active pattern - Page 491

EC2 instance suddenly refusing SSH connections and won't respond to ping

- by Chris

My instance was running fine and this morning I was able to access a Ruby on Rails app hosted on it. An hour later I suddenly wasn't able to access my site, my SSH connection attempts were refused and the server wasn't even responding to ping. I didn't change anything on my system during that hour and reboots aren't fixing it. I've never had any problems connecting or pinging the system before. Can someone please help? This is on my production system! OS: CentOS 5 AMI ID: ami-10b55379 Type: m1.small [] ~% ssh -v *****@meeteor.com OpenSSH_5.2p1, OpenSSL 0.9.8l 5 Nov 2009 debug1: Reading configuration data /etc/ssh_config debug1: Connecting to meeteor.com [184.73.235.191] port 22. debug1: connect to address 184.73.235.191 port 22: Connection refused ssh: connect to host meeteor.com port 22: Connection refused [] ~% ping meeteor.com PING meeteor.com (184.73.235.191): 56 data bytes Request timeout for icmp_seq 0 Request timeout for icmp_seq 1 Request timeout for icmp_seq 2 ^C --- meeteor.com ping statistics --- 4 packets transmitted, 0 packets received, 100.0% packet loss [] ~% ========= System Log ========= Restarting system. Linux version 2.6.16-xenU ([email protected]) (gcc version 4.0.1 20050727 (Red Hat 4.0.1-5)) #1 SMP Mon May 28 03:41:49 SAST 2007 BIOS-provided physical RAM map: Xen: 0000000000000000 - 000000006a400000 (usable) 980MB HIGHMEM available. 727MB LOWMEM available. NX (Execute Disable) protection: active IRQ lockup detection disabled Built 1 zonelists Kernel command line: root=/dev/sda1 ro 4 Enabling fast FPU save and restore... done. Enabling unmasked SIMD FPU exception support... done. Initializing CPU#0 PID hash table entries: 4096 (order: 12, 65536 bytes) Xen reported: 2599.998 MHz processor. Dentry cache hash table entries: 131072 (order: 7, 524288 bytes) Inode-cache hash table entries: 65536 (order: 6, 262144 bytes) Software IO TLB disabled vmalloc area: ee000000-f53fe000, maxmem 2d7fe000 Memory: 1718700k/1748992k available (1958k kernel code, 20948k reserved, 620k data, 144k init, 1003528k highmem) Checking if this processor honours the WP bit even in supervisor mode... Ok. Calibrating delay using timer specific routine.. 5202.30 BogoMIPS (lpj=26011526) Mount-cache hash table entries: 512 CPU: L1 I Cache: 64K (64 bytes/line), D cache 64K (64 bytes/line) CPU: L2 Cache: 1024K (64 bytes/line) Checking 'hlt' instruction... OK. Brought up 1 CPUs migration_cost=0 Grant table initialized NET: Registered protocol family 16 Brought up 1 CPUs xen_mem: Initialising balloon driver. highmem bounce pool size: 64 pages VFS: Disk quotas dquot_6.5.1 Dquot-cache hash table entries: 1024 (order 0, 4096 bytes) Initializing Cryptographic API io scheduler noop registered io scheduler anticipatory registered (default) io scheduler deadline registered io scheduler cfq registered i8042.c: No controller found. RAMDISK driver initialized: 16 RAM disks of 4096K size 1024 blocksize Xen virtual console successfully installed as tty1 Event-channel device installed. netfront: Initialising virtual ethernet driver. mice: PS/2 mouse device common for all mice md: md driver 0.90.3 MAX_MD_DEVS=256, MD_SB_DISKS=27 md: bitmap version 4.39 NET: Registered protocol family 2 Registering block device major 8 IP route cache hash table entries: 65536 (order: 6, 262144 bytes) TCP established hash table entries: 262144 (order: 9, 2097152 bytes) TCP bind hash table entries: 65536 (order: 7, 524288 bytes) TCP: Hash tables configured (established 262144 bind 65536) TCP reno registered TCP bic registered NET: Registered protocol family 1 NET: Registered protocol family 17 NET: Registered protocol family 15 Using IPI No-Shortcut mode md: Autodetecting RAID arrays. md: autorun ... md: ... autorun DONE. kjournald starting. Commit interval 5 seconds EXT3-fs: mounted filesystem with ordered data mode. VFS: Mounted root (ext3 filesystem) readonly. Freeing unused kernel memory: 144k freed *************************************************************** *************************************************************** ** WARNING: Currently emulating unsupported memory accesses ** ** in /lib/tls glibc libraries. The emulation is ** ** slow. To ensure full performance you should ** ** install a 'xen-friendly' (nosegneg) version of ** ** the library, or disable tls support by executing ** ** the following as root: ** ** mv /lib/tls /lib/tls.disabled ** ** Offending process: init (pid=1) ** *************************************************************** *************************************************************** Pausing... 5Pausing... 4Pausing... 3Pausing... 2Pausing... 1Continuing... INIT: version 2.86 booting Welcome to CentOS release 5.4 (Final) Press 'I' to enter interactive startup. Setting clock : Fri Oct 1 14:35:26 EDT 2010 [ OK ] Starting udev: [ OK ] Setting hostname localhost.localdomain: [ OK ] No devices found Setting up Logical Volume Management: [ OK ] Checking filesystems Checking all file systems. [/sbin/fsck.ext3 (1) -- /] fsck.ext3 -a /dev/sda1 /dev/sda1: clean, 275424/1310720 files, 1161123/2621440 blocks [ OK ] Remounting root filesystem in read-write mode: [ OK ] Mounting local filesystems: [ OK ] Enabling local filesystem quotas: [ OK ] Enabling /etc/fstab swaps: [ OK ] INIT: Entering runlevel: 4 Entering non-interactive startup Starting background readahead: [ OK ] Applying ip6tables firewall rules: modprobe: FATAL: Module ip6_tables not found. ip6tables-restore v1.3.5: ip6tables-restore: unable to initializetable 'filter' Error occurred at line: 3 Try `ip6tables-restore -h' or 'ip6tables-restore --help' for more information. [FAILED] Applying iptables firewall rules: [ OK ] Loading additional iptables modules: ip_conntrack_netbios_ns [ OK ] Bringing up loopback interface: [ OK ] Bringing up interface eth0: Determining IP information for eth0... done. [ OK ] Starting auditd: [FAILED] Starting irqbalance: [ OK ] Starting portmap: [ OK ] FATAL: Module lockd not found. Starting NFS statd: [ OK ] Starting RPC idmapd: FATAL: Module sunrpc not found. FATAL: Error running install command for sunrpc Error: RPC MTAB does not exist. Starting system message bus: [ OK ] Starting Bluetooth services:[ OK ] [ OK ] Can't open RFCOMM control socket: Address family not supported by protocol Mounting other filesystems: [ OK ] Starting PC/SC smart card daemon (pcscd): [ OK ] Starting hidd: Can't open HIDP control socket: Address family not supported by protocol [FAILED] Starting autofs: Starting automount: automount: test mount forbidden or incorrect kernel protocol version, kernel protocol version 5.00 or above required. [FAILED] [FAILED] Starting sshd: [ OK ] Starting cups: [ OK ] Starting sendmail: [ OK ] Starting sm-client: [ OK ] Starting console mouse services: no console device found[FAILED] Starting crond: [ OK ] Starting xfs: [ OK ] Starting anacron: [ OK ] Starting atd: [ OK ] % Total % Received % Xferd Average Speed Time Time Time Current Dload Upload Total Spent Left Speed 100 390 100 390 0 0 58130 0 --:--:-- --:--:-- --:--:-- 58130 100 390 100 390 0 0 56984 0 --:--:-- --:--:-- --:--:-- 0 Starting yum-updatesd: [ OK ] Starting Avahi daemon... [ OK ] Starting HAL daemon: [ OK ] Starting OSSEC: [ OK ] Starting smartd: [ OK ] c CentOS release 5.4 (Final) Kernel 2.6.16-xenU on an i686 domU-12-31-39-00-C4-97 login: INIT: Id "2" respawning too fast: disabled for 5 minutes INIT: Id "3" respawning too fast: disabled for 5 minutes INIT: Id "4" respawning too fast: disabled for 5 minutes INIT: Id "5" respawning too fast: disabled for 5 minutes INIT: Id "6" respawning too fast: disabled for 5 minutes

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solved: passenger(mod_rails) fails to start puppet master under nginx

- by Anadi Misra

On the server [root@bangvmpllDA02 logs]# ruby -v ruby 1.8.7 (2011-06-30 patchlevel 352) [x86_64-linux] [root@bangvmpllDA02 logs]# puppet --version 3.0.1 and [root@bangvmpllDA02 logs]# service nginx configtest nginx: the configuration file /apps/nginx/nginx.conf syntax is ok nginx: configuration file /apps/nginx/nginx.conf test is successful [root@bangvmpllDA02 logs]# service nginx status nginx (pid 25923 25921 25920 25917 25908) is running... [root@bangvmpllDA02 logs]# however none of my agents are able to connect to the master, they all fail with errors like so [amisr1@blramisr195602 ~]$ puppet agent --test --verbose --server bangvmpllda02.XXX.com Info: Creating a new SSL certificate request for blramisr195602.XXX.com Info: Certificate Request fingerprint (SHA256): 26:EB:08:1F:82:32:E4:03:7A:64:8E:30:A3:99:93:26:E6:66:B9:B0:49:B6:08:F9:67:CA:1B:0C:00:B9:1D:41 Error: Could not request certificate: Error 405 on SERVER: <html> <head><title>405 Not Allowed</title></head> <body bgcolor="white"> <center><h1>405 Not Allowed</h1></center> <hr><center>nginx</center> </body> </html> Exiting; failed to retrieve certificate and waitforcert is disabled when I check logs on puppet master [root@bangvmpllDA02 logs]# tail puppet_access.log [05/Dec/2012:17:45:18 +0530] "GET /production/certificate/ca? HTTP/1.1" 404 162 "-" "Ruby" [05/Dec/2012:18:32:23 +0530] "PUT /production/certificate_request/sl63anadi.XXX.com HTTP/1.1" 405 166 "-" "-" [05/Dec/2012:18:33:33 +0530] "GET /production/certificate/sl63anadi.XXX.com? HTTP/1.1" 404 162 "-" "-" [05/Dec/2012:18:33:33 +0530] "GET /production/certificate_request/sl63anadi.XXX.com? HTTP/1.1" 404 162 "-" "-" [05/Dec/2012:18:33:33 +0530] "PUT /production/certificate_request/sl63anadi.XXX.com HTTP/1.1" 405 166 "-" "-" and the error logs show that nginx is not really able to process the request well 2012/12/05 18:33:33 [error] 25920#0: *23 open() "/etc/puppet/rack/public/production/certificate/sl63anadi.XXX.com" failed (2: No such file or directory), client: 10.209.47.26, server: , request: "GET /production/certificate/sl63anadi.XXX.com? HTTP/1.1", host: "bangvmpllda02.XXX.com:8140" 2012/12/05 18:33:33 [error] 25920#0: *24 open() "/etc/puppet/rack/public/production/certificate_request/sl63anadi.XXX.com" failed (2: No such file or directory), client: 10.209.47.26, server: , request: "GET /production/certificate_request/sl63anadi.XXX.com? HTTP/1.1", host: "bangvmpllda02.XXX.com:8140" 2012/12/05 18:47:56 [error] 25923#0: *27 open() "/etc/puppet/rack/public/production/certificate/ca" failed (2: No such file or directory), client: 10.209.47.31, server: , request: "GET /production/certificate/ca? HTTP/1.1", host: "bangvmpllda02.XXX.com:8140" 2012/12/05 18:47:56 [error] 25923#0: *28 open() "/etc/puppet/rack/public/production/certificate_request/blramisr195602.XXX.com" failed (2: No such file or directory), client: 10.209.47.31, server: , request: "GET /production/certificate_request/blramisr195602.XXX.com? HTTP/1.1", host: "bangvmpllda02.XXX.com:8140" Passenger does not show any application groups either [root@bangvmpllDA02 nginx]# passenger-status ----------- General information ----------- max = 15 count = 0 active = 0 inactive = 0 Waiting on global queue: 0 ----------- Application groups ----------- [root@bangvmpllDA02 nginx]# here's my nginx configuration [root@bangvmpllDA02 logs]# cat ../nginx.conf user puppet; worker_processes 4; #error_log logs/error.log; #error_log logs/error.log notice; error_log logs/error.log info; #pid logs/nginx.pid; events { use epoll; worker_connections 1024; } http { include mime.types; default_type application/octet-stream; log_format main '$remote_addr - $remote_user [$time_local] "$request" ' '$status $body_bytes_sent "$http_referer" ' '"$http_user_agent" "$http_x_forwarded_for"'; access_log logs/access.log main; sendfile on; #tcp_nopush on; server_tokens off; #keepalive_timeout 0; keepalive_timeout 120; gzip on; gzip_http_version 1.1; gzip_disable "msie6"; gzip_vary on; gzip_min_length 1100; gzip_buffers 64 8k; gzip_comp_level 3; gzip_proxied any; gzip_types text/plain text/css application/x-javascript text/xml application/xml; server { listen 80; server_name bangvmpllda02.XXXX.com; charset utf-8; #access_log logs/http.access.log main; location / { root html; index index.html index.htm index.php; } #error_page 404 /404.html; # redirect server error pages to the static page /50x.html # error_page 500 502 503 504 /50x.html; location = /50x.html { root html; } # proxy the PHP scripts to Apache listening on 127.0.0.1:80 # #location ~ \.php$ { # proxy_pass http://127.0.0.1; #} # pass the PHP scripts to FastCGI server listening on 127.0.0.1:9000 # location ~ \.php$ { root html; fastcgi_pass unix:/var/run/php-fpm/php-fpm.sock; fastcgi_index index.php; fastcgi_param SCRIPT_FILENAME $document_root$fastcgi_script_name; fastcgi_param SCRIPT_NAME $fastcgi_script_name; include fastcgi_params; } # deny access to .htaccess files, if Apache's document root # concurs with nginx's one # location ~ /\.ht { access_log off; log_not_found off; deny all; } location ~* \.(jpg|jpeg|gif|png|css|js|ico|xml)$ { access_log off; log_not_found off; expires 2d; } } # Passenger needed for puppet passenger_root /usr/lib/ruby/gems/1.8/gems/passenger-3.0.18; passenger_ruby /usr/bin/ruby; passenger_max_pool_size 15; server { ssl on; listen 8140 default ssl; server_name bangvmpllda02.XXXX.com; passenger_enabled on; passenger_set_cgi_param HTTP_X_CLIENT_DN $ssl_client_s_dn; passenger_set_cgi_param HTTP_X_CLIENT_VERIFY $ssl_client_verify; passenger_min_instances 5; access_log logs/puppet_access.log; error_log logs/puppet_error.log; root /etc/puppet/rack/public; ssl_certificate /var/lib/puppet/ssl/certs/bangvmpllda02.XXX.com.pem; ssl_certificate_key /var/lib/puppet/ssl/private_keys/bangvmpllda02.XXX.com.pem; ssl_crl /var/lib/puppet/ssl/ca/ca_crl.pem; ssl_client_certificate /var/lib/puppet/ssl/certs/ca.pem; ssl_ciphers SSLv2:-LOW:-EXPORT:RC4+RSA; ssl_prefer_server_ciphers on; ssl_verify_client optional; ssl_verify_depth 1; ssl_session_cache shared:SSL:128m; ssl_session_timeout 5m; } } and the puppet.conf [main] # The Puppet log directory. # The default value is '$vardir/log'. logdir = /var/log/puppet # Where Puppet PID files are kept. # The default value is '$vardir/run'. rundir = /var/run/puppet dns_alt_names = devops.XXXX.com,devops confdir = /etc/puppet vardir = /var/lib/puppet storeconfigs = true storeconfigs_backend = puppetdb thin_storeconfigs = false async_storeconfigs = false ssl_client_header = SSL_CLIENT_S_D ssl_client_verify_header = SSL_CLIENT_VERIFY # Where SSL certificates are kept. # The default value is '$confdir/ssl'. ssldir = $vardir/ssl any ideas where am I going wrong? I checkthe directory permissions; /usr/share/puppet, /etc/puppet and /var/lib/puppet (and files inside them) are owned by puppet user. Solved The simple solution to my complicated problem was that I had placed the config.ru in wrong place moved it to /etc/puppet/rack , it was in /etc/puppet/rack/public Well!!! :-/

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Task scheduler ran a task twice

- by Ross Buggins

Update: This has now happened two days in a row. Update: XML of scheduled tasks and images now included. Two servers located in London, both Windows 2012, have a scheduled task set to run at 3pm Monday to Friday. This has been set up for the last 5 months without a problem. However, on Monday the 28th of August they both ran the scheduled task at 3pm and then again at 4pm. When it was first reported, I thought it was too much of a coincidence to be the day after the clocks had gone back an hour. However, I’m failing in being able to explain why it has happened and if it is related to the clock change at all. The relevant logs from one server (the logs for the other follow this pattern as well): Event 129 15:00:20 Task Scheduler launch task "\3pm", instance "C:\Program Files (x86)\PHP\v5.4\php.exe" with process ID 2388. Event 100 15:00:20 Task Scheduler started "{75a3590f-dec1-4dee-bd27-73d63a50a9d7}" instance of the "\3pm" task for user "x\y". Event 200 15:00:20 Task Scheduler launched action "C:\Program Files (x86)\PHP\v5.4\php.exe" in instance "{75a3590f-dec1-4dee-bd27-73d63a50a9d7}" of task "\3pm". Event 201 15:00:23 Task Scheduler successfully completed task "\3pm" , instance "{75a3590f-dec1-4dee-bd27-73d63a50a9d7}" , action "C:\Program Files (x86)\PHP\v5.4\php.exe" with return code 0. Event 129 16:00:20 Task Scheduler launch task "\3pm", instance "C:\Program Files (x86)\PHP\v5.4\php.exe" with process ID 1224. Event 100 16:00:20 Task Scheduler started "{3dd46ca9-c525-4796-86b5-5e513fd45f26}" instance of the "\3pm" task for user "x\y". Event 200 16:00:20 Task Scheduler launched action "C:\Program Files (x86)\PHP\v5.4\php.exe" in instance "{3dd46ca9-c525-4796-86b5-5e513fd45f26}" of task "\3pm". Event 201 16:00:23 Task Scheduler successfully completed task "\3pm" , instance "{3dd46ca9-c525-4796-86b5-5e513fd45f26}" , action "C:\Program Files (x86)\PHP\v5.4\php.exe" with return code 0. I've seen this question Scheduled task running twice from time to time which points to a bug at http://support.microsoft.com/kb/2461249 being the cause. However, this doesn't include Server 2012 in it's list of problem operating systems. I’m struggling to explain this, can anyone else? The XML export for the scheduled task is: <?xml version="1.0" encoding="UTF-16"?> <Task version="1.2" xmlns="http://schemas.microsoft.com/windows/2004/02/mit/task"> <RegistrationInfo> <Date>2013-04-16T14:04:17.4897806</Date> <Author>x\y</Author> </RegistrationInfo> <Triggers> <CalendarTrigger> <StartBoundary>2013-04-17T15:00:20</StartBoundary> <Enabled>true</Enabled> <ScheduleByWeek> <DaysOfWeek> <Monday /> <Tuesday /> <Wednesday /> <Thursday /> <Friday /> </DaysOfWeek> <WeeksInterval>1</WeeksInterval> </ScheduleByWeek> </CalendarTrigger> </Triggers> <Principals> <Principal id="Author"> <UserId>x\y</UserId> <LogonType>Password</LogonType> <RunLevel>LeastPrivilege</RunLevel> </Principal> </Principals> <Settings> <MultipleInstancesPolicy>IgnoreNew</MultipleInstancesPolicy> <DisallowStartIfOnBatteries>true</DisallowStartIfOnBatteries> <StopIfGoingOnBatteries>true</StopIfGoingOnBatteries> <AllowHardTerminate>true</AllowHardTerminate> <StartWhenAvailable>false</StartWhenAvailable> <RunOnlyIfNetworkAvailable>false</RunOnlyIfNetworkAvailable> <IdleSettings> <StopOnIdleEnd>true</StopOnIdleEnd> <RestartOnIdle>false</RestartOnIdle> </IdleSettings> <AllowStartOnDemand>true</AllowStartOnDemand> <Enabled>true</Enabled> <Hidden>false</Hidden> <RunOnlyIfIdle>false</RunOnlyIfIdle> <WakeToRun>false</WakeToRun> <ExecutionTimeLimit>P3D</ExecutionTimeLimit> <Priority>7</Priority> </Settings> <Actions Context="Author"> <Exec> <Command>"C:\Program Files (x86)\PHP\v5.4\php.exe"</Command> <Arguments>-f "c:\a.php"</Arguments> </Exec> </Actions> </Task> 29 October 17:00 - Update - Both servers have again run the scheduled task at 15:00 and 16:00. I've now updated the php file that is run by the scheduler to not actually do anything whilst I'm going through trying to solve this. I'm planning on restarting one server to see if this changes anything tomorrow. 30 October 08:25 - Update - When exporting the task XML I remembered that I hadn't included the fact that the scheduled task on second server was created by importing the XML of the task created on the first. The only difference between the two tasks is the path of the php file they are executing and the user account that they are running as.

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More interruptions than cpu context switches

- by Christopher Valles

I have a machine running Debian GNU/Linux 5.0.8 (lenny) 8 cores and 12Gb of RAM. We have one core permanently around 40% ~ 60% wait time and trying to spot what is happening I realized that we have more interruptions than cpu context switches. I found that the normal ratio between context switch and interruptions is around 10x more context switching than interruptions but on my server the values are completely different. backend1:~# vmstat -s 12330788 K total memory 12221676 K used memory 3668624 K active memory 6121724 K inactive memory 109112 K free memory 3929400 K buffer memory 4095536 K swap cache 4194296 K total swap 7988 K used swap 4186308 K free swap 44547459 non-nice user cpu ticks 702408 nice user cpu ticks 13346333 system cpu ticks 1607583668 idle cpu ticks 374043393 IO-wait cpu ticks 4144149 IRQ cpu ticks 3994255 softirq cpu ticks 0 stolen cpu ticks 4445557114 pages paged in 2910596714 pages paged out 128642 pages swapped in 267400 pages swapped out 3519307319 interrupts 2464686911 CPU context switches 1306744317 boot time 11555115 forks Any ideas if that is an issue? And in that case, how can I spot the cause and fix it? Update Following the instructions of the comments and focusing on the core stuck in wait I checked the processes attached to that core and below you can find the list: PID USER PR NI VIRT RES SHR S %CPU %MEM TIME+ P COMMAND 24 root RT -5 0 0 0 S 0 0.0 0:03.42 7 migration/7 25 root 15 -5 0 0 0 S 0 0.0 0:04.78 7 ksoftirqd/7 26 root RT -5 0 0 0 S 0 0.0 0:00.00 7 watchdog/7 34 root 15 -5 0 0 0 S 0 0.0 1:18.90 7 events/7 83 root 15 -5 0 0 0 S 0 0.0 1:10.68 7 kblockd/7 291 root 15 -5 0 0 0 S 0 0.0 0:00.00 7 aio/7 569 root 15 -5 0 0 0 S 0 0.0 0:00.00 7 ata/7 1545 root 15 -5 0 0 0 S 0 0.0 0:00.00 7 ksnapd 1644 root 15 -5 0 0 0 S 0 0.0 0:36.73 7 kjournald 1725 root 16 -4 16940 1152 488 S 0 0.0 0:00.00 7 udevd 2342 root 20 0 8828 1140 956 S 0 0.0 0:00.00 7 sh 2375 root 20 0 8848 1220 1016 S 0 0.0 0:00.00 7 locate 2421 root 30 10 8896 1268 1016 S 0 0.0 0:00.00 7 updatedb.findut 2430 root 30 10 58272 49m 616 S 0 0.4 0:17.44 7 sort 2431 root 30 10 3792 448 360 S 0 0.0 0:00.00 7 frcode 2682 root 15 -5 0 0 0 S 0 0.0 3:25.98 7 kjournald 2683 root 15 -5 0 0 0 S 0 0.0 0:00.64 7 kjournald 2687 root 15 -5 0 0 0 S 0 0.0 1:31.30 7 kjournald 3261 root 15 -5 0 0 0 S 0 0.0 2:30.56 7 kondemand/7 3364 root 20 0 3796 596 476 S 0 0.0 0:00.00 7 acpid 3575 root 20 0 8828 1140 956 S 0 0.0 0:00.00 7 sh 3597 root 20 0 8848 1216 1016 S 0 0.0 0:00.00 7 locate 3603 root 30 10 8896 1268 1016 S 0 0.0 0:00.00 7 updatedb.findut 3612 root 30 10 58272 49m 616 S 0 0.4 0:27.04 7 sort 3655 root 20 0 11056 2852 516 S 0 0.0 5:36.46 7 redis-server 3706 root 20 0 19832 1056 816 S 0 0.0 0:01.64 7 cron 3746 root 20 0 3796 580 484 S 0 0.0 0:00.00 7 getty 3748 root 20 0 3796 580 484 S 0 0.0 0:00.00 7 getty 7674 root 20 0 28376 1000 736 S 0 0.0 0:00.00 7 cron 7675 root 20 0 8828 1140 956 S 0 0.0 0:00.00 7 sh 7708 root 30 10 58272 49m 616 S 0 0.4 0:03.36 7 sort 22049 root 20 0 8828 1136 956 S 0 0.0 0:00.00 7 sh 22095 root 20 0 8848 1220 1016 S 0 0.0 0:00.00 7 locate 22099 root 30 10 8896 1264 1016 S 0 0.0 0:00.00 7 updatedb.findut 22108 root 30 10 58272 49m 616 S 0 0.4 0:44.55 7 sort 22109 root 30 10 3792 452 360 S 0 0.0 0:00.00 7 frcode 26927 root 20 0 8828 1140 956 S 0 0.0 0:00.00 7 sh 26947 root 20 0 8848 1216 1016 S 0 0.0 0:00.00 7 locate 26951 root 30 10 8896 1268 1016 S 0 0.0 0:00.00 7 updatedb.findut 26960 root 30 10 58272 49m 616 S 0 0.4 0:10.24 7 sort 26961 root 30 10 3792 452 360 S 0 0.0 0:00.00 7 frcode 27952 root 20 0 65948 3028 2400 S 0 0.0 0:00.00 7 sshd 30731 root 20 0 0 0 0 S 0 0.0 0:01.34 7 pdflush 31204 root 20 0 0 0 0 S 0 0.0 0:00.24 7 pdflush 21857 deploy 20 0 1227m 2240 868 S 0 0.0 2:44.22 7 nginx 21858 deploy 20 0 1228m 2784 868 S 0 0.0 2:42.45 7 nginx 21862 deploy 20 0 1228m 2732 868 S 0 0.0 2:43.90 7 nginx 21869 deploy 20 0 1228m 2840 868 S 0 0.0 2:44.14 7 nginx 27994 deploy 20 0 19372 2216 1380 S 0 0.0 0:00.00 7 bash 28493 deploy 20 0 331m 32m 16m S 4 0.3 0:00.40 7 apache2 21856 deploy 20 0 1228m 2844 868 S 0 0.0 2:43.64 7 nginx 3622 nobody 30 10 21156 10m 916 D 0 0.1 4:42.31 7 find 7716 nobody 30 10 12268 1280 888 D 0 0.0 0:43.50 7 find 22116 nobody 30 10 12612 1696 916 D 0 0.0 6:32.26 7 find 26968 nobody 30 10 12268 1284 888 D 0 0.0 1:56.92 7 find Update As suggested I take a look at /proc/interrupts and below the info there: CPU0 CPU1 CPU2 CPU3 CPU4 CPU5 CPU6 CPU7 0: 35 0 0 1469085485 0 0 0 0 IO-APIC-edge timer 1: 0 0 0 8 0 0 0 0 IO-APIC-edge i8042 8: 0 0 0 1 0 0 0 0 IO-APIC-edge rtc0 9: 0 0 0 0 0 0 0 0 IO-APIC-fasteoi acpi 12: 0 0 0 105 0 0 0 0 IO-APIC-edge i8042 16: 0 0 0 0 0 0 0 580212114 IO-APIC-fasteoi 3w-9xxx, uhci_hcd:usb1 18: 0 0 142 0 0 0 0 0 IO-APIC-fasteoi uhci_hcd:usb6, ehci_hcd:usb7 19: 9 0 0 0 0 0 0 0 IO-APIC-fasteoi uhci_hcd:usb3, uhci_hcd:usb5 21: 0 0 0 0 0 0 0 0 IO-APIC-fasteoi uhci_hcd:usb2 23: 0 0 0 0 0 0 0 0 IO-APIC-fasteoi uhci_hcd:usb4, ehci_hcd:usb8 1273: 0 0 1600400502 0 0 0 0 0 PCI-MSI-edge eth0 1274: 0 0 0 0 0 0 0 0 PCI-MSI-edge ahci NMI: 0 0 0 0 0 0 0 0 Non-maskable interrupts LOC: 214252181 69439018 317298553 21943690 72562482 56448835 137923978 407514738 Local timer interrupts RES: 27516446 16935944 26430972 44957009 24935543 19881887 57746906 24298747 Rescheduling interrupts CAL: 10655 10705 10685 10567 10689 10669 10667 396 function call interrupts TLB: 529548 462587 801138 596193 922202 747313 2027966 946594 TLB shootdowns TRM: 0 0 0 0 0 0 0 0 Thermal event interrupts THR: 0 0 0 0 0 0 0 0 Threshold APIC interrupts SPU: 0 0 0 0 0 0 0 0 Spurious interrupts ERR: 0 All the values seems more or less the same for all the cores but this one IO-APIC-fasteoi 3w-9xxx, uhci_hcd:usb1 only affects to the core 7 (the same with the wait time of 40% ~ 60%) could be something attached to the usb port causing the issue? Thanks in advanced

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VPN still working after rebooting without client - DrayTek client shows "No Connection"

- by HeavenCore

My home network is a simple router + pc's setup, nothing fancy - the router has DHCP enabled for 192.168.0.X (255.255.255.0) and my PC picks up the address 192.168.0.82. There are no devices on my local lan in the 192.168.1.x range. On my pc i have the DrayTek VPN client, and a company i do some work for has a DrayTek Vigor router. The VPN client establishes a VPN to that remote company using an IPSec Tunnel (PreShared Key - no encryption) Last night i shut down my pc with the VPN tunnel still connected, when i turned my computer on this morning i accidentally clicked an RDP shortcut to 192.168.1.2 (a host in the remote company) and to my amazement it connected?!? I checked and the DrayTek VPN client isnt running, and when i did run it, it clearly shows "Status: No connection". confused as to how my machine can still talk to this remote machine i tried a trace: C:\Users\HeavenCore>tracert 192.168.1.2 Tracing route to C4SERVERII [192.168.1.2] over a maximum of 30 hops: 1 * * * Request timed out. 2 * * * Request timed out. 3 * * * Request timed out. 4 * * * Request timed out. 5 * * * Request timed out. 6 * * * Request timed out. 7 * * * Request timed out. 8 * * * Request timed out. 9 * * * Request timed out. 10 * * * Request timed out. 11 * * * Request timed out. 12 15 ms 21 ms 32 ms C4SERVERII [192.168.1.2] Trace complete. No indication there as to how it's getting from my network to the remote host. with my network mask being 255.255.255.0 with ip 192.168.0.1 i dont even see how packets are routing to 192.168.1.1 - unless there was a static route in place, so i checked the route table: IPv4 Route Table =========================================================================== Active Routes: Network Destination Netmask Gateway Interface Metric 0.0.0.0 0.0.0.0 192.168.0.1 192.168.0.82 266 127.0.0.0 255.0.0.0 On-link 127.0.0.1 306 127.0.0.1 255.255.255.255 On-link 127.0.0.1 306 127.255.255.255 255.255.255.255 On-link 127.0.0.1 306 192.168.0.0 255.255.255.0 On-link 192.168.0.82 266 192.168.0.82 255.255.255.255 On-link 192.168.0.82 266 192.168.0.255 255.255.255.255 On-link 192.168.0.82 266 224.0.0.0 240.0.0.0 On-link 127.0.0.1 306 224.0.0.0 240.0.0.0 On-link 192.168.0.82 266 255.255.255.255 255.255.255.255 On-link 127.0.0.1 306 255.255.255.255 255.255.255.255 On-link 192.168.0.82 266 =========================================================================== Persistent Routes: Network Address Netmask Gateway Address Metric 0.0.0.0 0.0.0.0 192.168.0.1 Default =========================================================================== As far as i can see, nothing indicating how my packets are getting to 192.168.1.2??? To confirm i was on a different subnet i did an ipconfig /all: Ethernet adapter Local Area Connection: Connection-specific DNS Suffix . : Description . . . . . . . . . . . : Marvell Yukon 88E8056 PCI-E Gigabit Ether net Controller Physical Address. . . . . . . . . : 00-23-54-F3-4E-BA DHCP Enabled. . . . . . . . . . . : No Autoconfiguration Enabled . . . . : Yes IPv4 Address. . . . . . . . . . . : 192.168.0.82(Preferred) Subnet Mask . . . . . . . . . . . : 255.255.255.0 Default Gateway . . . . . . . . . : 192.168.0.1 DNS Servers . . . . . . . . . . . : 192.168.0.1 208.67.222.222 NetBIOS over Tcpip. . . . . . . . : Enabled Yet straight after confirming my ip and subnet as above i can go ahead and ping the remote machine: C:\Users\HeavenCore>ping 192.168.1.2 Pinging 192.168.1.2 with 32 bytes of data: Reply from 192.168.1.2: bytes=32 time=48ms TTL=127 Reply from 192.168.1.2: bytes=32 time=23ms TTL=127 Reply from 192.168.1.2: bytes=32 time=103ms TTL=127 Reply from 192.168.1.2: bytes=32 time=25ms TTL=127 Ping statistics for 192.168.1.2: Packets: Sent = 4, Received = 4, Lost = 0 (0% loss), Approximate round trip times in milli-seconds: Minimum = 23ms, Maximum = 103ms, Average = 49ms Also, note on the ping how the times are 35ms ish, this clearly shows the pings are to the remote host and not something on my local lan (all stuff on my local lan pings in 0ms) - plus i verified the host was actually the host via RDP. My Question: Can an IPSec tunnel stay up some how after a reboot without use of the VPN client? (well, i can clearly see that it can) - where in windows is there visibility of this? how does my machine know where to route the packets? I appreciate any insights & thoughts!

