Length + 1); if (property.Name == propertyNameToMatch) { property.SetValue(returnObject, Convert.ChangeType(collection.Get(key), property.PropertyType), null); } else if(property.GetValue(returnObject,null) == null) { property.SetValue(returnObject, RecurseType(collection, property.PropertyType, String.Concat(prefix, property.PropertyType.Name)), null); } } } } return returnObject; } catch (MissingMethodException) { //Quite a blunt way of dealing with Types without default constructor return null; } }   Another thing is performance, I have not measured this in anyway, it is as I say the first pass, so I hope this can be the start of a more perfected implementation.  I tested this out with a complex type of three levels, there is no intended logical meaning to the properties, they are simply for the purposes of example.  You could call this a spiking session, as from here on in, now I know what I am building I would take a more TDD approach.  OK, purists, why did I not do this from the start, well I didn’t, this was a brain dump and now I know what I am building I can. The console test and how I used with AutoMapper is as follows: static void Main(string[] args) { var collection = new NameValueCollection(); collection.Add("Name", "Andrew Rea"); collection.Add("Number", "1"); collection.Add("AddressLine1", "123 Street"); collection.Add("AddressNumber", "2"); collection.Add("AddressPostCodeCountry", "United Kingdom"); collection.Add("AddressPostCodeNumber", "3"); AutoMapper.Mapper.CreateMap<NameValueCollection, Person>() .ConvertUsing(x => { return(Person) RecurseType(x, typeof(Person), null); }); var person = AutoMapper.Mapper.Map<NameValueCollection, Person>(collection); Console.WriteLine(person.Name); Console.WriteLine(person.Number); Console.WriteLine(person.Address.Line1); Console.WriteLine(person.Address.Number); Console.WriteLine(person.Address.PostCode.Country); Console.WriteLine(person.Address.PostCode.Number); Console.ReadLine(); }   Notice the convention that I am using and that this method requires you do use.  Each property is prefixed with the constructed name of its parents combined.  This is the convention used by AutoMapper and it makes sense. I can also think of other uses for this including using with ASP.NET MVC ModelBinders for creating a complex type from the QueryString which is itself is a NameValueCollection. Hope this is of some help to people and I would welcome any code reviews you could give me. References: Agatha : http://code.google.com/p/agatha-rrsl/ AutoMapper : http://automapper.codeplex.com/   Cheers for now, Andrew   P.S. I will have the proposed solution for a more complete REST implementation for AGATHA very soon. 

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