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  • Starting/Stopping Custom PHP Chat Server Linux Service (CentOS)

    - by chad
    I have been trying all night to get this service working properly. I created this script from a template and am very new to bash coding. I wrote a fully functioning chat server in php which runs endlessly, but now want to make it a dedicated service. I want to do this so that it starts on server boot and boots back up if possible when there are any down-times with the server. The issue is that I need this thing to run in a detached screen so that I can monitor packet data or send server commands via SSH when need-be. The main problem that i'm having is that it needs to have its own PID when it starts so that I can stop/restart it when needed. I am the type who grinds on coding until I figure it out, but this is so new to me that it seems the learning curve here is very steep and frustrating. Below is my code if anybody can please help me with this one, i've gotten so tired I can't even concentrate any more :( #!/bin/sh # # chatserver # # chkconfig: 345 20 90 # description: chatServer Linux Service Daemon \ # for general server handling ### BEGIN INIT INFO # Provides: chatserver # Required-Start: $local_fs $network $named $syslog # Required-Stop: $local_fs $syslog # Default-Start: 3 4 5 # Default-Stop: 0 1 2 6 # Short-Description: This service maintains the chatServer # Description: chatServer Linux Service Daemon # for general server handling ### END INIT INFO # Source function library. . /etc/rc.d/init.d/functions exec="screen php -q /var/www/html/chatServer.php" prog="chatserver" config="/etc/sysconfig/$prog" pidfile="/var/run/chatserver.pid" [ -e /etc/sysconfig/$prog ] && . /etc/sysconfig/$prog lockfile=/var/lock/subsys/$prog start() { #$exec || exit 5 echo -n $"Starting $prog: " daemon $exec --name=$exec --pidfile=$pidfile retval=$? echo [ $retval -eq 0 ] && touch $lockfile return $retval } stop() { echo -n $"Stopping $prog: " killproc -p $pidfile rm -f $pidfile retval=$? echo [ $retval -eq 0 ] && rm -f $lockfile return $retval } restart() { stop start } reload() { restart } force_reload() { restart } rh_status() { # run checks to determine if the service is running or use generic status status $prog } rh_status_q() { rh_status >/dev/null 2>&1 } case "$1" in start) rh_status_q && exit 0 $1 ;; stop) rh_status_q || exit 0 $1 ;; restart) $1 ;; reload) rh_status_q || exit 7 $1 ;; force-reload) force_reload ;; status) rh_status ;; condrestart|try-restart) rh_status_q || exit 0 restart ;; *) echo $"Usage: $0 {start|stop|status|restart|condrestart|try-restart|reload|force-reload}" exit 2 esac exit $?

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  • Why can't I use SSL certs imported via Server Admin in a custom Apache install?

    - by morgant
    I've got a couple of Mac OS X 10.6.8 Server web servers that run a custom AMP255 (Apache 2.x, MySQL 5.x, and PHP 5.x) stack installed using MacPorts. We've got a lot of Mac OS X Server servers and generally install SSL certs via Server Admin and they "just work" in the built-in services, however, these web servers have always had SSL certs installed in a non-standard location and used only for Apache. Long story short, we're trying to standardize this part of our administration and install certs via Server Admin, but have run into the following issue: when the certs are installed via Server Admin and referenced in our Apache conf files, Apache then prompts for a password upon trying to start. It does not seem to be any password we know, certainly not the admin or keychain passwords! We've added the _www user to the certusers (mainly just to ensure they have the proper access to the private key in /etc/certificates/). So, with the custom installed certs we have the following files (basically just pasted in from the company we purchase our certs from): -rw-r--r-- 1 root admin 1395 Apr 10 11:22 *.domain.tld.ca -rw-r--r-- 1 root admin 1656 Apr 10 11:21 *.domain.tld.cert -rw-r--r-- 1 root admin 1680 Apr 10 11:22 *.domain.tld.key And the following in the VirtualHost in /opt/local/apache2/conf/extra/httpd-ssl.conf: SSLCertificateFile /path/to/certs/*.domain.tld.cert SSLCertificateKeyFile /path/to/certs/*.domain.tld.key SSLCACertificateFile /path/to/certs/*.domain.tld.ca This setup functions normally. If we use the certs installed via Server Admin, which both Server Admin & Keychain Assistant show as valid, they're installed in /etc/certificates/ as follows: -rw-r--r-- 1 root wheel 1655 Apr 9 13:44 *.domain.tld.SOMELONGHASH.cert.pem -rw-r--r-- 1 root wheel 4266 Apr 9 13:44 *.domain.tld.SOMELONGHASH.chain.pem -rw-r----- 1 root certusers 3406 Apr 9 13:44 *.domain.tld.SOMELONGHASH.concat.pem -rw-r----- 1 root certusers 1751 Apr 9 13:44 *.domain.tld.SOMELONGHASH.key.pem And if we replace the aforementioned lines in our httpd-ssl.conf with the following: SSLCertificateFile /etc/certificates/*.domain.tld.SOMELONGHASH.cert.pem SSLCertificateKeyFile /etc/certificates/*.domain.tld.SOMELONGHASH.key.pem SSLCertificateChainFile /etc/certificates/*.domain.tld.SOMELONGHASH.chain.pem This prompts for the unknown password. I have also tried httpd-ssl.conf configured as follows: SSLCertificateFile /etc/certificates/*.domain.tld.SOMELONGHASH.cert.pem SSLCertificateKeyFile /etc/certificates/*.domain.tld.SOMELONGHASH.key.pem SSLCertificateChainFile /etc/certificates/*.domain.tld.SOMELONGHASH.concat.pem And as: SSLCertificateFile /etc/certificates/*.domain.tld.SOMELONGHASH.cert.pem SSLCertificateKeyFile /etc/certificates/*.domain.tld.SOMELONGHASH.key.pem SSLCACertificateFile /etc/certificates/*.domain.tld.SOMELONGHASH.chain.pem We've verified that the certificate is configured to allow all applications access it (in Keychain Assistant). A diff of the /etc/certificates/*.domain.tld.SOMELONGHASH.key.pem & *.domain.tld.key files shows the former is encrypted and the latter is not, so we're assuming that Server Admin/Keychain Assistant is encrypting them for some reason. I know I can create an unencrypted key file as follows: sudo openssl rsa -in /etc/certificates/*.domain.tld.SOMELONGHASH.key.pem -out /etc/certificates/*.domain.tld.SOMELONGHASH.key.no_password.pem But, I can't do that without entering the password. I thought maybe I could export an unencrypted copy of the key from Keychain Admin, but I'm not seeing such an option (not to mention that the .pem options are greyed out in all export options). Any assistance would be greatly appreciated.

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  • Optimal setup for ASUS P6X58D Premium BIOS (no OC)?

    - by rumtscho
    Normally, I'd trust the mainboard manufacturer to choose the best options as defaults. But I had trouble with the board, because even with Quick Boot enabled, it booted twice as slowly as a Pentium 4 Celeron. Then I changed lots of options at once (most of them weren't explained in the manual, just mentioned with a single sentence) and the boot time is only marginally worse than the Pentium 4 (54 sec against 46 sec from button to pw entering screen). Now I don't know if I have turned something off which should have stayed on. I guess I even won't be able to boot from a CD now, because even though it is present in the boot sequence, I took off a timeout I think it needs to check whether there is a disk in the drive. The second reason is that I don't have an internal HDD, only a SSD. I forgot my sources blush but I am under the impression that today's BIOS and OS options are geared toward booting from a HDD, which is often less than optimal when one boots from a SSD, especially when there are functions which cause avoidable writing cycles, as a SSD wears out after too many writing cycles. Most of the things I've read concern the OS, but there are some BIOS-relevant options too. I am especially confused about the disk mode. The board supports AHCI, IDE-simulation and RAID, but of the different articles I've read, there is a proponent for each and no clear arguments for any. So can one tell me which options are important in general and which are important for a SSD-only system? I don't want to overclock the CPU, so you don't have to say anything about this (yes I know the board is meant for OC:)). I am thinking of overclocking the RAM, since they sold me 1600er heatsinked modules which are running at 1066 now, but I'm not sure yet about that. The rest of the system: i7-930, Intel X25-m G2, 6 GB RAM, GTS 250, some no-name Blue-ray ROM. 2 external HDDs over USB 2.0. Lots of other USB-connected hardware (12 devices I think), no SATA 3 drives (will disabling the controller have an impact on performance?), no LAN, only WiFi. Lucid Lynx 64 bit, no dual boot, no virtual installations. The main uses of the system are: managing and playing/showing all the media stored on the external disks, lots of image manipulation, some video editing, a bit of (non-demanding) gaming, rarely development. Lots of Internet surfing too, but this shouldn't have much impact on performance.

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  • Splunk is fantastically expensive: What are the alternatives? [closed]

    - by samsmith
    Possible Duplicate: Alternatives to Splunk? This has been discussed, but it has been several months, so it may be time to revisit it: Earlier discussion RE Splunk alternatives For the record, Splunk rocks. But the pricing is simply beyond what we can consider (When I spoke with Splunk today, the cost for a system to index 5gb/day of data is over $30,000.) That is more than we spend on SQL Server (by a large multiple), more than we spend on a rack of servers (by a multiple), etc. etc. The splunk sales team is correct (that for $30K we get more value and functionality than if we spend the same building our own system), but it doesn't matter. The splunk cost is simply too high (by a multiple). Soooooo, we are looking around! Is anyone out there building a splunk like system? Our basic need: Able to listen for syslog messages on multiple udp ports Able to index the incoming data in an async way Some kind of search engine Some kind of UI An API to the search engine (to embed in our console) We currently need to index 3-5gb/day, but need to be able to scale to 10gb/day or more. We do not need a lot of history (30 days is fine). We use Windows 2008 and 2003 servers. Thanks for your thoughts! UPDATE: We spent two weeks researching commercial and open source options. Our conclusion: Write our own (we are a software company... we know how to write things). We built a great system built on mongodb and .NET that gives us the functions we needed from MongoDB in about one engineering week. We have now completed our implementation. We use two Mongodb servers (master and slave), and are able to log and index any amount of log data (5gb/day, 15gb/day, etc), limited only by disk space. OBSERVATIONS: This space needs a solid solution that is $1000-3000 flat rate. The licensing models used by the commercial firms are based on a "milk the data center ops guys" models. That is their right (of course!), but it leaves a HUGE space open for someone to come in underneath them. My guess is that in another year or two there will be a good open source solution that will be really usable. Thank you all for your input (even if it was self promotion).