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Linux servers going into Halt when pressing Control-D in putty or exit in the shell

- by Itai Ganot

Since today at noon, there's a number of Linux CentOS servers which are going to Halt whenever i type exit or use Control-D to close the putty window. Did anyone encounter this weird behavior before? I've checked the aliases list on the servers and there is no alias regarding halt command. After the server came online i've checked the history and saw a "logout" command there but nothing which is related to Halt. At first, i thought it happens only from my computer but later i realized that it happens to everyone which types exit, logout or control+d. 2 of these server are our main iptables firewalls and so it's super critical, your assistance is much appreciated. It looks like that, and it only happens on servers with active IPTables: [root@srv1 bin]# ssh srv2 root@srv2's password: Last login: Sun Nov 11 17:19:41 2012 from 192.168.12.98 [root@srv2 ~]# vim /etc/crontab [root@srv2 ~]# exit logout Broadcast message from root (pts/1) (Tue Nov 13 10:44:04 2012): The system is going down for system halt NOW! Connection to srv2 closed. [root@srv1 bin]# In my troubleshooting steps i came across the command strace, and so i've opened two bash windows to one of the problematic server and i used strace -p PID_of_bash. When i typed in exit in the first shell it did go to halt, attached is the strace output, if you can check it out and tell me if you see anything suspicious i'd be more than thankful. RER, 0x2b0e45a8f2f0}, 8) = 0 rt_sigaction(SIGALRM, {0x4484f0, [HUP INT ILL TRAP ABRT BUS FPE USR1 SEGV USR2 PIPE ALRM TERM XCPU XFSZ VTALRM SYS], SA_RESTORER, 0x2b0e45a8f2f0}, {0x47c450, [], SA_RESTORER, 0x2b0e45a8f2f0}, 8) = 0 rt_sigaction(SIGTSTP, {0x1, [], SA_RESTORER, 0x2b0e45a8f2f0}, {0x1, [], SA_RESTORER, 0x2b0e45a8f2f0}, 8) = 0 rt_sigaction(SIGTTOU, {0x1, [], SA_RESTORER, 0x2b0e45a8f2f0}, {0x1, [], SA_RESTORER, 0x2b0e45a8f2f0}, 8) = 0 rt_sigaction(SIGTTIN, {0x1, [], SA_RESTORER, 0x2b0e45a8f2f0}, {0x1, [], SA_RESTORER, 0x2b0e45a8f2f0}, 8) = 0 rt_sigaction(SIGWINCH, {0x448370, [], SA_RESTORER, 0x2b0e45a8f2f0}, {0x47c410, [], SA_RESTORER|SA_RESTART, 0x2b0e45a8f2f0}, 8) = 0 socket(PF_NETLINK, SOCK_RAW, 9) = 3 sendmsg(3, {msg_name(12)={sa_family=AF_NETLINK, pid=0, groups=00000000}, msg_iov(2)=[{"\25\0\0\0d\4\1\0\0\0\0\0\0\0\0\0", 16}, {"exit\0", 5}], msg_controllen=0, msg_flags=0}, 0) = 21 close(3) = 0 rt_sigaction(SIGINT, {0x448700, [], SA_RESTORER, 0x2b0e45a8f2f0}, {0x448700, [], SA_RESTORER, 0x2b0e45a8f2f0}, 8) = 0 write(2, "logout\n", 7) = 7 write(2, "There are stopped jobs.\n", 24) = 24 rt_sigprocmask(SIG_BLOCK, NULL, [], 8) = 0 rt_sigaction(SIGINT, {0x448700, [], SA_RESTORER, 0x2b0e45a8f2f0}, {0x448700, [], SA_RESTORER, 0x2b0e45a8f2f0}, 8) = 0 rt_sigprocmask(SIG_BLOCK, NULL, [], 8) = 0 rt_sigprocmask(SIG_BLOCK, [INT CHLD], [], 8) = 0 pipe([3, 4]) = 0 clone(child_stack=0, flags=CLONE_CHILD_CLEARTID|CLONE_CHILD_SETTID|SIGCHLD, child_tidptr=0x2b0e45db6fe0) = 23717 setpgid(23717, 23717) = 0 rt_sigprocmask(SIG_SETMASK, [], NULL, 8) = 0 rt_sigprocmask(SIG_BLOCK, [CHLD], [], 8) = 0 close(3) = 0 close(4) = 0 rt_sigprocmask(SIG_BLOCK, [CHLD TSTP TTIN TTOU], [CHLD], 8) = 0 ioctl(255, TIOCSPGRP, [23717]) = 0 rt_sigprocmask(SIG_SETMASK, [CHLD], NULL, 8) = 0 rt_sigprocmask(SIG_SETMASK, [], NULL, 8) = 0 rt_sigprocmask(SIG_BLOCK, [CHLD], [], 8) = 0 wait4(-1, [{WIFEXITED(s) && WEXITSTATUS(s) == 0}], WSTOPPED|WCONTINUED, NULL) = 23717 rt_sigprocmask(SIG_BLOCK, [CHLD TSTP TTIN TTOU], [CHLD], 8) = 0 ioctl(255, TIOCSPGRP, [20458]) = 0 rt_sigprocmask(SIG_SETMASK, [CHLD], NULL, 8) = 0 ioctl(255, SNDCTL_TMR_TIMEBASE or TCGETS, {B38400 opost isig icanon echo ...}) = 0 ioctl(255, TIOCGWINSZ, {ws_row=53, ws_col=211, ws_xpixel=0, ws_ypixel=0}) = 0 rt_sigprocmask(SIG_SETMASK, [], NULL, 8) = 0 --- SIGCHLD (Child exited) @ 0 (0) --- wait4(-1, 0x7fff395da984, WNOHANG|WSTOPPED|WCONTINUED, NULL) = 0 rt_sigreturn(0x11) = 0 rt_sigprocmask(SIG_BLOCK, [CHLD TSTP TTIN TTOU], [], 8) = 0 ioctl(255, TIOCSPGRP, [20458]) = 0 rt_sigprocmask(SIG_SETMASK, [], NULL, 8) = 0 rt_sigaction(SIGINT, {0x448700, [], SA_RESTORER, 0x2b0e45a8f2f0}, {0x448700, [], SA_RESTORER, 0x2b0e45a8f2f0}, 8) = 0 rt_sigprocmask(SIG_BLOCK, [INT], [], 8) = 0 ioctl(0, TIOCGWINSZ, {ws_row=53, ws_col=211, ws_xpixel=0, ws_ypixel=0}) = 0 ioctl(0, TIOCSWINSZ, {ws_row=53, ws_col=211, ws_xpixel=0, ws_ypixel=0}) = 0 ioctl(0, SNDCTL_TMR_TIMEBASE or TCGETS, {B38400 opost isig icanon echo ...}) = 0 ioctl(0, SNDCTL_TMR_STOP or TCSETSW, {B38400 opost isig -icanon -echo ...}) = 0 rt_sigprocmask(SIG_SETMASK, [], NULL, 8) = 0 rt_sigprocmask(SIG_BLOCK, [INT QUIT ALRM TSTP TTIN TTOU], [], 8) = 0 rt_sigaction(SIGINT, {0x47c450, [], SA_RESTORER, 0x2b0e45a8f2f0}, {0x448700, [], SA_RESTORER, 0x2b0e45a8f2f0}, 8) = 0 rt_sigaction(SIGTERM, {0x47c450, [], SA_RESTORER, 0x2b0e45a8f2f0}, {0x1, [], SA_RESTORER, 0x2b0e45a8f2f0}, 8) = 0 rt_sigaction(SIGTERM, {0x1, [], SA_RESTORER, 0x2b0e45a8f2f0}, {0x47c450, [], SA_RESTORER, 0x2b0e45a8f2f0}, 8) = 0 rt_sigaction(SIGQUIT, {0x47c450, [], SA_RESTORER, 0x2b0e45a8f2f0}, {0x1, [], SA_RESTORER, 0x2b0e45a8f2f0}, 8) = 0 rt_sigaction(SIGQUIT, {0x1, [], SA_RESTORER, 0x2b0e45a8f2f0}, {0x47c450, [], SA_RESTORER, 0x2b0e45a8f2f0}, 8) = 0 rt_sigaction(SIGALRM, {0x47c450, [], SA_RESTORER, 0x2b0e45a8f2f0}, {0x4484f0, [HUP INT ILL TRAP ABRT BUS FPE USR1 SEGV USR2 PIPE ALRM TERM XCPU XFSZ VTALRM SYS], SA_RESTORER, 0x2b0e45a8f2f0}, 8) = 0 rt_sigaction(SIGTSTP, {0x47c450, [], SA_RESTORER, 0x2b0e45a8f2f0}, {0x1, [], SA_RESTORER, 0x2b0e45a8f2f0}, 8) = 0 rt_sigaction(SIGTSTP, {0x1, [], SA_RESTORER, 0x2b0e45a8f2f0}, {0x47c450, [], SA_RESTORER, 0x2b0e45a8f2f0}, 8) = 0 rt_sigaction(SIGTTOU, {0x47c450, [], SA_RESTORER, 0x2b0e45a8f2f0}, {0x1, [], SA_RESTORER, 0x2b0e45a8f2f0}, 8) = 0 rt_sigaction(SIGTTOU, {0x1, [], SA_RESTORER, 0x2b0e45a8f2f0}, {0x47c450, [], SA_RESTORER, 0x2b0e45a8f2f0}, 8) = 0 rt_sigaction(SIGTTIN, {0x47c450, [], SA_RESTORER, 0x2b0e45a8f2f0}, {0x1, [], SA_RESTORER, 0x2b0e45a8f2f0}, 8) = 0 rt_sigaction(SIGTTIN, {0x1, [], SA_RESTORER, 0x2b0e45a8f2f0}, {0x47c450, [], SA_RESTORER, 0x2b0e45a8f2f0}, 8) = 0 rt_sigprocmask(SIG_SETMASK, [], NULL, 8) = 0 rt_sigaction(SIGWINCH, {0x47c410, [], SA_RESTORER|SA_RESTART, 0x2b0e45a8f2f0}, {0x448370, [], SA_RESTORER, 0x2b0e45a8f2f0}, 8) = 0 write(2, "[root@g2-lga ~]# ", 17) = 17 rt_sigprocmask(SIG_BLOCK, NULL, [], 8) = 0 read(0, "e", 1) = 1 write(2, "e", 1) = 1 rt_sigprocmask(SIG_BLOCK, NULL, [], 8) = 0 read(0, "x", 1) = 1 write(2, "x", 1) = 1 rt_sigprocmask(SIG_BLOCK, NULL, [], 8) = 0 read(0, "i", 1) = 1 write(2, "i", 1) = 1 rt_sigprocmask(SIG_BLOCK, NULL, [], 8) = 0 read(0, "t", 1) = 1 write(2, "t", 1) = 1 rt_sigprocmask(SIG_BLOCK, NULL, [], 8) = 0 read(0, "\r", 1) = 1 write(2, "\n", 1) = 1 rt_sigprocmask(SIG_BLOCK, [INT], [], 8) = 0 ioctl(0, SNDCTL_TMR_STOP or TCSETSW, {B38400 opost isig icanon echo ...}) = 0 rt_sigprocmask(SIG_SETMASK, [], NULL, 8) = 0 rt_sigaction(SIGINT, {0x448700, [], SA_RESTORER, 0x2b0e45a8f2f0}, {0x47c450, [], SA_RESTORER, 0x2b0e45a8f2f0}, 8) = 0 rt_sigaction(SIGTERM, {0x1, [], SA_RESTORER, 0x2b0e45a8f2f0}, {0x1, [], SA_RESTORER, 0x2b0e45a8f2f0}, 8) = 0 rt_sigaction(SIGQUIT, {0x1, [], SA_RESTORER, 0x2b0e45a8f2f0}, {0x1, [], SA_RESTORER, 0x2b0e45a8f2f0}, 8) = 0 rt_sigaction(SIGALRM, {0x4484f0, [HUP INT ILL TRAP ABRT BUS FPE USR1 SEGV USR2 PIPE ALRM TERM XCPU XFSZ VTALRM SYS], SA_RESTORER, 0x2b0e45a8f2f0}, {0x47c450, [], SA_RESTORER, 0x2b0e45a8f2f0}, 8) = 0 rt_sigaction(SIGTSTP, {0x1, [], SA_RESTORER, 0x2b0e45a8f2f0}, {0x1, [], SA_RESTORER, 0x2b0e45a8f2f0}, 8) = 0 rt_sigaction(SIGTTOU, {0x1, [], SA_RESTORER, 0x2b0e45a8f2f0}, {0x1, [], SA_RESTORER, 0x2b0e45a8f2f0}, 8) = 0 rt_sigaction(SIGTTIN, {0x1, [], SA_RESTORER, 0x2b0e45a8f2f0}, {0x1, [], SA_RESTORER, 0x2b0e45a8f2f0}, 8) = 0 rt_sigaction(SIGWINCH, {0x448370, [], SA_RESTORER, 0x2b0e45a8f2f0}, {0x47c410, [], SA_RESTORER|SA_RESTART, 0x2b0e45a8f2f0}, 8) = 0 socket(PF_NETLINK, SOCK_RAW, 9) = 3 sendmsg(3, {msg_name(12)={sa_family=AF_NETLINK, pid=0, groups=00000000}, msg_iov(2)=[{"\25\0\0\0d\4\1\0\0\0\0\0\0\0\0\0", 16}, {"exit\0", 5}], msg_controllen=0, msg_flags=0}, 0) = 21 close(3) = 0 rt_sigaction(SIGINT, {0x448700, [], SA_RESTORER, 0x2b0e45a8f2f0}, {0x448700, [], SA_RESTORER, 0x2b0e45a8f2f0}, 8) = 0 write(2, "logout\n", 7) = 7 open("/root/.bash_logout", O_RDONLY) = 3 fstat(3, {st_mode=S_IFREG|0644, st_size=24, ...}) = 0 read(3, "# ~/.bash_logout\n\nclear\n", 24) = 24 close(3) = 0 rt_sigprocmask(SIG_BLOCK, NULL, [], 8) = 0 rt_sigprocmask(SIG_BLOCK, NULL, [], 8) = 0 rt_sigprocmask(SIG_BLOCK, NULL, [], 8) = 0 rt_sigprocmask(SIG_BLOCK, NULL, [], 8) = 0 stat(".", {st_mode=S_IFDIR|0750, st_size=12288, ...}) = 0 stat("/usr/kerberos/sbin/clear", 0x7fff395da960) = -1 ENOENT (No such file or directory) stat("/usr/kerberos/bin/clear", 0x7fff395da960) = -1 ENOENT (No such file or directory) stat("/usr/local/sbin/clear", 0x7fff395da960) = -1 ENOENT (No such file or directory) stat("/usr/local/bin/clear", 0x7fff395da960) = -1 ENOENT (No such file or directory) stat("/sbin/clear", 0x7fff395da960) = -1 ENOENT (No such file or directory) stat("/bin/clear", 0x7fff395da960) = -1 ENOENT (No such file or directory) stat("/usr/sbin/clear", 0x7fff395da960) = -1 ENOENT (No such file or directory) stat("/usr/bin/clear", {st_mode=S_IFREG|0755, st_size=12712, ...}) = 0 access("/usr/bin/clear", X_OK) = 0 access("/usr/bin/clear", R_OK) = 0 stat("/usr/bin/clear", {st_mode=S_IFREG|0755, st_size=12712, ...}) = 0 access("/usr/bin/clear", X_OK) = 0 access("/usr/bin/clear", R_OK) = 0 rt_sigprocmask(SIG_BLOCK, [INT CHLD], [], 8) = 0 pipe([3, 4]) = 0 clone(child_stack=0, flags=CLONE_CHILD_CLEARTID|CLONE_CHILD_SETTID|SIGCHLD, child_tidptr=0x2b0e45db6fe0) = 23726 setpgid(23726, 23726) = 0 rt_sigprocmask(SIG_SETMASK, [], NULL, 8) = 0 rt_sigprocmask(SIG_BLOCK, [CHLD], [], 8) = 0 close(3) = 0 close(4) = 0 rt_sigprocmask(SIG_BLOCK, [CHLD TSTP TTIN TTOU], [CHLD], 8) = 0 ioctl(255, TIOCSPGRP, [23726]) = 0 rt_sigprocmask(SIG_SETMASK, [CHLD], NULL, 8) = 0 rt_sigprocmask(SIG_SETMASK, [], NULL, 8) = 0 rt_sigprocmask(SIG_BLOCK, [CHLD], [], 8) = 0 wait4(-1, Broadcast message from root (pts/0) (Wed Nov 14 12:41:44 2012): The system is going down for system halt NOW! [{WIFEXITED(s) && WEXITSTATUS(s) == 0}], WSTOPPED|WCONTINUED, NULL) = 23726 rt_sigprocmask(SIG_BLOCK, [CHLD TSTP TTIN TTOU], [CHLD], 8) = 0 ioctl(255, TIOCSPGRP, [20458]) = 0 rt_sigprocmask(SIG_SETMASK, [CHLD], NULL, 8) = 0 ioctl(255, SNDCTL_TMR_TIMEBASE or TCGETS, {B38400 opost isig icanon echo ...}) = 0 ioctl(255, TIOCGWINSZ, {ws_row=53, ws_col=211, ws_xpixel=0, ws_ypixel=0}) = 0 rt_sigprocmask(SIG_SETMASK, [], NULL, 8) = 0 --- SIGCHLD (Child exited) @ 0 (0) --- wait4(-1, 0x7fff395da634, WNOHANG|WSTOPPED|WCONTINUED, NULL) = 0 rt_sigreturn(0x11) = 0 open("/etc/bash.bash_logout", O_RDONLY) = -1 ENOENT (No such file or directory) rt_sigprocmask(SIG_SETMASK, [], NULL, 8) = 0 rt_sigaction(SIGINT, {0x448700, [], SA_RESTORER, 0x2b0e45a8f2f0}, {0x448700, [], SA_RESTORER, 0x2b0e45a8f2f0}, 8) = 0 stat("/root/.bash_history", {st_mode=S_IFREG|0600, st_size=28900, ...}) = 0 open("/root/.bash_history", O_WRONLY|O_APPEND) = 3 write(3, "cd /etc/profile.d/\nls\nls -alrt\ng"..., 1120) = 1120 close(3) = 0 open("/root/.bash_history", O_RDONLY) = 3 fstat(3, {st_mode=S_IFREG|0600, st_size=30020, ...}) = 0 read(3, "history \nping g1-lga\nping g1-lga"..., 30020) = 30020 close(3) = 0 open("/root/.bash_history", O_WRONLY|O_TRUNC) = 3 write(3, "grep \"216.18\" *\nhistory \nexit\nvi"..., 27609) = 27609 close(3) = 0 kill(4294965658, SIGTERM) = 0 kill(4294965658, SIGCONT) = 0 --- SIGCHLD (Child exited) @ 0 (0) --- wait4(-1, [{WIFSIGNALED(s) && WTERMSIG(s) == SIGTERM}], WNOHANG|WSTOPPED|WCONTINUED, NULL) = 1638 wait4(-1, 0x7fff395dac34, WNOHANG|WSTOPPED|WCONTINUED, NULL) = -1 ECHILD (No child processes) rt_sigreturn(0x11) = 0 --- SIGCHLD (Child exited) @ 0 (0) --- wait4(-1, 0x7fff395dac34, WNOHANG|WSTOPPED|WCONTINUED, NULL) = -1 ECHILD (No child processes) rt_sigreturn(0xffffffffffffffff) = 0 rt_sigprocmask(SIG_BLOCK, [CHLD TSTP TTIN TTOU], [], 8) = 0 ioctl(255, TIOCSPGRP, [20458]) = 0 rt_sigprocmask(SIG_SETMASK, [], NULL, 8) = 0 setpgid(0, 20458) = -1 EPERM (Operation not permitted) exit_group(1) = ? Process 20458 detached [root@g2-lga ~]#

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Alternate way to create a clone of a UNIX System

- by Spirit

THE STORY: (If you don't like to read much, down below is the question :) ) Where I work we have two HP RP2470 servers same hardware same number of hard drives same everything :). One of them is a production server and runs HP-UX 11.00. The poor ba***rd hasn't been turned off for years and now I have to make a clone of it on the other server - just in case, for redundancy. The problem is simple (or not simple) as I have to make the the other server exactly the same. However the old version of OS (UX 11.00 is a history now) and the old software running on it, have made my task almost impossible. On the production server there is also a cloning/recover utility Ignite-UX. I tried many times to create a recovery tape with it. Then when I load the tape on the backup server, it succeeds with the loading of the tape (no errors no warnings) but on the next restart it fails to load the OS :S and drops into HP`s ISL prompt. --- THE QUESTION: Is there an alternate way to create a clone of the Unix System? The environment is: 1. 2x HP RP2470 Servers (non-Intel), same hardware, same number od HDDs (two each of them) same everything. 2. OS running: HP-UX 11.00 The production server has to be cloned without downtime - sadly :( as I hope that they will reconsider on this one For example (like on Windows platforms), if you try to copy an entire HDD with Windows inside on another HDD, and then put that HDD on another PC it will still work, as long as the hardware is the same. Can I do something like that with a Unix system? Can I somehow COPY the contents of the entire HDD, put those on another HDD, and then just load the HDD into the other server? (If you haven't read the story the servers are exactly the same) Will it work? Can it be done with ordinary commands like cp or dump or something like that? Does any one have a similar experience? --- UPDATE: 26.01.2012 NOTE: The update is related to "The Story". If you haven't read that part then you can skip this update. This is just a short update on the recover logs from the Ignite Tape.. someone with more exp. might notice something.. ... --- READING CONTENTS OF THE IGNITE TAPE --- --- OUTPUT OMITED --- ... ... x ./configure3, 413696 bytes, 808 tape blocks x ./monitor_bpr, 20480 bytes, 40 tape blocks * Download_mini-system: Complete * Loading_software: Begin * Installing boot area on disk. * Enabling swap areas. * Backing up LVM configuration for "vg00". * Processing the archive source (Recovery Archive). * Wed Jan 25 15:27:32 EST 2012: Starting archive load of the source (Recovery Archive). * Positioning the tape (/dev/rmt/0mn). * Archive extraction from tape is beginning. Please wait. * Wed Jan 25 15:39:52 EST 2012: Completed archive load of the source (Recovery Archive). * Executing user specified script: "/opt/ignite/data/scripts/os_arch_post_l". * Running in recovery mode (os_arch_post_l). * Running the ioinit command ("/sbin/ioinit -c") * Creating device files via the insf command. insf: Installing special files for sdisk instance 0 address 0/0/1/1.15.0 insf: Installing special files for sdisk instance 1 address 0/0/2/0.1.0 insf: Installing special files for sdisk instance 2 address 0/0/2/1.15.0 insf: Installing special files for stape instance 0 address 0/0/1/0.3.0 insf: Installing special files for btlan instance 0 address 0/0/0/0 insf: Installing special files for btlan instance 1 address 0/2/0/0 insf: Installing special files for pseudo driver dlpi insf: Installing special files for pseudo driver kepd insf: Installing special files for pseudo driver framebuf insf: Installing special files for pseudo driver sad * Running "/opt/upgrade/bin/tlinstall -v" and correcting transition link permissions. * Constructing the bootconf file. * Setting primary boot path to "0/0/1/1.15.0". * Executing: "/var/adm/sw/products/PHSS_20146/pfiles/iux_postload". * Executing: "/var/adm/sw/products/PHSS_25982/pfiles/iux_postload". NOTE: tlinstall is searching filesystem - please be patient NOTE: Successfully completed * Loading_software: Complete * Build_Kernel: Begin NOTE: Since the /stand/vmunix kernel is already in place, the kernel will not be re-built. Note that no mod_kernel directives will be processed. * Build_Kernel: Complete * Boot_From_Client_Disk: Begin * Rebooting machine as expected. NOTE: Rebooting system. sync'ing disks (0 buffers to flush): 0 buffers not flushed 0 buffers still dirty Closing open logical volumes... Done Console reset done. Boot device reset done. ********** VIRTUAL FRONT PANEL ********** System Boot detected ***************************************** LEDs: RUN ATTENTION FAULT REMOTE POWER FLASH OFF OFF ON ON LED State: Running non-OS code. (i.e. Boot or Diagnostics) ... ... ... --- SERVER IS PERFORMING POST SEQUENCE HERE --- --- OUTPUT OMITED --- ... ... ... ***************************************** ************ EARLY BOOT VFP ************* End of early boot detected ***************************************** Firmware Version 43.50 Duplex Console IO Dependent Code (IODC) revision 1 ------------------------------------------------------------------------------ (c) Copyright 1995-2002, Hewlett-Packard Company, All rights reserved ------------------------------------------------------------------------------ Processor Speed State CoProcessor State Cache Size Number State Inst Data --------- -------- --------------------- ----------------- ------------ 0 650 MHz Active Functional 750 KB 1.5 MB 1 650 MHz Idle Functional 750 KB 1.5 MB Central Bus Speed (in MHz) : 120 Available Memory : 2097152 KB Good Memory Required : 16140 KB Primary boot path: 0/0/1/1.15 Alternate boot path: 0/0/2/1.15 Console path: 0/0/4/1.643 Keyboard path: 0/0/4/0.0 Processor is starting autoboot process. To discontinue, press any key within 10 seconds. 10 seconds expired. Proceeding... Trying Primary Boot Path ------------------------ Booting... Boot IO Dependent Code (IODC) revision 1 HARD Booted. ISL Revision A.00.38 OCT 26, 1994 ISL booting hpux ISL>

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cannot access localhost using ip

- by Robert

I have done a small web development project using eclipse. It runs well when I try running it on browser with url localhost:8080/myproject/home.html. But if I want to access it on another machine (laptop, mobile, etc. using the same wifi) it is not possible; it is not able to connect. After Googling for a while found out that I have to use the IP address instead of 'localhost'. So I tried 10.0.0.4:8080/myproject/home.html, but still does not work. In fact i am unable to open that url on the same machine (where localhost:8080/myproject/home.html works fine). I also added a new Inbound rule in control panel firewall settings, allowing access to all ports for protocol TCP. Still have problem in running application with the url 10.0.0.4:8080/myproject/home.html (both on same machine as well as laptop and mobile). FYI i am using Eclipse Indigo, Apache tomcat 6.0 and server.xml file contents is as below: <?xml version="1.0" encoding="UTF-8"?> <Server port="8005" shutdown="SHUTDOWN">  <Listener SSLEngine="on" className="org.apache.catalina.core.AprLifecycleListener"/>  <Listener className="org.apache.catalina.core.JasperListener"/>  <Listener className="org.apache.catalina.core.JreMemoryLeakPreventionListener"/>  <Listener className="org.apache.catalina.mbeans.ServerLifecycleListener"/> <Listener className="org.apache.catalina.mbeans.GlobalResourcesLifecycleListener"/>  <GlobalNamingResources>  <Resource auth="Container" description="User database that can be updated and saved" factory="org.apache.catalina.users.MemoryUserDatabaseFactory" name="UserDatabase" pathname="conf/tomcat-users.xml" type="org.apache.catalina.UserDatabase"/> </GlobalNamingResources>  <Service name="Catalina">    <Connector port="8080" protocol="HTTP/1.1" address="10.0.0.4" connectionTimeout="20000" redirectPort="8443" />      <Connector port="8009" protocol="AJP/1.3" redirectPort="8443"/>   <Engine defaultHost="localhost" name="Catalina">      <Realm className="org.apache.catalina.realm.UserDatabaseRealm" resourceName="UserDatabase"/>  <Host appBase="webapps" autoDeploy="true" name="localhost" unpackWARs="true" xmlNamespaceAware="false" xmlValidation="false">     <Context docBase="myproject" path="/myproject" reloadable="true" source="org.eclipse.jst.jee.server:myproject"/></Host> </Engine> </Service> </Server>

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Integrating JavaScript Unit Tests with Visual Studio

