Search Results

Search found 8268 results on 331 pages for 'difference'.

Page 160/331 | < Previous Page | 156 157 158 159 160 161 162 163 164 165 166 167 | Next Page >

*this vs this in C++

- by m00st

I understand that this does but what is the difference between *this and this? Yes I have Googled and read over *this in my text book; but I just don't get it...

Read the article
[ASP.NET] <%# %> vs <%= %>

- by Bart

Hello I am wondering what is the difference between <%# ... % and <%= ... % and <%$ ... % ? I couldn't find anything abt it, cause it's impossible to find <%= in search engine;/ Does these constructions have name to look for it? Could you explain pls? Cheers

Read the article
MySQL Wildcards * and %

- by RussellDias

What is the difference between % and * wildcards in MySQL? In a query like so : "SELECT * FROM $table WHERE MATCH (message) AGAINST('$string*' IN BOOLEAN MODE)"

Read the article
Long pause when accessing DFS namespace

- by Matt

We've recently migrated our Windows network to use DFS for shared files. DFS is working well, except for one annoying problem: users experience a significant delay when they try to access a DFS namespace that they have not accessed for some time. I have tried to troubleshoot the issue but have not had any success so far, and I was hoping someone here may have some pointers to help resolve the problem. Firstly, some background on our network: The network uses a Windows 2008 functional level Active Directory domain with two Windows 2008 DCs and two DNS servers (one on each of the DCs). The network is DNS only - no WINS. All computers are located at the same site and connected by Gigabit Ethernet. We have approximately 20 Domain-based DFS namespaces in Windows 2008 mode, and each DFS namespace has two Windows 2008 DFS namespace servers (the same two servers for all namespaces). All namespace servers are in FQDN mode and all folder targets are specified using their FQDN. All computers are up-to-date with Service Packs and patches. The actual folder targets (i.e. the SMB shares our DFS folders point to) are scattered across several file and application servers, all running Windows 2008 bar two application servers which run Windows 2003 R2, with no replication setup at all (e.g. all DFS folders currently only have one folder target). Some more detail on the problem: The namespace access delay is generally 1 - 10 seconds long and seems to occur when a particular computer has not accessed the requested namespace for approximately five minutes or more. For example, if the user has not accessed \\domain.name\namespace1\ for more than five minutes and attempts to access \\domain.name\namespace1\ via Windows Explorer, the Explorer window will freeze for 1 - 10 seconds before finally resuming and displaying the folders that exist in \\domain.name\namespace1. If they then close the Explorer window and attempt to access \\domain.name\namespace1\ again within five minutes the contents will be displayed almost instantly - if they wait longer than five minutes it will go through the 1 - 10 second pause again. Once "inside" the namespace everything is nice and snappy, it's just the initial connection to the namespace that is slow. The browsing delays seem to affect all variants of Windows that we use (Windows 2008 x64 SP2, Windows 2003 R2 x86 SP2, Windows XP Pro x86 SP3) - it is possibly a bit worse in Windows XP / 2003 than in Windows 2008, but I'm not sure if the difference isn't just psychological. Accessing the underlying folder targets directly exhibits no delay at all - i.e. if the SMB shares pointed to by DFS are accessed directly (bypassing DFS) then there is no pause. During trouble-shooting I noticed that the "Cache duration" for all of our DFS roots is set to 300 seconds - 5 minutes. Given that this is the same amount of time required to trigger the pause I assume that this caching is somehow related, although I am unsure exactly what is cached on the client and hence what needs to be looked up again after 5 minutes have elapsed. In trying to resolve the problem I have already tried / checked the following (without success): Run dcdiag on both Domain Controllers - no problems found Done some basic DNS server checks without finding any problems - I don't know how to check the DNS servers in detail, but I would add that the network is not exhibiting any other strange behavior that may point to a DNS problem Disabled Anti-virus on clients and servers Removing one of the namespace servers from a couple of namespaces - no difference So that's where I'm up to - and I'm out of ideas. Can anyone suggest what may be causing the delays and/or what I should be trying next?

Read the article
Nginx + PHP-FPM = "Random" 502 Bad Gateway

- by david

I am running Nginx and proxying php requests via FastCGI to PHP-FPM for processing. I will randomly receive 502 Bad Gateway error pages - I can reproduce this issue by clicking around my PHP websites very rapidly/refreshing a page for a minute or two. When I get the 502 error page all I have to do is refresh the browser and the page refreshes properly. Here is my setup: nginx/0.7.64 PHP 5.3.2 (fpm-fcgi) (built: Apr 1 2010 06:42:04) Ubuntu 9.10 (Latest 2.6 Paravirt) I compiled PHP-FPM using this ./configure directive ./configure --enable-fpm --sysconfdir=/etc/php5/conf.d --with-config-file-path=/etc/php5/conf.d/php.ini --with-zlib --with-openssl --enable-zip --enable-exif --enable-ftp --enable-mbstring --enable-mbregex --enable-soap --enable-sockets --disable-cgi --with-curl --with-curlwrappers --with-gd --with-mcrypt --enable-memcache --with-mhash --with-jpeg-dir=/usr/local/lib --with-mysql=/usr/bin/mysql --with-mysqli=/usr/bin/mysql_config --enable-pdo --with-pdo-mysql=/usr/bin/mysql --with-pdo-sqlite --with-pspell --with-snmp --with-sqlite --with-tidy --with-xmlrpc --with-xsl My php-fpm.conf looks like this (the relevant parts): ... <value name="pm"> <value name="max_children">3</value> ... <value name="request_terminate_timeout">60s</value> <value name="request_slowlog_timeout">30s</value> <value name="slowlog">/var/log/php-fpm.log.slow</value> <value name="rlimit_files">1024</value> <value name="rlimit_core">0</value> <value name="chroot"></value> <value name="chdir"></value> <value name="catch_workers_output">yes</value> <value name="max_requests">500</value> ... I've tried increasing the max_children to 10 and it makes no difference. I've also tried setting it to 'dynamic' and setting max_children to 50, and start_server to '5' without any difference. I have tried using both 1 and 5 nginx worker processes. My fastcgi_params conf looks like: 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 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; fastcgi_param REDIRECT_STATUS 200; Nginx logs the error as: [error] 3947#0: *10530 connect() failed (111: Connection refused) while connecting to upstream, client: 68.40.xxx.xxx, server: www.domain.com, request: "GET /favicon.ico HTTP/1.1", upstream: "fastcgi://127.0.0.1:9000", host: "www.domain.com" PHP-FPM logs the follow at the time of the error: [NOTICE] pid 17161, fpm_unix_init_main(), line 255: getrlimit(nofile): max:1024, cur:1024 [NOTICE] pid 17161, fpm_event_init_main(), line 93: libevent: using epoll [NOTICE] pid 17161, fpm_init(), line 50: fpm is running, pid 17161 [DEBUG] pid 17161, fpm_children_make(), line 403: [pool default] child 17162 started [DEBUG] pid 17161, fpm_children_make(), line 403: [pool default] child 17163 started [DEBUG] pid 17161, fpm_children_make(), line 403: [pool default] child 17164 started [NOTICE] pid 17161, fpm_event_loop(), line 111: ready to handle connections My CPU usage maxes out around 10-15% when I recreate the issue. My Free mem (free -m) is 130MB I had this intermittent 502 Bad Gateway issue when in was using php5-cgi to service my php requests as well. Does anyone know how to fix this?

Read the article
Apache cyclic redirection problem

- by slicedlime

I have an extremely weird problem with one of my sites. I run a number of blogs off a single apache2 server with a shared wordpress install. Each site has a www.domain.com main domain, but a ServerAlias of domain.com. This works fine for all the blogs except one, which instead of redirecting to www.domain.com redirects to domain.com, causing a cyclic redirection. The configuration for each host looks like this: <VirtualHost *:80> ServerName www.domain.com ServerAlias domain.com DocumentRoot "/home/www/www.domain.com" <Directory "/home/www/www.domain.com"> Options MultiViews Indexes Includes FollowSymLinks ExecCGI AllowOverride All Order allow,deny Allow from all </Directory> </VirtualHost> As this didn't work, I tried a mod_rewrite rule for it, which still didn't redirect correctly. The weird thing here is that if i rewrite it to redirect to any other domain it will redirect correctly, even to another subdomain. So a rewrite rule for domain.com that redirects to foo.domain.com works, but not to www.domain.com. In the same way, trying to redirec to www.domain.com/foo/ ends me up with a redirection to domain.com/foo/. Even weirder, I tried setting up domain.com as a completely separate virtual host, and ran this php test script as index.php on it: <?php header('Location: http://www.domain.com/' . $_SERVER["request_uri"]); ?> Hitting domain.com still redirects to domain.com! Checking out the headers sent to the server verifies that I get exactly the redirect URL I wanted, except with the "www." stripped. The same script works like a charm if I replace www. with foo or redirect to any other domain. This has now foiled me for a long time. I've diffed the vhosts configs for a working domain and the faulty one, and the only difference is the domain name itself. I've diffed the .htaccess files for both sites, and the only difference is a path related to the sharing of wordpress installation for the blogs: php_value include_path ".:/home/www/www.domain.com/local/:/home/www/www.domain.com/" I searched through everything in /etc (including apache conf) for the domain name of the faulty host and found nothing weird, searched through everything in /etc for one of the working ones to make sure it didn't differ, I even went so far to check on the DNS setup of two domains to make sure there wasn't anything strange going on. Here's the response for the faulty one: user@localhost dir $ wget -S domain.com --2010-03-20 21:47:24-- http://domain.com/ Resolving domain.com... x.x.x.x Connecting to domain.com|x.x.x.x|:80... connected. HTTP request sent, awaiting response... HTTP/1.1 301 Moved Permanently Via: 1.1 ISA Connection: Keep-Alive Proxy-Connection: Keep-Alive Content-Length: 0 Date: Sat, 20 Mar 2010 20:47:24 GMT Location: http://domain.com/ Content-Type: text/html; charset=UTF-8 Server: Apache X-Powered-By: PHP/5.2.10-pl0-gentoo X-Pingback: http://domain.com/xmlrpc.php Keep-Alive: timeout=15, max=100 Location: http://domain.com/ [following] And a working one: user@localhost dir $ wget -S domain.com --2010-03-20 21:51:33-- http://domain.com/ Resolving domain.com... x.x.x.x Connecting to domain.com|x.x.x.x|:80... connected. HTTP request sent, awaiting response... HTTP/1.1 301 Moved Permanently Via: 1.1 ISA Connection: Keep-Alive Proxy-Connection: Keep-Alive Content-Length: 0 Date: Sat, 20 Mar 2010 20:51:33 GMT Location: http://www.domain.com/ Content-Type: text/html; charset=UTF-8 Server: Apache X-Powered-By: PHP/5.2.10-pl0-gentoo X-Pingback: http://www.domain.com/xmlrpc.php Keep-Alive: timeout=15, max=100 Location: http://www.domain.com/ [following] I'm stumped. I've had this problem for a long time, and I feel like I've tried everything. I can't see why the different domains would act differently under the same installation with the same settings. Help :(

Read the article
Intelligent "Subtraction" of one text logfile from another

- by Vi

Example: Application generates large text log file A with many different messages. It generates similarly large log file B when does not function correctly. I want to see what messages in file B are essentially new, i.e. to filter-out everything from A. Trivial prototype is: Sort | uniq both files Join files sort | uniq -c grep -v "^2" This produces symmetric difference and inconvenient. How to do it better? (including non-symmetric difference and preserving of messages order in B) Program should first analyse A and learn which messages are common, then analyse B showing with messages needs attention. Ideally it should automatically disregard things like timestamps, line numbers or other volatile things. Example. A: 0:00:00.234 Received buffer 0x324234 0:00:00.237 Processeed buffer 0x324234 0:00:00.238 Send buffer 0x324255 0:00:03.334 Received buffer 0x324255 0:00:03.337 Processeed buffer 0x324255 0:00:03.339 Send buffer 0x324255 0:00:05.171 Received buffer 0x32421A 0:00:05.173 Processeed buffer 0x32421A 0:00:05.178 Send buffer 0x32421A B: 0:00:00.134 Received buffer 0x324111 0:00:00.137 Processeed buffer 0x324111 0:00:00.138 Send buffer 0x324111 0:00:03.334 Received buffer 0x324222 0:00:03.337 Processeed buffer 0x324222 0:00:03.338 Error processing buffer 0x324222 0:00:03.339 Send buffer 0x3242222 0:00:05.271 Received buffer 0x3242FA 0:00:05.273 Processeed buffer 0x3242FA 0:00:05.278 Send buffer 0x3242FA 0:00:07.280 Send buffer 0x3242FA failed Result: 0:00:03.338 Error processing buffer 0x324222 0:00:07.280 Send buffer 0x3242FA failed One of ways of solving it can be something like that: Split each line to logical units: 0:00:00.134 Received buffer 0x324111,0:00:00.134,Received,buffer,0x324111,324111,Received buffer, \d:\d\d:\d\d\.\d\d\d, \d+:\d+:\d+.\d+, 0x[0-9A-F]{6}, ... It should find individual words, simple patterns in numbers, common layouts (e.g. "some date than text than number than text than end_of_line"), also handle combinations of above. As it is not easy task, user assistance (adding regexes with explicit "disregard that","make the main factor","don't split to parts","consider as date/number","take care of order/quantity of such messages" rules) should be supported (but not required) for it. Find recurring units and "categorize" lines, filter out too volatile things like timestamps, addresses or line numbers. Analyse the second file, find things that has new logical units (one-time or recurring), or anything that will "amaze" the system which has got used to the first file. Example of doing some bit of this manually: $ cat A | head -n 1 0:00:00.234 Received buffer 0x324234 $ cat A | egrep -v "Received buffer" | head -n 1 0:00:00.237 Processeed buffer 0x324234 $ cat A | egrep -v "Received buffer|Processeed buffer" | head -n 1 0:00:00.238 Send buffer 0x324255 $ cat A | egrep -v "Received buffer|Processeed buffer|Send buffer" | head -n 1 $ cat B | egrep -v "Received buffer|Processeed buffer|Send buffer" 0:00:03.338 Error processing buffer 0x324222 0:00:07.280 Send buffer 0x3242FA failed This is a boring thing (there are a lot of message types); also I can accidentally include some too broad pattern. Also it can't handle complicated things like interrelation between messages. I know that it is AI-related. May be there are already developed tools?