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  • Could I centralize batch files more efficiently?

    - by PeanutsMonkey
    I am new to the world of batch scripting so please forgive what may appear as basic questions. I am learning as I get assigned different jobs and I am a huge proponent of automation where possible. I have several batch files that perform several tasks. Each of these files had their paths hard-coded e.g. c:\temp. d:\data, etc in the batch file. Initially I moved these to a text file I could call from a batch file e.g. for /f "tokens=1,2 delims==" %%R in (config.txt) do ( if %%R==bdata set bdata=%%S if %%R==cdata set cdata=%%S ) The config.txt file contains these values bdata=c:\temp cdata=d:\data I realized that each time I would need to create a new variable, I would need to update the config.txt file as well the config.bat files. I decided I would move all the values to just the config.bat file as follows set bdata=c:\temp set cdata=d:\data I then updated each of the existing batch files to call the variables rather than the hard-coded paths. I also added the following lines of code to each batch file except config.bat. The only additional line added to the config.bat file is @echo off. @echo off setlocal enableextensions enabledelayedexpansion call config.bat I then have another batch file that centralizes calling all the batch files in sequence. The name of this batch file is start.bat. The reason I am using start /wait is because there have been instances of where the delete.bat runs before compress.bat has had an opportunity to finish. start /wait compress.bat start /wait validate.bat start /wait delete.bat Questions Is this the best way to centralize values and if not, what is a better way? Do I need to specify setlocal enableextensions enabledelayedexpansion in all the existing batch files? Do all the batch files have to have @echo off or is it sufficient for just the config.bat file? Is start /wait the best way to call multiple files? Can I pass values from one batch file to another using the said command? All the batch files have different functions e.g. move, delete, etc however use %%a or %%b. Is this okay? For example The validate.bat file has the code for %%a in (%bdata%\*.*) do if "%%~xa" == "" move /Y "%bdata%\%%~xa" "%bdata%\%done%" and the delete.bat file has the code for %%a in (%bdata%\*.*) do if "%%~xa" == ".txt" del "%%a"

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  • Explanation of the init.d/scripts Fedora

    - by Shahmir Javaid
    Below is a copy of vsftpd, i need some explanations of some of the scripts mentioned below in this script: #!/bin/bash # ### BEGIN INIT INFO # Provides: vsftpd # Required-Start: $local_fs $network $named $remote_fs $syslog # Required-Stop: $local_fs $network $named $remote_fs $syslog # Short-Description: Very Secure Ftp Daemon # Description: vsftpd is a Very Secure FTP daemon. It was written completely from # scratch ### END INIT INFO # vsftpd This shell script takes care of starting and stopping # standalone vsftpd. # # chkconfig: - 60 50 # description: Vsftpd is a ftp daemon, which is the program \ # that answers incoming ftp service requests. # processname: vsftpd # config: /etc/vsftpd/vsftpd.conf # Source function library. . /etc/rc.d/init.d/functions # Source networking configuration. . /etc/sysconfig/network RETVAL=0 prog="vsftpd" start() { # Start daemons. # Check that networking is up. [ ${NETWORKING} = "no" ] && exit 1 [ -x /usr/sbin/vsftpd ] || exit 1 if [ -d /etc/vsftpd ] ; then CONFS=`ls /etc/vsftpd/*.conf 2>/dev/null` [ -z "$CONFS" ] && exit 6 for i in $CONFS; do site=`basename $i .conf` echo -n $"Starting $prog for $site: " daemon /usr/sbin/vsftpd $i RETVAL=$? echo if [ $RETVAL -eq 0 ]; then touch /var/lock/subsys/$prog break else if [ -f /var/lock/subsys/$prog ]; then RETVAL=0 break fi fi done else RETVAL=1 fi return $RETVAL } stop() { # Stop daemons. echo -n $"Shutting down $prog: " killproc $prog RETVAL=$? echo [ $RETVAL -eq 0 ] && rm -f /var/lock/subsys/$prog return $RETVAL } # See how we were called. case "$1" in start) start ;; stop) stop ;; restart|reload) stop start RETVAL=$? ;; condrestart|try-restart|force-reload) if [ -f /var/lock/subsys/$prog ]; then stop start RETVAL=$? fi ;; status) status $prog RETVAL=$? ;; *) echo $"Usage: $0 {start|stop|restart|try-restart|force-reload|status}" exit 1 esac exit $RETVAL Question I What the hell is the difference between the && and || signs in the below commands, and is it just an easy way to do a simple if check or is it completely different to a if[..something..]; then ..something.. fi: # Check that networking is up. [ ${NETWORKING} = "no" ] && exit 1 [ -x /usr/sbin/vsftpd ] || exit 1 Question II i get what -eq and -gt is (equal to, greater than) but is there a simple website that explains what -x, -d and -f are? Any help would be apreciated Running Fedora 12 on my OS. Script copied from /etc/init.d/vsftpd Question III It says required starts are $local_fs $network $named $remote_fs $syslog but i cant see any where it checks for those.

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  • eAccelerator settings for PHP/Centos/Apache

    - by bobbyh
    I have eAccelerator installed on a server running Wordpress using PHP/Apache on CentOS. I am occassionally getting persistent "white pages", which presumably are PHP Fatal Errors (although these errors don't appear in my error_log). These "white pages" are sprinkled here and there throughout the site. They persist until I go to my eAccelerator control.php page and clear/clean/purge my caches, which suggests to me that I've configured eAccelerator improperly. Here are my current /etc/php.ini settings: memory_limit = 128M; eaccelerator.shm_size="64", where shm.size is "the amount of shared memory eAccelerator should allocate to cache PHP scripts" (see http://eaccelerator.net/wiki/Settings) eaccelerator.shm_max="0", where shm_max is "the maximum size a user can put in shared memory with functions like eaccelerator_put ... The default value is "0" which disables the limit" eaccelerator.shm_ttl="0" - "When eAccelerator doesn't have enough free shared memory to cache a new script it will remove all scripts from shared memory cache that haven't been accessed in at least shm_ttl seconds. By default this value is set to "0" which means that eAccelerator won't try to remove any old scripts from shared memory." eaccelerator.shm_prune_period="0" - "When eAccelerator doesn't have enough free shared memory to cache a script it tries to remove old scripts if the previous try was made more then "shm_prune_period" seconds ago. Default value is "0" which means that eAccelerator won't try to remove any old script from shared memory." eaccelerator.keys = "shm_only" - "These settings control the places eAccelerator may cache user content. ... 'shm_only' cache[s] data in shared memory" On my phpinfo page, it says: memory_limit 128M Version 0.9.5.3 and Caching Enabled true On my eAccelerator control.php page, it says 64 MB of total RAM available Memory usage 77.70% (49.73MB/ 64.00MB) 27.6 MB is used by cached scripts in the PHP opcode cache (I added up the file sizes myself) 22.1 MB is used by the cache keys, which is populated by the Wordpress object cache. My questions are: Is it true that there is only 36.4 MB of room in the eAccelerator cache for total "cache keys" (64 MB of total RAM minus whatever is taken by cached scripts, which is 27.6 MB at the moment)? What happens if my app tries to write more than 22.1 MB of cache keys to the eAccelerator memory cache? Does this cause eAccelerator to go crazy, like I've seen? If I change eaccelerator.shm_max to be equal to (say) 32 MB, would that avoid this problem? Do I also need to change shm_ttl and shm_prune_period to make eAccelerator respect the MB limit set by shm_max? Thanks! :-)

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  • Nagios shell script cannot be executed

    - by MeinAccount
    I'm trying to monitor GitLab with nagios. I've created the following command definition and shell script but when checking the service I'm receiving the following e-mail. How can I solve this? The file is executable. [...] nagios : 3 incorrect password attempts ; TTY=unknown ; PWD=/ ; USER=git ; COMMAND=/bin/bash -c /var/lib/nagios/custom_plugins/check_gitlab.sh Command definition: define command { command_name custom_check_gitlab command_line /var/lib/nagios/custom_plugins/check_gitlab.sh } Shell script: #! /bin/sh # [...] RAILS_ENV="production" # Script variable names should be lower-case not to conflict with internal /bin/sh variables such as PATH, EDITOR or SHELL. app_root="/home/git/gitlab" app_user="git" unicorn_conf="$app_root/config/unicorn.rb" pid_path="$app_root/tmp/pids" socket_path="$app_root/tmp/sockets" web_server_pid_path="$pid_path/unicorn.pid" sidekiq_pid_path="$pid_path/sidekiq.pid" ### Here ends user configuration ### # Switch to the app_user if it is not he/she who is running the script. if [ "$USER" != "$app_user" ]; then sudo -u "$app_user" -H -i $0 "$@"; exit; fi # Switch to the gitlab path, if it fails exit with an error. if ! cd "$app_root" ; then echo "Failed to cd into $app_root, exiting!"; exit 1 fi ### Init Script functions check_pids(){ if ! mkdir -p "$pid_path"; then echo "Could not create the path $pid_path needed to store the pids." exit 1 fi # If there exists a file which should hold the value of the Unicorn pid: read it. if [ -f "$web_server_pid_path" ]; then wpid=$(cat "$web_server_pid_path") else wpid=0 fi if [ -f "$sidekiq_pid_path" ]; then spid=$(cat "$sidekiq_pid_path") else spid=0 fi } # Checks whether the different parts of the service are already running or not. check_status(){ check_pids # If the web server is running kill -0 $wpid returns true, or rather 0. # Checks of *_status should only check for == 0 or != 0, never anything else. if [ $wpid -ne 0 ]; then kill -0 "$wpid" 2>/dev/null web_status="$?" else web_status="-1" fi if [ $spid -ne 0 ]; then kill -0 "$spid" 2>/dev/null sidekiq_status="$?" else sidekiq_status="-1" fi } check_pids check_status if [ "$web_status" != "0" -a "$sidekiq_status" != "0" ]; then echo "GitLab is not running." exit 2 fi if [ "$web_status" != "0" ]; then printf "The GitLab Unicorn webserver is \033[31mnot running\033[0m.\n" exit 1 fi if [ "$sidekiq_status" != "0" ]; then printf "The GitLab Sidekiq job dispatcher is \033[31mnot running\033[0m.\n" exit 1 fi if [ "$web_status" = "0" -a "$sidekiq_status" = "0" ]; then printf "GitLab and all it's components are \033[32mup and running\033[0m.\n" exit 0 fi

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  • How to change the Nginx default folder?