- by Stephen Walther

Modern ASP.NET web applications take full advantage of client-side JavaScript to provide better interactivity and responsiveness. If you are building an ASP.NET application in the right way, you quickly end up with lots and lots of JavaScript code. When writing server code, you should be writing unit tests. One big advantage of unit tests is that they provide you with a safety net that enable you to safely modify your existing code – for example, fix bugs, add new features, and make performance enhancements -- without breaking your existing code. Every time you modify your code, you can execute your unit tests to verify that you have not broken anything. For the same reason that you should write unit tests for your server code, you should write unit tests for your client code. JavaScript is just as susceptible to bugs as C#. There is no shortage of unit testing frameworks for JavaScript. Each of the major JavaScript libraries has its own unit testing framework. For example, jQuery has QUnit, Prototype has UnitTestJS, YUI has YUI Test, and Dojo has Dojo Objective Harness (DOH). The challenge is integrating a JavaScript unit testing framework with Visual Studio. Visual Studio and Visual Studio ALM provide fantastic support for server-side unit tests. You can easily view the results of running your unit tests in the Visual Studio Test Results window. You can set up a check-in policy which requires that all unit tests pass before your source code can be committed to the source code repository. In addition, you can set up Team Build to execute your unit tests automatically. Unfortunately, Visual Studio does not provide “out-of-the-box” support for JavaScript unit tests. MS Test, the unit testing framework included in Visual Studio, does not support JavaScript unit tests. As soon as you leave the server world, you are left on your own. The goal of this blog entry is to describe one approach to integrating JavaScript unit tests with MS Test so that you can execute your JavaScript unit tests side-by-side with your C# unit tests. The goal is to enable you to execute JavaScript unit tests in exactly the same way as server-side unit tests. You can download the source code described by this project by scrolling to the end of this blog entry. Rejected Approach: Browser Launchers One popular approach to executing JavaScript unit tests is to use a browser as a test-driver. When you use a browser as a test-driver, you open up a browser window to execute and view the results of executing your JavaScript unit tests. For example, QUnit – the unit testing framework for jQuery – takes this approach. The following HTML page illustrates how you can use QUnit to create a unit test for a function named addNumbers(). <!DOCTYPE HTML PUBLIC "-//W3C//DTD HTML 4.01 Transitional//EN" "http://www.w3.org/TR/html4/loose.dtd"> <html> <head> <title>Using QUnit</title> <link rel="stylesheet" href="http://github.com/jquery/qunit/raw/master/qunit/qunit.css" type="text/css" /> </head> <body> <h1 id="qunit-header">QUnit example</h1> <h2 id="qunit-banner"></h2> <div id="qunit-testrunner-toolbar"></div> <h2 id="qunit-userAgent"></h2> <ol id="qunit-tests"></ol> <div id="qunit-fixture">test markup, will be hidden</div> <script type="text/javascript" src="http://code.jquery.com/jquery-latest.js"></script> <script type="text/javascript" src="http://github.com/jquery/qunit/raw/master/qunit/qunit.js"></script> <script type="text/javascript"> // The function to test function addNumbers(a, b) { return a+b; } // The unit test test("Test of addNumbers", function () { equals(4, addNumbers(1,3), "1+3 should be 4"); }); </script> </body> </html> This test verifies that calling addNumbers(1,3) returns the expected value 4. When you open this page in a browser, you can see that this test does, in fact, pass. The idea is that you can quickly refresh this QUnit HTML JavaScript test driver page in your browser whenever you modify your JavaScript code. In other words, you can keep a browser window open and keep refreshing it over and over while you are developing your application. That way, you can know very quickly whenever you have broken your JavaScript code. While easy to setup, there are several big disadvantages to this approach to executing JavaScript unit tests: You must view your JavaScript unit test results in a different location than your server unit test results. The JavaScript unit test results appear in the browser and the server unit test results appear in the Visual Studio Test Results window. Because all of your unit test results don’t appear in a single location, you are more likely to introduce bugs into your code without noticing it. Because your unit tests are not integrated with Visual Studio – in particular, MS Test -- you cannot easily include your JavaScript unit tests when setting up check-in policies or when performing automated builds with Team Build. A more sophisticated approach to using a browser as a test-driver is to automate the web browser. Instead of launching the browser and loading the test code yourself, you use a framework to automate this process. There are several different testing frameworks that support this approach: · Selenium – Selenium is a very powerful framework for automating browser tests. You can create your tests by recording a Firefox session or by writing the test driver code in server code such as C#. You can learn more about Selenium at http://seleniumhq.org/. LTAF – The ASP.NET team uses the Lightweight Test Automation Framework to test JavaScript code in the ASP.NET framework. You can learn more about LTAF by visiting the project home at CodePlex: http://aspnet.codeplex.com/releases/view/35501 jsTestDriver – This framework uses Java to automate the browser. jsTestDriver creates a server which can be used to automate multiple browsers simultaneously. This project is located at http://code.google.com/p/js-test-driver/ TestSwam – This framework, created by John Resig, uses PHP to automate the browser. Like jsTestDriver, the framework creates a test server. You can open multiple browsers that are automated by the test server. Learn more about TestSwarm by visiting the following address: https://github.com/jeresig/testswarm/wiki Yeti – This is the framework introduced by Yahoo for automating browser tests. Yeti uses server-side JavaScript and depends on Node.js. Learn more about Yeti at http://www.yuiblog.com/blog/2010/08/25/introducing-yeti-the-yui-easy-testing-interface/ All of these frameworks are great for integration tests – however, they are not the best frameworks to use for unit tests. In one way or another, all of these frameworks depend on executing tests within the context of a “living and breathing” browser. If you create an ASP.NET Unit Test then Visual Studio will launch a web server before executing the unit test. Why is launching a web server so bad? It is not the worst thing in the world. However, it does introduce dependencies that prevent your code from being tested in isolation. One of the defining features of a unit test -- versus an integration test – is that a unit test tests code in isolation. Another problem with launching a web server when performing unit tests is that launching a web server can be slow. If you cannot execute your unit tests quickly, you are less likely to execute your unit tests each and every time you make a code change. You are much more likely to fall into the pit of failure. Launching a browser when performing a JavaScript unit test has all of the same disadvantages as launching a web server when performing an ASP.NET unit test. Instead of testing a unit of JavaScript code in isolation, you are testing JavaScript code within the context of a particular browser. Using the frameworks listed above for integration tests makes perfect sense. However, I want to consider a different approach for creating unit tests for JavaScript code. Using Server-Side JavaScript for JavaScript Unit Tests A completely different approach to executing JavaScript unit tests is to perform the tests outside of any browser. If you really want to test JavaScript then you should test JavaScript and leave the browser out of the testing process. There are several ways that you can execute JavaScript on the server outside the context of any browser: Rhino – Rhino is an implementation of JavaScript written in Java. The Rhino project is maintained by the Mozilla project. Learn more about Rhino at http://www.mozilla.org/rhino/ V8 – V8 is the open-source Google JavaScript engine written in C++. This is the JavaScript engine used by the Chrome web browser. You can download V8 and embed it in your project by visiting http://code.google.com/p/v8/ JScript – JScript is the JavaScript Script Engine used by Internet Explorer (up to but not including Internet Explorer 9), Windows Script Host, and Active Server Pages. Internet Explorer is still the most popular web browser. Therefore, I decided to focus on using the JScript Script Engine to execute JavaScript unit tests. Using the Microsoft Script Control There are two basic ways that you can pass JavaScript to the JScript Script Engine and execute the code: use the Microsoft Windows Script Interfaces or use the Microsoft Script Control. The difficult and proper way to execute JavaScript using the JScript Script Engine is to use the Microsoft Windows Script Interfaces. You can learn more about the Script Interfaces by visiting http://msdn.microsoft.com/en-us/library/t9d4xf28(VS.85).aspx The main disadvantage of using the Script Interfaces is that they are difficult to use from .NET. There is a great series of articles on using the Script Interfaces from C# located at http://www.drdobbs.com/184406028. I picked the easier alternative and used the Microsoft Script Control. The Microsoft Script Control is an ActiveX control that provides a higher level abstraction over the Window Script Interfaces. You can download the Microsoft Script Control from here: http://www.microsoft.com/downloads/en/details.aspx?FamilyID=d7e31492-2595-49e6-8c02-1426fec693ac After you download the Microsoft Script Control, you need to add a reference to it to your project. Select the Visual Studio menu option Project, Add Reference to open the Add Reference dialog. Select the COM tab and add the Microsoft Script Control 1.0. Using the Script Control is easy. You call the Script Control AddCode() method to add JavaScript code to the Script Engine. Next, you call the Script Control Run() method to run a particular JavaScript function. The reference documentation for the Microsoft Script Control is located at the MSDN website: http://msdn.microsoft.com/en-us/library/aa227633%28v=vs.60%29.aspx Creating the JavaScript Code to Test To keep things simple, let’s imagine that you want to test the following JavaScript function named addNumbers() which simply adds two numbers together: MvcApplication1\Scripts\Math.js function addNumbers(a, b) { return 5; } Notice that the addNumbers() method always returns the value 5. Right-now, it will not pass a good unit test. Create this file and save it in your project with the name Math.js in your MVC project’s Scripts folder (Save the file in your actual MVC application and not your MVC test application). Creating the JavaScript Test Helper Class To make it easier to use the Microsoft Script Control in unit tests, we can create a helper class. This class contains two methods: LoadFile() – Loads a JavaScript file. Use this method to load the JavaScript file being tested or the JavaScript file containing the unit tests. ExecuteTest() – Executes the JavaScript code. Use this method to execute a JavaScript unit test. Here’s the code for the JavaScriptTestHelper class: JavaScriptTestHelper.cs using System; using System.IO; using Microsoft.VisualStudio.TestTools.UnitTesting; using MSScriptControl; namespace MvcApplication1.Tests { public class JavaScriptTestHelper : IDisposable { private ScriptControl _sc; private TestContext _context; /// <summary> /// You need to use this helper with Unit Tests and not /// Basic Unit Tests because you need a Test Context /// </summary> /// <param name="testContext">Unit Test Test Context</param> public JavaScriptTestHelper(TestContext testContext) { if (testContext == null) { throw new ArgumentNullException("TestContext"); } _context = testContext; _sc = new ScriptControl(); _sc.Language = "JScript"; _sc.AllowUI = false; } /// <summary> /// Load the contents of a JavaScript file into the /// Script Engine. /// </summary> /// <param name="path">Path to JavaScript file</param> public void LoadFile(string path) { var fileContents = File.ReadAllText(path); _sc.AddCode(fileContents); } /// <summary> /// Pass the path of the test that you want to execute. /// </summary> /// <param name="testMethodName">JavaScript function name</param> public void ExecuteTest(string testMethodName) { dynamic result = null; try { result = _sc.Run(testMethodName, new object[] { }); } catch { var error = ((IScriptControl)_sc).Error; if (error != null) { var description = error.Description; var line = error.Line; var column = error.Column; var text = error.Text; var source = error.Source; if (_context != null) { var details = String.Format("{0} \r\nLine: {1} Column: {2}", source, line, column); _context.WriteLine(details); } } throw new AssertFailedException(error.Description); } } public void Dispose() { _sc = null; } } } Notice that the JavaScriptTestHelper class requires a Test Context to be instantiated. For this reason, you can use the JavaScriptTestHelper only with a Visual Studio Unit Test and not a Basic Unit Test (These are two different types of Visual Studio project items). Add the JavaScriptTestHelper file to your MVC test application (for example, MvcApplication1.Tests). Creating the JavaScript Unit Test Next, we need to create the JavaScript unit test function that we will use to test the addNumbers() function. Create a folder in your MVC test project named JavaScriptTests and add the following JavaScript file to this folder: MvcApplication1.Tests\JavaScriptTests\MathTest.js /// <reference path="JavaScriptUnitTestFramework.js"/> function testAddNumbers() { // Act var result = addNumbers(1, 3); // Assert assert.areEqual(4, result, "addNumbers did not return right value!"); } The testAddNumbers() function takes advantage of another JavaScript library named JavaScriptUnitTestFramework.js. This library contains all of the code necessary to make assertions. Add the following JavaScriptnitTestFramework.js to the same folder as the MathTest.js file: MvcApplication1.Tests\JavaScriptTests\JavaScriptUnitTestFramework.js var assert = { areEqual: function (expected, actual, message) { if (expected !== actual) { throw new Error("Expected value " + expected + " is not equal to " + actual + ". " + message); } } }; There is only one type of assertion supported by this file: the areEqual() assertion. Most likely, you would want to add additional types of assertions to this file to make it easier to write your JavaScript unit tests. Deploying the JavaScript Test Files This step is non-intuitive. When you use Visual Studio to run unit tests, Visual Studio creates a new folder and executes a copy of the files in your project. After you run your unit tests, your Visual Studio Solution will contain a new folder named TestResults that includes a subfolder for each test run. You need to configure Visual Studio to deploy your JavaScript files to the test run folder or Visual Studio won’t be able to find your JavaScript files when you execute your unit tests. You will get an error that looks something like this when you attempt to execute your unit tests: You can configure Visual Studio to deploy your JavaScript files by adding a Test Settings file to your Visual Studio Solution. It is important to understand that you need to add this file to your Visual Studio Solution and not a particular Visual Studio project. Right-click your Solution in the Solution Explorer window and select the menu option Add, New Item. Select the Test Settings item and click the Add button. After you create a Test Settings file for your solution, you can indicate that you want a particular folder to be deployed whenever you perform a test run. Select the menu option Test, Edit Test Settings to edit your test configuration file. Select the Deployment tab and select your MVC test project’s JavaScriptTest folder to deploy. Click the Apply button and the Close button to save the changes and close the dialog. Creating the Visual Studio Unit Test The very last step is to create the Visual Studio unit test (the MS Test unit test). Add a new unit test to your MVC test project by selecting the menu option Add New Item and selecting the Unit Test project item (Do not select the Basic Unit Test project item): The difference between a Basic Unit Test and a Unit Test is that a Unit Test includes a Test Context. We need this Test Context to use the JavaScriptTestHelper class that we created earlier. Enter the following test method for the new unit test: [TestMethod] public void TestAddNumbers() { var jsHelper = new JavaScriptTestHelper(this.TestContext); // Load JavaScript files jsHelper.LoadFile("JavaScriptUnitTestFramework.js"); jsHelper.LoadFile(@"..\..\..\MvcApplication1\Scripts\Math.js"); jsHelper.LoadFile("MathTest.js"); // Execute JavaScript Test jsHelper.ExecuteTest("testAddNumbers"); } This code uses the JavaScriptTestHelper to load three files: JavaScripUnitTestFramework.js – Contains the assert functions. Math.js – Contains the addNumbers() function from your MVC application which is being tested. MathTest.js – Contains the JavaScript unit test function. Next, the test method calls the JavaScriptTestHelper ExecuteTest() method to execute the testAddNumbers() JavaScript function. Running the Visual Studio JavaScript Unit Test After you complete all of the steps described above, you can execute the JavaScript unit test just like any other unit test. You can use the keyboard combination CTRL-R, CTRL-A to run all of the tests in the current Visual Studio Solution. Alternatively, you can use the buttons in the Visual Studio toolbar to run the tests: (Unfortunately, the Run All Impacted Tests button won’t work correctly because Visual Studio won’t detect that your JavaScript code has changed. Therefore, you should use either the Run Tests in Current Context or Run All Tests in Solution options instead.) The results of running the JavaScript tests appear side-by-side with the results of running the server tests in the Test Results window. For example, if you Run All Tests in Solution then you will get the following results: Notice that the TestAddNumbers() JavaScript test has failed. That is good because our addNumbers() function is hard-coded to always return the value 5. If you double-click the failing JavaScript test, you can view additional details such as the JavaScript error message and the line number of the JavaScript code that failed: Summary The goal of this blog entry was to explain an approach to creating JavaScript unit tests that can be easily integrated with Visual Studio and Visual Studio ALM. I described how you can use the Microsoft Script Control to execute JavaScript on the server. By taking advantage of the Microsoft Script Control, we were able to execute our JavaScript unit tests side-by-side with all of our other unit tests and view the results in the standard Visual Studio Test Results window. You can download the code discussed in this blog entry from here: http://StephenWalther.com/downloads/Blog/JavaScriptUnitTesting/JavaScriptUnitTests.zip Before running this code, you need to first install the Microsoft Script Control which you can download from here: http://www.microsoft.com/downloads/en/details.aspx?FamilyID=d7e31492-2595-49e6-8c02-1426fec693ac

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SimpleMembership, Membership Providers, Universal Providers and the new ASP.NET 4.5 Web Forms and ASP.NET MVC 4 templates

- by Jon Galloway

The ASP.NET MVC 4 Internet template adds some new, very useful features which are built on top of SimpleMembership. These changes add some great features, like a much simpler and extensible membership API and support for OAuth. However, the new account management features require SimpleMembership and won't work against existing ASP.NET Membership Providers. I'll start with a summary of top things you need to know, then dig into a lot more detail. Summary: SimpleMembership has been designed as a replacement for traditional the previous ASP.NET Role and Membership provider system SimpleMembership solves common problems people ran into with the Membership provider system and was designed for modern user / membership / storage needs SimpleMembership integrates with the previous membership system, but you can't use a MembershipProvider with SimpleMembership The new ASP.NET MVC 4 Internet application template AccountController requires SimpleMembership and is not compatible with previous MembershipProviders You can continue to use existing ASP.NET Role and Membership providers in ASP.NET 4.5 and ASP.NET MVC 4 - just not with the ASP.NET MVC 4 AccountController The existing ASP.NET Role and Membership provider system remains supported as is part of the ASP.NET core ASP.NET 4.5 Web Forms does not use SimpleMembership; it implements OAuth on top of ASP.NET Membership The ASP.NET Web Site Administration Tool (WSAT) is not compatible with SimpleMembership The following is the result of a few conversations with Erik Porter (PM for ASP.NET MVC) to make sure I had some the overall details straight, combined with a lot of time digging around in ILSpy and Visual Studio's assembly browsing tools. SimpleMembership: The future of membership for ASP.NET The ASP.NET Membership system was introduces with ASP.NET 2.0 back in 2005. It was designed to solve common site membership requirements at the time, which generally involved username / password based registration and profile storage in SQL Server. It was designed with a few extensibility mechanisms - notably a provider system (which allowed you override some specifics like backing storage) and the ability to store additional profile information (although the additional profile information was packed into a single column which usually required access through the API). While it's sometimes frustrating to work with, it's held up for seven years - probably since it handles the main use case (username / password based membership in a SQL Server database) smoothly and can be adapted to most other needs (again, often frustrating, but it can work). The ASP.NET Web Pages and WebMatrix efforts allowed the team an opportunity to take a new look at a lot of things - e.g. the Razor syntax started with ASP.NET Web Pages, not ASP.NET MVC. The ASP.NET Web Pages team designed SimpleMembership to (wait for it) simplify the task of dealing with membership. As Matthew Osborn said in his post Using SimpleMembership With ASP.NET WebPages: With the introduction of ASP.NET WebPages and the WebMatrix stack our team has really be focusing on making things simpler for the developer. Based on a lot of customer feedback one of the areas that we wanted to improve was the built in security in ASP.NET. So with this release we took that time to create a new built in (and default for ASP.NET WebPages) security provider. I say provider because the new stuff is still built on the existing ASP.NET framework. So what do we call this new hotness that we have created? Well, none other than SimpleMembership. SimpleMembership is an umbrella term for both SimpleMembership and SimpleRoles. Part of simplifying membership involved fixing some common problems with ASP.NET Membership. Problems with ASP.NET Membership ASP.NET Membership was very obviously designed around a set of assumptions: Users and user information would most likely be stored in a full SQL Server database or in Active Directory User and profile information would be optimized around a set of common attributes (UserName, Password, IsApproved, CreationDate, Comment, Role membership...) and other user profile information would be accessed through a profile provider Some problems fall out of these assumptions. Requires Full SQL Server for default cases The default, and most fully featured providers ASP.NET Membership providers (SQL Membership Provider, SQL Role Provider, SQL Profile Provider) require full SQL Server. They depend on stored procedure support, and they rely on SQL Server cache dependencies, they depend on agents for clean up and maintenance. So the main SQL Server based providers don't work well on SQL Server CE, won't work out of the box on SQL Azure, etc. Note: Cory Fowler recently let me know about these Updated ASP.net scripts for use with Microsoft SQL Azure which do support membership, personalization, profile, and roles. But the fact that we need a support page with a set of separate SQL scripts underscores the underlying problem. Aha, you say! Jon's forgetting the Universal Providers, a.k.a. System.Web.Providers! Hold on a bit, we'll get to those... Custom Membership Providers have to work with a SQL-Server-centric API If you want to work with another database or other membership storage system, you need to to inherit from the provider base classes and override a bunch of methods which are tightly focused on storing a MembershipUser in a relational database. It can be done (and you can often find pretty good ones that have already been written), but it's a good amount of work and often leaves you with ugly code that has a bunch of System.NotImplementedException fun since there are a lot of methods that just don't apply. Designed around a specific view of users, roles and profiles The existing providers are focused on traditional membership - a user has a username and a password, some specific roles on the site (e.g. administrator, premium user), and may have some additional "nice to have" optional information that can be accessed via an API in your application. This doesn't fit well with some modern usage patterns: In OAuth and OpenID, the user doesn't have a password Often these kinds of scenarios map better to user claims or rights instead of monolithic user roles For many sites, profile or other non-traditional information is very important and needs to come from somewhere other than an API call that maps to a database blob What would work a lot better here is a system in which you were able to define your users, rights, and other attributes however you wanted and the membership system worked with your model - not the other way around. Requires specific schema, overflow in blob columns I've already mentioned this a few times, but it bears calling out separately - ASP.NET Membership focuses on SQL Server storage, and that storage is based on a very specific database schema. SimpleMembership as a better membership system As you might have guessed, SimpleMembership was designed to address the above problems. Works with your Schema As Matthew Osborn explains in his Using SimpleMembership With ASP.NET WebPages post, SimpleMembership is designed to integrate with your database schema: All SimpleMembership requires is that there are two columns on your users table so that we can hook up to it – an “ID” column and a “username” column. The important part here is that they can be named whatever you want. For instance username doesn't have to be an alias it could be an email column you just have to tell SimpleMembership to treat that as the “username” used to log in. Matthew's example shows using a very simple user table named Users (it could be named anything) with a UserID and Username column, then a bunch of other columns he wanted in his app. Then we point SimpleMemberhip at that table with a one-liner: WebSecurity.InitializeDatabaseFile("SecurityDemo.sdf", "Users", "UserID", "Username", true); No other tables are needed, the table can be named anything we want, and can have pretty much any schema we want as long as we've got an ID and something that we can map to a username. Broaden database support to the whole SQL Server family While SimpleMembership is not database agnostic, it works across the SQL Server family. It continues to support full SQL Server, but it also works with SQL Azure, SQL Server CE, SQL Server Express, and LocalDB. Everything's implemented as SQL calls rather than requiring stored procedures, views, agents, and change notifications. Note that SimpleMembership still requires some flavor of SQL Server - it won't work with MySQL, NoSQL databases, etc. You can take a look at the code in WebMatrix.WebData.dll using a tool like ILSpy if you'd like to see why - there places where SQL Server specific SQL statements are being executed, especially when creating and initializing tables. It seems like you might be able to work with another database if you created the tables separately, but I haven't tried it and it's not supported at this point. Note: I'm thinking it would be possible for SimpleMembership (or something compatible) to run Entity Framework so it would work with any database EF supports. That seems useful to me - thoughts? Note: SimpleMembership has the same database support - anything in the SQL Server family - that Universal Providers brings to the ASP.NET Membership system. Easy to with Entity Framework Code First The problem with with ASP.NET Membership's system for storing additional account information is that it's the gate keeper. That means you're stuck with its schema and accessing profile information through its API. SimpleMembership flips that around by allowing you to use any table as a user store. That means you're in control of the user profile information, and you can access it however you'd like - it's just data. Let's look at a practical based on the AccountModel.cs class in an ASP.NET MVC 4 Internet project. Here I'm adding a Birthday property to the UserProfile class. [Table("UserProfile")] public class UserProfile { [Key] [DatabaseGeneratedAttribute(DatabaseGeneratedOption.Identity)] public int UserId { get; set; } public string UserName { get; set; } public DateTime Birthday { get; set; } } Now if I want to access that information, I can just grab the account by username and read the value. var context = new UsersContext(); var username = User.Identity.Name; var user = context.UserProfiles.SingleOrDefault(u => u.UserName == username); var birthday = user.Birthday; So instead of thinking of SimpleMembership as a big membership API, think of it as something that handles membership based on your user database. In SimpleMembership, everything's keyed off a user row in a table you define rather than a bunch of entries in membership tables that were out of your control. How SimpleMembership integrates with ASP.NET Membership Okay, enough sales pitch (and hopefully background) on why things have changed. How does this affect you? Let's start with a diagram to show the relationship (note: I've simplified by removing a few classes to show the important relationships): So SimpleMembershipProvider is an implementaiton of an ExtendedMembershipProvider, which inherits from MembershipProvider and adds some other account / OAuth related things. Here's what ExtendedMembershipProvider adds to MembershipProvider: The important thing to take away here is that a SimpleMembershipProvider is a MembershipProvider, but a MembershipProvider is not a SimpleMembershipProvider. This distinction is important in practice: you cannot use an existing MembershipProvider (including the Universal Providers found in System.Web.Providers) with an API that requires a SimpleMembershipProvider, including any of the calls in WebMatrix.WebData.WebSecurity or Microsoft.Web.WebPages.OAuth.OAuthWebSecurity. However, that's as far as it goes. Membership Providers still work if you're accessing them through the standard Membership API, and all of the core stuff - including the AuthorizeAttribute, role enforcement, etc. - will work just fine and without any change. Let's look at how that affects you in terms of the new templates. Membership in the ASP.NET MVC 4 project templates ASP.NET MVC 4 offers six Project Templates: Empty - Really empty, just the assemblies, folder structure and a tiny bit of basic configuration. Basic - Like Empty, but with a bit of UI preconfigured (css / images / bundling). Internet - This has both a Home and Account controller and associated views. The Account Controller supports registration and login via either local accounts and via OAuth / OpenID providers. Intranet - Like the Internet template, but it's preconfigured for Windows Authentication. Mobile - This is preconfigured using jQuery Mobile and is intended for mobile-only sites. Web API - This is preconfigured for a service backend built on ASP.NET Web API. Out of these templates, only one (the Internet template) uses SimpleMembership. ASP.NET MVC 4 Basic template The Basic template has configuration in place to use ASP.NET Membership with the Universal Providers. You can see that configuration in the ASP.NET MVC 4 Basic template's web.config: <profile defaultProvider="DefaultProfileProvider"> <providers> <add name="DefaultProfileProvider" type="System.Web.Providers.DefaultProfileProvider, System.Web.Providers, Version=1.0.0.0, Culture=neutral, PublicKeyToken=31bf3856ad364e35" connectionStringName="DefaultConnection" applicationName="/" /> </providers> </profile> <membership defaultProvider="DefaultMembershipProvider"> <providers> <add name="DefaultMembershipProvider" type="System.Web.Providers.DefaultMembershipProvider, System.Web.Providers, Version=1.0.0.0, Culture=neutral, PublicKeyToken=31bf3856ad364e35" connectionStringName="DefaultConnection" enablePasswordRetrieval="false" enablePasswordReset="true" requiresQuestionAndAnswer="false" requiresUniqueEmail="false" maxInvalidPasswordAttempts="5" minRequiredPasswordLength="6" minRequiredNonalphanumericCharacters="0" passwordAttemptWindow="10" applicationName="/" /> </providers> </membership> <roleManager defaultProvider="DefaultRoleProvider"> <providers> <add name="DefaultRoleProvider" type="System.Web.Providers.DefaultRoleProvider, System.Web.Providers, Version=1.0.0.0, Culture=neutral, PublicKeyToken=31bf3856ad364e35" connectionStringName="DefaultConnection" applicationName="/" /> </providers> </roleManager> <sessionState mode="InProc" customProvider="DefaultSessionProvider"> <providers> <add name="DefaultSessionProvider" type="System.Web.Providers.DefaultSessionStateProvider, System.Web.Providers, Version=1.0.0.0, Culture=neutral, PublicKeyToken=31bf3856ad364e35" connectionStringName="DefaultConnection" /> </providers> </sessionState> This means that it's business as usual for the Basic template as far as ASP.NET Membership works. ASP.NET MVC 4 Internet template The Internet template has a few things set up to bootstrap SimpleMembership: \Models\AccountModels.cs defines a basic user account and includes data annotations to define keys and such \Filters\InitializeSimpleMembershipAttribute.cs creates the membership database using the above model, then calls WebSecurity.InitializeDatabaseConnection which verifies that the underlying tables are in place and marks initialization as complete (for the application's lifetime) \Controllers\AccountController.cs makes heavy use of OAuthWebSecurity (for OAuth account registration / login / management) and WebSecurity. WebSecurity provides account management services for ASP.NET MVC (and Web Pages) WebSecurity can work with any ExtendedMembershipProvider. There's one in the box (SimpleMembershipProvider) but you can write your own. Since a standard MembershipProvider is not an ExtendedMembershipProvider, WebSecurity will throw exceptions if the default membership provider is a MembershipProvider rather than an ExtendedMembershipProvider. Practical example: Create a new ASP.NET MVC 4 application using the Internet application template Install the Microsoft ASP.NET Universal Providers for LocalDB NuGet package Run the application, click on Register, add a username and password, and click submit You'll get the following execption in AccountController.cs::Register: To call this method, the "Membership.Provider" property must be an instance of "ExtendedMembershipProvider". This occurs because the ASP.NET Universal Providers packages include a web.config transform that will update your web.config to add the Universal Provider configuration I showed in the Basic template example above. When WebSecurity tries to use the configured ASP.NET Membership Provider, it checks if it can be cast to an ExtendedMembershipProvider before doing anything else. So, what do you do? Options: If you want to use the new AccountController, you'll either need to use the SimpleMembershipProvider or another valid ExtendedMembershipProvider. This is pretty straightforward. If you want to use an existing ASP.NET Membership Provider in ASP.NET MVC 4, you can't use the new AccountController. You can do a few things: Replace the AccountController.cs and AccountModels.cs in an ASP.NET MVC 4 Internet project with one from an ASP.NET MVC 3 application (you of course won't have OAuth support). Then, if you want, you can go through and remove other things that were built around SimpleMembership - the OAuth partial view, the NuGet packages (e.g. the DotNetOpenAuthAuth package, etc.) Use an ASP.NET MVC 4 Internet application template and add in a Universal Providers NuGet package. Then copy in the AccountController and AccountModel classes. Create an ASP.NET MVC 3 project and upgrade it to ASP.NET MVC 4 using the steps shown in the ASP.NET MVC 4 release notes. None of these are particularly elegant or simple. Maybe we (or just me?) can do something to make this simpler - perhaps a NuGet package. However, this should be an edge case - hopefully the cases where you'd need to create a new ASP.NET but use legacy ASP.NET Membership Providers should be pretty rare. Please let me (or, preferably the team) know if that's an incorrect assumption. Membership in the ASP.NET 4.5 project template ASP.NET 4.5 Web Forms took a different approach which builds off ASP.NET Membership. Instead of using the WebMatrix security assemblies, Web Forms uses Microsoft.AspNet.Membership.OpenAuth assembly. I'm no expert on this, but from a bit of time in ILSpy and Visual Studio's (very pretty) dependency graphs, this uses a Membership Adapter to save OAuth data into an EF managed database while still running on top of ASP.NET Membership. Note: There may be a way to use this in ASP.NET MVC 4, although it would probably take some plumbing work to hook it up. How does this fit in with Universal Providers (System.Web.Providers)? Just to summarize: Universal Providers are intended for cases where you have an existing ASP.NET Membership Provider and you want to use it with another SQL Server database backend (other than SQL Server). It doesn't require agents to handle expired session cleanup and other background tasks, it piggybacks these tasks on other calls. Universal Providers are not really, strictly speaking, universal - at least to my way of thinking. They only work with databases in the SQL Server family. Universal Providers do not work with Simple Membership. The Universal Providers packages include some web config transforms which you would normally want when you're using them. What about the Web Site Administration Tool? Visual Studio includes tooling to launch the Web Site Administration Tool (WSAT) to configure users and roles in your application. WSAT is built to work with ASP.NET Membership, and is not compatible with Simple Membership. There are two main options there: Use the WebSecurity and OAuthWebSecurity API to manage the users and roles Create a web admin using the above APIs Since SimpleMembership runs on top of your database, you can update your users as you would any other data - via EF or even in direct database edits (in development, of course)

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Introducing the Earthquake Locator – A Bing Maps Silverlight Application, part 1