Read the article
Troubleshooting Website problems within the local network

- by HaydnWVN

Have an external website which opens fine on some PC's, yet seems to time out (or symptoms of timing out, but never actually does) on others. Seems to only affect (some) of our newer HP Pro 3305 MT Workstations. All of which are running Win7 32bit SP1 with all updates. Older PC's (Win7 32bit SP1 & WinXP) are unaffected. Using Google Chrome & Firefox makes no difference. Opening the website in IE9 Compatibility Mode has exactly the same symptoms. All PC's are on the same local network (Workgroup) using the same DNS server & gateway (inhouse) on the same internet connection, on the same subnet. There is no proxy server, no content filtering, no load balancing etc etc. Only group policy in effect (locally) is for Update scheduling. Local firewalls are all the same (Kaspersky WP4) and our external facing firewall has no IP specific settings. I have no control over the external website, traceroute shows the same destination on all PC's. It is a fairly popular website in our industry (Horticulture) and i'm not aware of any other people (even other sites within our sister companies) with the same problem. Update: Used Fiddler2 to monitor the HTTP request, seems its not getting fulfilled for some reason?! Request sent: GET http://www.rhs.org.uk/ HTTP/1.1 Host: www.rhs.org.uk Connection: keep-alive User-Agent: Mozilla/5.0 (Windows NT 6.1) AppleWebKit/536.11 (KHTML, like Gecko) Chrome/20.0.1132.47 Safari/536.11 Accept: text/html,application/xhtml+xml,application/xml;q=0.9,*/*;q=0.8 Accept-Encoding: gzip,deflate,sdch Accept-Language: en-GB,en-US;q=0.8,en;q=0.6 Accept-Charset: ISO-8859-1,utf-8;q=0.7,*;q=0.3 Log from Fiddler 2 of the request: This session is not yet complete. Press F5 to refresh when session is complete for updated statistics. Request Count: 1 Bytes Sent: 567 (headers:567; body:0) Bytes Received: 0 (headers:0; body:0) ACTUAL PERFORMANCE -------------- ClientConnected: 17:02:33.720 ClientBeginRequest: 17:02:39.118 GotRequestHeaders: 17:02:39.118 ClientDoneRequest: 17:02:39.118 Determine Gateway: 0ms DNS Lookup: 0ms TCP/IP Connect: 46ms HTTPS Handshake: 0ms ServerConnected: 17:02:39.165 FiddlerBeginRequest: 17:02:39.165 ServerGotRequest: 17:02:39.165 ServerBeginResponse: 00:00:00.000 GotResponseHeaders: 00:00:00.000 ServerDoneResponse: 00:00:00.000 ClientBeginResponse: 00:00:00.000 ClientDoneResponse: 00:00:00.000 RESPONSE BYTES (by Content-Type) -------------- ~headers~: 0 Log of a successful request from a working PC (done this morning, excuse the timestamps being different from above): Request Count: 1 Bytes Sent: 493 (headers:493; body:0) Bytes Received: 20,413 (headers:525; body:19,888) ACTUAL PERFORMANCE -------------- ClientConnected: 08:22:47.766 ClientBeginRequest: 08:22:47.766 GotRequestHeaders: 08:22:47.766 ClientDoneRequest: 08:22:47.766 Determine Gateway: 0ms DNS Lookup: 26ms TCP/IP Connect: 30ms HTTPS Handshake: 0ms ServerConnected: 08:22:47.828 FiddlerBeginRequest: 08:22:47.828 ServerGotRequest: 08:22:47.828 ServerBeginResponse: 08:22:48.905 GotResponseHeaders: 08:22:48.905 ServerDoneResponse: 08:22:48.905 ClientBeginResponse: 08:22:48.905 ClientDoneResponse: 08:22:48.905 Overall Elapsed: 00:00:01.1388020 RESPONSE BYTES (by Content-Type) -------------- text/html: 19,888 ~headers~: 525 So my question has evolved into: What is the difference between the 2 requests and how do I determine why 1 PC is not getting a reply to it's GET request?

Read the article
Ubuntu server loses exactly 5 minutes once in a while

- by Harold Smith

I noticed that my server, an Ubuntu server 12.04, was losing time. I figured the hardware clock was off or maybe dying due to a faulty CMOS battery. I installed NTP to ensure the drift would be corrected, but to no avail. During a day it would lose 20 minutes or so. To debug, I created a small cron job to check against a remote servers time, which I knew to be correct. The script calculates the difference in seconds between local and remote time. The result was interesting. It seems to be losing exactly 5 minutes several times during the day. Look at this log (difference from remote server noted in seconds): Tue Oct 23 03:30:02 CEST 2012: 284 Tue Oct 23 03:35:02 CEST 2012: 284 Tue Oct 23 03:40:01 CEST 2012: 285 Tue Oct 23 03:45:02 CEST 2012: 285 Tue Oct 23 03:50:02 CEST 2012: 285 Tue Oct 23 03:55:02 CEST 2012: 284 Tue Oct 23 04:00:02 CEST 2012: 284 Tue Oct 23 04:05:01 CEST 2012: 285 Tue Oct 23 04:10:01 CEST 2012: 285 Tue Oct 23 04:15:02 CEST 2012: 585 Tue Oct 23 04:20:02 CEST 2012: 584 Tue Oct 23 04:25:02 CEST 2012: 584 Tue Oct 23 04:30:02 CEST 2012: 584 Tue Oct 23 04:35:01 CEST 2012: 585 Tue Oct 23 04:40:01 CEST 2012: 585 Tue Oct 23 04:45:02 CEST 2012: 585 Tue Oct 23 04:50:02 CEST 2012: 584 Tue Oct 23 04:55:02 CEST 2012: 584 Tue Oct 23 05:00:02 CEST 2012: 584 Tue Oct 23 05:05:01 CEST 2012: 585 Tue Oct 23 05:10:01 CEST 2012: 585 Tue Oct 23 05:15:02 CEST 2012: 585 Tue Oct 23 05:20:02 CEST 2012: 584 Tue Oct 23 05:25:02 CEST 2012: 584 Tue Oct 23 05:30:02 CEST 2012: 584 Tue Oct 23 05:35:01 CEST 2012: 585 Tue Oct 23 05:40:01 CEST 2012: 585 Tue Oct 23 05:45:02 CEST 2012: 584 Tue Oct 23 05:50:02 CEST 2012: 584 Tue Oct 23 05:55:02 CEST 2012: 584 Tue Oct 23 06:00:02 CEST 2012: 584 Tue Oct 23 06:05:03 CEST 2012: 584 Tue Oct 23 06:10:02 CEST 2012: 584 Tue Oct 23 06:15:01 CEST 2012: 585 Tue Oct 23 06:20:02 CEST 2012: 584 Tue Oct 23 06:25:02 CEST 2012: 584 Tue Oct 23 06:30:02 CEST 2012: 584 Tue Oct 23 06:35:02 CEST 2012: 584 Tue Oct 23 06:40:02 CEST 2012: 584 Tue Oct 23 06:45:01 CEST 2012: 585 Tue Oct 23 06:50:02 CEST 2012: 584 Tue Oct 23 06:55:01 CEST 2012: 585 Tue Oct 23 07:00:02 CEST 2012: 584 Tue Oct 23 07:05:02 CEST 2012: 584 Tue Oct 23 07:10:02 CEST 2012: 584 Tue Oct 23 07:15:02 CEST 2012: 584 Tue Oct 23 07:20:02 CEST 2012: 584 Tue Oct 23 07:25:02 CEST 2012: 584 Tue Oct 23 07:30:01 CEST 2012: 585 Tue Oct 23 07:35:02 CEST 2012: 584 Tue Oct 23 07:40:02 CEST 2012: 584 Tue Oct 23 07:45:02 CEST 2012: 584 Tue Oct 23 07:50:02 CEST 2012: 584 Tue Oct 23 07:55:02 CEST 2012: 584 Tue Oct 23 08:00:01 CEST 2012: 585 Tue Oct 23 08:05:02 CEST 2012: 584 Tue Oct 23 08:10:02 CEST 2012: 584 Tue Oct 23 08:15:02 CEST 2012: 584 Tue Oct 23 08:20:02 CEST 2012: 584 Tue Oct 23 08:25:02 CEST 2012: 584 Tue Oct 23 08:30:01 CEST 2012: 585 Tue Oct 23 08:35:02 CEST 2012: 584 Tue Oct 23 08:40:02 CEST 2012: 584 Tue Oct 23 08:45:02 CEST 2012: 584 Tue Oct 23 08:50:02 CEST 2012: 584 Tue Oct 23 08:55:02 CEST 2012: 584 Tue Oct 23 09:00:02 CEST 2012: 584 Tue Oct 23 09:05:03 CEST 2012: 584 Tue Oct 23 09:10:02 CEST 2012: 584 Tue Oct 23 09:15:02 CEST 2012: 584 Tue Oct 23 09:20:02 CEST 2012: 584 Tue Oct 23 09:25:02 CEST 2012: 584 Tue Oct 23 09:30:01 CEST 2012: 584 Tue Oct 23 09:35:02 CEST 2012: 584 Tue Oct 23 09:40:02 CEST 2012: 584 Tue Oct 23 09:45:02 CEST 2012: 584 Tue Oct 23 09:50:02 CEST 2012: 584 Tue Oct 23 09:55:02 CEST 2012: 584 Tue Oct 23 10:00:01 CEST 2012: 584 Tue Oct 23 10:05:02 CEST 2012: 584 Tue Oct 23 10:10:07 CEST 2012: 584 Tue Oct 23 10:15:02 CEST 2012: 584 Tue Oct 23 10:20:02 CEST 2012: 884 Tue Oct 23 10:25:02 CEST 2012: 884 Tue Oct 23 10:30:02 CEST 2012: 883 Tue Oct 23 10:35:01 CEST 2012: 884 Tue Oct 23 10:40:02 CEST 2012: 884 Tue Oct 23 10:45:02 CEST 2012: 884 Tue Oct 23 10:50:02 CEST 2012: 884 Tue Oct 23 10:55:02 CEST 2012: 1184 Tue Oct 23 11:00:02 CEST 2012: 1183 Tue Oct 23 11:05:01 CEST 2012: 1184 Tue Oct 23 11:10:02 CEST 2012: 1184 Tue Oct 23 11:15:02 CEST 2012: 1184 Tue Oct 23 11:20:02 CEST 2012: 1184 This does not seem to be faulty CMOS battery in my opinion. But what do you think?

Read the article
Installing .NET Framework 4 Client Profile breaks Windows Update

- by Richard

I have a Samsung NC-10 netbook with a fresh install of Windows 7 Home Premium 32-bit (it only had 2GB). If Microsoft .NET Framework 4 Client Profile is installed on it, Windows Update will always return error code 8024402F ("Windows Update encountered an unknown error"). As soon as I uninstall it, Windows Update works just fine again. Out of the four computers in this house, only this netbook has the problem. My question is: How can I get the .NET Framework 4 Client Profile installed on my netbook and continue to have a functioning Windows Update? ----- More information ----- The hard-drive recently died on my netbook so I replaced it with a nice new SSD and did a fresh installation of Windows 7 Home Premium (SP1) - along with all the updates. At some point I found that, when I ran Windows Update, I was greeted with error code 8024402F ("Windows Update encountered an unknown error"). Looking in C:\Windows\WindowsUpdate.log, I saw the following issue: WARNING: ECP: Failed to validate cab file digest downloaded from http://download.windowsupdate.com/msdownload/update/software/dflt/2012/02/4913552_4a5c9563d1f58c77f30d0d5c9999e4b8bff3ab21.cab with error 0x80091007 WARNING: ECP: This roundtrip contained some optimized updates which failed. New Update count = 0, Old Count = 3 FATAL: ProcessCoreMetadata did not return any update to be committed WARNING: Sync of Updates: 0x8024402f WARNING: SyncServerUpdatesInternal failed: 0x8024402f When I downloaded the CAB from the URL listed and opened it, it contained a file called 4913552.txt. A search on Google suggested that it's related to Microsoft .NET Framework 4 Client Profile. Other people had reported problems with it breaking Windows Update, but they were running Windows XP. I tried the steps outlined on the Microsoft site for this error code, but it reported that there was nothing wrong. I also found this superuser question, I tried all the answers listed but none of them made any difference. My router doesn't block ActiveX, changing my internet settings in IE made no difference, assuming it was a corrupted update repository didn't do anything (except wipe my update history), my date and time was correct, switching to Google's DNS didn't work and neither did disabling IPv6. Figuring that this update was corrupted, I repaired it and nothing changed. In desperation I un-installed it and Windows Update started working again! Brilliant! I then downloaded the full version from the Microsoft website, installed it and, thankfully, Windows Update continued to work just fine. A week later I turn on my netbook and Windows Update is broken again with exactly the same error message and log entries as before. Repairing .NET Framework 4 Client Profile did nothing, removing it entirely solved the problem again. Thinking this might be some odd Windows installation issue, I formatted the hard-drive and re-installed Windows. Same problem as before - as soon as .NET Framework 4 Client Profile ended up on the netbook, Windows Update stopped working and reported error 8024402F. As soon as it was un-installed, everything worked just fine again. There are three other machines in this house and all of them have working Windows Update and this Client Profile. Does anyone know why it causes this netbook to break and, more importantly, how I can fix it?

Read the article
Box2D Joints in entity components system

- by Johnmph

I search a way to have Box2D joints in an entity component system, here is what i found : 1) Having the joints in Box2D/Body component as parameters, we have a joint array with an ID by joint and having in the other body component the same joint ID, like in this example : Entity1 - Box2D/Body component { Body => (body parameters), Joints => { Joint1 => (joint parameters), others joints... } } // Joint ID = Joint1 Entity2 - Box2D/Body component { Body => (body parameters), Joints => { Joint1 => (joint parameters), others joints... } } // Same joint ID than in Entity1 There are 3 problems with this solution : The first problem is the implementation of this solution, we must manage the joints ID to create joints and to know between which bodies they are connected. The second problem is the parameters of joint, where are they got ? on the Entity1 or Entity2 ? If they are the same parameters for the joint, there is no problem but if they are differents ? The third problem is that we can't limit number of bodies to 2 by joint (which is mandatory), a joint can only link 2 bodies, in this solution, nothing prevents to create more than 2 entities with for each a body component with the same joint ID, in this case, how we know the 2 bodies to joint and what to do with others bodies ? 2) Same solution than the first solution but by having entities ID instead of Joint ID, like in this example : Entity1 - Box2D/Body component { Body => (body parameters), Joints => { Entity2 => (joint parameters), others joints... } } Entity2 - Box2D/Body component { Body => (body parameters), Joints => { Entity1 => (joint parameters), others joints... } } With this solution, we fix the first problem of the first solution but we have always the two others problems. 3) Having a Box2D/Joint component which is inserted in the entities which contains the bodies to joint (we share the same joint component between entities with bodies to joint), like in this example : Entity1 - Box2D/Body component { Body => (body parameters) } - Box2D/Joint component { Joint => (Joint parameters) } // Shared, same as in Entity2 Entity2 - Box2D/Body component { Body => (body parameters) } - Box2D/Joint component { Joint => (joint parameters) } // Shared, same as in Entity1 There are 2 problems with this solution : The first problem is the same problem than in solution 1 and 2 : We can't limit number of bodies to 2 by joint (which is mandatory), a joint can only link 2 bodies, in this solution, nothing prevents to create more than 2 entities with for each a body component and the shared joint component, in this case, how we know the 2 bodies to joint and what to do with others bodies ? The second problem is that we can have only one joint by body because entity components system allows to have only one component of same type in an entity. So we can't put two Joint components in the same entity. 4) Having a Box2D/Joint component which is inserted in the entity which contains the first body component to joint and which has an entity ID parameter (this entity contains the second body to joint), like in this example : Entity1 - Box2D/Body component { Body => (body parameters) } - Box2D/Joint component { Entity2 => (Joint parameters) } // Entity2 is the entity ID which contains the other body to joint, the first body being in this entity Entity2 - Box2D/Body component { Body => (body parameters) } There are exactly the same problems that in the third solution, the only difference is that we can have two differents joints by entity instead of one (by putting one joint component in an entity and another joint component in another entity, each joint referencing to the other entity). 5) Having a Box2D/Joint component which take in parameter the two entities ID which contains the bodies to joint, this component can be inserted in any entity, like in this example : Entity1 - Box2D/Body component { Body => (body parameters) } Entity2 - Box2D/Body component { Body => (body parameters) } Entity3 - Box2D/Joint component { Joint => (Body1 => Entity1, Body2 => Entity2, others parameters of joint) } // Entity1 is the ID of the entity which have the first body to joint and Entity2 is the ID of the entity which have the second body to joint (This component can be in any entity, that doesn't matter) With this solution, we fix the problem of the body limitation by joint, we can only have two bodies per joint, which is correct. And we are not limited by number of joints per body, because we can create an another Box2D/Joint component, referencing to Entity1 and Entity2 and put this component in a new entity. The problem of this solution is : What happens if we change the Body1 or Body2 parameter of Joint component at runtime ? We need to add code to sync the Body1/Body2 parameters changes with the real joint object. 6) Same as solution 3 but in a better way : Having a Box2D/Joint component Box2D/Joint which is inserted in the entities which contains the bodies to joint, we share the same joint component between these entities BUT the difference is that we create a new entity to link the body component with the joint component, like in this example : Entity1 - Box2D/Body component { Body => (body parameters) } // Shared, same as in Entity3 Entity2 - Box2D/Body component { Body => (body parameters) } // Shared, same as in Entity4 Entity3 - Box2D/Body component { Body => (body parameters) } // Shared, same as in Entity1 - Box2D/Joint component { Joint => (joint parameters) } // Shared, same as in Entity4 Entity4 - Box2D/Body component { Body => (body parameters) } // Shared, same as in Entity2 - Box2D/Joint component { Joint => (joint parameters) } // Shared, same as in Entity3 With this solution, we fix the second problem of the solution 3, because we can create an Entity5 which will have the shared body component of Entity1 and an another joint component so we are no longer limited in the joint number per body. But the first problem of solution 3 remains, because we can't limit the number of entities which have the shared joint component. To resolve this problem, we can add a way to limit the number of share of a component, so for the Joint component, we limit the number of share to 2, because we can only joint 2 bodies per joint. This solution would be perfect because there is no need to add code to sync changes like in the solution 5 because we are notified by the entity components system when components / entities are added to/removed from the system. But there is a conception problem : How to know easily and quickly between which bodies the joint operates ? Because, there is no way to find easily an entity with a component instance. My question is : Which solution is the best ? Is there any other better solutions ? Sorry for the long text and my bad english.