    - by Ido Bukin
    I setup a server with Nginx and i set my Public_HTML in - /home/user/public_html/website.com/public And its always redirect to - /usr/local/nginx/html/ How can i change this ? Nginx.conf - user www-data www-data; worker_processes 4; events { worker_connections 1024; } http { include mime.types; default_type application/octet-stream; sendfile on; tcp_nopush on; tcp_nodelay off; keepalive_timeout 5; 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 /usr/local/nginx/sites-enabled/*; } /usr/local/nginx/sites-enabled/default - server { listen 80; server_name localhost; location / { root html; index index.php index.html index.htm; } # redirect server error pages to the static page /50x.html error_page 500 502 503 504 /50x.html; location = /50x.html { root html; } } /usr/local/nginx/sites-available/website.com - server { listen 80; server_name website.com; rewrite ^/(.*) http://www.website.com/$1 permanent; } server { listen 80; server_name www.website.com; access_log /home/user/public_html/website.com/log/access.log; error_log /home/user/public_html/website.com/log/error.log; location / { root /home/user/public_html/website.com/public/; index index.php index.html; } # pass the PHP scripts to FastCGI server listening on # 127.0.0.1:9000 location ~ \.php$ { fastcgi_pass 127.0.0.1:9000; fastcgi_index index.php; include /usr/local/nginx/conf/fastcgi_params; fastcgi_param SCRIPT_FILENAME /home/user/public_html/website.com/public/$fastcgi_script_name; } } The error message I get is Fatal error: require_once() [function.require]: Failed opening required '/usr/local/nginx/html/202-config/functions.php' the server try to find the file in the Nginx folder and not in my Public_Html

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  • How can I change how OS X's 'say' command pronounces a word?

    - by jwhitlock
    OS X's say command is useful for some tasks (such as Skype's 'notify me when a contact comes online), but it is pronouncing some names incorrectly. Is there a way to teach say to pronounce a word differently? For example, try: say "Hi, Joel Spolsky" The 'ol' sounds like 'ball' rather than 'old'. I'd like to add an exception that say "Pronounce Spolsky like this", rather than try to teach new linguistic rules. I bet there is a way since it can pronounce "iphone" as Apple wants. Update - After some research, here's what I've learned: Text-to-speech is split between turning the text to phonemes, and then the phonemes are turned into audio using a voice. Changing the voice doesn't effect the phonemes. The Speech Synthesis Manager has some functions for turning text to phonemes, and a method for registering a speech dictionary that will add new text-phoneme maps. However, Apple's speech dictionary must be in a binary form - I didn't find any plist XML. Using dtrace while running say, I found some interesting files opened in /System/Library/PrivateFrameworks/SpeechDictionary.framework/Resources. This is probably the speech dictionary, but they are all binary, except for Homophones, which is XML. Adding entries to Homophones does nothing - it is probably used in speech-to-text. They are also code signed by Apple - changing them may prevent some programs from working. PrefixDictionary CartNames CartLite SymbolDictionary Homophones There are ways to add text versions of application interface elements so VoiceOver works, a lot of which a developer gets for free, but there are tricky bits. The standard here appears to be to use a phonetic spelling as needed. My guesses are: say is a light layer of code on top of the Speech Synthesis Manager. It would be easy for the Apple devs to add a command line option to take the path to a speech dictionary plist for alternate phoneme mapping, but they didn't. It may be a useful open-source project to write a better say. Skype probably uses Speech Synthesis Manager directly, leaving no hooks to change the way my friend's names are pronounced, other than spelling them phonetically, which is silly. The easiest way to make a command line version of say is how JRobert suggested. Here's my quick implementation, using Doug Harris's spelling suggestion: #!/bin/sh echo $@ | tr '[A-Z]' '[a-z]' | sed "s/spolsky/spowlsky/g" | /usr/bin/say Finally, some fun command line stuff: # Apple is weird sqlite3 /System/Library/PrivateFrameworks/SpeechDictionary.framework/Resources/Tuples .dump # Get too much information about what files are being opened sudo dtrace -n 'syscall::open*:entry { printf("%s %s",execname,copyinstr(arg0)); }' # Just fun say -v bad "Joel Spolsky Spolsky Spolsky Spolsky Spolsky, Joel Spolsky Spolsky Spolsky Spolsky Spolsky" echo "scale=1000; 4*a(1)" | bc -l | say

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  • How does formatting works with a PowerShell function that returns a set of elements?

    - by Steve B
    If I write this small function : function Foo { Get-Process | % { $_ } } And if I run Foo It displays only a small subset of properties: PS C:\Users\Administrator> foo Handles NPM(K) PM(K) WS(K) VM(M) CPU(s) Id ProcessName ------- ------ ----- ----- ----- ------ -- ----------- 86 10 1680 412 31 0,02 5916 alg 136 10 2772 2356 78 0,06 3684 atieclxx 123 7 1780 1040 33 0,03 668 atiesrxx ... ... But even if only 8 columns are shown, there are plenty of other properties (as foo | gm is showing). What is causing this function to show only this 8 properties? I'm actually trying to build a similar function that is returning complex objects from a 3rd party .Net library. The library is flatting a 2 level hierarchy of objects : function Actual { $someDotnetObject.ACollectionProperty.ASecondLevelCollection | % { $_ } } This method is dumping the objects in a list form (one line per property). How can I control what is displayed, keeping the actual object available? I have tried this : function Actual { $someDotnetObject.ACollectionProperty.ASecondLevelCollection | % { $_ } | format-table Property1, Property2 } It shows in a console the expected table : Property1 Property2 --------- --------- ValA ValD ValB ValE ValC ValF But I lost my objects. Running Get-Member on the result shows : TypeName: Microsoft.PowerShell.Commands.Internal.Format.FormatStartData Name MemberType Definition ---- ---------- ---------- Equals Method bool Equals(System.Object obj) GetHashCode Method int GetHashCode() GetType Method type GetType() ToString Method string ToString() autosizeInfo Property Microsoft.PowerShell.Commands.Internal.Format.AutosizeInfo autosizeInfo {get;set;} ClassId2e4f51ef21dd47e99d3c952918aff9cd Property System.String ClassId2e4f51ef21dd47e99d3c952918aff9cd {get;} groupingEntry Property Microsoft.PowerShell.Commands.Internal.Format.GroupingEntry groupingEntry {get;set;} pageFooterEntry Property Microsoft.PowerShell.Commands.Internal.Format.PageFooterEntry pageFooterEntry {get;set;} pageHeaderEntry Property Microsoft.PowerShell.Commands.Internal.Format.PageHeaderEntry pageHeaderEntry {get;set;} shapeInfo Property Microsoft.PowerShell.Commands.Internal.Format.ShapeInfo shapeInfo {get;set;} TypeName: Microsoft.PowerShell.Commands.Internal.Format.GroupStartData Name MemberType Definition ---- ---------- ---------- Equals Method bool Equals(System.Object obj) GetHashCode Method int GetHashCode() GetType Method type GetType() ToString Method string ToString() ClassId2e4f51ef21dd47e99d3c952918aff9cd Property System.String ClassId2e4f51ef21dd47e99d3c952918aff9cd {get;} groupingEntry Property Microsoft.PowerShell.Commands.Internal.Format.GroupingEntry groupingEntry {get;set;} shapeInfo Property Microsoft.PowerShell.Commands.Internal.Format.ShapeInfo shapeInfo {get;set;} Instead of showing the 2nd level child object members. In this case, I can't pipe the result to functions waiting for this type of argument. How does Powershell is supposed to handle such scenario?

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  • Why did Intel drop the Itanium?

    - by Cole Johnson
    I was reading up on the history of the computer and I came along the IA-64 (Itanium) processors. They sounded really interesting and I was confused as to why Intel would decide to drop them. The ability to choose explicitly what 2 instructions you wanted to run in that cycle is a great idea, especially when writing your program in assembly, for example, a faster bootloader. The hundreds of registers should be convincing for any assembly programmer. You could essentially store all the functions variables in the registers if it doesn't call any other ones. The ability to do instructions like this: (qp) xor r1 = r2, r3 ; r1 = r2 XOR r3 (qp) xor r1 = (imm8), r3 ; r1 = (imm8) XOR r3 versus having to do: ; eax = r1 ; ebx = r2 ; ecx = r3 mov eax, ebx ; first put r2 into r1 xor eax, ecx ; then set r1 equivalent to r2 XOR r3 or ; SAME mov eax, (imm32) ; first put (imm32) into r1 xor eax, ecx ; then set r1 equivalent to (imm32) XOR r3 I heard it was because of no backwards x86 comparability, but couldn't thy be fixed by just adding the Pentium circuitry to it and just add a processor flag that would switch it to Itanium mode (like switching to Protected or Long mode) All the great things about it would have surly put them a giant leap ahead of AMD. Any ideas? Sadly this means you will need a very advanced compiler to do this. Or even one per specific model of the CPU. (E.g. a newer version of the Itanium with an extra feature would require different compiler). When I was working on a WinForms (target only had .NET 2.0) project in Visual Studio 2010, I had a compile target of IA-64. That means that there is a .NET runtime that was able to be compiled for IA-64 and a .NET runtime means Windows. Plus, Hamilton's answer mentions Windows NT. Having a full blown OS like Windows NT means that there is a compiler capable of generating IA-64 machine code.