- by Bobby Diaz

Update: Live demo and source code now available! The recent wave of earthquakes (no pun intended) being reported in the news got me wondering about the frequency and severity of earthquakes around the world. Since I’ve been doing a lot of Silverlight development lately, I decided to scratch my curiosity with a nice little Bing Maps application that will show the location and relative strength of recent seismic activity. Here is a list of technologies this application will utilize, so be sure to have everything downloaded and installed if you plan on following along. Silverlight 3 WCF RIA Services Bing Maps Silverlight Control * Managed Extensibility Framework (optional) MVVM Light Toolkit (optional) log4net (optional) * If you are new to Bing Maps or have not signed up for a Developer Account, you will need to visit www.bingmapsportal.com to request a Bing Maps key for your application. Getting Started We start out by creating a new Silverlight Application called EarthquakeLocator and specify that we want to automatically create the Web Application Project with RIA Services enabled. I cleaned up the web app by removing the Default.aspx and EarthquakeLocatorTestPage.html. Then I renamed the EarthquakeLocatorTestPage.aspx to Default.aspx and set it as my start page. I also set the development server to use a specific port, as shown below. RIA Services Next, I created a Services folder in the EarthquakeLocator.Web project and added a new Domain Service Class called EarthquakeService.cs. This is the RIA Services Domain Service that will provide earthquake data for our client application. I am not using LINQ to SQL or Entity Framework, so I will use the <empty domain service class> option. We will be pulling data from an external Atom feed, but this example could just as easily pull data from a database or another web service. This is an important distinction to point out because each scenario I just mentioned could potentially use a different Domain Service base class (i.e. LinqToSqlDomainService<TDataContext>). Now we can start adding Query methods to our EarthquakeService that pull data from the USGS web site. Here is the complete code for our service class: using System; using System.Collections.Generic; using System.IO; using System.Linq; using System.ServiceModel.Syndication; using System.Web.DomainServices; using System.Web.Ria; using System.Xml; using log4net; using EarthquakeLocator.Web.Model; namespace EarthquakeLocator.Web.Services { /// <summary> /// Provides earthquake data to client applications. /// </summary> [EnableClientAccess()] public class EarthquakeService : DomainService { private static readonly ILog log = LogManager.GetLogger(typeof(EarthquakeService)); // USGS Data Feeds: http://earthquake.usgs.gov/earthquakes/catalogs/ private const string FeedForPreviousDay = "http://earthquake.usgs.gov/earthquakes/catalogs/1day-M2.5.xml"; private const string FeedForPreviousWeek = "http://earthquake.usgs.gov/earthquakes/catalogs/7day-M2.5.xml"; /// <summary> /// Gets the earthquake data for the previous week. /// </summary> /// <returns>A queryable collection of <see cref="Earthquake"/> objects.</returns> public IQueryable<Earthquake> GetEarthquakes() { var feed = GetFeed(FeedForPreviousWeek); var list = new List<Earthquake>(); if ( feed != null ) { foreach ( var entry in feed.Items ) { var quake = CreateEarthquake(entry); if ( quake != null ) { list.Add(quake); } } } return list.AsQueryable(); } /// <summary> /// Creates an <see cref="Earthquake"/> object for each entry in the Atom feed. /// </summary> /// <param name="entry">The Atom entry.</param> /// <returns></returns> private Earthquake CreateEarthquake(SyndicationItem entry) { Earthquake quake = null; string title = entry.Title.Text; string summary = entry.Summary.Text; string point = GetElementValue<String>(entry, "point"); string depth = GetElementValue<String>(entry, "elev"); string utcTime = null; string localTime = null; string depthDesc = null; double? magnitude = null; double? latitude = null; double? longitude = null; double? depthKm = null; if ( !String.IsNullOrEmpty(title) && title.StartsWith("M") ) { title = title.Substring(2, title.IndexOf(',')-3).Trim(); magnitude = TryParse(title); } if ( !String.IsNullOrEmpty(point) ) { var values = point.Split(' '); if ( values.Length == 2 ) { latitude = TryParse(values[0]); longitude = TryParse(values[1]); } } if ( !String.IsNullOrEmpty(depth) ) { depthKm = TryParse(depth); if ( depthKm != null ) { depthKm = Math.Round((-1 * depthKm.Value) / 100, 2); } } if ( !String.IsNullOrEmpty(summary) ) { summary = summary.Replace("</p>", ""); var values = summary.Split( new string[] { "<p>" }, StringSplitOptions.RemoveEmptyEntries); if ( values.Length == 3 ) { var times = values[1].Split( new string[] { "<br>" }, StringSplitOptions.RemoveEmptyEntries); if ( times.Length > 0 ) { utcTime = times[0]; } if ( times.Length > 1 ) { localTime = times[1]; } depthDesc = values[2]; depthDesc = "Depth: " + depthDesc.Substring(depthDesc.IndexOf(":") + 2); } } if ( latitude != null && longitude != null ) { quake = new Earthquake() { Id = entry.Id, Title = entry.Title.Text, Summary = entry.Summary.Text, Date = entry.LastUpdatedTime.DateTime, Url = entry.Links.Select(l => Path.Combine(l.BaseUri.OriginalString, l.Uri.OriginalString)).FirstOrDefault(), Age = entry.Categories.Where(c => c.Label == "Age") .Select(c => c.Name).FirstOrDefault(), Magnitude = magnitude.GetValueOrDefault(), Latitude = latitude.GetValueOrDefault(), Longitude = longitude.GetValueOrDefault(), DepthInKm = depthKm.GetValueOrDefault(), DepthDesc = depthDesc, UtcTime = utcTime, LocalTime = localTime }; } return quake; } private T GetElementValue<T>(SyndicationItem entry, String name) { var el = entry.ElementExtensions.Where(e => e.OuterName == name).FirstOrDefault(); T value = default(T); if ( el != null ) { value = el.GetObject<T>(); } return value; } private double? TryParse(String value) { double d; if ( Double.TryParse(value, out d) ) { return d; } return null; } /// <summary> /// Gets the feed at the specified URL. /// </summary> /// <param name="url">The URL.</param> /// <returns>A <see cref="SyndicationFeed"/> object.</returns> public static SyndicationFeed GetFeed(String url) { SyndicationFeed feed = null; try { log.Debug("Loading RSS feed: " + url); using ( var reader = XmlReader.Create(url) ) { feed = SyndicationFeed.Load(reader); } } catch ( Exception ex ) { log.Error("Error occurred while loading RSS feed: " + url, ex); } return feed; } } } The only method that will be generated in the client side proxy class, EarthquakeContext, will be the GetEarthquakes() method. The reason being that it is the only public instance method and it returns an IQueryable<Earthquake> collection that can be consumed by the client application. GetEarthquakes() calls the static GetFeed(String) method, which utilizes the built in SyndicationFeed API to load the external data feed. You will need to add a reference to the System.ServiceModel.Web library in order to take advantage of the RSS/Atom reader. The API will also allow you to create your own feeds to serve up in your applications. Model I have also created a Model folder and added a new class, Earthquake.cs. The Earthquake object will hold the various properties returned from the Atom feed. Here is a sample of the code for that class. Notice the [Key] attribute on the Id property, which is required by RIA Services to uniquely identify the entity. using System; using System.Collections.Generic; using System.Linq; using System.Runtime.Serialization; using System.ComponentModel.DataAnnotations; namespace EarthquakeLocator.Web.Model { /// <summary> /// Represents an earthquake occurrence and related information. /// </summary> [DataContract] public class Earthquake { /// <summary> /// Gets or sets the id. /// </summary> /// <value>The id.</value> [Key] [DataMember] public string Id { get; set; } /// <summary> /// Gets or sets the title. /// </summary> /// <value>The title.</value> [DataMember] public string Title { get; set; } /// <summary> /// Gets or sets the summary. /// </summary> /// <value>The summary.</value> [DataMember] public string Summary { get; set; } // additional properties omitted } } View Model The recent trend to use the MVVM pattern for WPF and Silverlight provides a great way to separate the data and behavior logic out of the user interface layer of your client applications. I have chosen to use the MVVM Light Toolkit for the Earthquake Locator, but there are other options out there if you prefer another library. That said, I went ahead and created a ViewModel folder in the Silverlight project and added a EarthquakeViewModel class that derives from ViewModelBase. Here is the code: using System; using System.Collections.ObjectModel; using System.ComponentModel.Composition; using System.ComponentModel.Composition.Hosting; using Microsoft.Maps.MapControl; using GalaSoft.MvvmLight; using EarthquakeLocator.Web.Model; using EarthquakeLocator.Web.Services; namespace EarthquakeLocator.ViewModel { /// <summary> /// Provides data for views displaying earthquake information. /// </summary> public class EarthquakeViewModel : ViewModelBase { [Import] public EarthquakeContext Context; /// <summary> /// Initializes a new instance of the <see cref="EarthquakeViewModel"/> class. /// </summary> public EarthquakeViewModel() { var catalog = new AssemblyCatalog(GetType().Assembly); var container = new CompositionContainer(catalog); container.ComposeParts(this); Initialize(); } /// <summary> /// Initializes a new instance of the <see cref="EarthquakeViewModel"/> class. /// </summary> /// <param name="context">The context.</param> public EarthquakeViewModel(EarthquakeContext context) { Context = context; Initialize(); } private void Initialize() { MapCenter = new Location(20, -170); ZoomLevel = 2; } #region Private Methods private void OnAutoLoadDataChanged() { LoadEarthquakes(); } private void LoadEarthquakes() { var query = Context.GetEarthquakesQuery(); Context.Earthquakes.Clear(); Context.Load(query, (op) => { if ( !op.HasError ) { foreach ( var item in op.Entities ) { Earthquakes.Add(item); } } }, null); } #endregion Private Methods #region Properties private bool autoLoadData; /// <summary> /// Gets or sets a value indicating whether to auto load data. /// </summary> /// <value><c>true</c> if auto loading data; otherwise, <c>false</c>.</value> public bool AutoLoadData { get { return autoLoadData; } set { if ( autoLoadData != value ) { autoLoadData = value; RaisePropertyChanged("AutoLoadData"); OnAutoLoadDataChanged(); } } } private ObservableCollection<Earthquake> earthquakes; /// <summary> /// Gets the collection of earthquakes to display. /// </summary> /// <value>The collection of earthquakes.</value> public ObservableCollection<Earthquake> Earthquakes { get { if ( earthquakes == null ) { earthquakes = new ObservableCollection<Earthquake>(); } return earthquakes; } } private Location mapCenter; /// <summary> /// Gets or sets the map center. /// </summary> /// <value>The map center.</value> public Location MapCenter { get { return mapCenter; } set { if ( mapCenter != value ) { mapCenter = value; RaisePropertyChanged("MapCenter"); } } } private double zoomLevel; /// <summary> /// Gets or sets the zoom level. /// </summary> /// <value>The zoom level.</value> public double ZoomLevel { get { return zoomLevel; } set { if ( zoomLevel != value ) { zoomLevel = value; RaisePropertyChanged("ZoomLevel"); } } } #endregion Properties } } The EarthquakeViewModel class contains all of the properties that will be bound to by the various controls in our views. Be sure to read through the LoadEarthquakes() method, which handles calling the GetEarthquakes() method in our EarthquakeService via the EarthquakeContext proxy, and also transfers the loaded entities into the view model’s Earthquakes collection. Another thing to notice is what’s going on in the default constructor. I chose to use the Managed Extensibility Framework (MEF) for my composition needs, but you can use any dependency injection library or none at all. To allow the EarthquakeContext class to be discoverable by MEF, I added the following partial class so that I could supply the appropriate [Export] attribute: using System; using System.ComponentModel.Composition; namespace EarthquakeLocator.Web.Services { /// <summary> /// The client side proxy for the EarthquakeService class. /// </summary> [Export] public partial class EarthquakeContext { } } One last piece I wanted to point out before moving on to the user interface, I added a client side partial class for the Earthquake entity that contains helper properties that we will bind to later: using System; namespace EarthquakeLocator.Web.Model { /// <summary> /// Represents an earthquake occurrence and related information. /// </summary> public partial class Earthquake { /// <summary> /// Gets the location based on the current Latitude/Longitude. /// </summary> /// <value>The location.</value> public string Location { get { return String.Format("{0},{1}", Latitude, Longitude); } } /// <summary> /// Gets the size based on the Magnitude. /// </summary> /// <value>The size.</value> public double Size { get { return (Magnitude * 3); } } } } View Now the fun part! Usually, I would create a Views folder to place all of my View controls in, but I took the easy way out and added the following XAML code to the default MainPage.xaml file. Be sure to add the bing prefix associating the Microsoft.Maps.MapControl namespace after adding the assembly reference to your project. The MVVM Light Toolkit project templates come with a ViewModelLocator class that you can use via a static resource, but I am instantiating the EarthquakeViewModel directly in my user control. I am setting the AutoLoadData property to true as a way to trigger the LoadEarthquakes() method call. The MapItemsControl found within the <bing:Map> control binds its ItemsSource property to the Earthquakes collection of the view model, and since it is an ObservableCollection<T>, we get the automatic two way data binding via the INotifyCollectionChanged interface. <UserControl x:Class="EarthquakeLocator.MainPage" xmlns="http://schemas.microsoft.com/winfx/2006/xaml/presentation" xmlns:x="http://schemas.microsoft.com/winfx/2006/xaml" xmlns:d="http://schemas.microsoft.com/expression/blend/2008" xmlns:mc="http://schemas.openxmlformats.org/markup-compatibility/2006" xmlns:bing="clr-namespace:Microsoft.Maps.MapControl;assembly=Microsoft.Maps.MapControl" xmlns:vm="clr-namespace:EarthquakeLocator.ViewModel" mc:Ignorable="d" d:DesignWidth="640" d:DesignHeight="480" > <UserControl.Resources> <DataTemplate x:Key="EarthquakeTemplate"> <Ellipse Fill="Red" Stroke="Black" StrokeThickness="1" Width="{Binding Size}" Height="{Binding Size}" bing:MapLayer.Position="{Binding Location}" bing:MapLayer.PositionOrigin="Center"> <ToolTipService.ToolTip> <StackPanel> <TextBlock Text="{Binding Title}" FontSize="14" FontWeight="Bold" /> <TextBlock Text="{Binding UtcTime}" /> <TextBlock Text="{Binding LocalTime}" /> <TextBlock Text="{Binding DepthDesc}" /> </StackPanel> </ToolTipService.ToolTip> </Ellipse> </DataTemplate> </UserControl.Resources> <UserControl.DataContext> <vm:EarthquakeViewModel AutoLoadData="True" /> </UserControl.DataContext> <Grid x:Name="LayoutRoot"> <bing:Map x:Name="map" CredentialsProvider="--Your-Bing-Maps-Key--" Center="{Binding MapCenter, Mode=TwoWay}" ZoomLevel="{Binding ZoomLevel, Mode=TwoWay}"> <bing:MapItemsControl ItemsSource="{Binding Earthquakes}" ItemTemplate="{StaticResource EarthquakeTemplate}" /> </bing:Map> </Grid> </UserControl> The EarthquakeTemplate defines the Ellipse that will represent each earthquake, the Width and Height that are determined by the Magnitude, the Position on the map, and also the tooltip that will appear when we mouse over each data point. Running the application will give us the following result (shown with a tooltip example): That concludes this portion of our show but I plan on implementing additional functionality in later blog posts. Be sure to come back soon to see the next installments in this series. Enjoy! Additional Resources USGS Earthquake Data Feeds Brad Abrams shows how RIA Services and MVVM can work together

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Using Durandal to Create Single Page Apps

- by Stephen.Walther

A few days ago, I gave a talk on building Single Page Apps on the Microsoft Stack. In that talk, I recommended that people use Knockout, Sammy, and RequireJS to build their presentation layer and use the ASP.NET Web API to expose data from their server. After I gave the talk, several people contacted me and suggested that I investigate a new open-source JavaScript library named Durandal. Durandal stitches together Knockout, Sammy, and RequireJS to make it easier to use these technologies together. In this blog entry, I want to provide a brief walkthrough of using Durandal to create a simple Single Page App. I am going to demonstrate how you can create a simple Movies App which contains (virtual) pages for viewing a list of movies, adding new movies, and viewing movie details. The goal of this blog entry is to give you a sense of what it is like to build apps with Durandal. Installing Durandal First things first. How do you get Durandal? The GitHub project for Durandal is located here: https://github.com/BlueSpire/Durandal The Wiki — located at the GitHub project — contains all of the current documentation for Durandal. Currently, the documentation is a little sparse, but it is enough to get you started. Instead of downloading the Durandal source from GitHub, a better option for getting started with Durandal is to install one of the Durandal NuGet packages. I built the Movies App described in this blog entry by first creating a new ASP.NET MVC 4 Web Application with the Basic Template. Next, I executed the following command from the Package Manager Console: Install-Package Durandal.StarterKit As you can see from the screenshot of the Package Manager Console above, the Durandal Starter Kit package has several dependencies including: · jQuery · Knockout · Sammy · Twitter Bootstrap The Durandal Starter Kit package includes a sample Durandal application. You can get to the Starter Kit app by navigating to the Durandal controller. Unfortunately, when I first tried to run the Starter Kit app, I got an error because the Starter Kit is hard-coded to use a particular version of jQuery which is already out of date. You can fix this issue by modifying the App_Start\DurandalBundleConfig.cs file so it is jQuery version agnostic like this: bundles.Add( new ScriptBundle("~/scripts/vendor") .Include("~/Scripts/jquery-{version}.js") .Include("~/Scripts/knockout-{version}.js") .Include("~/Scripts/sammy-{version}.js") // .Include("~/Scripts/jquery-1.9.0.min.js") // .Include("~/Scripts/knockout-2.2.1.js") // .Include("~/Scripts/sammy-0.7.4.min.js") .Include("~/Scripts/bootstrap.min.js") ); The recommendation is that you create a Durandal app in a folder off your project root named App. The App folder in the Starter Kit contains the following subfolders and files: · durandal – This folder contains the actual durandal JavaScript library. · viewmodels – This folder contains all of your application’s view models. · views – This folder contains all of your application’s views. · main.js — This file contains all of the JavaScript startup code for your app including the client-side routing configuration. · main-built.js – This file contains an optimized version of your application. You need to build this file by using the RequireJS optimizer (unfortunately, before you can run the optimizer, you must first install NodeJS). For the purpose of this blog entry, I wanted to start from scratch when building the Movies app, so I deleted all of these files and folders except for the durandal folder which contains the durandal library. Creating the ASP.NET MVC Controller and View A Durandal app is built using a single server-side ASP.NET MVC controller and ASP.NET MVC view. A Durandal app is a Single Page App. When you navigate between pages, you are not navigating to new pages on the server. Instead, you are loading new virtual pages into the one-and-only-one server-side view. For the Movies app, I created the following ASP.NET MVC Home controller: public class HomeController : Controller { public ActionResult Index() { return View(); } } There is nothing special about the Home controller – it is as basic as it gets. Next, I created the following server-side ASP.NET view. This is the one-and-only server-side view used by the Movies app: @{ Layout = null; } <!DOCTYPE html> <html> <head> <title>Index</title> </head> <body> <div id="applicationHost"> Loading app.... </div> @Scripts.Render("~/scripts/vendor") <script type="text/javascript" src="~/App/durandal/amd/require.js" data-main="/App/main"></script> </body> </html> Notice that I set the Layout property for the view to the value null. If you neglect to do this, then the default ASP.NET MVC layout will be applied to the view and you will get the <!DOCTYPE> and opening and closing <html> tags twice. Next, notice that the view contains a DIV element with the Id applicationHost. This marks the area where virtual pages are loaded. When you navigate from page to page in a Durandal app, HTML page fragments are retrieved from the server and stuck in the applicationHost DIV element. Inside the applicationHost element, you can place any content which you want to display when a Durandal app is starting up. For example, you can create a fancy splash screen. I opted for simply displaying the text “Loading app…”: Next, notice the view above includes a call to the Scripts.Render() helper. This helper renders out all of the JavaScript files required by the Durandal library such as jQuery and Knockout. Remember to fix the App_Start\DurandalBundleConfig.cs as described above or Durandal will attempt to load an old version of jQuery and throw a JavaScript exception and stop working. Your application JavaScript code is not included in the scripts rendered by the Scripts.Render helper. Your application code is loaded dynamically by RequireJS with the help of the following SCRIPT element located at the bottom of the view: <script type="text/javascript" src="~/App/durandal/amd/require.js" data-main="/App/main"></script> The data-main attribute on the SCRIPT element causes RequireJS to load your /app/main.js JavaScript file to kick-off your Durandal app. Creating the Durandal Main.js File The Durandal Main.js JavaScript file, located in your App folder, contains all of the code required to configure the behavior of Durandal. Here’s what the Main.js file looks like in the case of the Movies app: require.config({ paths: { 'text': 'durandal/amd/text' } }); define(function (require) { var app = require('durandal/app'), viewLocator = require('durandal/viewLocator'), system = require('durandal/system'), router = require('durandal/plugins/router'); //>>excludeStart("build", true); system.debug(true); //>>excludeEnd("build"); app.start().then(function () { //Replace 'viewmodels' in the moduleId with 'views' to locate the view. //Look for partial views in a 'views' folder in the root. viewLocator.useConvention(); //configure routing router.useConvention(); router.mapNav("movies/show"); router.mapNav("movies/add"); router.mapNav("movies/details/:id"); app.adaptToDevice(); //Show the app by setting the root view model for our application with a transition. app.setRoot('viewmodels/shell', 'entrance'); }); }); There are three important things to notice about the main.js file above. First, notice that it contains a section which enables debugging which looks like this: //>>excludeStart(“build”, true); system.debug(true); //>>excludeEnd(“build”); This code enables debugging for your Durandal app which is very useful when things go wrong. When you call system.debug(true), Durandal writes out debugging information to your browser JavaScript console. For example, you can use the debugging information to diagnose issues with your client-side routes: (The funny looking //> symbols around the system.debug() call are RequireJS optimizer pragmas). The main.js file is also the place where you configure your client-side routes. In the case of the Movies app, the main.js file is used to configure routes for three page: the movies show, add, and details pages. //configure routing router.useConvention(); router.mapNav("movies/show"); router.mapNav("movies/add"); router.mapNav("movies/details/:id"); The route for movie details includes a route parameter named id. Later, we will use the id parameter to lookup and display the details for the right movie. Finally, the main.js file above contains the following line of code: //Show the app by setting the root view model for our application with a transition. app.setRoot('viewmodels/shell', 'entrance'); This line of code causes Durandal to load up a JavaScript file named shell.js and an HTML fragment named shell.html. I’ll discuss the shell in the next section. Creating the Durandal Shell You can think of the Durandal shell as the layout or master page for a Durandal app. The shell is where you put all of the content which you want to remain constant as a user navigates from virtual page to virtual page. For example, the shell is a great place to put your website logo and navigation links. The Durandal shell is composed from two parts: a JavaScript file and an HTML file. Here’s what the HTML file looks like for the Movies app: <h1>Movies App</h1> <div class="container-fluid page-host">  </div> And here is what the JavaScript file looks like: define(function (require) { var router = require('durandal/plugins/router'); return { router: router, activate: function () { return router.activate('movies/show'); } }; }); The JavaScript file contains the view model for the shell. This view model returns the Durandal router so you can access the list of configured routes from your shell. Notice that the JavaScript file includes a function named activate(). This function loads the movies/show page as the first page in the Movies app. If you want to create a different default Durandal page, then pass the name of a different age to the router.activate() method. Creating the Movies Show Page Durandal pages are created out of a view model and a view. The view model contains all of the data and view logic required for the view. The view contains all of the HTML markup for rendering the view model. Let’s start with the movies show page. The movies show page displays a list of movies. The view model for the show page looks like this: define(function (require) { var moviesRepository = require("repositories/moviesRepository"); return { movies: ko.observable(), activate: function() { this.movies(moviesRepository.listMovies()); } }; }); You create a view model by defining a new RequireJS module (see http://requirejs.org). You create a RequireJS module by placing all of your JavaScript code into an anonymous function passed to the RequireJS define() method. A RequireJS module has two parts. You retrieve all of the modules which your module requires at the top of your module. The code above depends on another RequireJS module named repositories/moviesRepository. Next, you return the implementation of your module. The code above returns a JavaScript object which contains a property named movies and a method named activate. The activate() method is a magic method which Durandal calls whenever it activates your view model. Your view model is activated whenever you navigate to a page which uses it. In the code above, the activate() method is used to get the list of movies from the movies repository and assign the list to the view model movies property. The HTML for the movies show page looks like this: <table> <thead> <tr> <th>Title</th><th>Director</th> </tr> </thead> <tbody data-bind="foreach:movies"> <tr> <td data-bind="text:title"></td> <td data-bind="text:director"></td> <td><a data-bind="attr:{href:'#/movies/details/'+id}">Details</a></td> </tr> </tbody> </table> <a href="#/movies/add">Add Movie</a> Notice that this is an HTML fragment. This fragment will be stuffed into the page-host DIV element in the shell.html file which is stuffed, in turn, into the applicationHost DIV element in the server-side MVC view. The HTML markup above contains data-bind attributes used by Knockout to display the list of movies (To learn more about Knockout, visit http://knockoutjs.com). The list of movies from the view model is displayed in an HTML table. Notice that the page includes a link to a page for adding a new movie. The link uses the following URL which starts with a hash: #/movies/add. Because the link starts with a hash, clicking the link does not cause a request back to the server. Instead, you navigate to the movies/add page virtually. Creating the Movies Add Page The movies add page also consists of a view model and view. The add page enables you to add a new movie to the movie database. Here’s the view model for the add page: define(function (require) { var app = require('durandal/app'); var router = require('durandal/plugins/router'); var moviesRepository = require("repositories/moviesRepository"); return { movieToAdd: { title: ko.observable(), director: ko.observable() }, activate: function () { this.movieToAdd.title(""); this.movieToAdd.director(""); this._movieAdded = false; }, canDeactivate: function () { if (this._movieAdded == false) { return app.showMessage('Are you sure you want to leave this page?', 'Navigate', ['Yes', 'No']); } else { return true; } }, addMovie: function () { // Add movie to db moviesRepository.addMovie(ko.toJS(this.movieToAdd)); // flag new movie this._movieAdded = true; // return to list of movies router.navigateTo("#/movies/show"); } }; }); The view model contains one property named movieToAdd which is bound to the add movie form. The view model also has the following three methods: 1. activate() – This method is called by Durandal when you navigate to the add movie page. The activate() method resets the add movie form by clearing out the movie title and director properties. 2. canDeactivate() – This method is called by Durandal when you attempt to navigate away from the add movie page. If you return false then navigation is cancelled. 3. addMovie() – This method executes when the add movie form is submitted. This code adds the new movie to the movie repository. I really like the Durandal canDeactivate() method. In the code above, I use the canDeactivate() method to show a warning to a user if they navigate away from the add movie page – either by clicking the Cancel button or by hitting the browser back button – before submitting the add movie form: The view for the add movie page looks like this: <form data-bind="submit:addMovie"> <fieldset> <legend>Add Movie</legend> <div> <label> Title: <input data-bind="value:movieToAdd.title" required /> </label> </div> <div> <label> Director: <input data-bind="value:movieToAdd.director" required /> </label> </div> <div> <input type="submit" value="Add" /> <a href="#/movies/show">Cancel</a> </div> </fieldset> </form> I am using Knockout to bind the movieToAdd property from the view model to the INPUT elements of the HTML form. Notice that the FORM element includes a data-bind attribute which invokes the addMovie() method from the view model when the HTML form is submitted. Creating the Movies Details Page You navigate to the movies details Page by clicking the Details link which appears next to each movie in the movies show page: The Details links pass the movie ids to the details page: #/movies/details/0 #/movies/details/1 #/movies/details/2 Here’s what the view model for the movies details page looks like: define(function (require) { var router = require('durandal/plugins/router'); var moviesRepository = require("repositories/moviesRepository"); return { movieToShow: { title: ko.observable(), director: ko.observable() }, activate: function (context) { // Grab movie from repository var movie = moviesRepository.getMovie(context.id); // Add to view model this.movieToShow.title(movie.title); this.movieToShow.director(movie.director); } }; }); Notice that the view model activate() method accepts a parameter named context. You can take advantage of the context parameter to retrieve route parameters such as the movie Id. In the code above, the context.id property is used to retrieve the correct movie from the movie repository and the movie is assigned to a property named movieToShow exposed by the view model. The movie details view displays the movieToShow property by taking advantage of Knockout bindings: <div> <h2 data-bind="text:movieToShow.title"></h2> directed by <span data-bind="text:movieToShow.director"></span> </div> Summary The goal of this blog entry was to walkthrough building a simple Single Page App using Durandal and to get a feel for what it is like to use this library. I really like how Durandal stitches together Knockout, Sammy, and RequireJS and establishes patterns for using these libraries to build Single Page Apps. Having a standard pattern which developers on a team can use to build new pages is super valuable. Once you get the hang of it, using Durandal to create new virtual pages is dead simple. Just define a new route, view model, and view and you are done. I also appreciate the fact that Durandal did not attempt to re-invent the wheel and that Durandal leverages existing JavaScript libraries such as Knockout, RequireJS, and Sammy. These existing libraries are powerful libraries and I have already invested a considerable amount of time in learning how to use them. Durandal makes it easier to use these libraries together without losing any of their power. Durandal has some additional interesting features which I have not had a chance to play with yet. For example, you can use the RequireJS optimizer to combine and minify all of a Durandal app’s code. Also, Durandal supports a way to create custom widgets (client-side controls) by composing widgets from a controller and view. You can download the code for the Movies app by clicking the following link (this is a Visual Studio 2012 project): Durandal Movie App