Read the article
SQL SERVER – Guest Post – Architecting Data Warehouse – Niraj Bhatt

- by pinaldave

Niraj Bhatt works as an Enterprise Architect for a Fortune 500 company and has an innate passion for building / studying software systems. He is a top rated speaker at various technical forums including Tech·Ed, MCT Summit, Developer Summit, and Virtual Tech Days, among others. Having run a successful startup for four years Niraj enjoys working on – IT innovations that can impact an enterprise bottom line, streamlining IT budgets through IT consolidation, architecture and integration of systems, performance tuning, and review of enterprise applications. He has received Microsoft MVP award for ASP.NET, Connected Systems and most recently on Windows Azure. When he is away from his laptop, you will find him taking deep dives in automobiles, pottery, rafting, photography, cooking and financial statements though not necessarily in that order. He is also a manager/speaker at BDOTNET, Asia’s largest .NET user group. Here is the guest post by Niraj Bhatt. As data in your applications grows it’s the database that usually becomes a bottleneck. It’s hard to scale a relational DB and the preferred approach for large scale applications is to create separate databases for writes and reads. These databases are referred as transactional database and reporting database. Though there are tools / techniques which can allow you to create snapshot of your transactional database for reporting purpose, sometimes they don’t quite fit the reporting requirements of an enterprise. These requirements typically are data analytics, effective schema (for an Information worker to self-service herself), historical data, better performance (flat data, no joins) etc. This is where a need for data warehouse or an OLAP system arises. A Key point to remember is a data warehouse is mostly a relational database. It’s built on top of same concepts like Tables, Rows, Columns, Primary keys, Foreign Keys, etc. Before we talk about how data warehouses are typically structured let’s understand key components that can create a data flow between OLTP systems and OLAP systems. There are 3 major areas to it: a) OLTP system should be capable of tracking its changes as all these changes should go back to data warehouse for historical recording. For e.g. if an OLTP transaction moves a customer from silver to gold category, OLTP system needs to ensure that this change is tracked and send to data warehouse for reporting purpose. A report in context could be how many customers divided by geographies moved from sliver to gold category. In data warehouse terminology this process is called Change Data Capture. There are quite a few systems that leverage database triggers to move these changes to corresponding tracking tables. There are also out of box features provided by some databases e.g. SQL Server 2008 offers Change Data Capture and Change Tracking for addressing such requirements. b) After we make the OLTP system capable of tracking its changes we need to provision a batch process that can run periodically and takes these changes from OLTP system and dump them into data warehouse. There are many tools out there that can help you fill this gap – SQL Server Integration Services happens to be one of them. c) So we have an OLTP system that knows how to track its changes, we have jobs that run periodically to move these changes to warehouse. The question though remains is how warehouse will record these changes? This structural change in data warehouse arena is often covered under something called Slowly Changing Dimension (SCD). While we will talk about dimensions in a while, SCD can be applied to pure relational tables too. SCD enables a database structure to capture historical data. This would create multiple records for a given entity in relational database and data warehouses prefer having their own primary key, often known as surrogate key. As I mentioned a data warehouse is just a relational database but industry often attributes a specific schema style to data warehouses. These styles are Star Schema or Snowflake Schema. The motivation behind these styles is to create a flat database structure (as opposed to normalized one), which is easy to understand / use, easy to query and easy to slice / dice. Star schema is a database structure made up of dimensions and facts. Facts are generally the numbers (sales, quantity, etc.) that you want to slice and dice. Fact tables have these numbers and have references (foreign keys) to set of tables that provide context around those facts. E.g. if you have recorded 10,000 USD as sales that number would go in a sales fact table and could have foreign keys attached to it that refers to the sales agent responsible for sale and to time table which contains the dates between which that sale was made. These agent and time tables are called dimensions which provide context to the numbers stored in fact tables. This schema structure of fact being at center surrounded by dimensions is called Star schema. A similar structure with difference of dimension tables being normalized is called a Snowflake schema. This relational structure of facts and dimensions serves as an input for another analysis structure called Cube. Though physically Cube is a special structure supported by commercial databases like SQL Server Analysis Services, logically it’s a multidimensional structure where dimensions define the sides of cube and facts define the content. Facts are often called as Measures inside a cube. Dimensions often tend to form a hierarchy. E.g. Product may be broken into categories and categories in turn to individual items. Category and Items are often referred as Levels and their constituents as Members with their overall structure called as Hierarchy. Measures are rolled up as per dimensional hierarchy. These rolled up measures are called Aggregates. Now this may seem like an overwhelming vocabulary to deal with but don’t worry it will sink in as you start working with Cubes and others. Let’s see few other terms that we would run into while talking about data warehouses. ODS or an Operational Data Store is a frequently misused term. There would be few users in your organization that want to report on most current data and can’t afford to miss a single transaction for their report. Then there is another set of users that typically don’t care how current the data is. Mostly senior level executives who are interesting in trending, mining, forecasting, strategizing, etc. don’t care for that one specific transaction. This is where an ODS can come in handy. ODS can use the same star schema and the OLAP cubes we saw earlier. The only difference is that the data inside an ODS would be short lived, i.e. for few months and ODS would sync with OLTP system every few minutes. Data warehouse can periodically sync with ODS either daily or weekly depending on business drivers. Data marts are another frequently talked about topic in data warehousing. They are subject-specific data warehouse. Data warehouses that try to span over an enterprise are normally too big to scope, build, manage, track, etc. Hence they are often scaled down to something called Data mart that supports a specific segment of business like sales, marketing, or support. Data marts too, are often designed using star schema model discussed earlier. Industry is divided when it comes to use of data marts. Some experts prefer having data marts along with a central data warehouse. Data warehouse here acts as information staging and distribution hub with spokes being data marts connected via data feeds serving summarized data. Others eliminate the need for a centralized data warehouse citing that most users want to report on detailed data. Reference: Pinal Dave (http://blog.SQLAuthority.com) Filed under: Best Practices, Business Intelligence, Data Warehousing, Database, Pinal Dave, PostADay, Readers Contribution, SQL, SQL Authority, SQL Query, SQL Server, SQL Tips and Tricks, T SQL, Technology

Read the article
LightDM will not start after stopping it

- by Sweeters

I am running Ubuntu 11.10 "Oneiric Ocelot", and in trying to install the nvidia CUDA developer drivers I switched to a virtual terminal (Ctrl-Alt-F5) and stopped lightdm (installation required that no X server instance be running) through sudo service lightdm stop. Re-starting lightdm with sudo service lightdm start did not work: A couple of * Starting [...] lines where displayed, but the process hanged. (I do not remember at which point, but I think it was * Starting System V runlevel compatibility. I manually rebooted my laptop, and ever since booting seems to hang, usually around the * Starting anac(h)ronistic cron [OK] log line (not consistently at that point, though). From that point on, I seem to be able to interact with my system only through a tty session (Ctrl-Alt-F1). I've tried purging and reinstalling both lightdm and gdm, as well as selecting both as the default display managers (through sudo dpkg-reconfigure [lightdm / gdm] or by manually editing /etc/X11/default-display-manager) through both apt-get and aptitude (that shouldn't make a difference anyway) after updating the packages, but the problem persists. Some of the responses I'm getting are the following: After running sudo dpkg-reconfigure lightdm (but not ... gdm) I get the following message: dpkg-maintscript-helper:warning: environment variable DPKG_MATINSCRIPT_NAME missing dpkg-maintscript-helper:warning: environment variable DPKG_MATINSCRIPT_PACKAGE missing After trying sudo service lightdm start or sudo start lightdm I get to see the boot loading screen again but nothing changes. If I go back to the tty shell I see lightdm start/running, process <num> but ps -e | grep lightdm gives no output. After trying sudo service gdm start or sudo starg gdm I get the gdm start/running, process <num> message, and gdm-binary is supposedly an active process, but all that happens is that the screen blinks a couple of times and nothing else. Other candidate solutions that I'd found on the web included running startx but when I try that I get an error output [...] Fatal server error: no screens found [...]. Moreover, I made sure that lightdm-gtk-greeter is installed but that did not help either. Please excuse my not including complete outputs/logs; I am writing this post from another computer and it's hard to manually copy the complete logs. Also, I've seen several posts that had to do with similar problems, but either there was no fix, or the one suggested did not work for me. In closing: Please help! I very much hope to avoid re-installing Ubuntu from scratch! :) Alex @mosi I did not manage to fix the NVIDIA kernel driver as per your instructions. I should perhaps mention that I'm on a Dell XPS15 laptop with an NVIDIA Optimus graphics card, and that I have bumblebee installed (which installs nvidia drivers during its installation, I believe). Issuing the mentioned commands I get the following: ~$uname -r 3.0.0-12-generic ~$lsmod | grep -i nvidia nvidia 11713772 0 ~$dmesg | grep -i nvidia [ 8.980041] nvidia: module license 'NVIDIA' taints kernel. [ 9.354860] nvidia 0000:01:00.0: power state changed by ACPI to D0 [ 9.354864] nvidia 0000:01:00.0: power state changed by ACPI to D0 [ 9.354868] nvidia 0000:01:00.0: enabling device (0006 -> 0007) [ 9.354873] nvidia 0000:01:00.0: PCI INT A -> GSI 16 (level, low) -> IRQ 16 [ 9.354879] nvidia 0000:01:00.0: setting latency timer to 64 [ 9.355052] NVRM: loading NVIDIA UNIX x86_64 Kernel Module 280.13 Wed Jul 27 16:53:56 PDT 2011 Also, running aptitude search nvidia gives me the following: p nvidia-173 - NVIDIA binary Xorg driver, kernel module a p nvidia-173-dev - NVIDIA binary Xorg driver development file p nvidia-173-updates - NVIDIA binary Xorg driver, kernel module a p nvidia-173-updates-dev - NVIDIA binary Xorg driver development file p nvidia-96 - NVIDIA binary Xorg driver, kernel module a p nvidia-96-dev - NVIDIA binary Xorg driver development file p nvidia-96-updates - NVIDIA binary Xorg driver, kernel module a p nvidia-96-updates-dev - NVIDIA binary Xorg driver development file p nvidia-cg-toolkit - Cg Toolkit - GPU Shader Authoring Language p nvidia-common - Find obsolete NVIDIA drivers i nvidia-current - NVIDIA binary Xorg driver, kernel module a p nvidia-current-dev - NVIDIA binary Xorg driver development file c nvidia-current-updates - NVIDIA binary Xorg driver, kernel module a p nvidia-current-updates-dev - NVIDIA binary Xorg driver development file i nvidia-settings - Tool of configuring the NVIDIA graphics dr p nvidia-settings-updates - Tool of configuring the NVIDIA graphics dr v nvidia-va-driver - v nvidia-va-driver - I've tried manually installing (sudo aptitude install <package>) packages nvidia-common and nvidia-settings-updates but to no avail. For example, sudo aptitude install nvidia-settings-updates returns the following log: Reading package lists... Building dependency tree... Reading state information... Reading extended state information... Initializing package states... Writing extended state information... No packages will be installed, upgraded, or removed. 0 packages upgraded, 0 newly installed, 0 to remove and 83 not upgraded. Need to get 0 B of archives. After unpacking 0 B will be used. Writing extended state information... Reading package lists... Building dependency tree... Reading state information... Reading extended state information... Initializing package states... Writing extended state information... The same happens with the Linux headers (i.e. I cannot seem to be able to install linux-headers-3.0.0-12-generic). The output of aptitude search linux-headers is as follows: v linux-headers - v linux-headers - v linux-headers-2.6 - i linux-headers-2.6.38-11 - Header files related to Linux kernel versi i linux-headers-2.6.38-11-generic - Linux kernel headers for version 2.6.38 on i A linux-headers-2.6.38-8 - Header files related to Linux kernel versi i A linux-headers-2.6.38-8-generic - Linux kernel headers for version 2.6.38 on v linux-headers-3 - v linux-headers-3.0 - v linux-headers-3.0 - i A linux-headers-3.0.0-12 - Header files related to Linux kernel versi p linux-headers-3.0.0-12-generic - Linux kernel headers for version 3.0.0 on p linux-headers-3.0.0-12-generic- - Linux kernel headers for version 3.0.0 on p linux-headers-3.0.0-12-server - Linux kernel headers for version 3.0.0 on p linux-headers-3.0.0-12-virtual - Linux kernel headers for version 3.0.0 on p linux-headers-generic - Generic Linux kernel headers p linux-headers-generic-pae - Generic Linux kernel headers v linux-headers-lbm - v linux-headers-lbm - v linux-headers-lbm-2.6 - v linux-headers-lbm-2.6 - p linux-headers-lbm-3.0.0-12-gene - Header files related to linux-backports-mo p linux-headers-lbm-3.0.0-12-gene - Header files related to linux-backports-mo p linux-headers-lbm-3.0.0-12-serv - Header files related to linux-backports-mo p linux-headers-server - Linux kernel headers on Server Equipment. p linux-headers-virtual - Linux kernel headers for virtual machines @heartsmagic I did try purging and reinstalling any nvidia driver packages, but it did not seem to make a difference, My xorg.conf file contains the following: # nvidia-xconfig: X configuration file generated by nvidia-xconfig # nvidia-xconfig: version 280.13 ([email protected]) Wed Jul 27 17:15:58 PDT 2011 Section "ServerLayout" Identifier "Layout0" Screen 0 "Screen0" 0 0 InputDevice "Keyboard0" "CoreKeyboard" InputDevice "Mouse0" "CorePointer" EndSection Section "Files" EndSection Section "InputDevice" # generated from default Identifier "Mouse0" Driver "mouse" Option "Protocol" "auto" Option "Device" "/dev/psaux" Option "Emulate3Buttons" "no" Option "ZAxisMapping" "4 5" EndSection Section "InputDevice" # generated from default Identifier "Keyboard0" Driver "kbd" EndSection Section "Monitor" Identifier "Monitor0" VendorName "Unknown" ModelName "Unknown" HorizSync 28.0 - 33.0 VertRefresh 43.0 - 72.0 Option "DPMS" EndSection Section "Device" Identifier "Device0" Driver "nvidia" VendorName "NVIDIA Corporation" EndSection Section "Screen" Identifier "Screen0" Device "Device0" Monitor "Monitor0" DefaultDepth 24 SubSection "Display" Depth 24 EndSubSection EndSection