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  • VPN Connection Causes Internal LAN Connection Loss with Server

    - by sleepisfortheweak
    I've tried configuring basic PPTP VPN at my small business using a number of different tutorials. As far as I can tell, the actual VPN connection worked fine, but upon connecting a client, the Server 'disappears' from the internal LAN. The RRAS service must be stopped before the connection is restored. My Setup: The network is simply a DSL Gateway/Router to the outside functioning as NAT/Firewall/DHCP. The server is a Win Server 2008 machine at fixed IP 192.168.1.200. The server has 1 NIC, so I used the 'custom' option when configuring RRAS. The RRAS settings should be default except that I've disabled ports for connection types I'm not using and reduced PPTP ports to 10. I've also created an address pool and disabled DHCP packet forwarding. The server only functions as a File Share and now a VPN Server. Local LAN computers all have mapped network shares to the server authenticated based on Local User/Group setup on the server. The Problem: The moment a client connects through VPN, the server 'disappears' from the local network. All mapped drives disconnect and there is no response to a ping 192.168.1.200. Even if the client disconnects, the server does not re-appear at that address until the RRAS service is stopped. I've Tried: Using an Address Pool inside and outside the local subnet. Using DCHP Relay Checking Inbound/Outbound filters (none enabled) The fact that nothing I've tried has had any effect, and that I can connect and successfully obtain an IP tells me that it's something more fundamental I'm missing. My gut tells me that it's something to do with the second IP address added by the VPN client somehow taking over the interface or traffic from the local LAN accidently getting routed to the VPN client instead of handled at the server once RRAS has become 'active' when a client connects. Hopefully this may be obvious to someone with real IT experience. I've been doing this a while and almost never been stumped. I'm starting to think it might actually be something tricky since my setup is pretty basic yet refuses to work. I'll be happy to include more info if this doesn't ring any bells right away for anyone. Thanks

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  • Setting up Virtual Hosts with Apache on Windows 2008 server for multiple sites. Complicated setup, including subversion

    - by Roeland
    I am setting up apache on my windows 2008 server at my home. It will serve 2 functions. Subversion hosting to allow me and some others to manage company documents with version control Local website hosting for web development. Will need to run several websites since I generally work on more then one site at a time. Heres what I have done so far. I set up subversion and apache 2.2 using some walk troughs. I changed the default port to 1337. (im a nerd) Using dyndns.com I created a domain to forward to my home ip which is dynamic. ( company.gotdns.org) I then went into my DNS for my company.com and added a record to point repo.company.com to company.gotdns.org At this point people who need access to my file repository can access by going to repo.company.com/repo which is good so far. My question comes on the next step, setting up virtual hosts with apache. Ideally I would like to have my local website be viewable by some others in the company from their homes. So, say I am working on site1, I would like to have them be able to view this by going site1.roeland.bythepixel.com. At the same time, I would like to have site10.wouter.bythepixel.com go to his local setup for site10. What I have done for this: I went into my DNS for company.com and added a record to point roeland.company.com to company.gotdns.org (which translates to my ip). I added code to my httpd-vhosts.conf (listed at bottom) I added code to my host file (listed at bottom) Hah, so of course this doenst work as excepted.. going to site1.roeland.bythepixel.com doesnt bring up my test1 site. Could anyone point me where I may be going wrong? Thanks! hosts: 127.0.0.1 localhost 127.0.0.1 sensenich.roeland.bythepixel.com ::1 localhost httpd-vhosts.conf: <VirtualHost *:80> ServerAdmin [email protected] DocumentRoot "F:/Current Projects/sensenich.com" ServerName sensenich.roeland.bythepixel.com ErrorLog "logs/sensenich.roeland.bythepixel.com-error.log" CustomLog "logs/sensenich.roeland.bythepixel.com-access.log" common </VirtualHost>

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  • Why does this service refuse to start on Windows server 2003?

    - by PenguinCoder
    We have a Windows 2003 server with Cebos MQ1 (ver. 7 and ver. GRI) products installed that have been operational for years. After installing Microsoft 2010 C++ Redistributable package needed for other development, the MQ1 GRI service now fails to start. Event logs showed that two additional updates (.NET4 and the 2010 C++ Redistributable SP2) where installed by the redistributable as well. As soon as we discovered the MQ1 service was not starting properly, we removed these three installed packages. However the service still does not start; the dialog that pops up states 'The service started then stopped. '. Event logs when we attempt to start the service show nothing; IE: No errors, crashes, failures, or other information related to this service. Executing the MQ1Serv.exe directly specifies an issue of 'Missing command line operation, must specify install, uninstall and company abbreviation.' sc query MQ1Service(GRI) shows a clean exit for the Win32ExitCode of 0x0. Attempting to reinstall the client or server software gives an error of 'The procedure entry point ReInitializeCriticalSection could not be located in the dynamic link library KERNEL32.dll.' at the 'Registering Libraries' stage. At this point, further research has stated that the required function is in URL.dll and to verify the library is not corrupted. Running an sfc /scannow on the server has replaced a few DLLS; including the URL.DLL to versions from 2005. This actually broke other applications which required a reinstall (one of them being IE 7). After reinstall and updates, url.dll version is 7.0.5730.13 (2009) and Kernel32.dll is version 5.2.3790.4480 (2009). The MQ1 GRI service still will not start, specifying the same error as previous 'Service started then stopped'. Running a disassembler on Kernel32.dll and Url.dll show no functions named ReinitializeCriticalSection. Attempting the reinstall of the MQ1 client and server as well as starting the service again, fails once more. However, setting the compatibility mode on the MQ1 client install exe to 'Windows 95' actually gets the program to install. Setting the compatibility mode on the MQ1 server service does not enable it to start. I have been researching this problem for nearly a week and besides the advice to scan and replace url.dll, have come to no successful conclusions. This service was operational prior to the 2010 C++ install, without any additional parameters or settings. After removing the C++ install and all servicepacks/updates it installed silently, still does not correct the issue of the MQ1 GRI service not starting. Q: Has anyone else run into this or similar issue while attempting to get a service initialized? What have I overlooked or what else can I try in order to get this service started??

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  • Need help recovering a corrupt SQL database

    - by user570079
    I have a very special case that I have been working on for several days. I have a very large SQL Server 2008 database (about 2 TB) that contains 500 filegroups to support very large partitioned tables. Recently we had a catastophic failure on one of the drive and lost several filegroups and the database became in-accessible. We have been doing filegroup backups on a daily basis, but due to other issues, we lost our most recent backup of the log and the primary filegroup. We have all the data backed up but the primary filegroup backup is old. There have been no schema changes since the primary filegroup backup, but the lsn's are now all out of sync and we cannot recover the data. I have tried everything I could think of (and have tried just about every trick and hack I could google) but I still end up at the same point where I get messages saying that the files for filegroup x do not match the primary filegroup. I am now at the point of trying to edit the system tables (we have a separate temporary environment to do this so we are not worried about corrupting any production databases). I have tried updated sys.sysdbreg, sys.sysbrickfiles, and sys.sysprufiles to try to trick SQL into thinking all the files are online, but a "Select * From OPENROWSET(TABLE DBPROP, 5)" shows a different database state from what I see in sys.sysdbreg. I am now thinking I need to somehow edit the headers of the actual data files to try to line up the lsn's with the primary. I appreciate any help anyone can give me here, but please do not respond with things like "you are not supposed to do edit mdf, ndf files...." or "see msdn article....", etc. This is an advanced emergency case and I need a real hack so we can just get to the data in this corrupt database and export to a fresh new database. I know there is a way to do this, but not knowing what the DBPROP system functions does (i.e. does it look at system tables or does it actually open the file) is keeping me from trying to figure out how to fool SQL into allowing me to read these files. Thanks for any help.

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  • bluetooth connection using pybluez

    - by srj0408
    I am working on bluetooth not exactly on bluetooth stack-development but to use bluetooth in one of my project. I had done all that before using some of the py-bluez commands like hciconfig, hcitool scan , then simple-agents and using serial module inside python. But that was quite random. We were able to connect only one specific device based on its bluetooth address and there was no facility of reconnection once the devices are disconnected. Now i want to try out this stuff in a sequential manner like this (i am doing that all on a RPI and for at present on ubuntu 12.04.) i) Store some names in a file along with some other information with respect to that device. ii) Run a script to find out the device in locality with those names and if any one if found, report that. For this step, i had taken a reference from BTBook , made available from MIT. Below is the script for the same, but that script only search for the single name. from bluetooth import * target_name = "XT1033" target_address = None nearby_devices = discover_devices() for address in nearby_devices: if target_name == lookup_name( address ): target_address = address break if target_address is not None: print "found target bluetooth device with address ", target_address connect_socket(target_address); else: print "could not find target bluetooth device nearby" iii) Connect the device using client sock. But i dont have any device on which i can write a simple python script. My client can be any device that will be publishing data. Now i came through a script in the same book, that actually connect to a client requesting permission to connect to server. from bluetooth import * port = 1 server_sock=BluetoothSocket( RFCOMM ) server_sock.bind(("",port)) server_sock.listen(1) client_sock, client_info = server_sock.accept() print "Accepted connection from ", client_info data = client_sock.recv(1024) print "received [%s]" % data client_sock.close() server_sock.close() here client_sock, client_info = server_sock.accept() provide the client address and port requested to be connected. Can i pass address obtained from the earlier script to this, so that it connect server to the client? iv) Then if client get disconnected, re-connect(a simple polling can be used.) All this stuff can be done using bash and py-bluez functions but i want to do that in a sequential manner.I am not a master in python but i can do some small stuff. Can any one guide me for the same or can direct me to more usefull resource through which i can continue my coding part after finding the "X", "Y" named devices.