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Toorcon14

- by danx

Toorcon 2012 Information Security Conference San Diego, CA, http://www.toorcon.org/ Dan Anderson, October 2012 It's almost Halloween, and we all know what that means—yes, of course, it's time for another Toorcon Conference! Toorcon is an annual conference for people interested in computer security. This includes the whole range of hackers, computer hobbyists, professionals, security consultants, press, law enforcement, prosecutors, FBI, etc. We're at Toorcon 14—see earlier blogs for some of the previous Toorcon's I've attended (back to 2003). This year's "con" was held at the Westin on Broadway in downtown San Diego, California. The following are not necessarily my views—I'm just the messenger—although I could have misquoted or misparaphrased the speakers. Also, I only reviewed some of the talks, below, which I attended and interested me. MalAndroid—the Crux of Android Infections, Aditya K. Sood Programming Weird Machines with ELF Metadata, Rebecca "bx" Shapiro Privacy at the Handset: New FCC Rules?, Valkyrie Hacking Measured Boot and UEFI, Dan Griffin You Can't Buy Security: Building the Open Source InfoSec Program, Boris Sverdlik What Journalists Want: The Investigative Reporters' Perspective on Hacking, Dave Maas & Jason Leopold Accessibility and Security, Anna Shubina Stop Patching, for Stronger PCI Compliance, Adam Brand McAfee Secure & Trustmarks — a Hacker's Best Friend, Jay James & Shane MacDougall MalAndroid—the Crux of Android Infections Aditya K. Sood, IOActive, Michigan State PhD candidate Aditya talked about Android smartphone malware. There's a lot of old Android software out there—over 50% Gingerbread (2.3.x)—and most have unpatched vulnerabilities. Of 9 Android vulnerabilities, 8 have known exploits (such as the old Gingerbread Global Object Table exploit). Android protection includes sandboxing, security scanner, app permissions, and screened Android app market. The Android permission checker has fine-grain resource control, policy enforcement. Android static analysis also includes a static analysis app checker (bouncer), and a vulnerablity checker. What security problems does Android have? User-centric security, which depends on the user to grant permission and make smart decisions. But users don't care or think about malware (the're not aware, not paranoid). All they want is functionality, extensibility, mobility Android had no "proper" encryption before Android 3.0 No built-in protection against social engineering and web tricks Alternative Android app markets are unsafe. Simply visiting some markets can infect Android Aditya classified Android Malware types as: Type A—Apps. These interact with the Android app framework. For example, a fake Netflix app. Or Android Gold Dream (game), which uploads user files stealthy manner to a remote location. Type K—Kernel. Exploits underlying Linux libraries or kernel Type H—Hybrid. These use multiple layers (app framework, libraries, kernel). These are most commonly used by Android botnets, which are popular with Chinese botnet authors What are the threats from Android malware? These incude leak info (contacts), banking fraud, corporate network attacks, malware advertising, malware "Hackivism" (the promotion of social causes. For example, promiting specific leaders of the Tunisian or Iranian revolutions. Android malware is frequently "masquerated". That is, repackaged inside a legit app with malware. To avoid detection, the hidden malware is not unwrapped until runtime. The malware payload can be hidden in, for example, PNG files. Less common are Android bootkits—there's not many around. What they do is hijack the Android init framework—alteering system programs and daemons, then deletes itself. For example, the DKF Bootkit (China). Android App Problems: no code signing! all self-signed native code execution permission sandbox — all or none alternate market places no robust Android malware detection at network level delayed patch process Programming Weird Machines with ELF Metadata Rebecca "bx" Shapiro, Dartmouth College, NH https://github.com/bx/elf-bf-tools @bxsays on twitter Definitions. "ELF" is an executable file format used in linking and loading executables (on UNIX/Linux-class machines). "Weird machine" uses undocumented computation sources (I think of them as unintended virtual machines). Some examples of "weird machines" are those that: return to weird location, does SQL injection, corrupts the heap. Bx then talked about using ELF metadata as (an uintended) "weird machine". Some ELF background: A compiler takes source code and generates a ELF object file (hello.o). A static linker makes an ELF executable from the object file. A runtime linker and loader takes ELF executable and loads and relocates it in memory. The ELF file has symbols to relocate functions and variables. ELF has two relocation tables—one at link time and another one at loading time: .rela.dyn (link time) and .dynsym (dynamic table). GOT: Global Offset Table of addresses for dynamically-linked functions. PLT: Procedure Linkage Tables—works with GOT. The memory layout of a process (not the ELF file) is, in order: program (+ heap), dynamic libraries, libc, ld.so, stack (which includes the dynamic table loaded into memory) For ELF, the "weird machine" is found and exploited in the loader. ELF can be crafted for executing viruses, by tricking runtime into executing interpreted "code" in the ELF symbol table. One can inject parasitic "code" without modifying the actual ELF code portions. Think of the ELF symbol table as an "assembly language" interpreter. It has these elements: instructions: Add, move, jump if not 0 (jnz) Think of symbol table entries as "registers" symbol table value is "contents" immediate values are constants direct values are addresses (e.g., 0xdeadbeef) move instruction: is a relocation table entry add instruction: relocation table "addend" entry jnz instruction: takes multiple relocation table entries The ELF weird machine exploits the loader by relocating relocation table entries. The loader will go on forever until told to stop. It stores state on stack at "end" and uses IFUNC table entries (containing function pointer address). The ELF weird machine, called "Brainfu*k" (BF) has: 8 instructions: pointer inc, dec, inc indirect, dec indirect, jump forward, jump backward, print. Three registers - 3 registers Bx showed example BF source code that implemented a Turing machine printing "hello, world". More interesting was the next demo, where bx modified ping. Ping runs suid as root, but quickly drops privilege. BF modified the loader to disable the library function call dropping privilege, so it remained as root. Then BF modified the ping -t argument to execute the -t filename as root. It's best to show what this modified ping does with an example: $ whoami bx $ ping localhost -t backdoor.sh # executes backdoor $ whoami root $ The modified code increased from 285948 bytes to 290209 bytes. A BF tool compiles "executable" by modifying the symbol table in an existing ELF executable. The tool modifies .dynsym and .rela.dyn table, but not code or data. Privacy at the Handset: New FCC Rules? "Valkyrie" (Christie Dudley, Santa Clara Law JD candidate) Valkyrie talked about mobile handset privacy. Some background: Senator Franken (also a comedian) became alarmed about CarrierIQ, where the carriers track their customers. Franken asked the FCC to find out what obligations carriers think they have to protect privacy. The carriers' response was that they are doing just fine with self-regulation—no worries! Carriers need to collect data, such as missed calls, to maintain network quality. But carriers also sell data for marketing. Verizon sells customer data and enables this with a narrow privacy policy (only 1 month to opt out, with difficulties). The data sold is not individually identifiable and is aggregated. But Verizon recommends, as an aggregation workaround to "recollate" data to other databases to identify customers indirectly. The FCC has regulated telephone privacy since 1934 and mobile network privacy since 2007. Also, the carriers say mobile phone privacy is a FTC responsibility (not FCC). FTC is trying to improve mobile app privacy, but FTC has no authority over carrier / customer relationships. As a side note, Apple iPhones are unique as carriers have extra control over iPhones they don't have with other smartphones. As a result iPhones may be more regulated. Who are the consumer advocates? Everyone knows EFF, but EPIC (Electrnic Privacy Info Center), although more obsecure, is more relevant. What to do? Carriers must be accountable. Opt-in and opt-out at any time. Carriers need incentive to grant users control for those who want it, by holding them liable and responsible for breeches on their clock. Location information should be added current CPNI privacy protection, and require "Pen/trap" judicial order to obtain (and would still be a lower standard than 4th Amendment). Politics are on a pro-privacy swing now, with many senators and the Whitehouse. There will probably be new regulation soon, and enforcement will be a problem, but consumers will still have some benefit. Hacking Measured Boot and UEFI Dan Griffin, JWSecure, Inc., Seattle, @JWSdan Dan talked about hacking measured UEFI boot. First some terms: UEFI is a boot technology that is replacing BIOS (has whitelisting and blacklisting). UEFI protects devices against rootkits. TPM - hardware security device to store hashs and hardware-protected keys "secure boot" can control at firmware level what boot images can boot "measured boot" OS feature that tracks hashes (from BIOS, boot loader, krnel, early drivers). "remote attestation" allows remote validation and control based on policy on a remote attestation server. Microsoft pushing TPM (Windows 8 required), but Google is not. Intel TianoCore is the only open source for UEFI. Dan has Measured Boot Tool at http://mbt.codeplex.com/ with a demo where you can also view TPM data. TPM support already on enterprise-class machines. UEFI Weaknesses. UEFI toolkits are evolving rapidly, but UEFI has weaknesses: assume user is an ally trust TPM implicitly, and attached to computer hibernate file is unprotected (disk encryption protects against this) protection migrating from hardware to firmware delays in patching and whitelist updates will UEFI really be adopted by the mainstream (smartphone hardware support, bank support, apathetic consumer support) You Can't Buy Security: Building the Open Source InfoSec Program Boris Sverdlik, ISDPodcast.com co-host Boris talked about problems typical with current security audits. "IT Security" is an oxymoron—IT exists to enable buiness, uptime, utilization, reporting, but don't care about security—IT has conflict of interest. There's no Magic Bullet ("blinky box"), no one-size-fits-all solution (e.g., Intrusion Detection Systems (IDSs)). Regulations don't make you secure. The cloud is not secure (because of shared data and admin access). Defense and pen testing is not sexy. Auditors are not solution (security not a checklist)—what's needed is experience and adaptability—need soft skills. Step 1: First thing is to Google and learn the company end-to-end before you start. Get to know the management team (not IT team), meet as many people as you can. Don't use arbitrary values such as CISSP scores. Quantitive risk assessment is a myth (e.g. AV*EF-SLE). Learn different Business Units, legal/regulatory obligations, learn the business and where the money is made, verify company is protected from script kiddies (easy), learn sensitive information (IP, internal use only), and start with low-hanging fruit (customer service reps and social engineering). Step 2: Policies. Keep policies short and relevant. Generic SANS "security" boilerplate policies don't make sense and are not followed. Focus on acceptable use, data usage, communications, physical security. Step 3: Implementation: keep it simple stupid. Open source, although useful, is not free (implementation cost). Access controls with authentication & authorization for local and remote access. MS Windows has it, otherwise use OpenLDAP, OpenIAM, etc. Application security Everyone tries to reinvent the wheel—use existing static analysis tools. Review high-risk apps and major revisions. Don't run different risk level apps on same system. Assume host/client compromised and use app-level security control. Network security VLAN != segregated because there's too many workarounds. Use explicit firwall rules, active and passive network monitoring (snort is free), disallow end user access to production environment, have a proxy instead of direct Internet access. Also, SSL certificates are not good two-factor auth and SSL does not mean "safe." Operational Controls Have change, patch, asset, & vulnerability management (OSSI is free). For change management, always review code before pushing to production For logging, have centralized security logging for business-critical systems, separate security logging from administrative/IT logging, and lock down log (as it has everything). Monitor with OSSIM (open source). Use intrusion detection, but not just to fulfill a checkbox: build rules from a whitelist perspective (snort). OSSEC has 95% of what you need. Vulnerability management is a QA function when done right: OpenVas and Seccubus are free. Security awareness The reality is users will always click everything. Build real awareness, not compliance driven checkbox, and have it integrated into the culture. Pen test by crowd sourcing—test with logging COSSP http://www.cossp.org/ - Comprehensive Open Source Security Project What Journalists Want: The Investigative Reporters' Perspective on Hacking Dave Maas, San Diego CityBeat Jason Leopold, Truthout.org The difference between hackers and investigative journalists: For hackers, the motivation varies, but method is same, technological specialties. For investigative journalists, it's about one thing—The Story, and they need broad info-gathering skills. J-School in 60 Seconds: Generic formula: Person or issue of pubic interest, new info, or angle. Generic criteria: proximity, prominence, timeliness, human interest, oddity, or consequence. Media awareness of hackers and trends: journalists becoming extremely aware of hackers with congressional debates (privacy, data breaches), demand for data-mining Journalists, use of coding and web development for Journalists, and Journalists busted for hacking (Murdock). Info gathering by investigative journalists include Public records laws. Federal Freedom of Information Act (FOIA) is good, but slow. California Public Records Act is a lot stronger. FOIA takes forever because of foot-dragging—it helps to be specific. Often need to sue (especially FBI). CPRA is faster, and requests can be vague. Dumps and leaks (a la Wikileaks) Journalists want: leads, protecting ourselves, our sources, and adapting tools for news gathering (Google hacking). Anonomity is important to whistleblowers. They want no digital footprint left behind (e.g., email, web log). They don't trust encryption, want to feel safe and secure. Whistleblower laws are very weak—there's no upside for whistleblowers—they have to be very passionate to do it. Accessibility and Security or: How I Learned to Stop Worrying and Love the Halting Problem Anna Shubina, Dartmouth College Anna talked about how accessibility and security are related. Accessibility of digital content (not real world accessibility). mostly refers to blind users and screenreaders, for our purpose. Accessibility is about parsing documents, as are many security issues. "Rich" executable content causes accessibility to fail, and often causes security to fail. For example MS Word has executable format—it's not a document exchange format—more dangerous than PDF or HTML. Accessibility is often the first and maybe only sanity check with parsing. They have no choice because someone may want to read what you write. Google, for example, is very particular about web browser you use and are bad at supporting other browsers. Uses JavaScript instead of links, often requiring mouseover to display content. PDF is a security nightmare. Executible format, embedded flash, JavaScript, etc. 15 million lines of code. Google Chrome doesn't handle PDF correctly, causing several security bugs. PDF has an accessibility checker and PDF tagging, to help with accessibility. But no PDF checker checks for incorrect tags, untagged content, or validates lists or tables. None check executable content at all. The "Halting Problem" is: can one decide whether a program will ever stop? The answer, in general, is no (Rice's theorem). The same holds true for accessibility checkers. Language-theoretic Security says complicated data formats are hard to parse and cannot be solved due to the Halting Problem. W3C Web Accessibility Guidelines: "Perceivable, Operable, Understandable, Robust" Not much help though, except for "Robust", but here's some gems: * all information should be parsable (paraphrasing) * if not parsable, cannot be converted to alternate formats * maximize compatibility in new document formats Executible webpages are bad for security and accessibility. They say it's for a better web experience. But is it necessary to stuff web pages with JavaScript for a better experience? A good example is The Drudge Report—it has hand-written HTML with no JavaScript, yet drives a lot of web traffic due to good content. A bad example is Google News—hidden scrollbars, guessing user input. Solutions: Accessibility and security problems come from same source Expose "better user experience" myth Keep your corner of Internet parsable Remember "Halting Problem"—recognize false solutions (checking and verifying tools) Stop Patching, for Stronger PCI Compliance Adam Brand, protiviti @adamrbrand, http://www.picfun.com/ Adam talked about PCI compliance for retail sales. Take an example: for PCI compliance, 50% of Brian's time (a IT guy), 960 hours/year was spent patching POSs in 850 restaurants. Often applying some patches make no sense (like fixing a browser vulnerability on a server). "Scanner worship" is overuse of vulnerability scanners—it gives a warm and fuzzy and it's simple (red or green results—fix reds). Scanners give a false sense of security. In reality, breeches from missing patches are uncommon—more common problems are: default passwords, cleartext authentication, misconfiguration (firewall ports open). Patching Myths: Myth 1: install within 30 days of patch release (but PCI Â§6.1 allows a "risk-based approach" instead). Myth 2: vendor decides what's critical (also PCI Â§6.1). But Â§6.2 requires user ranking of vulnerabilities instead. Myth 3: scan and rescan until it passes. But PCI Â§11.2.1b says this applies only to high-risk vulnerabilities. Adam says good recommendations come from NIST 800-40. Instead use sane patching and focus on what's really important. From NIST 800-40: Proactive: Use a proactive vulnerability management process: use change control, configuration management, monitor file integrity. Monitor: start with NVD and other vulnerability alerts, not scanner results. Evaluate: public-facing system? workstation? internal server? (risk rank) Decide:on action and timeline Test: pre-test patches (stability, functionality, rollback) for change control Install: notify, change control, tickets McAfee Secure & Trustmarks — a Hacker's Best Friend Jay James, Shane MacDougall, Tactical Intelligence Inc., Canada "McAfee Secure Trustmark" is a website seal marketed by McAfee. A website gets this badge if they pass their remote scanning. The problem is a removal of trustmarks act as flags that you're vulnerable. Easy to view status change by viewing McAfee list on website or on Google. "Secure TrustGuard" is similar to McAfee. Jay and Shane wrote Perl scripts to gather sites from McAfee and search engines. If their certification image changes to a 1x1 pixel image, then they are longer certified. Their scripts take deltas of scans to see what changed daily. The bottom line is change in TrustGuard status is a flag for hackers to attack your site. Entire idea of seals is silly—you're raising a flag saying if you're vulnerable.

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Using BPEL Performance Statistics to Diagnose Performance Bottlenecks

- by fip

Tuning performance of Oracle SOA 11G applications could be challenging. Because SOA is a platform for you to build composite applications that connect many applications and "services", when the overall performance is slow, the bottlenecks could be anywhere in the system: the applications/services that SOA connects to, the infrastructure database, or the SOA server itself.How to quickly identify the bottleneck becomes crucial in tuning the overall performance. Fortunately, the BPEL engine in Oracle SOA 11G (and 10G, for that matter) collects BPEL Engine Performance Statistics, which show the latencies of low level BPEL engine activities. The BPEL engine performance statistics can make it a bit easier for you to identify the performance bottleneck. Although the BPEL engine performance statistics are always available, the access to and interpretation of them are somewhat obscure in the early and current (PS5) 11G versions. This blog attempts to offer instructions that help you to enable, retrieve and interpret the performance statistics, before the future versions provides a more pleasant user experience. Overview of BPEL Engine Performance Statistics SOA BPEL has a feature of collecting some performance statistics and store them in memory. One MBean attribute, StatLastN, configures the size of the memory buffer to store the statistics. This memory buffer is a "moving window", in a way that old statistics will be flushed out by the new if the amount of data exceeds the buffer size. Since the buffer size is limited by StatLastN, impacts of statistics collection on performance is minimal. By default StatLastN=-1, which means no collection of performance data. Once the statistics are collected in the memory buffer, they can be retrieved via another MBean oracle.as.soainfra.bpel:Location=[Server Name],name=BPELEngine,type=BPELEngine.> My friend in Oracle SOA development wrote this simple 'bpelstat' web app that looks up and retrieves the performance data from the MBean and displays it in a human readable form. It does not have beautiful UI but it is fairly useful. Although in Oracle SOA 11.1.1.5 onwards the same statistics can be viewed via a more elegant UI under "request break down" at EM -> SOA Infrastructure -> Service Engines -> BPEL -> Statistics, some unsophisticated minds like mine may still prefer the simplicity of the 'bpelstat' JSP. One thing that simple JSP does do well is that you can save the page and send it to someone to further analyze Follows are the instructions of how to install and invoke the BPEL statistic JSP. My friend in SOA Development will soon blog about interpreting the statistics. Stay tuned. Step1: Enable BPEL Engine Statistics for Each SOA Servers via Enterprise Manager First st you need to set the StatLastN to some number as a way to enable the collection of BPEL Engine Performance Statistics EM Console -> soa-infra(Server Name) -> SOA Infrastructure -> SOA Administration -> BPEL Properties Click on "More BPEL Configuration Properties" Click on attribute "StatLastN", set its value to some integer number. Typically you want to set it 1000 or more. Step 2: Download and Deploy bpelstat.war File to Admin Server, Note: the WAR file contains a JSP that does NOT have any security restriction. You do NOT want to keep in your production server for a long time as it is a security hazard. Deactivate the war once you are done. Download the bpelstat.war to your local PC At WebLogic Console, Go to Deployments -> Install Click on the "upload your file(s)" Click the "Browse" button to upload the deployment to Admin Server Accept the uploaded file as the path, click next Check the default option "Install this deployment as an application" Check "AdminServer" as the target server Finish the rest of the deployment with default settings Console -> Deployments Check the box next to "bpelstat" application Click on the "Start" button. It will change the state of the app from "prepared" to "active" Step 3: Invoke the BPEL Statistic Tool The BPELStat tool merely call the MBean of BPEL server and collects and display the in-memory performance statics. You usually want to do that after some peak loads. Go to http://<admin-server-host>:<admin-server-port>/bpelstat Enter the correct admin hostname, port, username and password Enter the SOA Server Name from which you want to collect the performance statistics. For example, SOA_MS1, etc. Click Submit Keep doing the same for all SOA servers. Step 3: Interpret the BPEL Engine Statistics You will see a few categories of BPEL Statistics from the JSP Page. First it starts with the overall latency of BPEL processes, grouped by synchronous and asynchronous processes. Then it provides the further break down of the measurements through the life time of a BPEL request, which is called the "request break down". 1. Overall latency of BPEL processes The top of the page shows that the elapse time of executing the synchronous process TestSyncBPELProcess from the composite TestComposite averages at about 1543.21ms, while the elapse time of executing the asynchronous process TestAsyncBPELProcess from the composite TestComposite2 averages at about 1765.43ms. The maximum and minimum latency were also shown. Synchronous process statistics <statistics> <stats key="default/TestComposite!2.0.2-ScopedJMSOSB*soa_bfba2527-a9ba-41a7-95c5-87e49c32f4ff/TestSyncBPELProcess" min="1234" max="4567" average="1543.21" count="1000"> </stats> </statistics> Asynchronous process statistics <statistics> <stats key="default/TestComposite2!2.0.2-ScopedJMSOSB*soa_bfba2527-a9ba-41a7-95c5-87e49c32f4ff/TestAsyncBPELProcess" min="2234" max="3234" average="1765.43" count="1000"> </stats> </statistics> 2. Request break down Under the overall latency categorized by synchronous and asynchronous processes is the "Request breakdown". Organized by statistic keys, the Request breakdown gives finer grain performance statistics through the life time of the BPEL requests.It uses indention to show the hierarchy of the statistics. Request breakdown <statistics> <stats key="eng-composite-request" min="0" max="0" average="0.0" count="0"> <stats key="eng-single-request" min="22" max="606" average="258.43" count="277"> <stats key="populate-context" min="0" max="0" average="0.0" count="248"> Please note that in SOA 11.1.1.6, the statistics under Request breakdown is aggregated together cross all the BPEL processes based on statistic keys. It does not differentiate between BPEL processes. If two BPEL processes happen to have the statistic that share same statistic key, the statistics from two BPEL processes will be aggregated together. Keep this in mind when we go through more details below. 2.1 BPEL process activity latencies A very useful measurement in the Request Breakdown is the performance statistics of the BPEL activities you put in your BPEL processes: Assign, Invoke, Receive, etc. The names of the measurement in the JSP page directly come from the names to assign to each BPEL activity. These measurements are under the statistic key "actual-perform" Example 1: Follows is the measurement for BPEL activity "AssignInvokeCreditProvider_Input", which looks like the Assign activity in a BPEL process that assign an input variable before passing it to the invocation: <stats key="AssignInvokeCreditProvider_Input" min="1" max="8" average="1.9" count="153"> <stats key="sensor-send-activity-data" min="0" max="1" average="0.0" count="306"> </stats> <stats key="sensor-send-variable-data" min="0" max="0" average="0.0" count="153"> </stats> <stats key="monitor-send-activity-data" min="0" max="0" average="0.0" count="306"> </stats> </stats> Note: because as previously mentioned that the statistics cross all BPEL processes are aggregated together based on statistic keys, if two BPEL processes happen to name their Invoke activity the same name, they will show up at one measurement (i.e. statistic key). Example 2: Follows is the measurement of BPEL activity called "InvokeCreditProvider". You can not only see that by average it takes 3.31ms to finish this call (pretty fast) but also you can see from the further break down that most of this 3.31 ms was spent on the "invoke-service". <stats key="InvokeCreditProvider" min="1" max="13" average="3.31" count="153"> <stats key="initiate-correlation-set-again" min="0" max="0" average="0.0" count="153"> </stats> <stats key="invoke-service" min="1" max="13" average="3.08" count="153"> <stats key="prep-call" min="0" max="1" average="0.04" count="153"> </stats> </stats> <stats key="initiate-correlation-set" min="0" max="0" average="0.0" count="153"> </stats> <stats key="sensor-send-activity-data" min="0" max="0" average="0.0" count="306"> </stats> <stats key="sensor-send-variable-data" min="0" max="0" average="0.0" count="153"> </stats> <stats key="monitor-send-activity-data" min="0" max="0" average="0.0" count="306"> </stats> <stats key="update-audit-trail" min="0" max="2" average="0.03" count="153"> </stats> </stats> 2.2 BPEL engine activity latency Another type of measurements under Request breakdown are the latencies of underlying system level engine activities. These activities are not directly tied to a particular BPEL process or process activity, but they are critical factors in the overall engine performance. These activities include the latency of saving asynchronous requests to database, and latency of process dehydration. My friend Malkit Bhasin is working on providing more information on interpreting the statistics on engine activities on his blog (https://blogs.oracle.com/malkit/). I will update this blog once the information becomes available. Update on 2012-10-02: My friend Malkit Bhasin has published the detail interpretation of the BPEL service engine statistics at his blog http://malkit.blogspot.com/2012/09/oracle-bpel-engine-soa-suite.html.

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The broken Promise of the Mobile Web

- by Rick Strahl

High end mobile devices have been with us now for almost 7 years and they have utterly transformed the way we access information. Mobile phones and smartphones that have access to the Internet and host smart applications are in the hands of a large percentage of the population of the world. In many places even very remote, cell phones and even smart phones are a common sight. I’ll never forget when I was in India in 2011 I was up in the Southern Indian mountains riding an elephant out of a tiny local village, with an elephant herder in front riding atop of the elephant in front of us. He was dressed in traditional garb with the loin wrap and head cloth/turban as did quite a few of the locals in this small out of the way and not so touristy village. So we’re slowly trundling along in the forest and he’s lazily using his stick to guide the elephant and… 10 minutes in he pulls out his cell phone from his sash and starts texting. In the middle of texting a huge pig jumps out from the side of the trail and he takes a picture running across our path in the jungle! So yeah, mobile technology is very pervasive and it’s reached into even very buried and unexpected parts of this world. Apps are still King Apps currently rule the roost when it comes to mobile devices and the applications that run on them. If there’s something that you need on your mobile device your first step usually is to look for an app, not use your browser. But native app development remains a pain in the butt, with the requirement to have to support 2 or 3 completely separate platforms. There are solutions that try to bridge that gap. Xamarin is on a tear at the moment, providing their cross-device toolkit to build applications using C#. While Xamarin tools are impressive – and also *very* expensive – they only address part of the development madness that is app development. There are still specific device integration isssues, dealing with the different developer programs, security and certificate setups and all that other noise that surrounds app development. There’s also PhoneGap/Cordova which provides a hybrid solution that involves creating local HTML/CSS/JavaScript based applications, and then packaging them to run in a specialized App container that can run on most mobile device platforms using a WebView interface. This allows for using of HTML technology, but it also still requires all the set up, configuration of APIs, security keys and certification and submission and deployment process just like native applications – you actually lose many of the benefits that Web based apps bring. The big selling point of Cordova is that you get to use HTML have the ability to build your UI once for all platforms and run across all of them – but the rest of the app process remains in place. Apps can be a big pain to create and manage especially when we are talking about specialized or vertical business applications that aren’t geared at the mainstream market and that don’t fit the ‘store’ model. If you’re building a small intra department application you don’t want to deal with multiple device platforms and certification etc. for various public or corporate app stores. That model is simply not a good fit both from the development and deployment perspective. Even for commercial, big ticket apps, HTML as a UI platform offers many advantages over native, from write-once run-anywhere, to remote maintenance, single point of management and failure to having full control over the application as opposed to have the app store overloads censor you. In a lot of ways Web based HTML/CSS/JavaScript applications have so much potential for building better solutions based on existing Web technologies for the very same reasons a lot of content years ago moved off the desktop to the Web. To me the Web as a mobile platform makes perfect sense, but the reality of today’s Mobile Web unfortunately looks a little different… Where’s the Love for the Mobile Web? Yet here we are in the middle of 2014, nearly 7 years after the first iPhone was released and brought the promise of rich interactive information at your fingertips, and yet we still don’t really have a solid mobile Web platform. I know what you’re thinking: “But we have lots of HTML/JavaScript/CSS features that allows us to build nice mobile interfaces”. I agree to a point – it’s actually quite possible to build nice looking, rich and capable Web UI today. We have media queries to deal with varied display sizes, CSS transforms for smooth animations and transitions, tons of CSS improvements in CSS 3 that facilitate rich layout, a host of APIs geared towards mobile device features and lately even a number of JavaScript framework choices that facilitate development of multi-screen apps in a consistent manner. Personally I’ve been working a lot with AngularJs and heavily modified Bootstrap themes to build mobile first UIs and that’s been working very well to provide highly usable and attractive UI for typical mobile business applications. From the pure UI perspective things actually look very good. Not just about the UI But it’s not just about the UI - it’s also about integration with the mobile device. When it comes to putting all those pieces together into what amounts to a consolidated platform to build mobile Web applications, I think we still have a ways to go… there are a lot of missing pieces to make it all work together and integrate with the device more smoothly, and more importantly to make it work uniformly across the majority of devices. I think there are a number of reasons for this. Slow Standards Adoption HTML standards implementations and ratification has been dreadfully slow, and browser vendors all seem to pick and choose different pieces of the technology they implement. The end result is that we have a capable UI platform that’s missing some of the infrastructure pieces to make it whole on mobile devices. There’s lots of potential but what is lacking that final 10% to build truly compelling mobile applications that can compete favorably with native applications. Some of it is the fragmentation of browsers and the slow evolution of the mobile specific HTML APIs. A host of mobile standards exist but many of the standards are in the early review stage and they have been there stuck for long periods of time and seem to move at a glacial pace. Browser vendors seem even slower to implement them, and for good reason – non-ratified standards mean that implementations may change and vendor implementations tend to be experimental and likely have to be changed later. Neither Vendors or developers are not keen on changing standards. This is the typical chicken and egg scenario, but without some forward momentum from some party we end up stuck in the mud. It seems that either the standards bodies or the vendors need to carry the torch forward and that doesn’t seem to be happening quickly enough. Mobile Device Integration just isn’t good enough Current standards are not far reaching enough to address a number of the use case scenarios necessary for many mobile applications. While not every application needs to have access to all mobile device features, almost every mobile application could benefit from some integration with other parts of the mobile device platform. Integration with GPS, phone, media, messaging, notifications, linking and contacts system are benefits that are unique to mobile applications and could be widely used, but are mostly (with the exception of GPS) inaccessible for Web based applications today. Unfortunately trying to do most of this today only with a mobile Web browser is a losing battle. Aside from PhoneGap/Cordova’s app centric model with its own custom API accessing mobile device features and the token exception of the GeoLocation API, most device integration features are not widely supported by the current crop of mobile browsers. For example there’s no usable messaging API that allows access to SMS or contacts from HTML. Even obvious components like the Media Capture API are only implemented partially by mobile devices. There are alternatives and workarounds for some of these interfaces by using browser specific code, but that’s might ugly and something that I thought we were trying to leave behind with newer browser standards. But it’s not quite working out that way. It’s utterly perplexing to me that mobile standards like Media Capture and Streams, Media Gallery Access, Responsive Images, Messaging API, Contacts Manager API have only minimal or no traction at all today. Keep in mind we’ve had mobile browsers for nearly 7 years now, and yet we still have to think about how to get access to an image from the image gallery or the camera on some devices? Heck Windows Phone IE Mobile just gained the ability to upload images recently in the Windows 8.1 Update – that’s feature that HTML has had for 20 years! These are simple concepts and common problems that should have been solved a long time ago. It’s extremely frustrating to see build 90% of a mobile Web app with relative ease and then hit a brick wall for the remaining 10%, which often can be show stoppers. The remaining 10% have to do with platform integration, browser differences and working around the limitations that browsers and ‘pinned’ applications impose on HTML applications. The maddening part is that these limitations seem arbitrary as they could easily work on all mobile platforms. For example, SMS has a URL Moniker interface that sort of works on Android, works badly with iOS (only works if the address is already in the contact list) and not at all on Windows Phone. There’s no reason this shouldn’t work universally using the same interface – after all all phones have supported SMS since before the year 2000! But, it doesn’t have to be this way Change can happen very quickly. Take the GeoLocation API for example. Geolocation has taken off at the very beginning of the mobile device era and today it works well, provides the necessary security (a big concern for many mobile APIs), and is supported by just about all major mobile and even desktop browsers today. It handles security concerns via prompts to avoid unwanted access which is a model that would work for most other device APIs in a similar fashion. One time approval and occasional re-approval if code changes or caches expire. Simple and only slightly intrusive. It all works well, even though GeoLocation actually has some physical limitations, such as representing the current location when no GPS device is present. Yet this is a solved problem, where other APIs that are conceptually much simpler to implement have failed to gain any traction at all. Technically none of these APIs should be a problem to implement, but it appears that the momentum is just not there. Inadequate Web Application Linking and Activation Another important piece of the puzzle missing is the integration of HTML based Web applications. Today HTML based applications are not first class citizens on mobile operating systems. When talking about HTML based content there’s a big difference between content and applications. Content is great for search engine discovery and plain browser usage. Content is usually accessed intermittently and permanent linking is not so critical for this type of content. But applications have different needs. Applications need to be started up quickly and must be easily switchable to support a multi-tasking user workflow. Therefore, it’s pretty crucial that mobile Web apps are integrated into the underlying mobile OS and work with the standard task management features. Unfortunately this integration is not as smooth as it should be. It starts with actually trying to find mobile Web applications, to ‘installing’ them onto a phone in an easily accessible manner in a prominent position. The experience of discovering a Mobile Web ‘App’ and making it sticky is by no means as easy or satisfying. Today the way you’d go about this is: Open the browser Search for a Web Site in the browser with your search engine of choice Hope that you find the right site Hope that you actually find a site that works for your mobile device Click on the link and run the app in a fully chrome’d browser instance (read tiny surface area) Pin the app to the home screen (with all the limitations outline above) Hope you pointed at the right URL when you pinned Even for you and me as developers, there are a few steps in there that are painful and annoying, but think about the average user. First figuring out how to search for a specific site or URL? And then pinning the app and hopefully from the right location? You’ve probably lost more than half of your audience at that point. This experience sucks. For developers too this process is painful since app developers can’t control the shortcut creation directly. This problem often gets solved by crazy coding schemes, with annoying pop-ups that try to get people to create shortcuts via fancy animations that are both annoying and add overhead to each and every application that implements this sort of thing differently. And that’s not the end of it - getting the link onto the home screen with an application icon varies quite a bit between browsers. Apple’s non-standard meta tags are prominent and they work with iOS and Android (only more recent versions), but not on Windows Phone. Windows Phone instead requires you to create an actual screen or rather a partial screen be captured for a shortcut in the tile manager. Who had that brilliant idea I wonder? Surprisingly Chrome on recent Android versions seems to actually get it right – icons use pngs, pinning is easy and pinned applications properly behave like standalone apps and retain the browser’s active page state and content. Each of the platforms has a different way to specify icons (WP doesn’t allow you to use an icon image at all), and the most widely used interface in use today is a bunch of Apple specific meta tags that other browsers choose to support. The question is: Why is there no standard implementation for installing shortcuts across mobile platforms using an official format rather than a proprietary one? Then there’s iOS and the crazy way it treats home screen linked URLs using a crazy hybrid format that is neither as capable as a Web app running in Safari nor a WebView hosted application. Moving off the Web ‘app’ link when switching to another app actually causes the browser and preview it to ‘blank out’ the Web application in the Task View (see screenshot on the right). Then, when the ‘app’ is reactivated it ends up completely restarting the browser with the original link. This is crazy behavior that you can’t easily work around. In some situations you might be able to store the application state and restore it using LocalStorage, but for many scenarios that involve complex data sources (like say Google Maps) that’s not a possibility. The only reason for this screwed up behavior I can think of is that it is deliberate to make Web apps a pain in the butt to use and forcing users trough the App Store/PhoneGap/Cordova route. App linking and management is a very basic problem – something that we essentially have solved in every desktop browser – yet on mobile devices where it arguably matters a lot more to have easy access to web content we have to jump through hoops to have even a remotely decent linking/activation experience across browsers. Where’s the Money? It’s not surprising that device home screen integration and Mobile Web support in general is in such dismal shape – the mobile OS vendors benefit financially from App store sales and have little to gain from Web based applications that bypass the App store and the cash cow that it presents. On top of that, platform specific vendor lock-in of both end users and developers who have invested in hardware, apps and consumables is something that mobile platform vendors actually aspire to. Web based interfaces that are cross-platform are the anti-thesis of that and so again it’s no surprise that the mobile Web is on a struggling path. But – that may be changing. More and more we’re seeing operations shifting to services that are subscription based or otherwise collect money for usage, and that may drive more progress into the Web direction in the end . Nothing like the almighty dollar to drive innovation forward. Do we need a Mobile Web App Store? As much as I dislike moderated experiences in today’s massive App Stores, they do at least provide one single place to look for apps for your device. I think we could really use some sort of registry, that could provide something akin to an app store for mobile Web apps, to make it easier to actually find mobile applications. This could take the form of a specialized search engine, or maybe a more formal store/registry like structure. Something like apt-get/chocolatey for Web apps. It could be curated and provide at least some feedback and reviews that might help with the integrity of applications. Coupled to that could be a native application on each platform that would allow searching and browsing of the registry and then also handle installation in the form of providing the home screen linking, plus maybe an initial security configuration that determines what features are allowed access to for the app. I’m not holding my breath. In order for this sort of thing to take off and gain widespread appeal, a lot of coordination would be required. And in order to get enough traction it would have to come from a well known entity – a mobile Web app store from a no name source is unlikely to gain high enough usage numbers to make a difference. In a way this would eliminate some of the freedom of the Web, but of course this would also be an optional search path in addition to the standard open Web search mechanisms to find and access content today. Security Security is a big deal, and one of the perceived reasons why so many IT professionals appear to be willing to go back to the walled garden of deployed apps is that Apps are perceived as safe due to the official review and curation of the App stores. Curated stores are supposed to protect you from malware, illegal and misleading content. It doesn’t always work out that way and all the major vendors have had issues with security and the review process at some time or another. Security is critical, but I also think that Web applications in general pose less of a security threat than native applications, by nature of the sandboxed browser and JavaScript environments. Web applications run externally completely and in the HTML and JavaScript sandboxes, with only a very few controlled APIs allowing access to device specific features. And as discussed earlier – security for any device interaction can be granted the same for mobile applications through a Web browser, as they can for native applications either via explicit policies loaded from the Web, or via prompting as GeoLocation does today. Security is important, but it’s certainly solvable problem for Web applications even those that need to access device hardware. Security shouldn’t be a reason for Web apps to be an equal player in mobile applications. Apps are winning, but haven’t we been here before? So now we’re finding ourselves back in an era of installed app, rather than Web based and managed apps. Only it’s even worse today than with Desktop applications, in that the apps are going through a gatekeeper that charges a toll and censors what you can and can’t do in your apps. Frankly it’s a mystery to me why anybody would buy into this model and why it’s lasted this long when we’ve already been through this process. It’s crazy… It’s really a shame that this regression is happening. We have the technology to make mobile Web apps much more prominent, but yet we’re basically held back by what seems little more than bureaucracy, partisan bickering and self interest of the major parties involved. Back in the day of the desktop it was Internet Explorer’s 98+% market shareholding back the Web from improvements for many years – now it’s the combined mobile OS market in control of the mobile browsers. If mobile Web apps were allowed to be treated the same as native apps with simple ways to install and run them consistently and persistently, that would go a long way to making mobile applications much more usable and seriously viable alternatives to native apps. But as it is mobile apps have a severe disadvantage in placement and operation. There are a few bright spots in all of this. Mozilla’s FireFoxOs is embracing the Web for it’s mobile OS by essentially building every app out of HTML and JavaScript based content. It supports both packaged and certified package modes (that can be put into the app store), and Open Web apps that are loaded and run completely off the Web and can also cache locally for offline operation using a manifest. Open Web apps are treated as full class citizens in FireFoxOS and run using the same mechanism as installed apps. Unfortunately FireFoxOs is getting a slow start with minimal device support and specifically targeting the low end market. We can hope that this approach will change and catch on with other vendors, but that’s also an uphill battle given the conflict of interest with platform lock in that it represents. Recent versions of Android also seem to be working reasonably well with mobile application integration onto the desktop and activation out of the box. Although it still uses the Apple meta tags to find icons and behavior settings, everything at least works as you would expect – icons to the desktop on pinning, WebView based full screen activation, and reliable application persistence as the browser/app is treated like a real application. Hopefully iOS will at some point provide this same level of rudimentary Web app support. What’s also interesting to me is that Microsoft hasn’t picked up on the obvious need for a solid Web App platform. Being a distant third in the mobile OS war, Microsoft certainly has nothing to lose and everything to gain by using fresh ideas and expanding into areas that the other major vendors are neglecting. But instead Microsoft is trying to beat the market leaders at their own game, fighting on their adversary’s terms instead of taking a new tack. Providing a kick ass mobile Web platform that takes the lead on some of the proposed mobile APIs would be something positive that Microsoft could do to improve its miserable position in the mobile device market. Where are we at with Mobile Web? It sure sounds like I’m really down on the Mobile Web, right? I’ve built a number of mobile apps in the last year and while overall result and response has been very positive to what we were able to accomplish in terms of UI, getting that final 10% that required device integration dialed was an absolute nightmare on every single one of them. Big compromises had to be made and some features were left out or had to be modified for some devices. In two cases we opted to go the Cordova route in order to get the integration we needed, along with the extra pain involved in that process. Unless you’re not integrating with device features and you don’t care deeply about a smooth integration with the mobile desktop, mobile Web development is fraught with frustration. So, yes I’m frustrated! But it’s not for lack of wanting the mobile Web to succeed. I am still a firm believer that we will eventually arrive a much more functional mobile Web platform that allows access to the most common device features in a sensible way. It wouldn't be difficult for device platform vendors to make Web based applications first class citizens on mobile devices. But unfortunately it looks like it will still be some time before this happens. So, what’s your experience building mobile Web apps? Are you finding similar issues? Just giving up on raw Web applications and building PhoneGap apps instead? Completely skipping the Web and going native? Leave a comment for discussion. Resources Rick Strahl on DotNet Rocks talking about Mobile Web© Rick Strahl, West Wind Technologies, 2005-2014Posted in HTML5 Mobile Tweet !function(d,s,id){var js,fjs=d.getElementsByTagName(s)[0];if(!d.getElementById(id)){js=d.createElement(s);js.id=id;js.src="//platform.twitter.com/widgets.js";fjs.parentNode.insertBefore(js,fjs);}}(document,"script","twitter-wjs"); (function() { var po = document.createElement('script'); po.type = 'text/javascript'; po.async = true; po.src = 'https://apis.google.com/js/plusone.js'; var s = document.getElementsByTagName('script')[0]; s.parentNode.insertBefore(po, s); })();