Read the article
Prioritizing Product Features

- by Robert May

A very common task in Agile Environments is prioritization. Teams that are functioning well will prioritize new features, old features, the backlog, and any other source of stories for the team, and they’ll do it regularly. Not all teams are good at prioritizing according to the real return on investment that building stories will yield to the company. This is unfortunate. Too often, teams end up building features that are less valuable, and everyone seems to know it except perhaps the product owner! Most features built into software are never even used. Clearly, not much return for features that go unused. So how does a company avoid building features that add little value to the company? This is a tough question to answer, but usually, this prioritization starts at the top with the executives of the company. After all, they’re responsible for the overall vision of the company. Here’s what I recommend: Know your market. Know your customers and users. Know where you’re going and what you want to achieve. Implement the Vision Know Your Market We often see companies that don’t know their market. Personally, I’m surprised by this. These companies don’t know who their competitors are, don’t know what features make their product desirable in the market, and in many cases, get by with saying, “I’ve been doing this for XX years. I know what the market wants!” In many cases, they equate “marketing” with “advertising” and don’t understand the difference. This is almost never true. Good companies will spend significant amounts of time and money finding out who they’re competing against and what makes their competitors successful in the marketplace. Good companies understand that marketing involves more than just advertising. Often, marketing is mostly research and analysis, not sales. Until you understand your market, you cannot know what features will give you the best return on your investment dollar. Good companies have a marketing department and can answer the next important step which is to know your customers and your users. Know your Customers and Users First, note that I included both customers and users. They’re often not the same thing. Users use the product that you build. Customers buy the product that you build. It’s a subtle difference, but too often, I’ve seen companies that focus exclusively on one or the other and are not successful simply because they ignore an important part of the group. If your company is doing appropriate marketing, you know that these are two different aspects of your product and that both deserve attention to have a product that is successful in your target market. Your marketing department should be spending a lot of time understanding these personas and then conveying that information to the company. I’m always surprised when development teams think that they can build a product that people want to use without understanding the users of that product. Developers think differently than most people in the world. They know what the computer is doing. The computer isn’t magic to them. So when they assume that they know how to build something, they bring with them quite a bit of baggage. Never assume that you know your customer unless you’re regularly having interaction with them. Also, don’t just leave this to Marketing or Product Management. Take them time to get your developers out with the customers as well. Developers are very smart people, and often, seeing how someone uses their software inspires them to make a much better product. Very often, because the users and customers aren’t know, teams will spend a significant amount of time building apps that are super flexible and configurable so that any possible combination of feature can be used. This demonstrates a clear lack of understanding of the customer. Most configuration questions can quickly be answered by talking to the customer. In most cases, if your software requires significant setup and configuration before its usable, you probably don’t know your customers and users very well. Until you know your customers, you cannot know what features will be most valuable to your customers and you cannot build those features in a way that your customers can use. Know Where You’re Going and What You Want to Achieve Many companies suffer from not having a plan. Executives will tell the team to make them a plan. The team, not knowing their market and customers and users, will come up with a plan that doesn’t reflect reality and doesn’t consider ROI. Management then wonders why the product is doing poorly in the market place. Instead of leaving this up to the teams, as executives, work with Marketing to understand what broad categories of features will sell the most product in the marketplace. Then, once you’ve determined that, give this vision to the team and let them run with it. Revise the vision as needed, but avoid changing streams frequently. Sure, sometimes you need to, but often, executives will change priorities many times a month, leading to nothing more than confusion. If the team has a vision, they’ll be able to execute that vision far better than they could otherwise. By knowing what products are most important, you can set budgetary goals and guidelines that will help you achieve the vision that was created. Implement the Vision Creating the vision is often where the general executives stop participating in the plan. The team is responsible for implementing that vision. Executives should attend showcases and and should remain aware of the progress that the team is making towards meeting the vision, however. Once a broad vision has been created, the team should break that vision down into minimal market features (MMF). These MMFs should be sized using story points so that, using the team’s velocity, an estimated cost can be determined for each feature. The product management team should then try to quantify the relative value of the MMFs based on customer feedback and interviews. Once the value and cost of creating the feature is understood, a return on investment can be calculated. The features should then be prioritized with the MMF’s that have the highest value and lowest cost rising to the top of features to implement. Don’t let politics get in the way! Once the MMF’s have been prioritized, they should go through release planning to schedule them for implementation. Conclusion By having a good grasp on the strategy of the company, your Agile teams can be much more effective. Each and every story the team is implementing will roll up into features that matter to the company and provide ROI to them. The steps outlined in this post should be repeated on a regular basis. I recommend reviewing them at least once per quarter to make sure that the vision hasn’t shifted and that the teams are still working on what matters most to the company. Technorati Tags: Agile,Product Owner,ROI

Read the article
Parallel LINQ - PLINQ

- by nmarun

Turns out now with .net 4.0 we can run a query like a multi-threaded application. Say you want to query a collection of objects and return only those that meet certain conditions. Until now, we basically had one ‘control’ that iterated over all the objects in the collection, checked the condition on each object and returned if it passed. We obviously agree that if we can ‘break’ this task into smaller ones, assign each task to a different ‘control’ and ask all the controls to do their job - in-parallel, the time taken the finish the entire task will be much lower. Welcome to PLINQ. Let’s take some examples. I have the following method that uses our good ol’ LINQ. 1: private static void Linq(int lowerLimit, int upperLimit) 2: { 3: // populate an array with int values from lowerLimit to the upperLimit 4: var source = Enumerable.Range(lowerLimit, upperLimit); 5: 6: // Start a timer 7: Stopwatch stopwatch = new Stopwatch(); 8: stopwatch.Start(); 9: 10: // set the expectation => build the expression tree 11: var evenNumbers = from num in source 12: where IsDivisibleBy(num, 2) 13: select num; 14: 15: // iterate over and print the returned items 16: foreach (var number in evenNumbers) 17: { 18: Console.WriteLine(string.Format("** {0}", number)); 19: } 20: 21: stopwatch.Stop(); 22: 23: // check the metrics 24: Console.WriteLine(String.Format("Elapsed {0}ms", stopwatch.ElapsedMilliseconds)); 25: } I’ve added comments for the major steps, but the only thing I want to talk about here is the IsDivisibleBy() method. I know I could have just included the logic directly in the where clause. I called a method to add ‘delay’ to the execution of the query - to simulate a loooooooooong operation (will be easier to compare the results). 1: private static bool IsDivisibleBy(int number, int divisor) 2: { 3: // iterate over some database query 4: // to add time to the execution of this method; 5: // the TableB has around 10 records 6: for (int i = 0; i < 10; i++) 7: { 8: DataClasses1DataContext dataContext = new DataClasses1DataContext(); 9: var query = from b in dataContext.TableBs select b; 10: 11: foreach (var row in query) 12: { 13: // Do NOTHING (wish my job was like this) 14: } 15: } 16: 17: return number % divisor == 0; 18: } Now, let’s look at how to modify this to PLINQ. 1: private static void Plinq(int lowerLimit, int upperLimit) 2: { 3: // populate an array with int values from lowerLimit to the upperLimit 4: var source = Enumerable.Range(lowerLimit, upperLimit); 5: 6: // Start a timer 7: Stopwatch stopwatch = new Stopwatch(); 8: stopwatch.Start(); 9: 10: // set the expectation => build the expression tree 11: var evenNumbers = from num in source.AsParallel() 12: where IsDivisibleBy(num, 2) 13: select num; 14: 15: // iterate over and print the returned items 16: foreach (var number in evenNumbers) 17: { 18: Console.WriteLine(string.Format("** {0}", number)); 19: } 20: 21: stopwatch.Stop(); 22: 23: // check the metrics 24: Console.WriteLine(String.Format("Elapsed {0}ms", stopwatch.ElapsedMilliseconds)); 25: } That’s it, this is now in PLINQ format. Oh and if you haven’t found the difference, look line 11 a little more closely. You’ll see an extension method ‘AsParallel()’ added to the ‘source’ variable. Couldn’t be more simpler right? So this is going to improve the performance for us. Let’s test it. So in my Main method of the Console application that I’m working on, I make a call to both. 1: static void Main(string[] args) 2: { 3: // set lower and upper limits 4: int lowerLimit = 1; 5: int upperLimit = 20; 6: // call the methods 7: Console.WriteLine("Calling Linq() method"); 8: Linq(lowerLimit, upperLimit); 9: 10: Console.WriteLine(); 11: Console.WriteLine("Calling Plinq() method"); 12: Plinq(lowerLimit, upperLimit); 13: 14: Console.ReadLine(); // just so I get enough time to read the output 15: } YMMV, but here are the results that I got: It’s quite obvious from the above results that the Plinq() method is taking considerably less time than the Linq() version. I’m sure you’ve already noticed that the output of the Plinq() method is not in order. That’s because, each of the ‘control’s we sent to fetch the results, reported with values as and when they obtained them. This is something about parallel LINQ that one needs to remember – the collection cannot be guaranteed to be undisturbed. This could be counted as a negative about PLINQ (emphasize ‘could’). Nevertheless, if we want the collection to be sorted, we can use a SortedSet (.net 4.0) or build our own custom ‘sorter’. Either way we go, there’s a good chance we’ll end up with a better performance using PLINQ. And there’s another negative of PLINQ (depending on how you see it). This is regarding the CPU cycles. See the usage for Linq() method (used ResourceMonitor): I have dual CPU’s and see the height of the peak in the bottom two blocks and now compare to what happens when I run the Plinq() method. The difference is obvious. Higher usage, but for a shorter duration (width of the peak). Both these points make sense in both cases. Linq() runs for a longer time, but uses less resources whereas Plinq() runs for a shorter time and consumes more resources. Even after knowing all these, I’m still inclined towards PLINQ. PLINQ rocks! (no hard feelings LINQ)

Read the article
C#: LINQ vs foreach - Round 1.

- by James Michael Hare

So I was reading Peter Kellner's blog entry on Resharper 5.0 and its LINQ refactoring and thought that was very cool. But that raised a point I had always been curious about in my head -- which is a better choice: manual foreach loops or LINQ? The answer is not really clear-cut. There are two sides to any code cost arguments: performance and maintainability. The first of these is obvious and quantifiable. Given any two pieces of code that perform the same function, you can run them side-by-side and see which piece of code performs better. Unfortunately, this is not always a good measure. Well written assembly language outperforms well written C++ code, but you lose a lot in maintainability which creates a big techncial debt load that is hard to offset as the application ages. In contrast, higher level constructs make the code more brief and easier to understand, hence reducing technical cost. Now, obviously in this case we're not talking two separate languages, we're comparing doing something manually in the language versus using a higher-order set of IEnumerable extensions that are in the System.Linq library. Well, before we discuss any further, let's look at some sample code and the numbers. First, let's take a look at the for loop and the LINQ expression. This is just a simple find comparison: // find implemented via LINQ public static bool FindViaLinq(IEnumerable<int> list, int target) { return list.Any(item => item == target); } // find implemented via standard iteration public static bool FindViaIteration(IEnumerable<int> list, int target) { foreach (var i in list) { if (i == target) { return true; } } return false; } Okay, looking at this from a maintainability point of view, the Linq expression is definitely more concise (8 lines down to 1) and is very readable in intention. You don't have to actually analyze the behavior of the loop to determine what it's doing. So let's take a look at performance metrics from 100,000 iterations of these methods on a List<int> of varying sizes filled with random data. For this test, we fill a target array with 100,000 random integers and then run the exact same pseudo-random targets through both searches. List<T> On 100,000 Iterations Method Size Total (ms) Per Iteration (ms) % Slower Any 10 26 0.00046 30.00% Iteration 10 20 0.00023 - Any 100 116 0.00201 18.37% Iteration 100 98 0.00118 - Any 1000 1058 0.01853 16.78% Iteration 1000 906 0.01155 - Any 10,000 10,383 0.18189 17.41% Iteration 10,000 8843 0.11362 - Any 100,000 104,004 1.8297 18.27% Iteration 100,000 87,941 1.13163 - The LINQ expression is running about 17% slower for average size collections and worse for smaller collections. Presumably, this is due to the overhead of the state machine used to track the iterators for the yield returns in the LINQ expressions, which seems about right in a tight loop such as this. So what about other LINQ expressions? After all, Any() is one of the more trivial ones. I decided to try the TakeWhile() algorithm using a Count() to get the position stopped like the sample Pete was using in his blog that Resharper refactored for him into LINQ: // Linq form public static int GetTargetPosition1(IEnumerable<int> list, int target) { return list.TakeWhile(item => item != target).Count(); } // traditionally iterative form public static int GetTargetPosition2(IEnumerable<int> list, int target) { int count = 0; foreach (var i in list) { if(i == target) { break; } ++count; } return count; } Once again, the LINQ expression is much shorter, easier to read, and should be easier to maintain over time, reducing the cost of technical debt. So I ran these through the same test data: List<T> On 100,000 Iterations Method Size Total (ms) Per Iteration (ms) % Slower TakeWhile 10 41 0.00041 128% Iteration 10 18 0.00018 - TakeWhile 100 171 0.00171 88% Iteration 100 91 0.00091 - TakeWhile 1000 1604 0.01604 94% Iteration 1000 825 0.00825 - TakeWhile 10,000 15765 0.15765 92% Iteration 10,000 8204 0.08204 - TakeWhile 100,000 156950 1.5695 92% Iteration 100,000 81635 0.81635 - Wow! I expected some overhead due to the state machines iterators produce, but 90% slower? That seems a little heavy to me. So then I thought, well, what if TakeWhile() is not the right tool for the job? The problem is TakeWhile returns each item for processing using yield return, whereas our for-loop really doesn't care about the item beyond using it as a stop condition to evaluate. So what if that back and forth with the iterator state machine is the problem? Well, we can quickly create an (albeit ugly) lambda that uses the Any() along with a count in a closure (if a LINQ guru knows a better way PLEASE let me know!), after all , this is more consistent with what we're trying to do, we're trying to find the first occurence of an item and halt once we find it, we just happen to be counting on the way. This mostly matches Any(). // a new method that uses linq but evaluates the count in a closure. public static int TakeWhileViaLinq2(IEnumerable<int> list, int target) { int count = 0; list.Any(item => { if(item == target) { return true; } ++count; return false; }); return count; } Now how does this one compare? List<T> On 100,000 Iterations Method Size Total (ms) Per Iteration (ms) % Slower TakeWhile 10 41 0.00041 128% Any w/Closure 10 23 0.00023 28% Iteration 10 18 0.00018 - TakeWhile 100 171 0.00171 88% Any w/Closure 100 116 0.00116 27% Iteration 100 91 0.00091 - TakeWhile 1000 1604 0.01604 94% Any w/Closure 1000 1101 0.01101 33% Iteration 1000 825 0.00825 - TakeWhile 10,000 15765 0.15765 92% Any w/Closure 10,000 10802 0.10802 32% Iteration 10,000 8204 0.08204 - TakeWhile 100,000 156950 1.5695 92% Any w/Closure 100,000 108378 1.08378 33% Iteration 100,000 81635 0.81635 - Much better! It seems that the overhead of TakeAny() returning each item and updating the state in the state machine is drastically reduced by using Any() since Any() iterates forward until it finds the value we're looking for -- for the task we're attempting to do. So the lesson there is, make sure when you use a LINQ expression you're choosing the best expression for the job, because if you're doing more work than you really need, you'll have a slower algorithm. But this is true of any choice of algorithm or collection in general. Even with the Any() with the count in the closure it is still about 30% slower, but let's consider that angle carefully. For a list of 100,000 items, it was the difference between 1.01 ms and 0.82 ms roughly in a List<T>. That's really not that bad at all in the grand scheme of things. Even running at 90% slower with TakeWhile(), for the vast majority of my projects, an extra millisecond to save potential errors in the long term and improve maintainability is a small price to pay. And if your typical list is 1000 items or less we're talking only microseconds worth of difference. It's like they say: 90% of your performance bottlenecks are in 2% of your code, so over-optimizing almost never pays off. So personally, I'll take the LINQ expression wherever I can because they will be easier to read and maintain (thus reducing technical debt) and I can rely on Microsoft's development to have coded and unit tested those algorithm fully for me instead of relying on a developer to code the loop logic correctly. If something's 90% slower, yes, it's worth keeping in mind, but it's really not until you start get magnitudes-of-order slower (10x, 100x, 1000x) that alarm bells should really go off. And if I ever do need that last millisecond of performance? Well then I'll optimize JUST THAT problem spot. To me it's worth it for the readability, speed-to-market, and maintainability.