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  • Django CMS - not able to upload images through cmsplugin_filer_image

    - by Luke
    i have a problem with a local installation on django cms 2.3.3: i've installed it trough pip, in a separated virtualenv. next i followed the tutorial for settings.py configuration, i started the server. Then in the admin i created an page (home), and i've tried to add an image in the placeholder through the cmsplugin_filer_image, but the upload seems that doesn't work. here's my settings.py: # Django settings for cms1 project. # -*- coding: utf-8 -*- import os gettext = lambda s: s PROJECT_PATH = os.path.abspath(os.path.dirname(__file__)) DEBUG = True TEMPLATE_DEBUG = DEBUG ADMINS = ( # ('Your Name', '[email protected]'), ) MANAGERS = ADMINS DATABASES = { 'default': { 'ENGINE': 'django.db.backends.postgresql_psycopg2', # Add 'postgresql_psycopg2', 'mysql', 'sqlite3' or 'oracle'. 'NAME': 'cms1', # Or path to database file if using sqlite3. 'USER': 'cms', # Not used with sqlite3. 'PASSWORD': 'cms', # Not used with sqlite3. 'HOST': '', # Set to empty string for localhost. Not used with sqlite3. 'PORT': '', # Set to empty string for default. Not used with sqlite3. } } # Local time zone for this installation. Choices can be found here: # http://en.wikipedia.org/wiki/List_of_tz_zones_by_name # although not all choices may be available on all operating systems. # In a Windows environment this must be set to your system time zone. TIME_ZONE = 'Europe/Rome' # Language code for this installation. All choices can be found here: # http://www.i18nguy.com/unicode/language-identifiers.html LANGUAGE_CODE = 'it-it' SITE_ID = 1 # If you set this to False, Django will make some optimizations so as not # to load the internationalization machinery. USE_I18N = True # If you set this to False, Django will not format dates, numbers and # calendars according to the current locale. USE_L10N = True # If you set this to False, Django will not use timezone-aware datetimes. USE_TZ = True # Absolute filesystem path to the directory that will hold user-uploaded files. # Example: "/home/media/media.lawrence.com/media/" MEDIA_ROOT = os.path.join(PROJECT_PATH, "media") # URL that handles the media served from MEDIA_ROOT. Make sure to use a # trailing slash. # Examples: "http://media.lawrence.com/media/", "http://example.com/media/" MEDIA_URL = '/media/' # Absolute path to the directory static files should be collected to. # Don't put anything in this directory yourself; store your static files # in apps' "static/" subdirectories and in STATICFILES_DIRS. # Example: "/home/media/media.lawrence.com/static/" STATIC_ROOT = os.path.join(PROJECT_PATH, "static") STATIC_URL = "/static/" # Additional locations of static files STATICFILES_DIRS = ( os.path.join(PROJECT_PATH, "static_auto"), # Put strings here, like "/home/html/static" or "C:/www/django/static". # Always use forward slashes, even on Windows. # Don't forget to use absolute paths, not relative paths. ) # List of finder classes that know how to find static files in # various locations. STATICFILES_FINDERS = ( 'django.contrib.staticfiles.finders.FileSystemFinder', 'django.contrib.staticfiles.finders.AppDirectoriesFinder', # 'django.contrib.staticfiles.finders.DefaultStorageFinder', ) # Make this unique, and don't share it with anybody. SECRET_KEY = '^c2q3d8w)f#gk%5i)(#i*lwt%lm-!2=(*1d!1cf+rg&amp;-hqi_9u' # List of callables that know how to import templates from various sources. TEMPLATE_LOADERS = ( 'django.template.loaders.filesystem.Loader', 'django.template.loaders.app_directories.Loader', # 'django.template.loaders.eggs.Loader', ) MIDDLEWARE_CLASSES = ( 'django.middleware.common.CommonMiddleware', 'django.contrib.sessions.middleware.SessionMiddleware', 'django.middleware.csrf.CsrfViewMiddleware', 'django.contrib.auth.middleware.AuthenticationMiddleware', 'django.contrib.messages.middleware.MessageMiddleware', 'cms.middleware.multilingual.MultilingualURLMiddleware', 'cms.middleware.page.CurrentPageMiddleware', 'cms.middleware.user.CurrentUserMiddleware', 'cms.middleware.toolbar.ToolbarMiddleware', # Uncomment the next line for simple clickjacking protection: # 'django.middleware.clickjacking.XFrameOptionsMiddleware', ) ROOT_URLCONF = 'cms1.urls' # Python dotted path to the WSGI application used by Django's runserver. WSGI_APPLICATION = 'cms1.wsgi.application' TEMPLATE_DIRS = ( os.path.join(PROJECT_PATH, "templates"), # Put strings here, like "/home/html/django_templates" or "C:/www/django/templates". # Always use forward slashes, even on Windows. # Don't forget to use absolute paths, not relative paths. ) CMS_TEMPLATES = ( ('template_1.html', 'Template One'), ('template_2.html', 'Template Two'), ) TEMPLATE_CONTEXT_PROCESSORS = ( 'django.contrib.auth.context_processors.auth', 'django.core.context_processors.i18n', 'django.core.context_processors.request', 'django.core.context_processors.media', 'django.core.context_processors.static', 'cms.context_processors.media', 'sekizai.context_processors.sekizai', ) LANGUAGES = [ ('it', 'Italiano'), ('en', 'English'), ] INSTALLED_APPS = ( 'django.contrib.auth', 'django.contrib.contenttypes', 'django.contrib.sessions', 'django.contrib.sites', 'django.contrib.messages', 'django.contrib.staticfiles', 'cms', #django CMS itself 'mptt', #utilities for implementing a modified pre-order traversal tree 'menus', #helper for model independent hierarchical website navigation 'south', #intelligent schema and data migrations 'sekizai', #for javascript and css management #'cms.plugins.file', 'cms.plugins.flash', 'cms.plugins.googlemap', 'cms.plugins.link', #'cms.plugins.picture', 'cms.plugins.snippet', 'cms.plugins.teaser', 'cms.plugins.text', #'cms.plugins.video', 'cms.plugins.twitter', 'filer', 'cmsplugin_filer_file', 'cmsplugin_filer_folder', 'cmsplugin_filer_image', 'cmsplugin_filer_teaser', 'cmsplugin_filer_video', 'easy_thumbnails', 'PIL', # Uncomment the next line to enable the admin: 'django.contrib.admin', # Uncomment the next line to enable admin documentation: # 'django.contrib.admindocs', ) # A sample logging configuration. The only tangible logging # performed by this configuration is to send an email to # the site admins on every HTTP 500 error when DEBUG=False. # See http://docs.djangoproject.com/en/dev/topics/logging for # more details on how to customize your logging configuration. LOGGING = { 'version': 1, 'disable_existing_loggers': False, 'filters': { 'require_debug_false': { '()': 'django.utils.log.RequireDebugFalse' } }, 'handlers': { 'mail_admins': { 'level': 'ERROR', 'filters': ['require_debug_false'], 'class': 'django.utils.log.AdminEmailHandler' } }, 'loggers': { 'django.request': { 'handlers': ['mail_admins'], 'level': 'ERROR', 'propagate': True, }, } } when i try to upload an image, in the clipboard section i don't have the thumbnail, but just an 'undefined' message: and this is the runserver console while trying to upload: [20/Oct/2012 15:15:56] "POST /admin/filer/clipboard/operations/upload/?qqfile=29708_1306856312320_7706073_n.jpg HTTP/1.1" 500 248133 [20/Oct/2012 15:15:56] "GET /it/admin/filer/folder/unfiled_images/undefined HTTP/1.1" 301 0 [20/Oct/2012 15:15:56] "GET /it/admin/filer/folder/unfiled_images/undefined/ HTTP/1.1" 404 1739 Also, this is project filesystem: cms1 +-- cms1 ¦   +-- __init__.py ¦   +-- __init__.pyc ¦   +-- media ¦   ¦   +-- filer_public ¦   ¦   +-- 2012 ¦   ¦   +-- 10 ¦   ¦   +-- 20 ¦   ¦   +-- 29708_1306856312320_7706073_n_1.jpg ¦   ¦   +-- 29708_1306856312320_7706073_n_2.jpg ¦   ¦   +-- 29708_1306856312320_7706073_n_3.jpg ¦   ¦   +-- 29708_1306856312320_7706073_n_4.jpg ¦   ¦   +-- 29708_1306856312320_7706073_n_5.jpg ¦   ¦   +-- 29708_1306856312320_7706073_n_6.jpg ¦   ¦   +-- 29708_1306856312320_7706073_n_7.jpg ¦   ¦   +-- 29708_1306856312320_7706073_n.jpg ¦   ¦   +-- torrent-client-macosx.jpg ¦   +-- settings.py ¦   +-- settings.pyc ¦   +-- static ¦   +-- static_auto ¦   +-- static_manual ¦   +-- templates ¦   ¦   +-- base.html ¦   ¦   +-- template_1.html ¦   ¦   +-- template_2.html ¦   +-- urls.py ¦   +-- urls.pyc ¦   +-- wsgi.py ¦   +-- wsgi.pyc +-- manage.py So files are uploaded, but they are not accessible to cms. there's a similar question here, but doens't help me so much. It would be very helpful any help on this issue to me. Thanks, luke

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  • What does Ruby have that Python doesn't, and vice versa?