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Setting up a local AI server - easy with Solaris 11

- by Stefan Hinker

Many things are new in Solaris 11, Autoinstall is one of them. If, like me, you've known Jumpstart for the last 2 centuries or so, you'll have to start from scratch. Well, almost, as the concepts are similar, and it's not all that difficult. Just new. I wanted to have an AI server that I could use for demo purposes, on the train if need be. That answers the question of hardware requirements: portable. But let's start at the beginning. First, you need an OS image, of course. In the new world of Solaris 11, it is now called a repository. The original can be downloaded from the Solaris 11 page at Oracle. What you want is the "Oracle Solaris 11 11/11 Repository Image", which comes in two parts that can be combined using cat. MD5 checksums for these (and all other downloads from that page) are available closer to the top of the page. With that, building the repository is quick and simple: # zfs create -o mountpoint=/export/repo rpool/ai/repo # zfs create rpool/ai/repo/s11 # mount -o ro -F hsfs /tmp/sol-11-1111-repo-full.iso /mnt # rsync -aP /mnt/repo /export/repo/s11 # umount /mnt # pkgrepo rebuild -s /export/repo/sol11/repo # zfs snapshot rpool/ai/repo/sol11@fcs # pkgrepo info -s /export/repo/sol11/repo PUBLISHER PACKAGES STATUS UPDATED solaris 4292 online 2012-03-12T20:47:15.378639Z That's all there's to it. Let's make a snapshot, just to be on the safe side. You never know when one will come in handy. To use this repository, you could just add it as a file-based publisher: # pkg set-publisher -g file:///export/repo/sol11/repo solaris In case I'd want to access this repository through a (virtual) network, i'll now quickly activate the repository-service: # svccfg -s application/pkg/server \ setprop pkg/inst_root=/export/repo/sol11/repo # svccfg -s application/pkg/server setprop pkg/readonly=true # svcadm refresh application/pkg/server # svcadm enable application/pkg/server That's all you need - now point your browser to http://localhost/ to view your beautiful repository-server. Step 1 is done. All of this, by the way, is nicely documented in the README file that's contained in the repository image. Of course, we already have updates to the original release. You can find them in MOS in the Oracle Solaris 11 Support Repository Updates (SRU) Index. You can simply add these to your existing repository or create separate repositories for each SRU. The individual SRUs are self-sufficient and incremental - SRU4 includes all updates from SRU2 and SRU3. With ZFS, you can also get both: A full repository with all updates and at the same time incremental ones up to each of the updates: # mount -o ro -F hsfs /tmp/sol-11-1111-sru4-05-incr-repo.iso /mnt # pkgrecv -s /mnt/repo -d /export/repo/sol11/repo '*' # umount /mnt # pkgrepo rebuild -s /export/repo/sol11/repo # zfs snapshot rpool/ai/repo/sol11@sru4 # zfs set snapdir=visible rpool/ai/repo/sol11 # svcadm restart svc:/application/pkg/server:default The normal repository is now updated to SRU4. Thanks to the ZFS snapshots, there is also a valid repository of Solaris 11 11/11 without the update located at /export/repo/sol11/.zfs/snapshot/fcs . If you like, you can also create another repository service for each update, running on a separate port. But now lets continue with the AI server. Just a little bit of reading in the dokumentation makes it clear that we will need to run a DHCP server for this. Since I already have one active (for my SunRay installation) and since it's a good idea to have these kinds of services separate anyway, I decided to create this in a Zone. So, let's create one first: # zfs create -o mountpoint=/export/install rpool/ai/install # zfs create -o mountpoint=/zones rpool/zones # zonecfg -z ai-server zonecfg:ai-server> create create: Using system default template 'SYSdefault' zonecfg:ai-server> set zonepath=/zones/ai-server zonecfg:ai-server> add dataset zonecfg:ai-server:dataset> set name=rpool/ai/install zonecfg:ai-server:dataset> set alias=install zonecfg:ai-server:dataset> end zonecfg:ai-server> commit zonecfg:ai-server> exit # zoneadm -z ai-server install # zoneadm -z ai-server boot ; zlogin -C ai-server Give it a hostname and IP address at first boot, and there's the Zone. For a publisher for Solaris packages, it will be bound to the "System Publisher" from the Global Zone. The /export/install filesystem, of course, is intended to be used by the AI server. Let's configure it now: #zlogin ai-server root@ai-server:~# pkg install install/installadm root@ai-server:~# installadm create-service -n x86-fcs -a i386 \ -s pkg://solaris/install-image/[email protected],5.11-0.175.0.0.0.2.1482 \ -d /export/install/fcs -i 192.168.2.20 -c 3 With that, the core AI server is already done. What happened here? First, I installed the AI server software. IPS makes that nice and easy. If necessary, it'll also pull in the required DHCP-Server and anything else that might be missing. Watch out for that DHCP server software. In Solaris 11, there are two different versions. There's the one you might know from Solaris 10 and earlier, and then there's a new one from ISC. The latter is the one we need for AI. The SMF service names of both are very similar. The "old" one is "svc:/network/dhcp-server:default". The ISC-server comes with several SMF-services. We at least need "svc:/network/dhcp/server:ipv4". The command "installadm create-service" creates the installation-service. It's called "x86-fcs", serves the "i386" architecture and gets its boot image from the repository of the system publisher, using version 5.11,5.11-0.175.0.0.0.2.1482, which is Solaris 11 11/11. (The option "-a i386" in this example is optional, since the installserver itself runs on a x86 machine.) The boot-environment for clients is created in /export/install/fcs and the DHCP-server is configured for 3 IP-addresses starting at 192.168.2.20. This configuration is stored in a very human readable form in /etc/inet/dhcpd4.conf. An AI-service for SPARC systems could be created in the very same way, using "-a sparc" as the architecture option. Now we would be ready to register and install the first client. It would be installed with the default "solaris-large-server" using the publisher "http://pkg.oracle.com/solaris/release" and would query it's configuration interactively at first boot. This makes it very clear that an AI-server is really only a boot-server. The true source of packets to install can be different. Since I don't like these defaults for my demo setup, I did some extra config work for my clients. The configuration of a client is controlled by manifests and profiles. The manifest controls which packets are installed and how the filesystems are layed out. In that, it's very much like the old "rules.ok" file in Jumpstart. Profiles contain additional configuration like root passwords, primary user account, IP addresses, keyboard layout etc. Hence, profiles are very similar to the old sysid.cfg file. The easiest way to get your hands on a manifest is to ask the AI server we just created to give us it's default one. Then modify that to our liking and give it back to the installserver to use: root@ai-server:~# mkdir -p /export/install/configs/manifests root@ai-server:~# cd /export/install/configs/manifests root@ai-server:~# installadm export -n x86-fcs -m orig_default \ -o orig_default.xml root@ai-server:~# cp orig_default.xml s11-fcs.small.local.xml root@ai-server:~# vi s11-fcs.small.local.xml root@ai-server:~# more s11-fcs.small.local.xml <!DOCTYPE auto_install SYSTEM "file:///usr/share/install/ai.dtd.1"> <auto_install> <ai_instance name="S11 Small fcs local"> <target> <logical> <zpool name="rpool" is_root="true"> <filesystem name="export" mountpoint="/export"/> <filesystem name="export/home"/> <be name="solaris"/> </zpool> </logical> </target> <software type="IPS"> <destination> <image>  <facet set="false">facet.locale.*</facet> <facet set="true">facet.locale.de</facet> <facet set="true">facet.locale.de_DE</facet> <facet set="true">facet.locale.en</facet> <facet set="true">facet.locale.en_US</facet> </image> </destination> <source> <publisher name="solaris"> <origin name="http://192.168.2.12/"/> </publisher> </source>  <software_data action="install"> <name>pkg:/[email protected],5.11-0.175.0.0.0.2.0</name> <name>pkg:/group/system/solaris-small-server</name> </software_data> </software> </ai_instance> </auto_install> root@ai-server:~# installadm create-manifest -n x86-fcs -d \ -f ./s11-fcs.small.local.xml root@ai-server:~# installadm list -m -n x86-fcs Manifest Status Criteria -------- ------ -------- S11 Small fcs local Default None orig_default Inactive None The major points in this new manifest are: Install "solaris-small-server" Install a few locales less than the default. I'm not that fluid in French or Japanese... Use my own package service as publisher, running on IP address 192.168.2.12 Install the initial release of Solaris 11: pkg:/[email protected],5.11-0.175.0.0.0.2.0 Using a similar approach, I'll create a default profile interactively and use it as a template for a few customized building blocks, each defining a part of the overall system configuration. The modular approach makes it easy to configure numerous clients later on: root@ai-server:~# mkdir -p /export/install/configs/profiles root@ai-server:~# cd /export/install/configs/profiles root@ai-server:~# sysconfig create-profile -o default.xml root@ai-server:~# cp default.xml general.xml; cp default.xml mars.xml root@ai-server:~# cp default.xml user.xml root@ai-server:~# vi general.xml mars.xml user.xml root@ai-server:~# more general.xml mars.xml user.xml :::::::::::::: general.xml :::::::::::::: <!DOCTYPE service_bundle SYSTEM "/usr/share/lib/xml/dtd/service_bundle.dtd.1"> <service_bundle type="profile" name="sysconfig"> <service version="1" type="service" name="system/timezone"> <instance enabled="true" name="default"> <property_group type="application" name="timezone"> <propval type="astring" name="localtime" value="Europe/Berlin"/> </property_group> </instance> </service> <service version="1" type="service" name="system/environment"> <instance enabled="true" name="init"> <property_group type="application" name="environment"> <propval type="astring" name="LANG" value="C"/> </property_group> </instance> </service> <service version="1" type="service" name="system/keymap"> <instance enabled="true" name="default"> <property_group type="system" name="keymap"> <propval type="astring" name="layout" value="US-English"/> </property_group> </instance> </service> <service version="1" type="service" name="system/console-login"> <instance enabled="true" name="default"> <property_group type="application" name="ttymon"> <propval type="astring" name="terminal_type" value="vt100"/> </property_group> </instance> </service> <service version="1" type="service" name="network/physical"> <instance enabled="true" name="default"> <property_group type="application" name="netcfg"> <propval type="astring" name="active_ncp" value="DefaultFixed"/> </property_group> </instance> </service> <service version="1" type="service" name="system/name-service/switch"> <property_group type="application" name="config"> <propval type="astring" name="default" value="files"/> <propval type="astring" name="host" value="files dns"/> <propval type="astring" name="printer" value="user files"/> </property_group> <instance enabled="true" name="default"/> </service> <service version="1" type="service" name="system/name-service/cache"> <instance enabled="true" name="default"/> </service> <service version="1" type="service" name="network/dns/client"> <property_group type="application" name="config"> <property type="net_address" name="nameserver"> <net_address_list> <value_node value="192.168.2.1"/> </net_address_list> </property> </property_group> <instance enabled="true" name="default"/> </service> </service_bundle> :::::::::::::: mars.xml :::::::::::::: <!DOCTYPE service_bundle SYSTEM "/usr/share/lib/xml/dtd/service_bundle.dtd.1"> <service_bundle type="profile" name="sysconfig"> <service version="1" type="service" name="network/install"> <instance enabled="true" name="default"> <property_group type="application" name="install_ipv4_interface"> <propval type="astring" name="address_type" value="static"/> <propval type="net_address_v4" name="static_address" value="192.168.2.100/24"/> <propval type="astring" name="name" value="net0/v4"/> <propval type="net_address_v4" name="default_route" value="192.168.2.1"/> </property_group> <property_group type="application" name="install_ipv6_interface"> <propval type="astring" name="stateful" value="yes"/> <propval type="astring" name="stateless" value="yes"/> <propval type="astring" name="address_type" value="addrconf"/> <propval type="astring" name="name" value="net0/v6"/> </property_group> </instance> </service> <service version="1" type="service" name="system/identity"> <instance enabled="true" name="node"> <property_group type="application" name="config"> <propval type="astring" name="nodename" value="mars"/> </property_group> </instance> </service> </service_bundle> :::::::::::::: user.xml :::::::::::::: <!DOCTYPE service_bundle SYSTEM "/usr/share/lib/xml/dtd/service_bundle.dtd.1"> <service_bundle type="profile" name="sysconfig"> <service version="1" type="service" name="system/config-user"> <instance enabled="true" name="default"> <property_group type="application" name="root_account"> <propval type="astring" name="login" value="root"/> <propval type="astring" name="password" value="noIWillNotTellYouMyPasswordNotEvenEncrypted"/> <propval type="astring" name="type" value="role"/> </property_group> <property_group type="application" name="user_account"> <propval type="astring" name="login" value="stefan"/> <propval type="astring" name="password" value="noIWillNotTellYouMyPasswordNotEvenEncrypted"/> <propval type="astring" name="type" value="normal"/> <propval type="astring" name="description" value="Stefan Hinker"/> <propval type="count" name="uid" value="12345"/> <propval type="count" name="gid" value="10"/> <propval type="astring" name="shell" value="/usr/bin/bash"/> <propval type="astring" name="roles" value="root"/> <propval type="astring" name="profiles" value="System Administrator"/> <propval type="astring" name="sudoers" value="ALL=(ALL) ALL"/> </property_group> </instance> </service> </service_bundle> root@ai-server:~# installadm create-profile -n x86-fcs -f general.xml root@ai-server:~# installadm create-profile -n x86-fcs -f user.xml root@ai-server:~# installadm create-profile -n x86-fcs -f mars.xml \ -c ipv4=192.168.2.100 root@ai-server:~# installadm list -p Service Name Profile ------------ ------- x86-fcs general.xml mars.xml user.xml root@ai-server:~# installadm list -n x86-fcs -p Profile Criteria ------- -------- general.xml None mars.xml ipv4 = 192.168.2.100 user.xml None Here's the idea behind these files: "general.xml" contains settings valid for all my clients. Stuff like DNS servers, for example, which in my case will always be the same. "user.xml" only contains user definitions. That is, a root password and a primary user.Both of these profiles will be valid for all clients (for now). "mars.xml" defines network settings for an individual client. This profile is associated with an IP-Address. For this to work, I'll have to tweak the DHCP-settings in the next step: root@ai-server:~# installadm create-client -e 08:00:27:AA:3D:B1 -n x86-fcs root@ai-server:~# vi /etc/inet/dhcpd4.conf root@ai-server:~# tail -5 /etc/inet/dhcpd4.conf host 080027AA3DB1 { hardware ethernet 08:00:27:AA:3D:B1; fixed-address 192.168.2.100; filename "01080027AA3DB1"; } This completes the client preparations. I manually added the IP-Address for mars to /etc/inet/dhcpd4.conf. This is needed for the "mars.xml" profile. Disabling arbitrary DHCP-replies will shut up this DHCP server, making my life in a shared environment a lot more peaceful ;-)Now, I of course want this installation to be completely hands-off. For this to work, I'll need to modify the grub boot menu for this client slightly. You can find it in /etc/netboot. "installadm create-client" will create a new boot menu for every client, identified by the client's MAC address. The template for this can be found in a subdirectory with the name of the install service, /etc/netboot/x86-fcs in our case. If you don't want to change this manually for every client, modify that template to your liking instead. root@ai-server:~# cd /etc/netboot root@ai-server:~# cp menu.lst.01080027AA3DB1 menu.lst.01080027AA3DB1.org root@ai-server:~# vi menu.lst.01080027AA3DB1 root@ai-server:~# diff menu.lst.01080027AA3DB1 menu.lst.01080027AA3DB1.org 1,2c1,2 < default=1 < timeout=10 --- > default=0 > timeout=30 root@ai-server:~# more menu.lst.01080027AA3DB1 default=1 timeout=10 min_mem64=0 title Oracle Solaris 11 11/11 Text Installer and command line kernel$ /x86-fcs/platform/i86pc/kernel/$ISADIR/unix -B install_media=htt p://$serverIP:5555//export/install/fcs,install_service=x86-fcs,install_svc_addre ss=$serverIP:5555 module$ /x86-fcs/platform/i86pc/$ISADIR/boot_archive title Oracle Solaris 11 11/11 Automated Install kernel$ /x86-fcs/platform/i86pc/kernel/$ISADIR/unix -B install=true,inst all_media=http://$serverIP:5555//export/install/fcs,install_service=x86-fcs,inst all_svc_address=$serverIP:5555,livemode=text module$ /x86-fcs/platform/i86pc/$ISADIR/boot_archive Now just boot the client off the network using PXE-boot. For my demo purposes, that's a client from VirtualBox, of course. That's all there's to it. And despite the fact that this blog entry is a little longer - that wasn't that hard now, was it?

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Syncing Data with a Server using Silverlight and HTTP Polling Duplex