Read the article
SQL SERVER – Weekly Series – Memory Lane – #052

- by Pinal Dave

Let us continue with the final episode of the Memory Lane Series. Here is the list of selected articles of SQLAuthority.com across all these years. Instead of just listing all the articles I have selected a few of my most favorite articles and have listed them here with additional notes below it. Let me know which one of the following is your favorite article from memory lane. 2007 Set Server Level FILLFACTOR Using T-SQL Script Specifies a percentage that indicates how full the Database Engine should make the leaf level of each index page during index creation or alteration. fillfactor must be an integer value from 1 to 100. The default is 0. Limitation of Online Index Rebuld Operation Online operation means when online operations are happening in the database are in normal operational condition, the processes which are participating in online operations does not require exclusive access to the database. Get Permissions of My Username / Userlogin on Server / Database A few days ago, I was invited to one of the largest database company. I was asked to review database schema and propose changes to it. There was special username or user logic was created for me, so I can review their database. I was very much interested to know what kind of permissions I was assigned per server level and database level. I did not feel like asking Sr. DBA the question about permissions. Simple Example of WHILE Loop With CONTINUE and BREAK Keywords This question is one of those questions which is very simple and most of the users get it correct, however few users find it confusing for the first time. I have tried to explain the usage of simple WHILE loop in the first example. BREAK keyword will exit the stop the while loop and control is moved to the next statement after the while loop. CONTINUE keyword skips all the statement after its execution and control is sent to the first statement of while loop. Forced Parameterization and Simple Parameterization – T-SQL and SSMS When the PARAMETERIZATION option is set to FORCED, any literal value that appears in a SELECT, INSERT, UPDATE or DELETE statement is converted to a parameter during query compilation. When the PARAMETERIZATION database option is SET to SIMPLE, the SQL Server query optimizer may choose to parameterize the queries. 2008 Transaction and Local Variables – Swap Variables – Update All At Once Concept Summary : Transaction have no effect over memory variables. When UPDATE statement is applied over any table (physical or memory) all the updates are applied at one time together when the statement is committed. First of all I suggest that you read the article listed above about the effect of transaction on local variant. As seen there local variables are independent of any transaction effect. Simulate INNER JOIN using LEFT JOIN statement – Performance Analysis Just a day ago, while I was working with JOINs I find one interesting observation, which has prompted me to create following example. Before we continue further let me make very clear that INNER JOIN should be used where it cannot be used and simulating INNER JOIN using any other JOINs will degrade the performance. If there are scopes to convert any OUTER JOIN to INNER JOIN it should be done with priority. 2009 Introduction to Business Intelligence – Important Terms & Definitions Business intelligence (BI) is a broad category of application programs and technologies for gathering, storing, analyzing, and providing access to data from various data sources, thus providing enterprise users with reliable and timely information and analysis for improved decision making. Difference Between Candidate Keys and Primary Key Candidate Key – A Candidate Key can be any column or a combination of columns that can qualify as unique key in database. There can be multiple Candidate Keys in one table. Each Candidate Key can qualify as Primary Key. Primary Key – A Primary Key is a column or a combination of columns that uniquely identify a record. Only one Candidate Key can be Primary Key. 2010 Taking Multiple Backup of Database in Single Command – Mirrored Database Backup I recently had a very interesting experience. In one of my recent consultancy works, I was told by our client that they are going to take the backup of the database and will also a copy of it at the same time. I expressed that it was surely possible if they were going to use a mirror command. In addition, they told me that whenever they take two copies of the database, the size of the database, is always reduced. Now this was something not clear to me, I said it was not possible and so I asked them to show me the script. Corrupted Backup File and Unsuccessful Restore The CTO, who was also present at the location, got very upset with this situation. He then asked when the last successful restore test was done. As expected, the answer was NEVER.There were no successful restore tests done before. During that time, I was present and I could clearly see the stress, confusion, carelessness and anger around me. I did not appreciate the feeling and I was pretty sure that no one in there wanted the atmosphere like me. 2011 TRACEWRITE – Wait Type – Wait Related to Buffer and Resolution SQL Trace is a SQL Server database engine technology which monitors specific events generated when various actions occur in the database engine. When any event is fired it goes through various stages as well various routes. One of the routes is Trace I/O Provider, which sends data to its final destination either as a file or rowset. DATEDIFF – Accuracy of Various Dateparts If you want to have accuracy in seconds, you need to use a different approach. In the first example, the accurate method is to find the number of seconds first and then divide it by 60 to convert it in minutes. Dedicated Access Control for SQL Server Express Edition http://www.youtube.com/watch?v=1k00z82u4OI Book Signing at SQLPASS 2012 Who I Am And How I Got Here – True Story as Blog Post If there was a shortcut to success – I want to know. I learnt SQL Server hard way and I am still learning. There are so many things, I have to learn. There is not enough time to learn everything which we want to learn. I am constantly working on it every day. I welcome you to join my journey as well. Please join me in my journey to learn SQL Server – more the merrier. Vacation, Travel and Study – A New Concept Even those who have advanced degrees and went to college for years, or even decades, find studying hard. There is a difference between studying for a career and studying for a certification. At least to get a degree there is a variety of subjects, with labs, exams, and practice problems to make things more interesting. Order By Numeric Values Formatted as String We have a table which has a column containing alphanumeric data. The data always has first as an integer and later part as a string. The business need is to order the data based on the first part of the alphanumeric data which is an integer. Now the problem is that no matter how we use ORDER BY the result is not produced as expected. Let us understand this with an example. Resolving SQL Server Connection Errors – SQL in Sixty Seconds #030 – Video One of the most famous errors related to SQL Server is about connecting to SQL Server itself. Here is how it goes, most of the time developers have worked with SQL Server and knows pretty much every error which they face during development language. However, hardly they install fresh SQL Server. As the installation of the SQL Server is a rare occasion unless you are a DBA who is responsible for such an instance – the error faced during installations are pretty rare as well. http://www.youtube.com/watch?v=1k00z82u4OI Reference: Pinal Dave (http://blog.sqlauthority.com) Filed under: Memory Lane, PostADay, SQL, SQL Authority, SQL Query, SQL Server, SQL Tips and Tricks, T SQL, Technology

Read the article
Is RTD Stateless or Stateful?

- by [email protected]

Yes. A stateless service is one where each request is an independent transaction that can be processed by any of the servers in a cluster. A stateful service is one where state is kept in a server's memory from transaction to transaction, thus necessitating the proper routing of requests to the right server. The main advantage of stateless systems is simplicity of design. The main advantage of stateful systems is performance. I'm often asked whether RTD is a stateless or stateful service, so I wanted to clarify this issue in depth so that RTD's architecture will be properly understood. The short answer is: "RTD can be configured as a stateless or stateful service." The performance difference between stateless and stateful systems can be very significant, and while in a call center implementation it may be reasonable to use a pure stateless configuration, a web implementation that produces thousands of requests per second is practically impossible with a stateless configuration. RTD's performance is orders of magnitude better than most competing systems. RTD was architected from the ground up to achieve this performance. Features like automatic and dynamic compression of prediction models, automatic translation of metadata to machine code, lack of interpreted languages, and separation of model building from decisioning contribute to achieving this performance level. Because of this focus on performance we decided to have RTD's default configuration work in a stateful manner. By being stateful RTD requests are typically handled in a few milliseconds when repeated requests come to the same session. Now, those readers that have participated in implementations of RTD know that RTD's architecture is also focused on reducing Total Cost of Ownership (TCO) with features like automatic model building, automatic time windows, automatic maintenance of database tables, automatic evaluation of data mining models, automatic management of models partitioned by channel, geography, etcetera, and hot swapping of configurations. How do you reconcile the need for a low TCO and the need for performance? How do you get the performance of a stateful system with the simplicity of a stateless system? The answer is that you make the system behave like a stateless system to the exterior, but you let it automatically take advantage of situations where being stateful is better. For example, one of the advantages of stateless systems is that you can route a message to any server in a cluster, without worrying about sending it to the same server that was handling the session in previous messages. With an RTD stateful configuration you can still route the message to any server in the cluster, so from the point of view of the configuration of other systems, it is the same as a stateless service. The difference though comes in performance, because if the message arrives to the right server, RTD can serve it without any external access to the session's state, thus tremendously reducing processing time. In typical implementations it is not rare to have high percentages of messages routed directly to the right server, while those that are not, are easily handled by forwarding the messages to the right server. This architecture usually provides the best of both worlds with performance and simplicity of configuration. Configuring RTD as a pure stateless service A pure stateless configuration requires session data to be persisted at the end of handling each and every message and reloading that data at the beginning of handling any new message. This is of course, the root of the inefficiency of these configurations. This is also the reason why many "stateless" implementations actually do keep state to take advantage of a request coming back to the same server. Nevertheless, if the implementation requires a pure stateless decision service, this is easy to configure in RTD. The way to do it is: Mark every Integration Point to Close the session at the end of processing the message In the Session entity persist the session data on closing the session In the session entity check if a persisted version exists and load it An excellent solution for persisting the session data is Oracle Coherence, which provides a high performance, distributed cache that minimizes the performance impact of persisting and reloading the session. Alternatively, the session can be persisted to a local database. An interesting feature of the RTD stateless configuration is that it can cope with serializing concurrent requests for the same session. For example, if a web page produces two requests to the decision service, these requests could come concurrently to the decision services and be handled by different servers. Most stateless implementation would have the two requests step onto each other when saving the state, or fail one of the messages. When properly configured, RTD will make one message wait for the other before processing. A Word on Context Using the context of a customer interaction typically significantly increases lift. For example, offer success in a call center could double if the context of the call is taken into account. For this reason, it is important to utilize the contextual information in decision making. To make the contextual information available throughout a session it needs to be persisted. When there is a well defined owner for the information then there is no problem because in case of a session restart, the information can be easily retrieved. If there is no official owner of the information, then RTD can be configured to persist this information. Once again, RTD provides flexibility to ensure high performance when it is adequate to allow for some loss of state in the rare cases of server failure. For example, in a heavy use web site that serves 1000 pages per second the navigation history may be stored in the in memory session. In such sites it is typical that there is no OLTP that stores all the navigation events, therefore if an RTD server were to fail, it would be possible for the navigation to that point to be lost (note that a new session would be immediately established in one of the other servers). In most cases the loss of this navigation information would be acceptable as it would happen rarely. If it is desired to save this information, RTD would persist it every time the visitor navigates to a new page. Note that this practice is preferred whether RTD is configured in a stateless or stateful manner.