    - by Lennart Regebro
    There is a lot of discussions of Python vs Ruby, and I all find them completely unhelpful, because they all turn around why feature X sucks in language Y, or that claim language Y doesn't have X, although in fact it does. I also know exactly why I prefer Python, but that's also subjective, and wouldn't help anybody choosing, as they might not have the same tastes in development as I do. It would therefore be interesting to list the differences, objectively. So no "Python's lambdas sucks". Instead explain what Ruby's lambdas can do that Python's can't. No subjectivity. Example code is good! Don't have several differences in one answer, please. And vote up the ones you know are correct, and down those you know are incorrect (or are subjective). Also, differences in syntax is not interesting. We know Python does with indentation what Ruby does with brackets and ends, and that @ is called self in Python. UPDATE: This is now a community wiki, so we can add the big differences here. Ruby has a class reference in the class body In Ruby you have a reference to the class (self) already in the class body. In Python you don't have a reference to the class until after the class construction is finished. An example: class Kaka puts self end self in this case is the class, and this code would print out "Kaka". There is no way to print out the class name or in other ways access the class from the class definition body in Python. All classes are mutable in Ruby This lets you develop extensions to core classes. Here's an example of a rails extension: class String def starts_with?(other) head = self[0, other.length] head == other end end Ruby has Perl-like scripting features Ruby has first class regexps, $-variables, the awk/perl line by line input loop and other features that make it more suited to writing small shell scripts that munge text files or act as glue code for other programs. Ruby has first class continuations Thanks to the callcc statement. In Python you can create continuations by various techniques, but there is no support built in to the language. Ruby has blocks With the "do" statement you can create a multi-line anonymous function in Ruby, which will be passed in as an argument into the method in front of do, and called from there. In Python you would instead do this either by passing a method or with generators. Ruby: amethod { |here| many=lines+of+code goes(here) } Python: def function(here): many=lines+of+code goes(here) amethod(function) Interestingly, the convenience statement in Ruby for calling a block is called "yield", which in Python will create a generator. Ruby: def themethod yield 5 end themethod do |foo| puts foo end Python: def themethod(): yield 5 for foo in themethod: print foo Although the principles are different, the result is strikingly similar. Python has built-in generators (which are used like Ruby blocks, as noted above) Python has support for generators in the language. In Ruby you could use the generator module that uses continuations to create a generator from a block. Or, you could just use a block/proc/lambda! Moreover, in Ruby 1.9 Fibers are, and can be used as, generators. docs.python.org has this generator example: def reverse(data): for index in range(len(data)-1, -1, -1): yield data[index] Contrast this with the above block examples. Python has flexible name space handling In Ruby, when you import a file with require, all the things defined in that file will end up in your global namespace. This causes namespace pollution. The solution to that is Rubys modules. But if you create a namespace with a module, then you have to use that namespace to access the contained classes. In Python, the file is a module, and you can import its contained names with from themodule import *, thereby polluting the namespace if you want. But you can also import just selected names with from themodule import aname, another or you can simply import themodule and then access the names with themodule.aname. If you want more levels in your namespace you can have packages, which are directories with modules and an __init__.py file. Python has docstrings Docstrings are strings that are attached to modules, functions and methods and can be introspected at runtime. This helps for creating such things as the help command and automatic documentation. def frobnicate(bar): """frobnicate takes a bar and frobnicates it >>> bar = Bar() >>> bar.is_frobnicated() False >>> frobnicate(bar) >>> bar.is_frobnicated() True """ Python has more libraries Python has a vast amount of available modules and bindings for libraries. Python has multiple inheritance Ruby does not ("on purpose" -- see Ruby's website, see here how it's done in Ruby). It does reuse the module concept as a sort of abstract classes. Python has list/dict comprehensions Python: res = [x*x for x in range(1, 10)] Ruby: res = (0..9).map { |x| x * x } Python: >>> (x*x for x in range(10)) <generator object <genexpr> at 0xb7c1ccd4> >>> list(_) [0, 1, 4, 9, 16, 25, 36, 49, 64, 81] Ruby: p = proc { |x| x * x } (0..9).map(&p) Python: >>> {x:str(y*y) for x,y in {1:2, 3:4}.items()} {1: '4', 3: '16'} Ruby: >> Hash[{1=>2, 3=>4}.map{|x,y| [x,(y*y).to_s]}] => {1=>"4", 3=>"16"} Python has decorators Things similar to decorators can be created in Ruby, and it can also be argued that they aren't as necessary as in Python.

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  • AGENT: The World's Smartest Watch

    - by Rob Chartier
    AGENT: The World's Smartest Watch by Secret Labs + House of Horology Disclaimer: Most if not all of this content has been gleaned from the comments on the Kickstarter project page and comments section. Any discrepancies between this post and any documentation on agentwatches.com, kickstarter.com, etc.., those official sites take precedence. Overview The next generation smartwatch with brand-new technology. World-class developer tools, unparalleled battery life, Qi wireless charging. Kickstarter Page, Comments Funding period : May 21, 2013 - Jun 20, 2013 MSRP : $249 Other Urls http://www.agentwatches.com/ https://www.facebook.com/agentwatches http://twitter.com/agentwatches http://pinterest.com/agentwatches/ http://paper.li/robchartier/1371234640 Developer Story The first official launch of the preview SDK and emulator will happen on 20-Jun-2013.  All development will be done in Visual Studio 2012, using the .NET Micro Framework SDK 2.3.  The SDK will ship with the first round of the expected API for developers along with an emulator. With that said, there is no need to wait for the SDK.  You can download the tooling now and get started with Apps and Faces immediately.  The only thing that you will not be able to work with is the API; but for example, watch faces, you can start building the basic face rendering with the Bitmap graphics drawing in the .NET Micro Framework.   Does it look good? Before we dig into any more of the gory details, here are a few photos of the current available prototype models.   The watch on the tiny QI Charter   If you wander too far away from your phone, your watch will let you know with a vibration and a message, all but one button will dismiss the message.   An app showing the premium weather data!   Nice stitching on the straps, leather and silicon will be available, along with a few lengths to choose from (short, regular, long lengths). On to those gory details…. Hardware Specs Processor 120MHz ARM Cortex-M4 processor (ATSAM4SD32) with secondary AVR co-processor Flash & RAM 2MB of onboard flash and 160KB of RAM 1/4 of the onboard flash will be used by the OS The flash is permanent (non-volatile) storage. Bluetooth Bluetooth 4.0 BD/EDR + LE Bluetooth 4.0 is backwards compatible with Bluetooth 2.1, so classic Bluetooth functions (BD/EDR, SPP/AVRCP/PBAP/etc.) will work fine. Sensors 3D Accelerometer (Motion) ST LSM303DLHC Ambient Light Sensor Hardware power metering Vibration Motor (You can pulse it to create vibration patterns, not sure about the vibration strength - driven with PWM) No piezo/speaker or microphone. Other QI Wireless Charging, no NFC, no wall adapter included Custom LED Backlight No GPS in the watch. It uses the GPS in your phone. AGENT watch apps are deployed and debugged wirelessly from your PC via Bluetooth. RoHS, Pb-free Battery Expected to use a CR2430-sized rechargeable battery – replaceable (Mouser, Amazon) Estimated charging time from empty is 2 hours with provided charger 7 Days typical with Bluetooth on, 30 days with Bluetooth off (watch-face only mode) The battery should last at least 2 years, with 100s of charge cycles. Physical dimensions Roughly 38mm top-to-bottom on the front face 35mm left-to-right on the front face and around 12mm in depth 22mm strap Two ~1/16" hex screws to attach the watch pin The top watchcase material candidates are PVD stainless steel, brushed matte ceramic, and high-quality polycarbonate (TBD). The glass lens is mineral glass, Anti-glare glass lens Strap options Leather and silicon straps will be available Expected to have three sizes Display 1.28" Sharp Memory Display The display stays on 100% of the time. Dimensions: 128x128 pixels Buttons Custom "Pusher" buttons, they will not make noise like a mouse click, and are very durable. The top-left button activates the backlight; bottom-left changes apps; three buttons on the right are up/select/down and can be used for custom purposes by apps. Backup reset procedure is currently activated by holding the home/menu button and the top-right user button for about ten seconds Device Support Android 2.3 or newer iPhone 4S or newer Windows Phone 8 or newer Heart Rate monitors - Bluetooth SPP or Bluetooth LE (GATT) is what you'll want the heart monitor to support. Almost limitless Bluetooth device support! Internationalization & Localization Full UTF8 Support from the ground up. AGENT's user interface is in English. Your content (caller ID, music tracks, notifications) will be in your native language. We have a plan to cover most major character sets, with Latin characters pre-loaded on the watch. Simplified Chinese will be available Feature overview Phone lost alert Caller ID Music Control (possible volume control) Wireless Charging Timer Stopwatch Vibrating Alarm (possibly custom vibrations for caller id) A few default watch faces Airplane mode (by demand or low power) Can be turned off completely Customizable 3rd party watch faces, applications which can be loaded over bluetooth. Sample apps that maybe installed Weather Sample Apps not installed Exercise App Other Possible Skype integration over Bluetooth. They will provide an AGENT app for your smartphone (iPhone, Android, Windows Phone). You'll be able to use it to load apps onto the watch.. You will be able to cancel phone calls. With compatible phones you can also answer, end, etc. They are adopting the standard hands-free profile to provide these features and caller ID.