- by dwahlin

Many applications have the need to stay in-sync with data provided by a service. Although web applications typically rely on standard polling techniques to check if data has changed, Silverlight provides several interesting options for keeping an application in-sync that rely on server “push” technologies. A few years back I wrote several blog posts covering different “push” technologies available in Silverlight that rely on sockets or HTTP Polling Duplex. We recently had a project that looked like it could benefit from pushing data from a server to one or more clients so I thought I’d revisit the subject and provide some updates to the original code posted. If you’ve worked with AJAX before in Web applications then you know that until browsers fully support web sockets or other duplex (bi-directional communication) technologies that it’s difficult to keep applications in-sync with a server without relying on polling. The problem with polling is that you have to check for changes on the server on a timed-basis which can often be wasteful and take up unnecessary resources. With server “push” technologies, data can be pushed from the server to the client as it changes. Once the data is received, the client can update the user interface as appropriate. Using “push” technologies allows the client to listen for changes from the data but stay 100% focused on client activities as opposed to worrying about polling and asking the server if anything has changed. Silverlight provides several options for pushing data from a server to a client including sockets, TCP bindings and HTTP Polling Duplex. Each has its own strengths and weaknesses as far as performance and setup work with HTTP Polling Duplex arguably being the easiest to setup and get going. In this article I’ll demonstrate how HTTP Polling Duplex can be used in Silverlight 4 applications to push data and show how you can create a WCF server that provides an HTTP Polling Duplex binding that a Silverlight client can consume. What is HTTP Polling Duplex? Technologies that allow data to be pushed from a server to a client rely on duplex functionality. Duplex (or bi-directional) communication allows data to be passed in both directions. A client can call a service and the server can call the client. HTTP Polling Duplex (as its name implies) allows a server to communicate with a client without forcing the client to constantly poll the server. It has the benefit of being able to run on port 80 making setup a breeze compared to the other options which require specific ports to be used and cross-domain policy files to be exposed on port 943 (as with sockets and TCP bindings). Having said that, if you’re looking for the best speed possible then sockets and TCP bindings are the way to go. But, they’re not the only game in town when it comes to duplex communication. The first time I heard about HTTP Polling Duplex (initially available in Silverlight 2) I wasn’t exactly sure how it was any better than standard polling used in AJAX applications. I read the Silverlight SDK, looked at various resources and generally found the following definition unhelpful as far as understanding the actual benefits that HTTP Polling Duplex provided: "The Silverlight client periodically polls the service on the network layer, and checks for any new messages that the service wants to send on the callback channel. The service queues all messages sent on the client callback channel and delivers them to the client when the client polls the service." Although the previous definition explained the overall process, it sounded as if standard polling was used. Fortunately, Microsoft’s Scott Guthrie provided me with a more clear definition several years back that explains the benefits provided by HTTP Polling Duplex quite well (used with his permission): "The [HTTP Polling Duplex] duplex support does use polling in the background to implement notifications – although the way it does it is different than manual polling. It initiates a network request, and then the request is effectively “put to sleep” waiting for the server to respond (it doesn’t come back immediately). The server then keeps the connection open but not active until it has something to send back (or the connection times out after 90 seconds – at which point the duplex client will connect again and wait). This way you are avoiding hitting the server repeatedly – but still get an immediate response when there is data to send." After hearing Scott’s definition the light bulb went on and it all made sense. A client makes a request to a server to check for changes, but instead of the request returning immediately, it parks itself on the server and waits for data. It’s kind of like waiting to pick up a pizza at the store. Instead of calling the store over and over to check the status, you sit in the store and wait until the pizza (the request data) is ready. Once it’s ready you take it back home (to the client). This technique provides a lot of efficiency gains over standard polling techniques even though it does use some polling of its own as a request is initially made from a client to a server. So how do you implement HTTP Polling Duplex in your Silverlight applications? Let’s take a look at the process by starting with the server. Creating an HTTP Polling Duplex WCF Service Creating a WCF service that exposes an HTTP Polling Duplex binding is straightforward as far as coding goes. Add some one way operations into an interface, create a client callback interface and you’re ready to go. The most challenging part comes into play when configuring the service to properly support the necessary binding and that’s more of a cut and paste operation once you know the configuration code to use. To create an HTTP Polling Duplex service you’ll need to expose server-side and client-side interfaces and reference the System.ServiceModel.PollingDuplex assembly (located at C:\Program Files (x86)\Microsoft SDKs\Silverlight\v4.0\Libraries\Server on my machine) in the server project. For the demo application I upgraded a basketball simulation service to support the latest polling duplex assemblies. The service simulates a simple basketball game using a Game class and pushes information about the game such as score, fouls, shots and more to the client as the game changes over time. Before jumping too far into the game push service, it’s important to discuss two interfaces used by the service to communicate in a bi-directional manner. The first is called IGameStreamService and defines the methods/operations that the client can call on the server (see Listing 1). The second is IGameStreamClient which defines the callback methods that a server can use to communicate with a client (see Listing 2). [ServiceContract(Namespace = "Silverlight", CallbackContract = typeof(IGameStreamClient))] public interface IGameStreamService { [OperationContract(IsOneWay = true)] void GetTeamData(); } Listing 1. The IGameStreamService interface defines server operations that can be called on the server. [ServiceContract] public interface IGameStreamClient { [OperationContract(IsOneWay = true)] void ReceiveTeamData(List<Team> teamData); [OperationContract(IsOneWay = true, AsyncPattern=true)] IAsyncResult BeginReceiveGameData(GameData gameData, AsyncCallback callback, object state); void EndReceiveGameData(IAsyncResult result); } Listing 2. The IGameStreamClient interfaces defines client operations that a server can call. The IGameStreamService interface is decorated with the standard ServiceContract attribute but also contains a value for the CallbackContract property. This property is used to define the interface that the client will expose (IGameStreamClient in this example) and use to receive data pushed from the service. Notice that each OperationContract attribute in both interfaces sets the IsOneWay property to true. This means that the operation can be called and passed data as appropriate, however, no data will be passed back. Instead, data will be pushed back to the client as it’s available. Looking through the IGameStreamService interface you can see that the client can request team data whereas the IGameStreamClient interface allows team and game data to be received by the client. One interesting point about the IGameStreamClient interface is the inclusion of the AsyncPattern property on the BeginReceiveGameData operation. I initially created this operation as a standard one way operation and it worked most of the time. However, as I disconnected clients and reconnected new ones game data wasn’t being passed properly. After researching the problem more I realized that because the service could take up to 7 seconds to return game data, things were getting hung up. By setting the AsyncPattern property to true on the BeginReceivedGameData operation and providing a corresponding EndReceiveGameData operation I was able to get around this problem and get everything running properly. I’ll provide more details on the implementation of these two methods later in this post. Once the interfaces were created I moved on to the game service class. The first order of business was to create a class that implemented the IGameStreamService interface. Since the service can be used by multiple clients wanting game data I added the ServiceBehavior attribute to the class definition so that I could set its InstanceContextMode to InstanceContextMode.Single (in effect creating a Singleton service object). Listing 3 shows the game service class as well as its fields and constructor. [ServiceBehavior(ConcurrencyMode = ConcurrencyMode.Multiple, InstanceContextMode = InstanceContextMode.Single)] public class GameStreamService : IGameStreamService { object _Key = new object(); Game _Game = null; Timer _Timer = null; Random _Random = null; Dictionary<string, IGameStreamClient> _ClientCallbacks = new Dictionary<string, IGameStreamClient>(); static AsyncCallback _ReceiveGameDataCompleted = new AsyncCallback(ReceiveGameDataCompleted); public GameStreamService() { _Game = new Game(); _Timer = new Timer { Enabled = false, Interval = 2000, AutoReset = true }; _Timer.Elapsed += new ElapsedEventHandler(_Timer_Elapsed); _Timer.Start(); _Random = new Random(); }} Listing 3. The GameStreamService implements the IGameStreamService interface which defines a callback contract that allows the service class to push data back to the client. By implementing the IGameStreamService interface, GameStreamService must supply a GetTeamData() method which is responsible for supplying information about the teams that are playing as well as individual players. GetTeamData() also acts as a client subscription method that tracks clients wanting to receive game data. Listing 4 shows the GetTeamData() method. public void GetTeamData() { //Get client callback channel var context = OperationContext.Current; var sessionID = context.SessionId; var currClient = context.GetCallbackChannel<IGameStreamClient>(); context.Channel.Faulted += Disconnect; context.Channel.Closed += Disconnect; IGameStreamClient client; if (!_ClientCallbacks.TryGetValue(sessionID, out client)) { lock (_Key) { _ClientCallbacks[sessionID] = currClient; } } currClient.ReceiveTeamData(_Game.GetTeamData()); //Start timer which when fired sends updated score information to client if (!_Timer.Enabled) { _Timer.Enabled = true; } } Listing 4. The GetTeamData() method subscribes a given client to the game service and returns. The key the line of code in the GetTeamData() method is the call to GetCallbackChannel<IGameStreamClient>(). This method is responsible for accessing the calling client’s callback channel. The callback channel is defined by the IGameStreamClient interface shown earlier in Listing 2 and used by the server to communicate with the client. Before passing team data back to the client, GetTeamData() grabs the client’s session ID and checks if it already exists in the _ClientCallbacks dictionary object used to track clients wanting callbacks from the server. If the client doesn’t exist it adds it into the collection. It then pushes team data from the Game class back to the client by calling ReceiveTeamData(). Since the service simulates a basketball game, a timer is then started if it’s not already enabled which is then used to randomly send data to the client. When the timer fires, game data is pushed down to the client. Listing 5 shows the _Timer_Elapsed() method that is called when the timer fires as well as the SendGameData() method used to send data to the client. void _Timer_Elapsed(object sender, ElapsedEventArgs e) { int interval = _Random.Next(3000, 7000); lock (_Key) { _Timer.Interval = interval; _Timer.Enabled = false; } SendGameData(_Game.GetGameData()); } private void SendGameData(GameData gameData) { var cbs = _ClientCallbacks.Where(cb => ((IContextChannel)cb.Value).State == CommunicationState.Opened); for (int i = 0; i < cbs.Count(); i++) { var cb = cbs.ElementAt(i).Value; try { cb.BeginReceiveGameData(gameData, _ReceiveGameDataCompleted, cb); } catch (TimeoutException texp) { //Log timeout error } catch (CommunicationException cexp) { //Log communication error } } lock (_Key) _Timer.Enabled = true; } private static void ReceiveGameDataCompleted(IAsyncResult result) { try { ((IGameStreamClient)(result.AsyncState)).EndReceiveGameData(result); } catch (CommunicationException) { // empty } catch (TimeoutException) { // empty } } LIsting 5. _Timer_Elapsed is used to simulate time in a basketball game. When _Timer_Elapsed() fires the SendGameData() method is called which iterates through the clients wanting to be notified of changes. As each client is identified, their respective BeginReceiveGameData() method is called which ultimately pushes game data down to the client. Recall that this method was defined in the client callback interface named IGameStreamClient shown earlier in Listing 2. Notice that BeginReceiveGameData() accepts _ReceiveGameDataCompleted as its second parameter (an AsyncCallback delegate defined in the service class) and passes the client callback as the third parameter. The initial version of the sample application had a standard ReceiveGameData() method in the client callback interface. However, sometimes the client callbacks would work properly and sometimes they wouldn’t which was a little baffling at first glance. After some investigation I realized that I needed to implement an asynchronous pattern for client callbacks to work properly since 3 – 7 second delays are occurring as a result of the timer. Once I added the BeginReceiveGameData() and ReceiveGameDataCompleted() methods everything worked properly since each call was handled in an asynchronous manner. The final task that had to be completed to get the server working properly with HTTP Polling Duplex was adding configuration code into web.config. In the interest of brevity I won’t post all of the code here since the sample application includes everything you need. However, Listing 6 shows the key configuration code to handle creating a custom binding named pollingDuplexBinding and associate it with the service’s endpoint. <bindings> <customBinding> <binding name="pollingDuplexBinding"> <binaryMessageEncoding /> <pollingDuplex maxPendingSessions="2147483647" maxPendingMessagesPerSession="2147483647" inactivityTimeout="02:00:00" serverPollTimeout="00:05:00"/> <httpTransport /> </binding> </customBinding> </bindings> <services> <service name="GameService.GameStreamService" behaviorConfiguration="GameStreamServiceBehavior"> <endpoint address="" binding="customBinding" bindingConfiguration="pollingDuplexBinding" contract="GameService.IGameStreamService"/> <endpoint address="mex" binding="mexHttpBinding" contract="IMetadataExchange" /> </service> </services> Listing 6. Configuring an HTTP Polling Duplex binding in web.config and associating an endpoint with it. Calling the Service and Receiving “Pushed” Data Calling the service and handling data that is pushed from the server is a simple and straightforward process in Silverlight. Since the service is configured with a MEX endpoint and exposes a WSDL file, you can right-click on the Silverlight project and select the standard Add Service Reference item. After the web service proxy is created you may notice that the ServiceReferences.ClientConfig file only contains an empty configuration element instead of the normal configuration elements created when creating a standard WCF proxy. You can certainly update the file if you want to read from it at runtime but for the sample application I fed the service URI directly to the service proxy as shown next: var address = new EndpointAddress("http://localhost.:5661/GameStreamService.svc"); var binding = new PollingDuplexHttpBinding(); _Proxy = new GameStreamServiceClient(binding, address); _Proxy.ReceiveTeamDataReceived += _Proxy_ReceiveTeamDataReceived; _Proxy.ReceiveGameDataReceived += _Proxy_ReceiveGameDataReceived; _Proxy.GetTeamDataAsync(); This code creates the proxy and passes the endpoint address and binding to use to its constructor. It then wires the different receive events to callback methods and calls GetTeamDataAsync(). Calling GetTeamDataAsync() causes the server to store the client in the server-side dictionary collection mentioned earlier so that it can receive data that is pushed. As the server-side timer fires and game data is pushed to the client, the user interface is updated as shown in Listing 7. Listing 8 shows the _Proxy_ReceiveGameDataReceived() method responsible for handling the data and calling UpdateGameData() to process it. Listing 7. The Silverlight interface. Game data is pushed from the server to the client using HTTP Polling Duplex. void _Proxy_ReceiveGameDataReceived(object sender, ReceiveGameDataReceivedEventArgs e) { UpdateGameData(e.gameData); } private void UpdateGameData(GameData gameData) { //Update Score this.tbTeam1Score.Text = gameData.Team1Score.ToString(); this.tbTeam2Score.Text = gameData.Team2Score.ToString(); //Update ball visibility if (gameData.Action != ActionsEnum.Foul) { if (tbTeam1.Text == gameData.TeamOnOffense) { AnimateBall(this.BB1, this.BB2); } else //Team 2 { AnimateBall(this.BB2, this.BB1); } } if (this.lbActions.Items.Count > 9) this.lbActions.Items.Clear(); this.lbActions.Items.Add(gameData.LastAction); if (this.lbActions.Visibility == Visibility.Collapsed) this.lbActions.Visibility = Visibility.Visible; } private void AnimateBall(Image onBall, Image offBall) { this.FadeIn.Stop(); Storyboard.SetTarget(this.FadeInAnimation, onBall); Storyboard.SetTarget(this.FadeOutAnimation, offBall); this.FadeIn.Begin(); } Listing 8. As the server pushes game data, the client’s _Proxy_ReceiveGameDataReceived() method is called to process the data. In a real-life application I’d go with a ViewModel class to handle retrieving team data, setup data bindings and handle data that is pushed from the server. However, for the sample application I wanted to focus on HTTP Polling Duplex and keep things as simple as possible. Summary Silverlight supports three options when duplex communication is required in an application including TCP bindins, sockets and HTTP Polling Duplex. In this post you’ve seen how HTTP Polling Duplex interfaces can be created and implemented on the server as well as how they can be consumed by a Silverlight client. HTTP Polling Duplex provides a nice way to “push” data from a server while still allowing the data to flow over port 80 or another port of your choice. Sample Application Download

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Automatic Standby Recreation for Data Guard

- by pablo.boixeda(at)oracle.com

Hi,Unfortunately sometimes a Standby Instance needs to be recreated. This can happen for many reasons such as lost archive logs, standby data files, failover, among others.This is why we wanted to have one script to recreate standby instances in an easy way.This script recreates the standby considering some prereqs:-Database Version should be at least 11gR1-Dummy instance started on the standby node (Seeking to improve this so it won't be needed)-Broker configuration hasn't been removed-In our case we have two TNSNAMES files, one for the Standby creation (using SID) and the other one for production using service names (including broker service name)-Some environment variables set up by the environment db script (like ORACLE_HOME, PATH...)-The directory tree should not have been modified in the stanby hostWe are currently using it on our 11gR2 Data Guard tests.Any improvements will be welcome! Normal 0 21 false false false ES 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:0cm 5.4pt 0cm 5.4pt; mso-para-margin-top:0cm; mso-para-margin-right:0cm; mso-para-margin-bottom:10.0pt; mso-para-margin-left:0cm; line-height:115%; 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;} #!/bin/ksh ### NOMBRE / VERSION ### recrea_dg.sh v.1.00 ### ### DESCRIPCION ### reacreacion de la Standby ### ### DEVUELVE ### 0 Creacion de STANDBY correcta ### 1 Fallo ### ### NOTAS ### Este shell script NO DEBE MODIFICARSE. ### Todas las variables y constantes necesarias se toman del entorno. ### ### MODIFICADO POR: FECHA: COMENTARIOS: ### --------------- ---------- ------------------------------------- ### Oracle 15/02/2011 Creacion. ### ### ### Cargar entorno ### V_ADMIN_DIR=`dirname $0` . ${V_ADMIN_DIR}/entorno_bd.sh 1>>/dev/null if [ $? -ne 0 ] then echo "Error Loading the environment." exit 1 fi V_RET=0 V_DATE=`/bin/date` V_DATE_F=`/bin/date +%Y%m%d_%H%M%S` V_LOGFILE=${V_TRAZAS}/recrea_dg_${V_DATE_F}.log exec 4>&1 tee ${V_FICH_LOG} >&4 |& exec 1>&p 2>&1 ### ### Variables para Recrear el Data Guard ### V_DB_BR=`echo ${V_DB_NAME}|tr '[:lower:]' '[:upper:]'` if [ "${ORACLE_SID}" = "${V_DB_NAME}01" ] then V_LOCAL_BR=${V_DB_BR}'01' V_REMOTE_BR=${V_DB_BR}'02' else V_LOCAL_BR=${V_DB_BR}'02' V_REMOTE_BR=${V_DB_BR}'01' fi echo " Getting local instance ROLE ${ORACLE_SID} ..." sqlplus -s /nolog 1>>/dev/null 2>&1 <<-! whenever sqlerror exit 1 connect / as sysdba variable salida number declare v_database_role v\$database.database_role%type; begin select database_role into v_database_role from v\$database; :salida := case v_database_role when 'PRIMARY' then 2 when 'PHYSICAL STANDBY' then 3 else 4 end; end; / exit :salida ! case $? in 1) echo " ERROR: Cannot get instance ROLE ." | tee -a ${V_LOGFILE} 2>&1 V_RET=1 ;; 2) echo " Local Instance with PRIMARY role." | tee -a ${V_LOGFILE} 2>&1 V_DB_ROLE_LCL=PRIMARY ;; 3) echo " Local Instance with PHYSICAL STANDBY role." | tee -a ${V_LOGFILE} 2>&1 V_DB_ROLE_LCL=STANDBY ;; *) echo " ERROR: UNKNOWN ROLE." | tee -a ${V_LOGFILE} 2>&1 V_RET=1 ;; esac if [ "${V_DB_ROLE_LCL}" = "PRIMARY" ] then echo "####################################################################" | tee -a ${V_LOGFILE} 2>&1 echo "${V_DATE} - Reacreating STANDBY Instance." | tee -a ${V_LOGFILE} 2>&1 echo "" | tee -a ${V_LOGFILE} 2>&1 echo "DATAFILES, CONTROL FILES, REDO LOGS and ARCHIVE LOGS in standby instance ${V_REMOTE_BR} will be removed" | tee -a ${V_LOGFILE} 2>&1 echo "" | tee -a ${V_LOGFILE} 2>&1 V_PRIMARY=${V_LOCAL_BR} V_STANDBY=${V_REMOTE_BR} fi if [ "${V_DB_ROLE_LCL}" = "STANDBY" ] then echo "####################################################################" | tee -a ${V_LOGFILE} 2>&1 echo "${V_DATE} - Reacreating STANDBY Instance." | tee -a ${V_LOGFILE} 2>&1 echo "" | tee -a ${V_LOGFILE} 2>&1 echo "DATAFILES, CONTROL FILES, REDO LOGS and ARCHIVE LOGS in standby instance ${V_LOCAL_BR} will be removed" | tee -a ${V_LOGFILE} 2>&1 echo "" | tee -a ${V_LOGFILE} 2>&1 V_PRIMARY=${V_REMOTE_BR} V_STANDBY=${V_LOCAL_BR} fi # Cargamos las variables de los hosts # Cargamos las variables de los hosts PRY_HOST=`sqlplus /nolog << EOF | grep KEEP | sed 's/KEEP//;s/[ ]//g' connect sys/${V_DB_PWD}@${V_PRIMARY} as sysdba select 'KEEP',host_name from v\\$instance; EOF` SBY_HOST=`sqlplus /nolog << EOF | grep KEEP | sed 's/KEEP//;s/[ ]//g' connect sys/${V_DB_PWD}@${V_STANDBY} as sysdba select 'KEEP',host_name from v\\$instance; EOF` echo "el HOST primary es: ${PRY_HOST}" | tee -a ${V_LOGFILE} 2>&1 echo "el HOST standby es: ${SBY_HOST}" | tee -a ${V_LOGFILE} 2>&1 echo "" | tee -a ${V_LOGFILE} 2>&1 ## ## Paramos la instancia STANDBY ## V_DATE=`/bin/date` echo "${V_DATE} - Shutting down Standby instance" | tee -a ${V_LOGFILE} 2>&1 echo "" | tee -a ${V_LOGFILE} 2>&1 echo "********************************************************************************" | tee -a ${V_LOGFILE} 2>&1 ## ## Paramos la instancia STANDBY ## SBY_STATUS=`sqlplus /nolog << EOF | grep KEEP | sed 's/KEEP//;s/[ ]//g' connect sys/${V_DB_PWD}@${V_STANDBY} as sysdba select 'KEEP',status from v\\$instance; EOF` if [ ${SBY_STATUS} = 'STARTED' ] || [ ${SBY_STATUS} = 'MOUNTED' ] || [ ${SBY_STATUS} = 'OPEN' ] then echo "${V_DATE} - Standby instance shutdown in progress..." | tee -a ${V_LOGFILE} 2>&1 echo "" | tee -a ${V_LOGFILE} 2>&1 echo "********************************************************************************" | tee -a ${V_LOGFILE} 2>&1 sqlplus -s /nolog 1>>/dev/null 2>&1 <<-! whenever sqlerror exit 1 connect sys/${V_DB_PWD}@${V_STANDBY} as sysdba shutdown abort ! fi V_DATE=`/bin/date` echo "" echo "${V_DATE} - Standby instance stopped" | tee -a ${V_LOGFILE} 2>&1 echo "" | tee -a ${V_LOGFILE} 2>&1 echo "********************************************************************************" | tee -a ${V_LOGFILE} 2>&1 ## ## Eliminamos los ficheros de la base de datos ## V_SBY_SID=`echo ${V_STANDBY}|tr '[:upper:]' '[:lower:]'` V_PRY_SID=`echo ${V_PRIMARY}|tr '[:upper:]' '[:lower:]'` ssh ${SBY_HOST} rm /opt/oracle/db/db${V_DB_NAME}/${V_SBY_SID}/data/*.dbf ssh ${SBY_HOST} rm /opt/oracle/db/db${V_DB_NAME}/${V_SBY_SID}/arch/*.arc ssh ${SBY_HOST} rm /opt/oracle/db/db${V_DB_NAME}/${V_SBY_SID}/ctl/*.ctl ssh ${SBY_HOST} rm /opt/oracle/db/db${V_DB_NAME}/${V_SBY_SID}/redo/*.ctl ssh ${SBY_HOST} rm /opt/oracle/db/db${V_DB_NAME}/${V_SBY_SID}/redo/*.rdo ## ## Startup nomount stby instance ## V_DATE=`/bin/date` echo "" | tee -a ${V_LOGFILE} 2>&1 echo "${V_DATE} - Starting DUMMY Standby Instance " | tee -a ${V_LOGFILE} 2>&1 echo "" | tee -a ${V_LOGFILE} 2>&1 echo "********************************************************************************" | tee -a ${V_LOGFILE} 2>&1 ssh ${SBY_HOST} touch /home/oracle/init_dg.ora ssh ${SBY_HOST} 'echo "DB_NAME='${V_DB_NAME}'">>/home/oracle/init_dg.ora' ssh ${SBY_HOST} touch /home/oracle/start_dummy.sh ssh ${SBY_HOST} 'echo "ORACLE_HOME=/opt/oracle/db/db'${V_DB_NAME}'/soft/db11.2.0.2 ">>/home/oracle/start_dummy.sh' ssh ${SBY_HOST} 'echo "export ORACLE_HOME">>/home/oracle/start_dummy.sh' ssh ${SBY_HOST} 'echo "PATH=\$ORACLE_HOME/bin:\$PATH">>/home/oracle/start_dummy.sh' ssh ${SBY_HOST} 'echo "export PATH">>/home/oracle/start_dummy.sh' ssh ${SBY_HOST} 'echo "ORACLE_SID='${V_SBY_SID}'">>/home/oracle/start_dummy.sh' ssh ${SBY_HOST} 'echo "export ORACLE_SID">>/home/oracle/start_dummy.sh' ssh ${SBY_HOST} 'echo "sqlplus -s /nolog <<-!" >>/home/oracle/start_dummy.sh' ssh ${SBY_HOST} 'echo " whenever sqlerror exit 1 ">>/home/oracle/start_dummy.sh' ssh ${SBY_HOST} 'echo " connect / as sysdba ">>/home/oracle/start_dummy.sh' ssh ${SBY_HOST} 'echo " startup nomount pfile='\''/home/oracle/init_dg.ora'\''">>/home/oracle/start_dummy.sh' ssh ${SBY_HOST} 'echo "! ">>/home/oracle/start_dummy.sh' ssh ${SBY_HOST} 'chmod 744 /home/oracle/start_dummy.sh' ssh ${SBY_HOST} 'sh /home/oracle/start_dummy.sh' ssh ${SBY_HOST} 'rm /home/oracle/start_dummy.sh' ssh ${SBY_HOST} 'rm /home/oracle/init_dg.ora' ## ## TNSNAMES change, specific for RMAN duplicate ## V_DATE=`/bin/date` echo "" | tee -a ${V_LOGFILE} 2>&1 echo "${V_DATE} - Setting up TNSNAMES in PRIMARY host " | tee -a ${V_LOGFILE} 2>&1 echo "" | tee -a ${V_LOGFILE} 2>&1 echo "********************************************************************************" | tee -a ${V_LOGFILE} 2>&1 ssh ${PRY_HOST} 'cp /opt/oracle/db/db'${V_DB_NAME}'/soft/db11.2.0.2/network/admin/tnsnames.ora.inst /opt/oracle/db/db'${V_DB_NAME}'/soft/db11.2.0.2/network/admin/tnsnames.ora' V_DATE=`/bin/date` echo "" | tee -a ${V_LOGFILE} 2>&1 echo "${V_DATE} - Starting STANDBY creation with RMAN.. " | tee -a ${V_LOGFILE} 2>&1 echo "" | tee -a ${V_LOGFILE} 2>&1 echo "********************************************************************************" | tee -a ${V_LOGFILE} 2>&1 rman<<-! >>${V_LOGFILE} connect target sys/${V_DB_PWD}@${V_PRIMARY} connect auxiliary sys/${V_DB_PWD}@${V_STANDBY} run { allocate channel prmy1 type disk; allocate channel prmy2 type disk; allocate channel prmy3 type disk; allocate channel prmy4 type disk; allocate auxiliary channel stby type disk; duplicate target database for standby from active database dorecover spfile parameter_value_convert '${V_PRY_SID}','${V_SBY_SID}' set control_files='/opt/oracle/db/db${V_DB_NAME}/${V_SBY_SID}/ctl/control01.ctl','/opt/oracle/db/db${V_DB_NAME}/${V_SBY_SID}/redo/control02.ctl' set db_file_name_convert='/opt/oracle/db/db${V_DB_NAME}/${V_PRY_SID}/','/opt/oracle/db/db${V_DB_NAME}/${V_SBY_SID}/' set log_file_name_convert='/opt/oracle/db/db${V_DB_NAME}/${V_PRY_SID}/','/opt/oracle/db/db${V_DB_NAME}/${V_SBY_SID}/' set 'db_unique_name'='${V_SBY_SID}' set log_archive_config='DG_CONFIG=(${V_PRIMARY},${V_STANDBY})' set fal_client='${V_STANDBY}' set fal_server='${V_PRIMARY}' set log_archive_dest_1='LOCATION=/opt/oracle/db/db${V_DB_NAME}/${V_SBY_SID}/arch DB_UNIQUE_NAME=${V_SBY_SID} MANDATORY VALID_FOR=(ALL_LOGFILES,ALL_ROLES)' set log_archive_dest_2='SERVICE="${V_PRIMARY}"','SYNC AFFIRM DB_UNIQUE_NAME=${V_PRY_SID} DELAY=0 MAX_FAILURE=0 REOPEN=300 REGISTER VALID_FOR=(ONLINE_LOGFILES,PRIMARY_ROLE)' nofilenamecheck ; } ! V_DATE=`/bin/date` if [ $? -ne 0 ] then echo "" echo "${V_DATE} - Error creating STANDBY instance" echo "" echo "********************************************************************************" else echo "" echo "${V_DATE} - STANDBY instance created SUCCESSFULLY " echo "" echo "********************************************************************************" fi sqlplus -s /nolog 1>>/dev/null 2>&1 <<-! whenever sqlerror exit 1 connect sys/${V_DB_PWD}@${V_STANDBY} as sysdba alter system set local_listener='(ADDRESS=(PROTOCOL=TCP)(HOST=${SBY_HOST})(PORT=1544))' scope=both; alter system set service_names='${V_DB_NAME}.eu.roca.net,${V_SBY_SID}.eu.roca.net,${V_SBY_SID}_DGMGRL.eu.roca.net' scope=both; alter database recover managed standby database using current logfile disconnect from session; alter system set dg_broker_start=true scope=both; ! ## ## TNSNAMES change, back to Production Mode ## V_DATE=`/bin/date` echo " " | tee -a ${V_LOGFILE} 2>&1 echo "${V_DATE} - Restoring TNSNAMES in PRIMARY " | tee -a ${V_LOGFILE} 2>&1 echo "" | tee -a ${V_LOGFILE} 2>&1 echo "********************************************************************************" | tee -a ${V_LOGFILE} 2>&1 ssh ${PRY_HOST} 'cp /opt/oracle/db/db'${V_DB_NAME}'/soft/db11.2.0.2/network/admin/tnsnames.ora.prod /opt/oracle/db/db'${V_DB_NAME}'/soft/db11.2.0.2/network/admin/tnsnames.ora' echo "" | tee -a ${V_LOGFILE} 2>&1 echo "${V_DATE} - Waiting for media recovery before check the DATA GUARD Broker" | tee -a ${V_LOGFILE} 2>&1 echo "" | tee -a ${V_LOGFILE} 2>&1 echo "********************************************************************************" | tee -a ${V_LOGFILE} 2>&1 sleep 200 dgmgrl <<-! | grep SUCCESS 1>/dev/null 2>&1 connect ${V_DB_USR}/${V_DB_PWD}@${V_STANDBY} show configuration verbose; ! if [ $? -ne 0 ] ; then echo " ERROR: El status del Broker no es SUCCESS" | tee -a ${V_LOGFILE} 2>&1 ; V_RET=1 else echo " DATA GUARD OK " | tee -a ${V_LOGFILE} 2>&1 ; Normal 0 21 false false false ES 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:0cm 5.4pt 0cm 5.4pt; mso-para-margin-top:0cm; mso-para-margin-right:0cm; mso-para-margin-bottom:10.0pt; mso-para-margin-left:0cm; line-height:115%; 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;} V_RET=0 fi Hope it helps.