Read the article
Running Solaris 11 as a control domain on a T2000

- by jsavit

There is increased adoption of Oracle Solaris 11, and many customers are deploying it on systems that previously ran Solaris 10. That includes older T1-processor based systems like T1000 and T2000. Even though they are old (from 2005) and don't have the performance of current SPARC servers, they are still functional, stable servers that customers continue to operate. One reason to install Solaris 11 on them is that older machines are attractive for testing OS upgrades before updating current, production systems. Normally this does not present a challenge, because Solaris 11 runs on any T-series or M-series SPARC server. One scenario adds a complication: running Solaris 11 in a control domain on a T1000 or T2000 hosting logical domains. Solaris 11 pre-installed Oracle VM Server for SPARC incompatible with T1 Unlike Solaris 10, Solaris 11 comes with Oracle VM Server for SPARC preinstalled. The ldomsmanager package contains the logical domains manager for Oracle VM Server for SPARC 2.2, which requires a SPARC T2, T2+, T3, or T4 server. It does not work with T1-processor systems, which are only supported by LDoms Manager 1.2 and earlier. The following screenshot shows what happens (bold font) if you try to use Oracle VM Server for SPARC 2.x commands in a Solaris 11 control domain. The commands were issued in a control domain on a T2000 that previously ran Solaris 10. We also display the version of the logical domains manager installed in Solaris 11: root@t2000 psrinfo -vp The physical processor has 4 virtual processors (0-3) UltraSPARC-T1 (chipid 0, clock 1200 MHz) # prtconf|grep T SUNW,Sun-Fire-T200 # ldm -V Failed to connect to logical domain manager: Connection refused # pkg info ldomsmanager Name: system/ldoms/ldomsmanager Summary: Logical Domains Manager Description: LDoms Manager - Virtualization for SPARC T-Series Category: System/Virtualization State: Installed Publisher: solaris Version: 2.2.0.0 Build Release: 5.11 Branch: 0.175.0.8.0.3.0 Packaging Date: May 25, 2012 10:20:48 PM Size: 2.86 MB FMRI: pkg://solaris/system/ldoms/[email protected],5.11-0.175.0.8.0.3.0:20120525T222048Z The 2.2 version of the logical domains manager will have to be removed, and 1.2 installed, in order to use this as a control domain. Preparing to change - create a new boot environment Before doing anything else, lets create a new boot environment: # beadm list BE Active Mountpoint Space Policy Created -- ------ ---------- ----- ------ ------- solaris NR / 2.14G static 2012-09-25 10:32 # beadm create solaris-1 # beadm activate solaris-1 # beadm list BE Active Mountpoint Space Policy Created -- ------ ---------- ----- ------ ------- solaris N / 4.82M static 2012-09-25 10:32 solaris-1 R - 2.14G static 2012-09-29 11:40 # init 0 Normally an init 6 to reboot would have been sufficient, but in the next step I reset the system anyway in order to put the system in factory default mode for a "clean" domain configuration. Preparing to change - reset to factory default There was a leftover domain configuration on the T2000, so I reset it to the factory install state. Since the ldm command is't working yet, it can't be done from the control domain, so I did it by logging onto to the service processor: $ ssh -X admin@t2000-sc Copyright (c) 2010, Oracle and/or its affiliates. All rights reserved. Oracle Advanced Lights Out Manager CMT v1.7.9 Please login: admin Please Enter password: ******** sc> showhost Sun-Fire-T2000 System Firmware 6.7.10 2010/07/14 16:35 Host flash versions: OBP 4.30.4.b 2010/07/09 13:48 Hypervisor 1.7.3.c 2010/07/09 15:14 POST 4.30.4.b 2010/07/09 14:24 sc> bootmode config="factory-default" sc> poweroff Are you sure you want to power off the system [y/n]? y SC Alert: SC Request to Power Off Host. SC Alert: Host system has shut down. sc> poweron SC Alert: Host System has Reset At this point I rebooted into the new Solaris 11 boot environment, and Solaris commands showed it was running on the factory default configuration of a single domain owning all 32 CPUs and 32GB of RAM (that's what it looked like in 2005.) # psrinfo -vp The physical processor has 8 cores and 32 virtual processors (0-31) The core has 4 virtual processors (0-3) The core has 4 virtual processors (4-7) The core has 4 virtual processors (8-11) The core has 4 virtual processors (12-15) The core has 4 virtual processors (16-19) The core has 4 virtual processors (20-23) The core has 4 virtual processors (24-27) The core has 4 virtual processors (28-31) UltraSPARC-T1 (chipid 0, clock 1200 MHz) # prtconf|grep Mem Memory size: 32640 Megabytes Note that the older processor has 4 virtual CPUs per core, while current processors have 8 per core. Remove ldomsmanager 2.2 and install the 1.2 version The Solaris 11 pkg command is now used to remove the 2.2 version that shipped with Solaris 11: # pkg uninstall ldomsmanager Packages to remove: 1 Create boot environment: No Create backup boot environment: No Services to change: 2 PHASE ACTIONS Removal Phase 130/130 PHASE ITEMS Package State Update Phase 1/1 Package Cache Update Phase 1/1 Image State Update Phase 2/2 Finally, LDoms 1.2 installed via its install script, the same way it was done years ago: # unzip LDoms-1_2-Integration-10.zip # cd LDoms-1_2-Integration-10/Install/ # ./install-ldm Welcome to the LDoms installer. You are about to install the Logical Domains Manager package that will enable you to create, destroy and control other domains on your system. Given the capabilities of the LDoms domain manager, you can now change the security configuration of this Solaris instance using the Solaris Security Toolkit. ... ... normal install messages omitted ... The Solaris Security Toolkit applies to Solaris 10, and cannot be used in Solaris 11 (in which several things hardened by the Toolkit are already hardened by default), so answer b in the choice below: You are about to install the Logical Domains Manager package that will enable you to create, destroy and control other domains on your system. Given the capabilities of the LDoms domain manager, you can now change the security configuration of this Solaris instance using the Solaris Security Toolkit. Select a security profile from this list: a) Hardened Solaris configuration for LDoms (recommended) b) Standard Solaris configuration c) Your custom-defined Solaris security configuration profile Enter a, b, or c [a]: b ... other install messages omitted for brevity... After install I ensure that the necessary services are enabled, and verify the version of the installed LDoms Manager: # svcs ldmd STATE STIME FMRI online 22:00:36 svc:/ldoms/ldmd:default # svcs vntsd STATE STIME FMRI disabled Aug_19 svc:/ldoms/vntsd:default # ldm -V Logical Domain Manager (v 1.2-debug) Hypervisor control protocol v 1.3 Using Hypervisor MD v 1.1 System PROM: Hypervisor v. 1.7.3. @(#)Hypervisor 1.7.3.c 2010/07/09 15:14\015 OpenBoot v. 4.30.4. @(#)OBP 4.30.4.b 2010/07/09 13:48 Set up control domain and domain services At this point we have a functioning LDoms 1.2 environment that can be configured in the usual fashion. One difference is that LDoms 1.2 behavior had 'delayed configuration mode (as expected) during initial configuration before rebooting the control domain. Another minor difference with a Solaris 11 control domain is that you define virtual switches using the 'vanity name' of the network interface, rather than the hardware driver name as in Solaris 10. # ldm list ------------------------------------------------------------------------------ Notice: the LDom Manager is running in configuration mode. Configuration and resource information is displayed for the configuration under construction; not the current active configuration. The configuration being constructed will only take effect after it is downloaded to the system controller and the host is reset. ------------------------------------------------------------------------------ NAME STATE FLAGS CONS VCPU MEMORY UTIL UPTIME primary active -n-c-- SP 32 32640M 3.2% 4d 2h 50m # ldm add-vdiskserver primary-vds0 primary # ldm add-vconscon port-range=5000-5100 primary-vcc0 primary # ldm add-vswitch net-dev=net0 primary-vsw0 primary # ldm set-mau 2 primary # ldm set-vcpu 8 primary # ldm set-memory 4g primary # ldm add-config initial # ldm list-spconfig factory-default initial [current] That's it, really. After reboot, we are ready to install guest domains. Summary - new wine in old bottles This example shows that (new) Solaris 11 can be installed on (old) T2000 servers and used as a control domain. The main activity is to remove the preinstalled Oracle VM Server for 2.2 and install Logical Domains 1.2 - the last version of LDoms to support T1-processor systems. I tested Solaris 10 and Solaris 11 guest domains running on this server and they worked without any surprises. This is a viable way to get further into Solaris 11 adoption, even on older T-series equipment.

Read the article
Nashorn, the rhino in the room

- by costlow

Nashorn is a new runtime within JDK 8 that allows developers to run code written in JavaScript and call back and forth with Java. One advantage to the Nashorn scripting engine is that is allows for quick prototyping of functionality or basic shell scripts that use Java libraries. The previous JavaScript runtime, named Rhino, was introduced in JDK 6 (released 2006, end of public updates Feb 2013). Keeping tradition amongst the global developer community, "Nashorn" is the German word for rhino. The Java platform and runtime is an intentional home to many languages beyond the Java language itself. OpenJDK’s Da Vinci Machine helps coordinate work amongst language developers and tool designers and has helped different languages by introducing the Invoke Dynamic instruction in Java 7 (2011), which resulted in two major benefits: speeding up execution of dynamic code, and providing the groundwork for Java 8’s lambda executions. Many of these improvements are discussed at the JVM Language Summit, where language and tool designers get together to discuss experiences and issues related to building these complex components. There are a number of benefits to running JavaScript applications on JDK 8’s Nashorn technology beyond writing scripts quickly: Interoperability with Java and JavaScript libraries. Scripts do not need to be compiled. Fast execution and multi-threading of JavaScript running in Java’s JRE. The ability to remotely debug applications using an IDE like NetBeans, Eclipse, or IntelliJ (instructions on the Nashorn blog). Automatic integration with Java monitoring tools, such as performance, health, and SIEM. In the remainder of this blog post, I will explain how to use Nashorn and the benefit from those features. Nashorn execution environment The Nashorn scripting engine is included in all versions of Java SE 8, both the JDK and the JRE. Unlike Java code, scripts written in nashorn are interpreted and do not need to be compiled before execution. Developers and users can access it in two ways: Users running JavaScript applications can call the binary directly:jre8/bin/jjs This mechanism can also be used in shell scripts by specifying a shebang like #!/usr/bin/jjs Developers can use the API and obtain a ScriptEngine through:ScriptEngine engine = new ScriptEngineManager().getEngineByName("nashorn"); When using a ScriptEngine, please understand that they execute code. Avoid running untrusted scripts or passing in untrusted/unvalidated inputs. During compilation, consider isolating access to the ScriptEngine and using Type Annotations to only allow @Untainted String arguments. One noteworthy difference between JavaScript executed in or outside of a web browser is that certain objects will not be available. For example when run outside a browser, there is no access to a document object or DOM tree. Other than that, all syntax, semantics, and capabilities are present. Examples of Java and JavaScript The Nashorn script engine allows developers of all experience levels the ability to write and run code that takes advantage of both languages. The specific dialect is ECMAScript 5.1 as identified by the User Guide and its standards definition through ECMA international. In addition to the example below, Benjamin Winterberg has a very well written Java 8 Nashorn Tutorial that provides a large number of code samples in both languages. Basic Operations A basic Hello World application written to run on Nashorn would look like this: #!/usr/bin/jjs print("Hello World"); The first line is a standard script indication, so that Linux or Unix systems can run the script through Nashorn. On Windows where scripts are not as common, you would run the script like: jjs helloWorld.js. Receiving Arguments In order to receive program arguments your jjs invocation needs to use the -scripting flag and a double-dash to separate which arguments are for jjs and which are for the script itself:jjs -scripting print.js -- "This will print" #!/usr/bin/jjs var whatYouSaid = $ARG.length==0 ? "You did not say anything" : $ARG[0] print(whatYouSaid); Interoperability with Java libraries (including 3rd party dependencies) Another goal of Nashorn was to allow for quick scriptable prototypes, allowing access into Java types and any libraries. Resources operate in the context of the script (either in-line with the script or as separate threads) so if you open network sockets and your script terminates, those sockets will be released and available for your next run. Your code can access Java types the same as regular Java classes. The “import statements” are written somewhat differently to accommodate for language. There is a choice of two styles: For standard classes, just name the class: var ServerSocket = java.net.ServerSocket For arrays or other items, use Java.type: var ByteArray = Java.type("byte[]")You could technically do this for all. The same technique will allow your script to use Java types from any library or 3rd party component and quickly prototype items. Building a user interface One major difference between JavaScript inside and outside of a web browser is the availability of a DOM object for rendering views. When run outside of the browser, JavaScript has full control to construct the entire user interface with pre-fabricated UI controls, charts, or components. The example below is a variation from the Nashorn and JavaFX guide to show how items work together. Nashorn has a -fx flag to make the user interface components available. With the example script below, just specify: jjs -fx -scripting fx.js -- "My title" #!/usr/bin/jjs -fx var Button = javafx.scene.control.Button; var StackPane = javafx.scene.layout.StackPane; var Scene = javafx.scene.Scene; var clickCounter=0; $STAGE.title = $ARG.length>0 ? $ARG[0] : "You didn't provide a title"; var button = new Button(); button.text = "Say 'Hello World'"; button.onAction = myFunctionForButtonClicking; var root = new StackPane(); root.children.add(button); $STAGE.scene = new Scene(root, 300, 250); $STAGE.show(); function myFunctionForButtonClicking(){ var text = "Click Counter: " + clickCounter; button.setText(text); clickCounter++; print(text); } For a more advanced post on using Nashorn to build a high-performing UI, see JavaFX with Nashorn Canvas example. Interoperable with frameworks like Node, Backbone, or Facebook React The major benefit of any language is the interoperability gained by people and systems that can read, write, and use it for interactions. Because Nashorn is built for the ECMAScript specification, developers familiar with JavaScript frameworks can write their code and then have system administrators deploy and monitor the applications the same as any other Java application. A number of projects are also running Node applications on Nashorn through Project Avatar and the supported modules. In addition to the previously mentioned Nashorn tutorial, Benjamin has also written a post about Using Backbone.js with Nashorn. To show the multi-language power of the Java Runtime, there is another interesting example that unites Facebook React and Clojure on JDK 8’s Nashorn. Summary Nashorn provides a simple and fast way of executing JavaScript applications and bridging between the best of each language. By making the full range of Java libraries to JavaScript applications, and the quick prototyping style of JavaScript to Java applications, developers are free to work as they see fit. Software Architects and System Administrators can take advantage of one runtime and leverage any work that they have done to tune, monitor, and certify their systems. Additional information is available within: The Nashorn Users’ Guide Java Magazine’s article "Next Generation JavaScript Engine for the JVM." The Nashorn team’s primary blog or a very helpful collection of Nashorn links.