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  • How to Use Windows’ Advanced Search Features: Everything You Need to Know

    - by Chris Hoffman
    You should never have to hunt down a lost file on modern versions of Windows — just perform a quick search. You don’t even have to wait for a cartoon dog to find your files, like on Windows XP. The Windows search indexer is constantly running in the background to make quick local searches possible. This enables the kind of powerful search features you’d use on Google or Bing — but for your local files. Controlling the Indexer By default, the Windows search indexer watches everything under your user folder — that’s C:\Users\NAME. It reads all these files, creating an index of their names, contents, and other metadata. Whenever they change, it notices and updates its index. The index allows you to quickly find a file based on the data in the index. For example, if you want to find files that contain the word “beluga,” you can perform a search for “beluga” and you’ll get a very quick response as Windows looks up the word in its search index. If Windows didn’t use an index, you’d have to sit and wait as Windows opened every file on your hard drive, looked to see if the file contained the word “beluga,” and moved on. Most people shouldn’t have to modify this indexing behavior. However, if you store your important files in other folders — maybe you store your important data a separate partition or drive, such as at D:\Data — you may want to add these folders to your index. You can also choose which types of files you want to index, force Windows to rebuild the index entirely, pause the indexing process so it won’t use any system resources, or move the index to another location to save space on your system drive. To open the Indexing Options window, tap the Windows key on your keyboard, type “index”, and click the Indexing Options shortcut that appears. Use the Modify button to control the folders that Windows indexes or the Advanced button to control other options. To prevent Windows from indexing entirely, click the Modify button and uncheck all the included locations. You could also disable the search indexer entirely from the Programs and Features window. Searching for Files You can search for files right from your Start menu on Windows 7 or Start screen on Windows 8. Just tap the Windows key and perform a search. If you wanted to find files related to Windows, you could perform a search for “Windows.” Windows would show you files that are named Windows or contain the word Windows. From here, you can just click a file to open it. On Windows 7, files are mixed with other types of search results. On Windows 8 or 8.1, you can choose to search only for files. If you want to perform a search without leaving the desktop in Windows 8.1, press Windows Key + S to open a search sidebar. You can also initiate searches directly from Windows Explorer — that’s File Explorer on Windows 8. Just use the search box at the top-right of the window. Windows will search the location you’ve browsed to. For example, if you’re looking for a file related to Windows and know it’s somewhere in your Documents library, open the Documents library and search for Windows. Using Advanced Search Operators On Windows 7, you’ll notice that you can add “search filters” form the search box, allowing you to search by size, date modified, file type, authors, and other metadata. On Windows 8, these options are available from the Search Tools tab on the ribbon. These filters allow you to narrow your search results. If you’re a geek, you can use Windows’ Advanced Query Syntax to perform advanced searches from anywhere, including the Start menu or Start screen. Want to search for “windows,” but only bring up documents that don’t mention Microsoft? Search for “windows -microsoft”. Want to search for all pictures of penguins on your computer, whether they’re PNGs, JPEGs, or any other type of picture file? Search for “penguin kind:picture”. We’ve looked at Windows’ advanced search operators before, so check out our in-depth guide for more information. The Advanced Query Syntax gives you access to options that aren’t available in the graphical interface. Creating Saved Searches Windows allows you to take searches you’ve made and save them as a file. You can then quickly perform the search later by double-clicking the file. The file functions almost like a virtual folder that contains the files you specify. For example, let’s say you wanted to create a saved search that shows you all the new files created in your indexed folders within the last week. You could perform a search for “datecreated:this week”, then click the Save search button on the toolbar or ribbon. You’d have a new virtual folder you could quickly check to see your recent files. One of the best things about Windows search is that it’s available entirely from the keyboard. Just press the Windows key, start typing the name of the file or program you want to open, and press Enter to quickly open it. Windows 8 made this much more obnoxious with its non-unified search, but unified search is finally returning with Windows 8.1.     

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  • Security in OBIEE 11g, Part 2

    - by Rob Reynolds
    Continuing the series on OBIEE 11g, our guest blogger this week is Pravin Janardanam. Here is Part 2 of his overview of Security in OBIEE 11g. OBIEE 11g Security Overview, Part 2 by Pravin Janardanam In my previous blog on Security, I discussed the OBIEE 11g changes regarding Authentication mechanism, RPD protection and encryption. This blog will include a discussion about OBIEE 11g Authorization and other Security aspects. Authorization: Authorization in 10g was achieved using a combination of Users, Groups and association of privileges and object permissions to users and Groups. Two keys changes to Authorization in OBIEE 11g are: Application Roles Policies / Permission Groups Application Roles are introduced in OBIEE 11g. An application role is specific to the application. They can be mapped to other application roles defined in the same application scope and also to enterprise users or groups, and they are used in authorization decisions. Application roles in 11g take the place of Groups in 10g within OBIEE application. In OBIEE 10g, any changes to corporate LDAP groups require a corresponding change to Groups and their permission assignment. In OBIEE 11g, Application roles provide insulation between permission definitions and corporate LDAP Groups. Permissions are defined at Application Role level and changes to LDAP groups just require a reassignment of the Group to the Application Roles. Permissions and privileges are assigned to Application Roles and users in OBIEE 11g compared to Groups and Users in 10g. The diagram below shows the relationship between users, groups and application roles. Note that the Groups shown in the diagram refer to LDAP Groups (WebLogic Groups by default) and not OBIEE application Groups. The following screenshot compares the permission windows from Admin tool in 10g vs 11g. Note that the Groups in the OBIEE 10g are replaced with Application Roles in OBIEE 11g. The same is applicable to OBIEE web catalog objects.    The default Application Roles available after OBIEE 11g installation are BIAdministrator, BISystem, BIConsumer and BIAuthor. Application policies are the authorization policies that an application relies upon for controlling access to its resources. An Application Role is defined by the Application Policy. The following screenshot shows the policies defined for BIAdministrator and BISystem Roles. Note that the permission for impersonation is granted to BISystem Role. In OBIEE 10g, the permission to manage repositories and Impersonation were assigned to “Administrators” group with no control to separate these permissions in the Administrators group. Hence user “Administrator” also had the permission to impersonate. In OBI11g, BIAdministrator does not have the permission to impersonate. This gives more flexibility to have multiple users perform different administrative functions. Application Roles, Policies, association of Policies to application roles and association of users and groups to application roles are managed using Fusion Middleware Enterprise Manager (FMW EM). They reside in the policy store, identified by the system-jazn-data.xml file. The screenshots below show where they are created and managed in FMW EM. The following screenshot shows the assignment of WebLogic Groups to Application Roles. The following screenshot shows the assignment of Permissions to Application Roles (Application Policies). Note: Object level permission association to Applications Roles resides in the RPD for repository objects. Permissions and Privilege for web catalog objects resides in the OBIEE Web Catalog. Wherever Groups were used in the web catalog and RPD has been replaced with Application roles in OBIEE 11g. Following are the tools used in OBIEE 11g Security Administration: ·       Users and Groups are managed in Oracle WebLogic Administration console (by default). If WebLogic is integrated with other LDAP products, then Users and Groups needs to managed using the interface provide by the respective LDAP vendor – New in OBIEE 11g ·       Application Roles and Application Policies are managed in Oracle Enterprise Manager - Fusion Middleware Control – New in OBIEE 11g ·       Repository object permissions are managed in OBIEE Administration tool – Same as 10g but the assignment is to Application Roles instead of Groups ·       Presentation Services Catalog Permissions and Privileges are managed in OBI Application administration page - Same as 10g but the assignment is to Application Roles instead of Groups Credential Store: Credential Store is a single consolidated service provider to store and manage the application credentials securely. The credential store contains credentials that either user supplied or system generated. Credential store in OBIEE 10g is file based and is managed using cryptotools utility. In 11g, Credential store can be managed directly from the FMW Enterprise Manager and is stored in cwallet.sso file. By default, the Credential Store stores password for deployed RPDs, BI Publisher data sources and BISystem user. In addition, Credential store can be LDAP based but only Oracle Internet Directory is supported right now. As you can see OBIEE security is integrated with Oracle Fusion Middleware security architecture. This provides a common security framework for all components of Business Intelligence and Fusion Middleware applications.

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  • C# 5 Async, Part 1: Simplifying Asynchrony – That for which we await

    - by Reed
    Today’s announcement at PDC of the future directions C# is taking excite me greatly.  The new Visual Studio Async CTP is amazing.  Asynchronous code – code which frustrates and demoralizes even the most advanced of developers, is taking a huge leap forward in terms of usability.  This is handled by building on the Task functionality in .NET 4, as well as the addition of two new keywords being added to the C# language: async and await. This core of the new asynchronous functionality is built upon three key features.  First is the Task functionality in .NET 4, and based on Task and Task<TResult>.  While Task was intended to be the primary means of asynchronous programming with .NET 4, the .NET Framework was still based mainly on the Asynchronous Pattern and the Event-based Asynchronous Pattern. The .NET Framework added functionality and guidance for wrapping existing APIs into a Task based API, but the framework itself didn’t really adopt Task or Task<TResult> in any meaningful way.  The CTP shows that, going forward, this is changing. One of the three key new features coming in C# is actually a .NET Framework feature.  Nearly every asynchronous API in the .NET Framework has been wrapped into a new, Task-based method calls.  In the CTP, this is done via as external assembly (AsyncCtpLibrary.dll) which uses Extension Methods to wrap the existing APIs.  However, going forward, this will be handled directly within the Framework.  This will have a unifying effect throughout the .NET Framework.  This is the first building block of the new features for asynchronous programming: Going forward, all asynchronous operations will work via a method that returns Task or Task<TResult> The second key feature is the new async contextual keyword being added to the language.  The async keyword is used to declare an asynchronous function, which is a method that either returns void, a Task, or a Task<T>. Inside the asynchronous function, there must be at least one await expression.  This is a new C# keyword (await) that is used to automatically take a series of statements and break it up to potentially use discontinuous evaluation.  This is done by using await on any expression that evaluates to a Task or Task<T>. For example, suppose we want to download a webpage as a string.  There is a new method added to WebClient: Task<string> WebClient.DownloadStringTaskAsync(Uri).  Since this returns a Task<string> we can use it within an asynchronous function.  Suppose, for example, that we wanted to do something similar to my asynchronous Task example – download a web page asynchronously and check to see if it supports XHTML 1.0, then report this into a TextBox.  This could be done like so: private async void button1_Click(object sender, RoutedEventArgs e) { string url = "http://reedcopsey.com"; string content = await new WebClient().DownloadStringTaskAsync(url); this.textBox1.Text = string.Format("Page {0} supports XHTML 1.0: {1}", url, content.Contains("XHTML 1.0")); } .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; } Let’s walk through what’s happening here, step by step.  By adding the async contextual keyword to the method definition, we are able to use the await keyword on our WebClient.DownloadStringTaskAsync method call. When the user clicks this button, the new method (Task<string> WebClient.DownloadStringTaskAsync(string)) is called, which returns a Task<string>.  By adding the await keyword, the runtime will call this method that returns Task<string>, and execution will return to the caller at this point.  This means that our UI is not blocked while the webpage is downloaded.  Instead, the UI thread will “await” at this point, and let the WebClient do it’s thing asynchronously. When the WebClient finishes downloading the string, the user interface’s synchronization context will automatically be used to “pick up” where it left off, and the Task<string> returned from DownloadStringTaskAsync is automatically unwrapped and set into the content variable.  At this point, we can use that and set our text box content. There are a couple of key points here: Asynchronous functions are declared with the async keyword, and contain one or more await expressions In addition to the obvious benefits of shorter, simpler code – there are some subtle but tremendous benefits in this approach.  When the execution of this asynchronous function continues after the first await statement, the initial synchronization context is used to continue the execution of this function.  That means that we don’t have to explicitly marshal the call that sets textbox1.Text back to the UI thread – it’s handled automatically by the language and framework!  Exception handling around asynchronous method calls also just works. I’d recommend every C# developer take a look at the documentation on the new Asynchronous Programming for C# and Visual Basic page, download the Visual Studio Async CTP, and try it out.