Read the article
Automating deployments with the SQL Compare command line

- by Jonathan Hickford

In my previous article, “Five Tips to Get Your Organisation Releasing Software Frequently” I looked at how teams can automate processes to speed up release frequency. In this post, I’m looking specifically at automating deployments using the SQL Compare command line. SQL Compare compares SQL Server schemas and deploys the differences. It works very effectively in scenarios where only one deployment target is required – source and target databases are specified, compared, and a change script is automatically generated and applied. But if multiple targets exist, and pressure to increase the frequency of releases builds, this solution quickly becomes unwieldy. This is where SQL Compare’s command line comes into its own. I’ve put together a PowerShell script that loops through the Servers table and pulls out the server and database, these are then passed to sqlcompare.exe to be used as target parameters. In the example the source database is a scripts folder, a folder structure of scripted-out database objects used by both SQL Source Control and SQL Compare. The script can easily be adapted to use schema snapshots. -- Create a DeploymentTargets database and a Servers table CREATE DATABASE DeploymentTargets GO USE DeploymentTargets GO CREATE TABLE [dbo].[Servers]( [id] [int] IDENTITY(1,1) NOT NULL, [serverName] [nvarchar](50) NULL, [environment] [nvarchar](50) NULL, [databaseName] [nvarchar](50) NULL, CONSTRAINT [PK_Servers] PRIMARY KEY CLUSTERED ([id] ASC) ) GO -- Now insert your target server and database details INSERT INTO dbo.Servers ( serverName , environment , databaseName) VALUES ( N'myserverinstance' , N'myenvironment1' , N'mydb1') INSERT INTO dbo.Servers ( serverName , environment , databaseName) VALUES ( N'myserverinstance' , N'myenvironment2' , N'mydb2') Here’s the PowerShell script you can adapt for yourself as well. # We're holding the server names and database names that we want to deploy to in a database table. # We need to connect to that server to read these details $serverName = "" $databaseName = "DeploymentTargets" $authentication = "Integrated Security=SSPI" #$authentication = "User Id=xxx;PWD=xxx" # If you are using database authentication instead of Windows authentication. # Path to the scripts folder we want to deploy to the databases $scriptsPath = "SimpleTalk" # Path to SQLCompare.exe $SQLComparePath = "C:\Program Files (x86)\Red Gate\SQL Compare 10\sqlcompare.exe" # Create SQL connection string, and connection $ServerConnectionString = "Data Source=$serverName;Initial Catalog=$databaseName;$authentication" $ServerConnection = new-object system.data.SqlClient.SqlConnection($ServerConnectionString); # Create a Dataset to hold the DataTable $dataSet = new-object "System.Data.DataSet" "ServerList" # Create a query $query = "SET NOCOUNT ON;" $query += "SELECT serverName, environment, databaseName " $query += "FROM dbo.Servers; " # Create a DataAdapter to populate the DataSet with the results $dataAdapter = new-object "System.Data.SqlClient.SqlDataAdapter" ($query, $ServerConnection) $dataAdapter.Fill($dataSet) | Out-Null # Close the connection $ServerConnection.Close() # Populate the DataTable $dataTable = new-object "System.Data.DataTable" "Servers" $dataTable = $dataSet.Tables[0] #For every row in the DataTable $dataTable | FOREACH-OBJECT { "Server Name: $($_.serverName)" "Database Name: $($_.databaseName)" "Environment: $($_.environment)" # Compare the scripts folder to the database and synchronize the database to match # NB. Have set SQL Compare to abort on medium level warnings. $arguments = @("/scripts1:$($scriptsPath)", "/server2:$($_.serverName)", "/database2:$($_.databaseName)", "/AbortOnWarnings:Medium") # + @("/sync" ) # Commented out the 'sync' parameter for safety, write-host $arguments & $SQLComparePath $arguments "Exit Code: $LASTEXITCODE" # Some interesting variations # Check that every database matches a folder. # For example this might be a pre-deployment step to validate everything is at the same baseline state. # Or a post deployment script to validate the deployment worked. # An exit code of 0 means the databases are identical. # # $arguments = @("/scripts1:$($scriptsPath)", "/server2:$($_.serverName)", "/database2:$($_.databaseName)", "/Assertidentical") # Generate a report of the difference between the folder and each database. Generate a SQL update script for each database. # For example use this after the above to generate upgrade scripts for each database # Examine the warnings and the HTML diff report to understand how the script will change objects # #$arguments = @("/scripts1:$($scriptsPath)", "/server2:$($_.serverName)", "/database2:$($_.databaseName)", "/ScriptFile:update_$($_.environment+"_"+$_.databaseName).sql", "/report:update_$($_.environment+"_"+$_.databaseName).html" , "/reportType:Interactive", "/showWarnings", "/include:Identical") } It’s worth noting that the above example generates the deployment scripts dynamically. This approach should be problem-free for the vast majority of changes, but it is still good practice to review and test a pre-generated deployment script prior to deployment. An alternative approach would be to pre-generate a single deployment script using SQL Compare, and run this en masse to multiple targets programmatically using sqlcmd, or using a tool like SQL Multi Script. You can use the /ScriptFile, /report, and /showWarnings flags to generate change scripts, difference reports and any warnings. See the commented out example in the PowerShell: #$arguments = @("/scripts1:$($scriptsPath)", "/server2:$($_.serverName)", "/database2:$($_.databaseName)", "/ScriptFile:update_$($_.environment+"_"+$_.databaseName).sql", "/report:update_$($_.environment+"_"+$_.databaseName).html" , "/reportType:Interactive", "/showWarnings", "/include:Identical") There is a drawback of running a pre-generated deployment script; it assumes that a given database target hasn’t drifted from its expected state. Often there are (rightly or wrongly) many individuals within an organization who have permissions to alter the production database, and changes can therefore be made outside of the prescribed development processes. The consequence is that at deployment time, the applied script has been validated against a target that no longer represents reality. The solution here would be to add a check for drift prior to running the deployment script. This is achieved by using sqlcompare.exe to compare the target against the expected schema snapshot using the /Assertidentical flag. Should this return any differences (sqlcompare.exe Exit Code 79), a drift report is outputted instead of executing the deployment script. See the commented out example. # $arguments = @("/scripts1:$($scriptsPath)", "/server2:$($_.serverName)", "/database2:$($_.databaseName)", "/Assertidentical") Any checks and processes that should be undertaken prior to a manual deployment, should also be happen during an automated deployment. You might think about triggering backups prior to deployment – even better, automate the verification of the backup too. You can use SQL Compare’s command line interface along with PowerShell to automate multiple actions and checks that you need in your deployment process. Automation is a practical solution where multiple targets and a higher release cadence come into play. As we know, with great power comes great responsibility – responsibility to ensure that the necessary checks are made so deployments remain trouble-free. (The code sample supplied in this post automates the simple dynamic deployment case – if you are considering more advanced automation, e.g. the drift checks, script generation, deploying to large numbers of targets and backup/verification, please email me at [email protected] for further script samples or if you have further questions)

Read the article
Securing an ADF Application using OES11g: Part 1

- by user12587121

Future releases of the Oracle stack should allow ADF applications to be secured natively with Oracle Entitlements Server (OES). In a sequence of postings here I explore one way to achive this with the current technology, namely OES 11.1.1.5 and ADF 11.1.1.6. ADF Security Basics ADF Bascis The Application Development Framework (ADF) is Oracle’s preferred technology for developing GUI based Java applications. It can be used to develop a UI for Swing applications or, more typically in the Oracle stack, for Web and J2EE applications. ADF is based on and extends the Java Server Faces (JSF) technology. To get an idea, Oracle provides an online demo to showcase ADF components. ADF can be used to develop just the UI part of an application, where, for example, the data access layer is implemented using some custom Java beans or EJBs. However ADF also has it’s own data access layer, ADF Business Components (ADF BC) that will allow rapid integration of data from data bases and Webservice interfaces to the ADF UI component. In this way ADF helps implement the MVC approach to building applications with UI and data components. The canonical tutorial for ADF is to open JDeveloper, define a connection to a database, drag and drop a table from the database view to a UI page, build and deploy. One has an application up and running very quickly with the ability to quickly integrate changes to, for example, the DB schema. ADF allows web pages to be created graphically and components like tables, forms, text fields, graphs and so on to be easily added to a page. On top of JSF Oracle have added drag and drop tooling with JDeveloper and declarative binding of the UI to the data layer, be it database, WebService or Java beans. An important addition is the bounded task flow which is a reusable set of pages and transitions. ADF adds some steps to the page lifecycle defined in JSF and adds extra widgets including powerful visualizations. It is worth pointing out that the Oracle Web Center product (portal, content management and so on) is based on and extends ADF. ADF Security ADF comes with it’s own security mechanism that is exposed by JDeveloper at development time and in the WLS Console and Enterprise Manager (EM) at run time. The security elements that need to be addressed in an ADF application are: authentication, authorization of access to web pages, task-flows, components within the pages and data being returned from the model layer. One typically relies on WLS to handle authentication and because of this users and groups will also be handled by WLS. Typically in a Dev environment, users and groups are stored in the WLS embedded LDAP server. One has a choice when enabling ADF security (Application->Secure->Configure ADF Security) about whether to turn on ADF authorization checking or not: In the case where authorization is enabled for ADF one defines a set of roles in which we place users and then we grant access to these roles to the different ADF elements (pages or task flows or elements in a page). An important notion here is the difference between Enterprise Roles and Application Roles. The idea behind an enterprise role is that is defined in terms of users and LDAP groups from the WLS identity store. “Enterprise” in the sense that these are things available for use to all applications that use that store. The other kind of role is an Application Role and the idea is that a given application will make use of Enterprise roles and users to build up a set of roles for it’s own use. These application roles will be available only to that application. The general idea here is that the enterprise roles are relatively static (for example an Employees group in the LDAP directory) while application roles are more dynamic, possibly depending on time, location, accessed resource and so on. One of the things that OES adds that is that we can define these dynamic membership conditions in Role Mapping Policies. To make this concrete, here is how, at design time in Jdeveloper, one assigns these rights in Jdeveloper, which puts them into a file called jazn-data.xml: When the ADF app is deployed to a WLS this JAZN security data is pushed to the system-jazn-data.xml file of the WLS deployment for the policies and application roles and to the WLS backing LDAP for the users and enterprise roles. Note the difference here: after deploying the application we will see the users and enterprise roles show up in the WLS LDAP server. But the policies and application roles are defined in the system-jazn-data.xml file. Consult the embedded WLS LDAP server to manage users and enterprise roles by going to the domain console and then Security Realms->myrealm->Users and Groups: For production environments (or in future to share this data with OES) one would then perform the operation of “reassociating” this security policy and application role data to a DB schema (or an LDAP). This is done in the EM console by reassociating the Security Provider. This blog posting has more explanations and references on this reassociation process. If ADF Authentication and Authorization are enabled then the Security Policies for a deployed application can be managed in EM. Our goal is to be able to manage security policies for the applicaiton rather via OES and it's console. Security Requirements for an ADF Application With this package tour of ADF security we can see that to secure an ADF application with we would expect to be able to take care of at least the following items: Authentication, including a user and user-group store Authorization for page access Authorization for bounded Task Flow access. A bounded task flow has only one point of entry and so if we protect that entry point by calling to OES then all the pages in the flow are protected. Authorization for viewing data coming from the data access layer In the next posting we will describe a sample ADF application and required security policies. References ADF Dev Guide: Fusion Middleware Fusion Developer's Guide for Oracle Application Development Framework: Enabling ADF Security in a Fusion Web Application Oracle tutorial on securing a sample ADF application, appears to require ADF 11.1.2 Normal 0 false false false EN-US X-NONE X-NONE MicrosoftInternetExplorer4 /* Style Definitions */ table.MsoNormalTable {mso-style-name:"Table Normal"; mso-tstyle-rowband-size:0; mso-tstyle-colband-size:0; mso-style-noshow:yes; mso-style-priority:99; mso-style-qformat:yes; mso-style-parent:""; mso-padding-alt:0in 5.4pt 0in 5.4pt; mso-para-margin:0in; mso-para-margin-bottom:.0001pt; mso-pagination:widow-orphan; font-size:10.0pt; font-family:"Calibri","sans-serif"; mso-bidi-font-family:"Times New Roman";} Normal 0 false false false EN-US X-NONE X-NONE MicrosoftInternetExplorer4 /* Style Definitions */ table.MsoNormalTable {mso-style-name:"Table Normal"; mso-tstyle-rowband-size:0; mso-tstyle-colband-size:0; mso-style-noshow:yes; mso-style-priority:99; mso-style-qformat:yes; mso-style-parent:""; mso-padding-alt:0in 5.4pt 0in 5.4pt; mso-para-margin:0in; mso-para-margin-bottom:.0001pt; mso-pagination:widow-orphan; font-size:11.0pt; font-family:"Calibri","sans-serif"; mso-ascii-font-family:Calibri; mso-ascii-theme-font:minor-latin; mso-fareast-font-family:"Times New Roman"; mso-fareast-theme-font:minor-fareast; mso-hansi-font-family:Calibri; mso-hansi-theme-font:minor-latin; mso-bidi-font-family:"Times New Roman"; mso-bidi-theme-font:minor-bidi;}

Read the article
Experiences with learning Chinese

- by Greg Low

I've had a few friends asking me about learning Chinese and what I've found works and doesn't work. I was answering a question on a mailing list today and I thought I should post this info where it might be useful to many. The question that was initially asked was whether Rosetta Stone was useful but I've provided much more info on learning the language here. I’ve used Rosetta Stone with Chinese but it’s really hard to know whether to recommend it or not. Rosetta Stone works the same way in all languages. They show you photos and then let you both see and hear the target language and get you to work out what they’re talking about. The thinking is that that’s how children learn. However, at first, I found it very frustrating. I’d be staring at photos trying to work out what they were really trying to get at. Sometimes it’s far from obvious. I could not have survived without Google Translate open at the same time. The other weird thing is that the photos are from a mixture of countries. While that’s good in a way, it also means that they are endlessly showing pictures of something that would never happen in the target language and culture. For any language, constant interaction with a speaker of the target language is needed. Rosetta Stone has a “Studio” option. That’s the best part of the program. In my case, it lets me connect around twice a week to a live online class from Beijing. Classes usually have the teacher plus two to four students. You get some Studio access with the initial packages but need to purchase it for ongoing use. I find it very inexpensive. It seems to work out to about $70 (AUD/USD) for six months. That’s a real bargain. The other downside to Rosetta Stone is that they tend to teach very formal language, but as with other languages, that’s not how the locals speak. It might have been correct at one point but no-one actually says that. As an example, Rosetta Stone teach Gonggòng qìche (pronounced roughly like “gong gong chee chure” for bus. Most of my friends from areas like Taiwan would just say Gongche. Google Translate says Zongxiàn (pronounced somewhat like “dzong sheean”) instead. Mind you, the Rosetta Stone option isn't really as bad as "omnibus"; it's more like saying "public bus". If you say the option they provide, people would understand you. I also listen to ChinesePod in the car. They also have SpanishPod. Each podcast is about five minutes of spoken conversation. It is very good for providing current language. Another resource I use is local Meetup groups. Most cities have these and for a variety of languages. It’s way less structured (just standard conversation) but good for getting interaction. The obvious challenge for Asian languages is reading/writing. The input editors for Chinese that are part of Windows are excellent. Many of my Chinese friends speak fluently but cannot read or write. I was determined to learn to do both. For writing, I’m talking about on a computer, not with a pen. (Mind you, I can barely write English with a pen nowadays). When using Rosetta Stone, you can choose to have the Chinese words displayed in pinyin (Wo xihuan xuéxí zhongguó) or in Chinese characters (???????) or both. This year, I’ve been forcing myself to just use the Chinese characters. I use a pinyin input editor in Windows though, as it’s very fast. (The character recognition input in the iPad is also amazing). Notice from the example that I provided above that the pronunciation of the pinyin isn’t that obvious to us at first either. Since changing to only using characters, I find I can now read many more Chinese characters fluently. It’s a major challenge though. I can read about 300 now and yet you need around 2,500 to be able to read a newspaper fairly well. Tones are a major issue for some Asian languages. Mandarin has four tones (plus a neutral tone) and there is a major difference in meaning between two words that are spelled the same in pinyin but with different tones. For example, Ma (3rd tone?) is a horse, Ma (1st tone?) is like “mom”, and ma (neutral tone?) is a question mark and so on. Clearly you don’t want to mix these up. As in English, they also have words that do sound the same but mean different things in different contexts. What’s interesting is that even though we see two words that differ only by tone as very similar, to a native speaker, if you say the right words with the wrong tone, you might as well have said a completely different word. My wife’s dialect of Chinese has eight tones. It’s much worse. The reason I’m so keen to learn to read/write Chinese is that even though the different dialects are pronounced so differently that speakers of one dialect often cannot understand another dialect, the writing is generally the same. The only difference is that many years ago, the Chinese government created a simplified set of characters for some of the most commonly used ones. Older Chinese and most Cantonese speakers often struggle with the simplified characters. This is the simplified form of “three apples”: ???? This is the traditional form of the same words: ???? Note that two of the characters are the same but the middle two are quite different. For most languages, the best thing is to watch current movies in the target language but to watch them with the target language as subtitles, not your native language. You want to know what they actually said, not what it roughly means (which is what the English subtitle would give you). The difficulty with Asian languages like Chinese is that you have the added challenge of understanding the subtitles when they are written in the target language. I wish there were Mandarin Chinese movies with pinyin subtitles. For learning to read characters, I also recommend HanCard on the iPad. It is targeted at the HSK language proficiency levels. (I’m intending to take the first HSK exam as soon as I’m ready). Hope that info helps someone get started.