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  • SQL SERVER – Database Dynamic Caching by Automatic SQL Server Performance Acceleration

    - by pinaldave
    My second look at SafePeak’s new version (2.1) revealed to me few additional interesting features. For those of you who hadn’t read my previous reviews SafePeak and not familiar with it, here is a quick brief: SafePeak is in business of accelerating performance of SQL Server applications, as well as their scalability, without making code changes to the applications or to the databases. SafePeak performs database dynamic caching, by caching in memory result sets of queries and stored procedures while keeping all those cache correct and up to date. Cached queries are retrieved from the SafePeak RAM in microsecond speed and not send to the SQL Server. The application gets much faster results (100-500 micro seconds), the load on the SQL Server is reduced (less CPU and IO) and the application or the infrastructure gets better scalability. SafePeak solution is hosted either within your cloud servers, hosted servers or your enterprise servers, as part of the application architecture. Connection of the application is done via change of connection strings or adding reroute line in the c:\windows\system32\drivers\etc\hosts file on all application servers. For those who would like to learn more on SafePeak architecture and how it works, I suggest to read this vendor’s webpage: SafePeak Architecture. More interesting new features in SafePeak 2.1 In my previous review of SafePeak new I covered the first 4 things I noticed in the new SafePeak (check out my article “SQLAuthority News – SafePeak Releases a Major Update: SafePeak version 2.1 for SQL Server Performance Acceleration”): Cache setup and fine-tuning – a critical part for getting good caching results Database templates Choosing which database to cache Monitoring and analysis options by SafePeak Since then I had a chance to play with SafePeak some more and here is what I found. 5. Analysis of SQL Performance (present and history): In SafePeak v.2.1 the tools for understanding of performance became more comprehensive. Every 15 minutes SafePeak creates and updates various performance statistics. Each query (or a procedure execute) that arrives to SafePeak gets a SQL pattern, and after it is used again there are statistics for such pattern. An important part of this product is that it understands the dependencies of every pattern (list of tables, views, user defined functions and procs). From this understanding SafePeak creates important analysis information on performance of every object: response time from the database, response time from SafePeak cache, average response time, percent of traffic and break down of behavior. One of the interesting things this behavior column shows is how often the object is actually pdated. The break down analysis allows knowing the above information for: queries and procedures, tables, views, databases and even instances level. The data is show now on all arriving queries, both read queries (that can be cached), but also any types of updates like DMLs, DDLs, DCLs, and even session settings queries. The stats are being updated every 15 minutes and SafePeak dashboard allows going back in time and investigating what happened within any time frame. 6. Logon trigger, for making sure nothing corrupts SafePeak cache data If you have an application with many parts, many servers many possible locations that can actually update the database, or the SQL Server is accessible to many DBAs or software engineers, each can access some database directly and do some changes without going thru SafePeak – this can create a potential corruption of the data stored in SafePeak cache. To make sure SafePeak cache is correct it needs to get all updates to arrive to SafePeak, and if a DBA will access the database directly and do some changes, for example, then SafePeak will simply not know about it and will not clean SafePeak cache. In the new version, SafePeak brought a new feature called “Logon Trigger” to solve the above challenge. By special click of a button SafePeak can deploy a special server logon trigger (with a CLR object) on your SQL Server that actually monitors all connections and informs SafePeak on any connection that is coming not from SafePeak. In SafePeak dashboard there is an interface that allows to control which logins can be ignored based on login names and IPs, while the rest will invoke cache cleanup of SafePeak and actually locks SafePeak cache until this connection will not be closed. Important to note, that this does not interrupt any logins, only informs SafePeak on such connection. On the Dashboard screen in SafePeak you will be able to see those connections and then decide what to do with them. Configuration of this feature in SafePeak dashboard can be done here: Settings -> SQL instances management -> click on instance -> Logon Trigger tab. Other features: 7. User management ability to grant permissions to someone without changing its configuration and only use SafePeak as performance analysis tool. 8. Better reports for analysis of performance using 15 minute resolution charts. 9. Caching of client cursors 10. Support for IPv6 Summary SafePeak is a great SQL Server performance acceleration solution for users who want immediate results for sites with performance, scalability and peak spikes challenges. Especially if your apps are packaged or 3rd party, since no code changes are done. SafePeak can significantly increase response times, by reducing network roundtrip to the database, decreasing CPU resource usage, eliminating I/O and storage access. SafePeak team provides a free fully functional trial www.safepeak.com/download and actually provides a one-on-one assistance during such trial. Reference: Pinal Dave (http://blog.SQLAuthority.com) Filed under: About Me, Pinal Dave, PostADay, SQL, SQL Authority, SQL Performance, SQL Query, SQL Server, SQL Tips and Tricks, SQL Utility, T SQL, Technology

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  • MySQL 5.5 brings in new ways to authenticate users

    - by Georgi Kodinov
    Ever wanted to use your server's OS for authenticating MySQL users ? Or the corporate LDAP repository ? Unfortunately options like the above are plentiful nowadays. And providing hard-coded support for protocol X or service Y is not the best possible idea. MySQL 5.5 has taken the step into the right direction by providing an infrastructure allowing one to make the server understand different authentication protocols by creating a set of simple plugins (one for the client and one for the server). So now you can easily extend MySQL to search for and authenticate users in your favorite user directory. In fact the API supplied is so versatile that we took the possibility to re-design the current "native" authentication mechanism into a built-in always-on plugin ! OK, let me give you an example: Imagine we have a bunch of users defined in your OS, e.g. we have a user joro with his respective password. And we have a MySQL instance running on the same computer. It would not be unexpected to need to let joro access and/or modify MySQL data. The first step is to define him as a MySQL user. And there's a problem right there : MySQL's CREATE USER joro@localhost IDENTIFIED BY 'joros_password' statement needs a password. And this is a password in no way related to the password that joro have set up in the OS. What's worse : if joro changes his OS password this will in no way be reflected in MySQL. So he'll need to change his MySQL password in a separate step. Not very convenient, specially when you have a lot of users. This is a laborious setup for joro's DBA as well : he'll have to disable his access in both MySQL and the OS should he decides that joro's out of the "nice" list. Now mysql 5.5 to the rescue: Imagine that the smart DBA has created a MySQL server plugin that will check if the name of the user logging in is a valid and enabled OS name and if the password supplied to the mysql client matches the OS and has called this plugin 'auth_os'. Now all that's left to do is to define joro as a MySQL user that will be authenticated externally. This is done by the following command : CREATE USER 'joro'@'localhost' IDENTIFIED WITH 'auth_os'; Now joro can login to MySQL using his current OS password. Note : joro is still a valid MySQL user, so you can grant privileges to him just like you would for all other users. What's better: you can have users that authenticate using different mechanisms in the same server. So you can e.g. safely experiment with external authentication for selected users while keeping your current user base operational. What happens under the hood when joro logs in ? The server will find out by the user definition that it needs to use a non-default authentication and will ask the client to "switch" to using the appropriate client-side plugin (if of course the client is not already using it). If the client can't do this (e.g. because it's an old client or doesn't have the necessary plugin available) the server will reject the login. Otherwise the server will let the server-side plugin decide (while possibly talking to the client side plugin and the OS user directory) if this is a valid login or not. If it is the login process will continue as usual, while if it's not the login will get rejected. There's a lot more that MySQL 5.5 can do for you than just the simple case above. Stay tuned for more advanced use cases like mapping groups of external users to a single MySQL user (so you won't have to have 1-to-1 mapping between your external user directory and your mysql user repository) or ways to control the process as a DBA. Or you can simply skip ahead and read the relevant topics from MySQL's excellent online documentation. Or take a look at the example plugins in plugin/auth. Or take a look at the test suite in mysql-test/t/plugin_auth.test. Changelog entry: http://dev.mysql.com/doc/refman/5.5/en/news-5-5-7.html Primary new sections: Pluggable authentication Proxy users Client plugin C API functions Revised sections: New PROXY privilege New proxies_priv grant table Passwords might be external New external_user and proxy_user system variables New --default-auth and --plugin-dir mysql options New MYSQL_DEFAULT_AUTH and MYSQL_PLUGIN_DIR options for mysql_options() CREATE USER has IDENTIFIED WITH clause to specify auth plugin GRANT has PROXY privilege, IDENTIFIED WITH clause to specify auth plugin The data structure for writing client plugins

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