Read the article
Why does Google mark one e-mail as spam while does not the other?

- by nKn

I've a Postfix installation which works fine, I don't get any trouble with mails sent through a mail client (in my case, Thunderbird or RoundCube) when the To: address is a GMail account. However, I recently needed to use the PHPMailer tool to send some e-mails to some GMail accounts, so I configured an account to be used via SASL authentication + TLS. I don't mean mass mailing, just 2-3 mails. If I send the e-mail from the Thunderbird or RoundCube clients, the mail is not marked as spam. However, if I use PHPMailer, it always gets catalogued as spam. So I compared both headers and I just can't find the reason why the second is marked as spam while the first one is just ok. The first header sent from a mail client which is not marked as spam: Delivered-To: [email protected] Received: by 10.76.153.102 with SMTP id vf6csp230573oab; Tue, 19 Aug 2014 11:08:19 -0700 (PDT) X-Received: by 10.60.23.39 with SMTP id j7mr45544050oef.20.1408471699715; Tue, 19 Aug 2014 11:08:19 -0700 (PDT) Return-Path: <[email protected]> Received: from mail.mydomain.com (X.ip-92-222-X.eu. [92.222.X.X]) by mx.google.com with ESMTPS id t5si27115082oej.10.2014.08.19.11.08.18 for <[email protected]> (version=TLSv1.2 cipher=ECDHE-RSA-AES128-GCM-SHA256 bits=128/128); Tue, 19 Aug 2014 11:08:19 -0700 (PDT) Received-SPF: pass (google.com: domain of [email protected] designates 92.222.X.X as permitted sender) client-ip=92.222.X.X; Authentication-Results: mx.google.com; spf=pass (google.com: domain of [email protected] designates 92.222.X.X as permitted sender) [email protected]; dkim=pass (test mode) [email protected] Received: by mail.mydomain.com (Postfix, from userid 111) id D8F69120293D; Tue, 19 Aug 2014 19:08:17 +0100 (BST) DKIM-Signature: v=1; a=rsa-sha256; c=relaxed/relaxed; d=mydomain.com; s=mail; t=1408471697; bh=wKMX9gkQ7tCLv8ezrG5t4bICm/SSLQsNfTdZMToksWw=; h=Date:From:To:Subject:From; b=qRNcYVdmk+n3D1uuv0FInTx7/LzH2ojck9DgCmabFPvfke233lkojUOjezCUGx7iV DL8EayZ28mzzzHpB7ETeMzop/5OS3BmvFtGKVD9gzc78cDIFXTDoRFAnkRWDR2IOxI SOn5tiyODTFpkbDgJOndzQ6qL5K0S9ASNGCZrNL4= X-Spam-Checker-Version: SpamAssassin 3.4.0 (2014-02-07) on vpsX.ovh.net X-Spam-Level: X-Spam-Status: No, score=-1.0 required=3.0 tests=ALL_TRUSTED,T_DKIM_INVALID autolearn=ham autolearn_force=no version=3.4.0 Received: from [192.168.1.111] (unknown [77.231.X.X]) (using TLSv1 with cipher ECDHE-RSA-AES128-SHA (128/128 bits)) (No client certificate requested) (Authenticated sender: [email protected]) by mail.mydomain.com (Postfix) with ESMTPSA id 910341202624 for <[email protected]>; Tue, 19 Aug 2014 19:08:17 +0100 (BST) DKIM-Signature: v=1; a=rsa-sha256; c=relaxed/relaxed; d=mydomain.com; s=mail; t=1408471697; bh=wKMX9gkQ7tCLv8ezrG5t4bICm/SSLQsNfTdZMToksWw=; h=Date:From:To:Subject:From; b=qRNcYVdmk+n3D1uuv0FInTx7/LzH2ojck9DgCmabFPvfke233lkojUOjezCUGx7iV DL8EayZ28mzzzHpB7ETeMzop/5OS3BmvFtGKVD9gzc78cDIFXTDoRFAnkRWDR2IOxI SOn5tiyODTFpkbDgJOndzQ6qL5K0S9ASNGCZrNL4= Message-ID: <[email protected]> Date: Tue, 19 Aug 2014 19:08:24 +0100 From: My Name <[email protected]> User-Agent: Mozilla/5.0 (Windows NT 6.1; WOW64; rv:24.0) Gecko/20100101 Thunderbird/24.6.0 MIME-Version: 1.0 To: My other account <[email protected]> Subject: . Content-Type: text/plain; charset=ISO-8859-1; format=flowed Content-Transfer-Encoding: 7bit . The second header sent from PHPMailer which is always marked as spam: Delivered-To: [email protected] Received: by 10.76.153.102 with SMTP id vf6csp230832oab; Tue, 19 Aug 2014 11:12:10 -0700 (PDT) X-Received: by 10.60.121.67 with SMTP id li3mr44086252oeb.17.1408471930520; Tue, 19 Aug 2014 11:12:10 -0700 (PDT) Return-Path: <[email protected]> Received: from mail.mydomain.com (X.ip-92-222-X.eu. [92.222.X.X]) by mx.google.com with ESMTPS id w8si27103806obn.30.2014.08.19.11.12.10 for <[email protected]> (version=TLSv1.2 cipher=ECDHE-RSA-AES128-GCM-SHA256 bits=128/128); Tue, 19 Aug 2014 11:12:10 -0700 (PDT) Received-SPF: pass (google.com: domain of [email protected] designates 92.222.X.X as permitted sender) client-ip=92.222.X.X; Authentication-Results: mx.google.com; spf=pass (google.com: domain of [email protected] designates 92.222.X.X as permitted sender) [email protected]; dkim=pass (test mode) [email protected] Received: by mail.mydomain.com (Postfix, from userid 111) id 1999D120293D; Tue, 19 Aug 2014 19:12:09 +0100 (BST) DKIM-Signature: v=1; a=rsa-sha256; c=relaxed/relaxed; d=mydomain.com; s=mail; t=1408471929; bh=N1JuHq1S+8GrjHcEK3xn8P1JS+ygEBv5LKe0BiXuVJo=; h=Date:To:From:Reply-to:Subject:From; b=K7tcPyArzSTY91VEw6mAAFtDurSGwgTLGkfUZdC5mqsg0g/1LzmZkgwdjj4NdJa6M E2kDz3dwYN8FcZmbampJYFXxj4NQVtSnzjiWV40rpfOFqD2rXDGNIyB2QOjBZZ4WK3 7s4lyoJ/BrdQH4en8ctLVsDHed/KpHD4iGFEl67E= X-Spam-Checker-Version: SpamAssassin 3.4.0 (2014-02-07) on vpsX.ovh.net X-Spam-Level: X-Spam-Status: No, score=-1.0 required=3.0 tests=ALL_TRUSTED,T_DKIM_INVALID autolearn=ham autolearn_force=no version=3.4.0 Received: from rpi.mydomain.com (unknown [77.231.X.X]) (using TLSv1 with cipher ECDHE-RSA-AES256-SHA (256/256 bits)) (No client certificate requested) (Authenticated sender: [email protected]) by mail.mydomain.com (Postfix) with ESMTPSA id B42AF1202624 for <[email protected]>; Tue, 19 Aug 2014 19:12:08 +0100 (BST) DKIM-Signature: v=1; a=rsa-sha256; c=relaxed/relaxed; d=mydomain.com; s=mail; t=1408471928; bh=N1JuHq1S+8GrjHcEK3xn8P1JS+ygEBv5LKe0BiXuVJo=; h=Date:To:From:Reply-to:Subject:From; b=iXPM0tS36swudPTT4FOHHtPi5Ll6LbR60kNqCinZ8utcWoFE31SFTpoMEq5aCM5ux wQMdFiN8c6vkjRGabmvqFTTIbwJsrToHo/4+Lt5HEBoQQE2Y3T+xGmnmGAHCS6stKB yb7SVmtrIAsVtSMKA8VYIbmu2oYqV3afYt7g0OMQ= Date: Tue, 19 Aug 2014 20:12:07 +0200 To: [email protected] From: Trying another account <[email protected]> Reply-to: Trying another account <[email protected]> Subject: . Message-ID: <[email protected]> X-Priority: 3 X-Mailer: PHPMailer 5.1 (phpmailer.sourceforge.net) MIME-Version: 1.0 Content-Transfer-Encoding: 8bit Content-Type: text/plain; charset="UTF-8" . I also tried: Adding a User-Agent header to match the first one. Removing the X-Mailer header. No one of them made a difference. Is there some significant difference which is making the second e-mail to be marked as spam by Google?

Read the article
Why Is Vertical Resolution Monitor Resolution so Often a Multiple of 360?

- by Jason Fitzpatrick

Stare at a list of monitor resolutions long enough and you might notice a pattern: many of the vertical resolutions, especially those of gaming or multimedia displays, are multiples of 360 (720, 1080, 1440, etc.) But why exactly is this the case? Is it arbitrary or is there something more at work? Today’s Question & Answer session comes to us courtesy of SuperUser—a subdivision of Stack Exchange, a community-driven grouping of Q&A web sites. The Question SuperUser reader Trojandestroy recently noticed something about his display interface and needs answers: YouTube recently added 1440p functionality, and for the first time I realized that all (most?) vertical resolutions are multiples of 360. Is this just because the smallest common resolution is 480×360, and it’s convenient to use multiples? (Not doubting that multiples are convenient.) And/or was that the first viewable/conveniently sized resolution, so hardware (TVs, monitors, etc) grew with 360 in mind? Taking it further, why not have a square resolution? Or something else unusual? (Assuming it’s usual enough that it’s viewable). Is it merely a pleasing-the-eye situation? So why have the display be a multiple of 360? The Answer SuperUser contributor User26129 offers us not just an answer as to why the numerical pattern exists but a history of screen design in the process: Alright, there are a couple of questions and a lot of factors here. Resolutions are a really interesting field of psychooptics meeting marketing. First of all, why are the vertical resolutions on youtube multiples of 360. This is of course just arbitrary, there is no real reason this is the case. The reason is that resolution here is not the limiting factor for Youtube videos – bandwidth is. Youtube has to re-encode every video that is uploaded a couple of times, and tries to use as little re-encoding formats/bitrates/resolutions as possible to cover all the different use cases. For low-res mobile devices they have 360×240, for higher res mobile there’s 480p, and for the computer crowd there is 360p for 2xISDN/multiuser landlines, 720p for DSL and 1080p for higher speed internet. For a while there were some other codecs than h.264, but these are slowly being phased out with h.264 having essentially ‘won’ the format war and all computers being outfitted with hardware codecs for this. Now, there is some interesting psychooptics going on as well. As I said: resolution isn’t everything. 720p with really strong compression can and will look worse than 240p at a very high bitrate. But on the other side of the spectrum: throwing more bits at a certain resolution doesn’t magically make it better beyond some point. There is an optimum here, which of course depends on both resolution and codec. In general: the optimal bitrate is actually proportional to the resolution. So the next question is: what kind of resolution steps make sense? Apparently, people need about a 2x increase in resolution to really see (and prefer) a marked difference. Anything less than that and many people will simply not bother with the higher bitrates, they’d rather use their bandwidth for other stuff. This has been researched quite a long time ago and is the big reason why we went from 720×576 (415kpix) to 1280×720 (922kpix), and then again from 1280×720 to 1920×1080 (2MP). Stuff in between is not a viable optimization target. And again, 1440P is about 3.7MP, another ~2x increase over HD. You will see a difference there. 4K is the next step after that. Next up is that magical number of 360 vertical pixels. Actually, the magic number is 120 or 128. All resolutions are some kind of multiple of 120 pixels nowadays, back in the day they used to be multiples of 128. This is something that just grew out of LCD panel industry. LCD panels use what are called line drivers, little chips that sit on the sides of your LCD screen that control how bright each subpixel is. Because historically, for reasons I don’t really know for sure, probably memory constraints, these multiple-of-128 or multiple-of-120 resolutions already existed, the industry standard line drivers became drivers with 360 line outputs (1 per subpixel). If you would tear down your 1920×1080 screen, I would be putting money on there being 16 line drivers on the top/bottom and 9 on one of the sides. Oh hey, that’s 16:9. Guess how obvious that resolution choice was back when 16:9 was ‘invented’. Then there’s the issue of aspect ratio. This is really a completely different field of psychology, but it boils down to: historically, people have believed and measured that we have a sort of wide-screen view of the world. Naturally, people believed that the most natural representation of data on a screen would be in a wide-screen view, and this is where the great anamorphic revolution of the ’60s came from when films were shot in ever wider aspect ratios. Since then, this kind of knowledge has been refined and mostly debunked. Yes, we do have a wide-angle view, but the area where we can actually see sharply – the center of our vision – is fairly round. Slightly elliptical and squashed, but not really more than about 4:3 or 3:2. So for detailed viewing, for instance for reading text on a screen, you can utilize most of your detail vision by employing an almost-square screen, a bit like the screens up to the mid-2000s. However, again this is not how marketing took it. Computers in ye olden days were used mostly for productivity and detailed work, but as they commoditized and as the computer as media consumption device evolved, people didn’t necessarily use their computer for work most of the time. They used it to watch media content: movies, television series and photos. And for that kind of viewing, you get the most ‘immersion factor’ if the screen fills as much of your vision (including your peripheral vision) as possible. Which means widescreen. But there’s more marketing still. When detail work was still an important factor, people cared about resolution. As many pixels as possible on the screen. SGI was selling almost-4K CRTs! The most optimal way to get the maximum amount of pixels out of a glass substrate is to cut it as square as possible. 1:1 or 4:3 screens have the most pixels per diagonal inch. But with displays becoming more consumery, inch-size became more important, not amount of pixels. And this is a completely different optimization target. To get the most diagonal inches out of a substrate, you want to make the screen as wide as possible. First we got 16:10, then 16:9 and there have been moderately successful panel manufacturers making 22:9 and 2:1 screens (like Philips). Even though pixel density and absolute resolution went down for a couple of years, inch-sizes went up and that’s what sold. Why buy a 19″ 1280×1024 when you can buy a 21″ 1366×768? Eh… I think that about covers all the major aspects here. There’s more of course; bandwidth limits of HDMI, DVI, DP and of course VGA played a role, and if you go back to the pre-2000s, graphics memory, in-computer bandwdith and simply the limits of commercially available RAMDACs played an important role. But for today’s considerations, this is about all you need to know. Have something to add to the explanation? Sound off in the the comments. Want to read more answers from other tech-savvy Stack Exchange users? Check out the full discussion thread here.

Read the article

Search Results

Search found 8268 results on 331 pages for 'difference'.

Page 160/331 | < Previous Page | 156 157 158 159 160 161 162 163 164 165 166 167 | Next Page >

- by m00st

- by Bart

- by RussellDias

- by Matt

- by david

- by slicedlime

- by Vi

- by HaydnWVN

- by Harold Smith

- by Richard

- by Johnmph

- by pinaldave

- by Sweeters

- by Robert May

- by nmarun

- by James Michael Hare

- by Pinal Dave

- by [email protected]

- by jsavit

- by costlow

- by Jonathan Hickford

- by user12587121

- by Greg Low

- by nKn

- by Jason Fitzpatrick

< Previous Page | 156 157 158 159 160 161 162 163 164 165 166 167 | Next Page >