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Understanding G1 GC Logs

- by poonam

The purpose of this post is to explain the meaning of GC logs generated with some tracing and diagnostic options for G1 GC. We will take a look at the output generated with PrintGCDetails which is a product flag and provides the most detailed level of information. Along with that, we will also look at the output of two diagnostic flags that get enabled with -XX:+UnlockDiagnosticVMOptions option - G1PrintRegionLivenessInfo that prints the occupancy and the amount of space used by live objects in each region at the end of the marking cycle and G1PrintHeapRegions that provides detailed information on the heap regions being allocated and reclaimed. We will be looking at the logs generated with JDK 1.7.0_04 using these options. Option -XX:+PrintGCDetails Here's a sample log of G1 collection generated with PrintGCDetails. 0.522: [GC pause (young), 0.15877971 secs] [Parallel Time: 157.1 ms] [GC Worker Start (ms): 522.1 522.2 522.2 522.2 Avg: 522.2, Min: 522.1, Max: 522.2, Diff: 0.1] [Ext Root Scanning (ms): 1.6 1.5 1.6 1.9 Avg: 1.7, Min: 1.5, Max: 1.9, Diff: 0.4] [Update RS (ms): 38.7 38.8 50.6 37.3 Avg: 41.3, Min: 37.3, Max: 50.6, Diff: 13.3] [Processed Buffers : 2 2 3 2 Sum: 9, Avg: 2, Min: 2, Max: 3, Diff: 1] [Scan RS (ms): 9.9 9.7 0.0 9.7 Avg: 7.3, Min: 0.0, Max: 9.9, Diff: 9.9] [Object Copy (ms): 106.7 106.8 104.6 107.9 Avg: 106.5, Min: 104.6, Max: 107.9, Diff: 3.3] [Termination (ms): 0.0 0.0 0.0 0.0 Avg: 0.0, Min: 0.0, Max: 0.0, Diff: 0.0] [Termination Attempts : 1 4 4 6 Sum: 15, Avg: 3, Min: 1, Max: 6, Diff: 5] [GC Worker End (ms): 679.1 679.1 679.1 679.1 Avg: 679.1, Min: 679.1, Max: 679.1, Diff: 0.1] [GC Worker (ms): 156.9 157.0 156.9 156.9 Avg: 156.9, Min: 156.9, Max: 157.0, Diff: 0.1] [GC Worker Other (ms): 0.3 0.3 0.3 0.3 Avg: 0.3, Min: 0.3, Max: 0.3, Diff: 0.0] [Clear CT: 0.1 ms] [Other: 1.5 ms] [Choose CSet: 0.0 ms] [Ref Proc: 0.3 ms] [Ref Enq: 0.0 ms] [Free CSet: 0.3 ms] [Eden: 12M(12M)->0B(10M) Survivors: 0B->2048K Heap: 13M(64M)->9739K(64M)] [Times: user=0.59 sys=0.02, real=0.16 secs] This is the typical log of an Evacuation Pause (G1 collection) in which live objects are copied from one set of regions (young OR young+old) to another set. It is a stop-the-world activity and all the application threads are stopped at a safepoint during this time. This pause is made up of several sub-tasks indicated by the indentation in the log entries. Here's is the top most line that gets printed for the Evacuation Pause. 0.522: [GC pause (young), 0.15877971 secs] This is the highest level information telling us that it is an Evacuation Pause that started at 0.522 secs from the start of the process, in which all the regions being evacuated are Young i.e. Eden and Survivor regions. This collection took 0.15877971 secs to finish. Evacuation Pauses can be mixed as well. In which case the set of regions selected include all of the young regions as well as some old regions. 1.730: [GC pause (mixed), 0.32714353 secs] Let's take a look at all the sub-tasks performed in this Evacuation Pause. [Parallel Time: 157.1 ms] Parallel Time is the total elapsed time spent by all the parallel GC worker threads. The following lines correspond to the parallel tasks performed by these worker threads in this total parallel time, which in this case is 157.1 ms. [GC Worker Start (ms): 522.1 522.2 522.2 522.2Avg: 522.2, Min: 522.1, Max: 522.2, Diff: 0.1] The first line tells us the start time of each of the worker thread in milliseconds. The start times are ordered with respect to the worker thread ids – thread 0 started at 522.1ms and thread 1 started at 522.2ms from the start of the process. The second line tells the Avg, Min, Max and Diff of the start times of all of the worker threads. [Ext Root Scanning (ms): 1.6 1.5 1.6 1.9 Avg: 1.7, Min: 1.5, Max: 1.9, Diff: 0.4] This gives us the time spent by each worker thread scanning the roots (globals, registers, thread stacks and VM data structures). Here, thread 0 took 1.6ms to perform the root scanning task and thread 1 took 1.5 ms. The second line clearly shows the Avg, Min, Max and Diff of the times spent by all the worker threads. [Update RS (ms): 38.7 38.8 50.6 37.3 Avg: 41.3, Min: 37.3, Max: 50.6, Diff: 13.3] Update RS gives us the time each thread spent in updating the Remembered Sets. Remembered Sets are the data structures that keep track of the references that point into a heap region. Mutator threads keep changing the object graph and thus the references that point into a particular region. We keep track of these changes in buffers called Update Buffers. The Update RS sub-task processes the update buffers that were not able to be processed concurrently, and updates the corresponding remembered sets of all regions. [Processed Buffers : 2 2 3 2Sum: 9, Avg: 2, Min: 2, Max: 3, Diff: 1] This tells us the number of Update Buffers (mentioned above) processed by each worker thread. [Scan RS (ms): 9.9 9.7 0.0 9.7 Avg: 7.3, Min: 0.0, Max: 9.9, Diff: 9.9] These are the times each worker thread had spent in scanning the Remembered Sets. Remembered Set of a region contains cards that correspond to the references pointing into that region. This phase scans those cards looking for the references pointing into all the regions of the collection set. [Object Copy (ms): 106.7 106.8 104.6 107.9 Avg: 106.5, Min: 104.6, Max: 107.9, Diff: 3.3] These are the times spent by each worker thread copying live objects from the regions in the Collection Set to the other regions. [Termination (ms): 0.0 0.0 0.0 0.0 Avg: 0.0, Min: 0.0, Max: 0.0, Diff: 0.0] Termination time is the time spent by the worker thread offering to terminate. But before terminating, it checks the work queues of other threads and if there are still object references in other work queues, it tries to steal object references, and if it succeeds in stealing a reference, it processes that and offers to terminate again. [Termination Attempts : 1 4 4 6 Sum: 15, Avg: 3, Min: 1, Max: 6, Diff: 5] This gives the number of times each thread has offered to terminate. [GC Worker End (ms): 679.1 679.1 679.1 679.1 Avg: 679.1, Min: 679.1, Max: 679.1, Diff: 0.1] These are the times in milliseconds at which each worker thread stopped. [GC Worker (ms): 156.9 157.0 156.9 156.9 Avg: 156.9, Min: 156.9, Max: 157.0, Diff: 0.1] These are the total lifetimes of each worker thread. [GC Worker Other (ms): 0.3 0.3 0.3 0.3Avg: 0.3, Min: 0.3, Max: 0.3, Diff: 0.0] These are the times that each worker thread spent in performing some other tasks that we have not accounted above for the total Parallel Time. [Clear CT: 0.1 ms] This is the time spent in clearing the Card Table. This task is performed in serial mode. [Other: 1.5 ms] Time spent in the some other tasks listed below. The following sub-tasks (which individually may be parallelized) are performed serially. [Choose CSet: 0.0 ms] Time spent in selecting the regions for the Collection Set. [Ref Proc: 0.3 ms] Total time spent in processing Reference objects. [Ref Enq: 0.0 ms] Time spent in enqueuing references to the ReferenceQueues. [Free CSet: 0.3 ms] Time spent in freeing the collection set data structure. [Eden: 12M(12M)->0B(13M) Survivors: 0B->2048K Heap: 14M(64M)->9739K(64M)] This line gives the details on the heap size changes with the Evacuation Pause. This shows that Eden had the occupancy of 12M and its capacity was also 12M before the collection. After the collection, its occupancy got reduced to 0 since everything is evacuated/promoted from Eden during a collection, and its target size grew to 13M. The new Eden capacity of 13M is not reserved at this point. This value is the target size of the Eden. Regions are added to Eden as the demand is made and when the added regions reach to the target size, we start the next collection. Similarly, Survivors had the occupancy of 0 bytes and it grew to 2048K after the collection. The total heap occupancy and capacity was 14M and 64M receptively before the collection and it became 9739K and 64M after the collection. Apart from the evacuation pauses, G1 also performs concurrent-marking to build the live data information of regions. 1.416: [GC pause (young) (initial-mark), 0.62417980 secs] ….... 2.042: [GC concurrent-root-region-scan-start] 2.067: [GC concurrent-root-region-scan-end, 0.0251507] 2.068: [GC concurrent-mark-start] 3.198: [GC concurrent-mark-reset-for-overflow] 4.053: [GC concurrent-mark-end, 1.9849672 sec] 4.055: [GC remark 4.055: [GC ref-proc, 0.0000254 secs], 0.0030184 secs] [Times: user=0.00 sys=0.00, real=0.00 secs] 4.088: [GC cleanup 117M->106M(138M), 0.0015198 secs] [Times: user=0.00 sys=0.00, real=0.00 secs] 4.090: [GC concurrent-cleanup-start] 4.091: [GC concurrent-cleanup-end, 0.0002721] The first phase of a marking cycle is Initial Marking where all the objects directly reachable from the roots are marked and this phase is piggy-backed on a fully young Evacuation Pause. 2.042: [GC concurrent-root-region-scan-start] This marks the start of a concurrent phase that scans the set of root-regions which are directly reachable from the survivors of the initial marking phase. 2.067: [GC concurrent-root-region-scan-end, 0.0251507] End of the concurrent root region scan phase and it lasted for 0.0251507 seconds. 2.068: [GC concurrent-mark-start] Start of the concurrent marking at 2.068 secs from the start of the process. 3.198: [GC concurrent-mark-reset-for-overflow] This indicates that the global marking stack had became full and there was an overflow of the stack. Concurrent marking detected this overflow and had to reset the data structures to start the marking again. 4.053: [GC concurrent-mark-end, 1.9849672 sec] End of the concurrent marking phase and it lasted for 1.9849672 seconds. 4.055: [GC remark 4.055: [GC ref-proc, 0.0000254 secs], 0.0030184 secs] This corresponds to the remark phase which is a stop-the-world phase. It completes the left over marking work (SATB buffers processing) from the previous phase. In this case, this phase took 0.0030184 secs and out of which 0.0000254 secs were spent on Reference processing. 4.088: [GC cleanup 117M->106M(138M), 0.0015198 secs] Cleanup phase which is again a stop-the-world phase. It goes through the marking information of all the regions, computes the live data information of each region, resets the marking data structures and sorts the regions according to their gc-efficiency. In this example, the total heap size is 138M and after the live data counting it was found that the total live data size dropped down from 117M to 106M. 4.090: [GC concurrent-cleanup-start] This concurrent cleanup phase frees up the regions that were found to be empty (didn't contain any live data) during the previous stop-the-world phase. 4.091: [GC concurrent-cleanup-end, 0.0002721] Concurrent cleanup phase took 0.0002721 secs to free up the empty regions. Option -XX:G1PrintRegionLivenessInfo Now, let's look at the output generated with the flag G1PrintRegionLivenessInfo. This is a diagnostic option and gets enabled with -XX:+UnlockDiagnosticVMOptions. G1PrintRegionLivenessInfo prints the live data information of each region during the Cleanup phase of the concurrent-marking cycle. 26.896: [GC cleanup ### PHASE Post-Marking @ 26.896### HEAP committed: 0x02e00000-0x0fe00000 reserved: 0x02e00000-0x12e00000 region-size: 1048576 Cleanup phase of the concurrent-marking cycle started at 26.896 secs from the start of the process and this live data information is being printed after the marking phase. Committed G1 heap ranges from 0x02e00000 to 0x0fe00000 and the total G1 heap reserved by JVM is from 0x02e00000 to 0x12e00000. Each region in the G1 heap is of size 1048576 bytes. ### type address-range used prev-live next-live gc-eff### (bytes) (bytes) (bytes) (bytes/ms) This is the header of the output that tells us about the type of the region, address-range of the region, used space in the region, live bytes in the region with respect to the previous marking cycle, live bytes in the region with respect to the current marking cycle and the GC efficiency of that region. ### FREE 0x02e00000-0x02f00000 0 0 0 0.0 This is a Free region. ### OLD 0x02f00000-0x03000000 1048576 1038592 1038592 0.0 Old region with address-range from 0x02f00000 to 0x03000000. Total used space in the region is 1048576 bytes, live bytes as per the previous marking cycle are 1038592 and live bytes with respect to the current marking cycle are also 1038592. The GC efficiency has been computed as 0. ### EDEN 0x03400000-0x03500000 20992 20992 20992 0.0 This is an Eden region. ### HUMS 0x0ae00000-0x0af00000 1048576 1048576 1048576 0.0### HUMC 0x0af00000-0x0b000000 1048576 1048576 1048576 0.0### HUMC 0x0b000000-0x0b100000 1048576 1048576 1048576 0.0### HUMC 0x0b100000-0x0b200000 1048576 1048576 1048576 0.0### HUMC 0x0b200000-0x0b300000 1048576 1048576 1048576 0.0### HUMC 0x0b300000-0x0b400000 1048576 1048576 1048576 0.0### HUMC 0x0b400000-0x0b500000 1001480 1001480 1001480 0.0 These are the continuous set of regions called Humongous regions for storing a large object. HUMS (Humongous starts) marks the start of the set of humongous regions and HUMC (Humongous continues) tags the subsequent regions of the humongous regions set. ### SURV 0x09300000-0x09400000 16384 16384 16384 0.0 This is a Survivor region. ### SUMMARY capacity: 208.00 MB used: 150.16 MB / 72.19 % prev-live: 149.78 MB / 72.01 % next-live: 142.82 MB / 68.66 % At the end, a summary is printed listing the capacity, the used space and the change in the liveness after the completion of concurrent marking. In this case, G1 heap capacity is 208MB, total used space is 150.16MB which is 72.19% of the total heap size, live data in the previous marking was 149.78MB which was 72.01% of the total heap size and the live data as per the current marking is 142.82MB which is 68.66% of the total heap size. Option -XX:+G1PrintHeapRegions G1PrintHeapRegions option logs the regions related events when regions are committed, allocated into or are reclaimed. COMMIT/UNCOMMIT events G1HR COMMIT [0x6e900000,0x6ea00000]G1HR COMMIT [0x6ea00000,0x6eb00000] Here, the heap is being initialized or expanded and the region (with bottom: 0x6eb00000 and end: 0x6ec00000) is being freshly committed. COMMIT events are always generated in order i.e. the next COMMIT event will always be for the uncommitted region with the lowest address. G1HR UNCOMMIT [0x72700000,0x72800000]G1HR UNCOMMIT [0x72600000,0x72700000] Opposite to COMMIT. The heap got shrunk at the end of a Full GC and the regions are being uncommitted. Like COMMIT, UNCOMMIT events are also generated in order i.e. the next UNCOMMIT event will always be for the committed region with the highest address. GC Cycle events G1HR #StartGC 7G1HR CSET 0x6e900000G1HR REUSE 0x70500000G1HR ALLOC(Old) 0x6f800000G1HR RETIRE 0x6f800000 0x6f821b20G1HR #EndGC 7 This shows start and end of an Evacuation pause. This event is followed by a GC counter tracking both evacuation pauses and Full GCs. Here, this is the 7th GC since the start of the process. G1HR #StartFullGC 17G1HR UNCOMMIT [0x6ed00000,0x6ee00000]G1HR POST-COMPACTION(Old) 0x6e800000 0x6e854f58G1HR #EndFullGC 17 Shows start and end of a Full GC. This event is also followed by the same GC counter as above. This is the 17th GC since the start of the process. ALLOC events G1HR ALLOC(Eden) 0x6e800000 The region with bottom 0x6e800000 just started being used for allocation. In this case it is an Eden region and allocated into by a mutator thread. G1HR ALLOC(StartsH) 0x6ec00000 0x6ed00000G1HR ALLOC(ContinuesH) 0x6ed00000 0x6e000000 Regions being used for the allocation of Humongous object. The object spans over two regions. G1HR ALLOC(SingleH) 0x6f900000 0x6f9eb010 Single region being used for the allocation of Humongous object. G1HR COMMIT [0x6ee00000,0x6ef00000]G1HR COMMIT [0x6ef00000,0x6f000000]G1HR COMMIT [0x6f000000,0x6f100000]G1HR COMMIT [0x6f100000,0x6f200000]G1HR ALLOC(StartsH) 0x6ee00000 0x6ef00000G1HR ALLOC(ContinuesH) 0x6ef00000 0x6f000000G1HR ALLOC(ContinuesH) 0x6f000000 0x6f100000G1HR ALLOC(ContinuesH) 0x6f100000 0x6f102010 Here, Humongous object allocation request could not be satisfied by the free committed regions that existed in the heap, so the heap needed to be expanded. Thus new regions are committed and then allocated into for the Humongous object. G1HR ALLOC(Old) 0x6f800000 Old region started being used for allocation during GC. G1HR ALLOC(Survivor) 0x6fa00000 Region being used for copying old objects into during a GC. Note that Eden and Humongous ALLOC events are generated outside the GC boundaries and Old and Survivor ALLOC events are generated inside the GC boundaries. Other Events G1HR RETIRE 0x6e800000 0x6e87bd98 Retire and stop using the region having bottom 0x6e800000 and top 0x6e87bd98 for allocation. Note that most regions are full when they are retired and we omit those events to reduce the output volume. A region is retired when another region of the same type is allocated or we reach the start or end of a GC(depending on the region). So for Eden regions: For example: 1. ALLOC(Eden) Foo2. ALLOC(Eden) Bar3. StartGC At point 2, Foo has just been retired and it was full. At point 3, Bar was retired and it was full. If they were not full when they were retired, we will have a RETIRE event: 1. ALLOC(Eden) Foo2. RETIRE Foo top3. ALLOC(Eden) Bar4. StartGC G1HR CSET 0x6e900000 Region (bottom: 0x6e900000) is selected for the Collection Set. The region might have been selected for the collection set earlier (i.e. when it was allocated). However, we generate the CSET events for all regions in the CSet at the start of a GC to make sure there's no confusion about which regions are part of the CSet. G1HR POST-COMPACTION(Old) 0x6e800000 0x6e839858 POST-COMPACTION event is generated for each non-empty region in the heap after a full compaction. A full compaction moves objects around, so we don't know what the resulting shape of the heap is (which regions were written to, which were emptied, etc.). To deal with this, we generate a POST-COMPACTION event for each non-empty region with its type (old/humongous) and the heap boundaries. At this point we should only have Old and Humongous regions, as we have collapsed the young generation, so we should not have eden and survivors. POST-COMPACTION events are generated within the Full GC boundary. G1HR CLEANUP 0x6f400000G1HR CLEANUP 0x6f300000G1HR CLEANUP 0x6f200000 These regions were found empty after remark phase of Concurrent Marking and are reclaimed shortly afterwards. G1HR #StartGC 5G1HR CSET 0x6f400000G1HR CSET 0x6e900000G1HR REUSE 0x6f800000 At the end of a GC we retire the old region we are allocating into. Given that its not full, we will carry on allocating into it during the next GC. This is what REUSE means. In the above case 0x6f800000 should have been the last region with an ALLOC(Old) event during the previous GC and should have been retired before the end of the previous GC. G1HR ALLOC-FORCE(Eden) 0x6f800000 A specialization of ALLOC which indicates that we have reached the max desired number of the particular region type (in this case: Eden), but we decided to allocate one more. Currently it's only used for Eden regions when we extend the young generation because we cannot do a GC as the GC-Locker is active. G1HR EVAC-FAILURE 0x6f800000 During a GC, we have failed to evacuate an object from the given region as the heap is full and there is no space left to copy the object. This event is generated within GC boundaries and exactly once for each region from which we failed to evacuate objects. When Heap Regions are reclaimed ? It is also worth mentioning when the heap regions in the G1 heap are reclaimed. All regions that are in the CSet (the ones that appear in CSET events) are reclaimed at the end of a GC. The exception to that are regions with EVAC-FAILURE events. All regions with CLEANUP events are reclaimed. After a Full GC some regions get reclaimed (the ones from which we moved the objects out). But that is not shown explicitly, instead the non-empty regions that are left in the heap are printed out with the POST-COMPACTION events.

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JMS Step 5 - How to Create an 11g BPEL Process Which Reads a Message Based on an XML Schema from a JMS Queue

- by John-Brown.Evans

JMS Step 5 - How to Create an 11g BPEL Process Which Reads a Message Based on an XML Schema from a JMS Queue .jblist{list-style-type:disc;margin:0;padding:0;padding-left:0pt;margin-left:36pt} ol{margin:0;padding:0} .c12_5{vertical-align:top;width:468pt;border-style:solid;background-color:#f3f3f3;border-color:#000000;border-width:1pt;padding:5pt 5pt 5pt 5pt} .c8_5{vertical-align:top;border-style:solid;border-color:#000000;border-width:1pt;padding:5pt 5pt 0pt 5pt} .c10_5{vertical-align:top;width:207pt;border-style:solid;border-color:#000000;border-width:1pt;padding:5pt 5pt 5pt 5pt} .c14_5{vertical-align:top;border-style:solid;border-color:#000000;border-width:1pt;padding:0pt 5pt 0pt 5pt} .c21_5{background-color:#ffffff} .c18_5{color:#1155cc;text-decoration:underline} .c16_5{color:#666666;font-size:12pt} .c5_5{background-color:#f3f3f3;font-weight:bold} .c19_5{color:inherit;text-decoration:inherit} .c3_5{height:11pt;text-align:center} .c11_5{font-weight:bold} .c20_5{background-color:#00ff00} .c6_5{font-style:italic} .c4_5{height:11pt} .c17_5{background-color:#ffff00} .c0_5{direction:ltr} .c7_5{font-family:"Courier New"} .c2_5{border-collapse:collapse} .c1_5{line-height:1.0} .c13_5{background-color:#f3f3f3} .c15_5{height:0pt} .c9_5{text-align:center} .title{padding-top:24pt;line-height:1.15;text-align:left;color:#000000;font-size:36pt;font-family:"Arial";font-weight:bold;padding-bottom:6pt} .subtitle{padding-top:18pt;line-height:1.15;text-align:left;color:#666666;font-style:italic;font-size:24pt;font-family:"Georgia";padding-bottom:4pt} li{color:#000000;font-size:10pt;font-family:"Arial"} p{color:#000000;font-size:10pt;margin:0;font-family:"Arial"} h1{padding-top:0pt;line-height:1.15;text-align:left;color:#888;font-size:24pt;font-family:"Arial";font-weight:normal} h2{padding-top:0pt;line-height:1.15;text-align:left;color:#888;font-size:18pt;font-family:"Arial";font-weight:normal} h3{padding-top:0pt;line-height:1.15;text-align:left;color:#888;font-size:14pt;font-family:"Arial";font-weight:normal} h4{padding-top:0pt;line-height:1.15;text-align:left;color:#888;font-size:12pt;font-family:"Arial";font-weight:normal} h5{padding-top:0pt;line-height:1.15;text-align:left;color:#888;font-size:11pt;font-family:"Arial";font-weight:normal} h6{padding-top:0pt;line-height:1.15;text-align:left;color:#888;font-size:10pt;font-family:"Arial";font-weight:normal} Welcome to another post in the series of blogs which demonstrates how to use JMS queues in a SOA context. The previous posts were: JMS Step 1 - How to Create a Simple JMS Queue in Weblogic Server 11g JMS Step 2 - Using the QueueSend.java Sample Program to Send a Message to a JMS Queue JMS Step 3 - Using the QueueReceive.java Sample Program to Read a Message from a JMS Queue JMS Step 4 - How to Create an 11g BPEL Process Which Writes a Message Based on an XML Schema to a JMS Queue Today we will create a BPEL process which will read (dequeue) the message from the JMS queue, which we enqueued in the last example. The JMS adapter will dequeue the full XML payload from the queue. 1. Recap and Prerequisites In the previous examples, we created a JMS Queue, a Connection Factory and a Connection Pool in the WebLogic Server Console. Then we designed and deployed a BPEL composite, which took a simple XML payload and enqueued it to the JMS queue. In this example, we will read that same message from the queue, using a JMS adapter and a BPEL process. As many of the configuration steps required to read from that queue were done in the previous samples, this one will concentrate on the new steps. A summary of the required objects is listed below. To find out how to create them please see the previous samples. They also include instructions on how to verify the objects are set up correctly. WebLogic Server Objects Object Name Type JNDI Name TestConnectionFactory Connection Factory jms/TestConnectionFactory TestJMSQueue JMS Queue jms/TestJMSQueue eis/wls/TestQueue Connection Pool eis/wls/TestQueue Schema XSD File The following XSD file is used for the message format. It was created in the previous example and will be copied to the new process. stringPayload.xsd <?xml version="1.0" encoding="windows-1252" ?> <xsd:schema xmlns:xsd="http://www.w3.org/2001/XMLSchema" xmlns="http://www.example.org" targetNamespace="http://www.example.org" elementFormDefault="qualified"> <xsd:element name="exampleElement" type="xsd:string"> </xsd:element> </xsd:schema> JMS Message After executing the previous samples, the following XML message should be in the JMS queue located at jms/TestJMSQueue: <?xml version="1.0" encoding="UTF-8" ?><exampleElement xmlns="http://www.example.org">Test Message</exampleElement> JDeveloper Connection You will need a valid Application Server Connection in JDeveloper pointing to the SOA server which the process will be deployed to. 2. Create a BPEL Composite with a JMS Adapter Partner Link In the previous example, we created a composite in JDeveloper called JmsAdapterWriteSchema. In this one, we will create a new composite called JmsAdapterReadSchema. There are probably many ways of incorporating a JMS adapter into a SOA composite for incoming messages. One way is design the process in such a way that the adapter polls for new messages and when it dequeues one, initiates a SOA or BPEL instance. This is possibly the most common use case. Other use cases include mid-flow adapters, which are activated from within the BPEL process. In this example we will use a polling adapter, because it is the most simple to set up and demonstrate. But it has one disadvantage as a demonstrative model. When a polling adapter is active, it will dequeue all messages as soon as they reach the queue. This makes it difficult to monitor messages we are writing to the queue, because they will disappear from the queue as soon as they have been enqueued. To work around this, we will shut down the composite after deploying it and restart it as required. (Another solution for this would be to pause the consumption for the queue and resume consumption again if needed. This can be done in the WLS console JMS-Modules -> queue -> Control -> Consumption -> Pause/Resume.) We will model the composite as a one-way incoming process. Usually, a BPEL process will do something useful with the message after receiving it, such as passing it to a database or file adapter, a human workflow or external web service. But we only want to demonstrate how to dequeue a JMS message using BPEL and a JMS adapter, so we won’t complicate the design with further activities. However, we do want to be able to verify that we have read the message correctly, so the BPEL process will include a small piece of embedded java code, which will print the message to standard output, so we can view it in the SOA server’s log file. Alternatively, you can view the instance in the Enterprise Manager and verify the message. The following steps are all executed in JDeveloper. Create the project in the same JDeveloper application used for the previous examples or create a new one. Create a SOA Project Create a new project and choose SOA Tier > SOA Project as its type. Name it JmsAdapterReadSchema. When prompted for the composite type, choose Empty Composite. Create a JMS Adapter Partner Link In the composite editor, drag a JMS adapter over from the Component Palette to the left-hand swim lane, under Exposed Services. This will start the JMS Adapter Configuration Wizard. Use the following entries: Service Name: JmsAdapterRead Oracle Enterprise Messaging Service (OEMS): Oracle WebLogic JMS AppServer Connection: Use an application server connection pointing to the WebLogic server on which the JMS queue and connection factory mentioned under Prerequisites above are located. Adapter Interface > Interface: Define from operation and schema (specified later) Operation Type: Consume Message Operation Name: Consume_message Consume Operation Parameters Destination Name: Press the Browse button, select Destination Type: Queues, then press Search. Wait for the list to populate, then select the entry for TestJMSQueue , which is the queue created in a previous example. JNDI Name: The JNDI name to use for the JMS connection. As in the previous example, this is probably the most common source of error. This is the JNDI name of the JMS adapter’s connection pool created in the WebLogic Server and which points to the connection factory. JDeveloper does not verify the value entered here. If you enter a wrong value, the JMS adapter won’t find the queue and you will get an error message at runtime, which is very difficult to trace. In our example, this is the value eis/wls/TestQueue . (See the earlier step on how to create a JMS Adapter Connection Pool in WebLogic Server for details.) Messages/Message SchemaURL: We will use the XSD file created during the previous example, in the JmsAdapterWriteSchema project to define the format for the incoming message payload and, at the same time, demonstrate how to import an existing XSD file into a JDeveloper project. Press the magnifying glass icon to search for schema files. In the Type Chooser, press the Import Schema File button. Select the magnifying glass next to URL to search for schema files. Navigate to the location of the JmsAdapterWriteSchema project > xsd and select the stringPayload.xsd file. Check the “Copy to Project” checkbox, press OK and confirm the following Localize Files popup. Now that the XSD file has been copied to the local project, it can be selected from the project’s schema files. Expand Project Schema Files > stringPayload.xsd and select exampleElement: string . Press Next and Finish, which will complete the JMS Adapter configuration.Save the project. Create a BPEL Component Drag a BPEL Process from the Component Palette (Service Components) to the Components section of the composite designer. Name it JmsAdapterReadSchema and select Template: Define Service Later and press OK. Wire the JMS Adapter to the BPEL Component Now wire the JMS adapter to the BPEL process, by dragging the arrow from the adapter to the BPEL process. A Transaction Properties popup will be displayed. Set the delivery mode to async.persist. This completes the steps at the composite level. 3 . Complete the BPEL Process Design Invoke the BPEL Flow via the JMS Adapter Open the BPEL component by double-clicking it in the design view of the composite.xml, or open it from the project navigator by selecting the JmsAdapterReadSchema.bpel file. This will display the BPEL process in the design view. You should see the JmsAdapterRead partner link in the left-hand swim lane. Drag a Receive activity onto the BPEL flow diagram, then drag a wire (left-hand yellow arrow) from it to the JMS adapter. This will open the Receive activity editor. Auto-generate the variable by pressing the green “+” button and check the “Create Instance” checkbox. This will result in a BPEL instance being created when a new JMS message is received. At this point it would actually be OK to compile and deploy the composite and it would pick up any messages from the JMS queue. In fact, you can do that to test it, if you like. But it is very rudimentary and would not be doing anything useful with the message. Also, you could only verify the actual message payload by looking at the instance’s flow in the Enterprise Manager. There are various other possibilities; we could pass the message to another web service, write it to a file using a file adapter or to a database via a database adapter etc. But these will all introduce unnecessary complications to our sample. So, to keep it simple, we will add a small piece of Java code to the BPEL process which will write the payload to standard output. This will be written to the server’s log file, which will be easy to monitor. Add a Java Embedding Activity First get the full name of the process’s input variable, as this will be needed for the Java code. Go to the Structure pane and expand Variables > Process > Variables. Then expand the input variable, for example, "Receive1_Consume_Message_InputVariable > body > ns2:exampleElement”, and note variable’s name and path, if they are different from this one. Drag a Java Embedding activity from the Component Palette (Oracle Extensions) to the BPEL flow, after the Receive activity, then open it to edit. Delete the example code and replace it with the following, replacing the variable parts with those in your sample, if necessary.: System.out.println("JmsAdapterReadSchema process picked up a message"); oracle.xml.parser.v2.XMLElement inputPayload = (oracle.xml.parser.v2.XMLElement)getVariableData( "Receive1_Consume_Message_InputVariable", "body", "/ns2:exampleElement"); String inputString = inputPayload.getFirstChild().getNodeValue(); System.out.println("Input String is " + inputPayload.getFirstChild().getNodeValue()); Tip. If you are not sure of the exact syntax of the input variable, create an Assign activity in the BPEL process and copy the variable to another, temporary one. Then check the syntax created by the BPEL designer. This completes the BPEL process design in JDeveloper. Save, compile and deploy the process to the SOA server. 3. Test the Composite Shut Down the JmsAdapterReadSchema Composite After deploying the JmsAdapterReadSchema composite to the SOA server it is automatically activated. If there are already any messages in the queue, the adapter will begin polling them. To ease the testing process, we will deactivate the process first Log in to the Enterprise Manager (Fusion Middleware Control) and navigate to SOA > soa-infra (soa_server1) > default (or wherever you deployed your composite to) and click on JmsAdapterReadSchema [1.0] . Press the Shut Down button to disable the composite and confirm the following popup. Monitor Messages in the JMS Queue In a separate browser window, log in to the WebLogic Server Console and navigate to Services > Messaging > JMS Modules > TestJMSModule > TestJMSQueue > Monitoring. This is the location of the JMS queue we created in an earlier sample (see the prerequisites section of this sample). Check whether there are any messages already in the queue. If so, you can dequeue them using the QueueReceive Java program created in an earlier sample. This will ensure that the queue is empty and doesn’t contain any messages in the wrong format, which would cause the JmsAdapterReadSchema to fail. Send a Test Message In the Enterprise Manager, navigate to the JmsAdapterWriteSchema created earlier, press Test and send a test message, for example “Message from JmsAdapterWriteSchema”. Confirm that the message was written correctly to the queue by verifying it via the queue monitor in the WLS Console. Monitor the SOA Server’s Output A program deployed on the SOA server will write its standard output to the terminal window in which the server was started, unless this has been redirected to somewhere else, for example to a file. If it has not been redirected, go to the terminal session in which the server was started, otherwise open and monitor the file to which it was redirected. Re-Enable the JmsAdapterReadSchema Composite In the Enterprise Manager, navigate to the JmsAdapterReadSchema composite again and press Start Up to re-enable it. This should cause the JMS adapter to dequeue the test message and the following output should be written to the server’s standard output: JmsAdapterReadSchema process picked up a message. Input String is Message from JmsAdapterWriteSchema Note that you can also monitor the payload received by the process, by navigating to the the JmsAdapterReadSchema’s Instances tab in the Enterprise Manager. Then select the latest instance and view the flow of the BPEL component. The Receive activity will contain and display the dequeued message too. 4 . Troubleshooting This sample demonstrates how to dequeue an XML JMS message using a BPEL process and no additional functionality. For example, it doesn’t contain any error handling. Therefore, any errors in the payload will result in exceptions being written to the log file or standard output. If you get any errors related to the payload, such as Message handle error ... ORABPEL-09500 ... XPath expression failed to execute. An error occurs while processing the XPath expression; the expression is /ns2:exampleElement. ... etc. check that the variable used in the Java embedding part of the process was entered correctly. Possibly follow the tip mentioned in previous section. If this doesn’t help, you can delete the Java embedding part and simply verify the message via the flow diagram in the Enterprise Manager. Or use a different method, such as writing it to a file via a file adapter. This concludes this example. In the next post, we will begin with an AQ JMS example, which uses JMS to write to an Advanced Queue stored in the database. Best regards John-Brown Evans Oracle Technology Proactive Support Delivery

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Using the ASP.NET Cache to cache data in a Model or Business Object layer, without a dependency on System.Web in the layer - Part One.

- by Rhames

ASP.NET applications can make use of the System.Web.Caching.Cache object to cache data and prevent repeated expensive calls to a database or other store. However, ideally an application should make use of caching at the point where data is retrieved from the database, which typically is inside a Business Objects or Model layer. One of the key features of using a UI pattern such as Model-View-Presenter (MVP) or Model-View-Controller (MVC) is that the Model and Presenter (or Controller) layers are developed without any knowledge of the UI layer. Introducing a dependency on System.Web into the Model layer would break this independence of the Model from the View. This article gives a solution to this problem, using dependency injection to inject the caching implementation into the Model layer at runtime. This allows caching to be used within the Model layer, without any knowledge of the actual caching mechanism that will be used. Create a sample application to use the caching solution Create a test SQL Server database This solution uses a SQL Server database with the same Sales data used in my previous post on calculating running totals. The advantage of using this data is that it gives nice slow queries that will exaggerate the effect of using caching! To create the data, first create a new SQL database called CacheSample. Next run the following script to create the Sale table and populate it: USE CacheSample GO CREATE TABLE Sale(DayCount smallint, Sales money) CREATE CLUSTERED INDEX ndx_DayCount ON Sale(DayCount) go INSERT Sale VALUES (1,120) INSERT Sale VALUES (2,60) INSERT Sale VALUES (3,125) INSERT Sale VALUES (4,40) DECLARE @DayCount smallint, @Sales money SET @DayCount = 5 SET @Sales = 10 WHILE @DayCount < 5000 BEGIN INSERT Sale VALUES (@DayCount,@Sales) SET @DayCount = @DayCount + 1 SET @Sales = @Sales + 15 END Next create a stored procedure to calculate the running total, and return a specified number of rows from the Sale table, using the following script: USE [CacheSample] GO SET ANSI_NULLS ON GO SET QUOTED_IDENTIFIER ON GO -- ============================================= -- Author: Robin -- Create date: -- Description: -- ============================================= CREATE PROCEDURE [dbo].[spGetRunningTotals] -- Add the parameters for the stored procedure here @HighestDayCount smallint = null AS BEGIN -- SET NOCOUNT ON added to prevent extra result sets from -- interfering with SELECT statements. SET NOCOUNT ON; IF @HighestDayCount IS NULL SELECT @HighestDayCount = MAX(DayCount) FROM dbo.Sale DECLARE @SaleTbl TABLE (DayCount smallint, Sales money, RunningTotal money) DECLARE @DayCount smallint, @Sales money, @RunningTotal money SET @RunningTotal = 0 SET @DayCount = 0 DECLARE rt_cursor CURSOR FOR SELECT DayCount, Sales FROM Sale ORDER BY DayCount OPEN rt_cursor FETCH NEXT FROM rt_cursor INTO @DayCount,@Sales WHILE @@FETCH_STATUS = 0 AND @DayCount <= @HighestDayCount BEGIN SET @RunningTotal = @RunningTotal + @Sales INSERT @SaleTbl VALUES (@DayCount,@Sales,@RunningTotal) FETCH NEXT FROM rt_cursor INTO @DayCount,@Sales END CLOSE rt_cursor DEALLOCATE rt_cursor SELECT DayCount, Sales, RunningTotal FROM @SaleTbl END GO Create the Sample ASP.NET application In Visual Studio create a new solution and add a class library project called CacheSample.BusinessObjects and an ASP.NET web application called CacheSample.UI. The CacheSample.BusinessObjects project will contain a single class to represent a Sale data item, with all the code to retrieve the sales from the database included in it for simplicity (normally I would at least have a separate Repository or other object that is responsible for retrieving data, and probably a data access layer as well, but for this sample I want to keep it simple). The C# code for the Sale class is shown below: using System; using System.Collections.Generic; using System.Data; using System.Data.SqlClient; namespace CacheSample.BusinessObjects { public class Sale { public Int16 DayCount { get; set; } public decimal Sales { get; set; } public decimal RunningTotal { get; set; } public static IEnumerable<Sale> GetSales(int? highestDayCount) { List<Sale> sales = new List<Sale>(); SqlParameter highestDayCountParameter = new SqlParameter("@HighestDayCount", SqlDbType.SmallInt); if (highestDayCount.HasValue) highestDayCountParameter.Value = highestDayCount; else highestDayCountParameter.Value = DBNull.Value; string connectionStr = System.Configuration.ConfigurationManager .ConnectionStrings["CacheSample"].ConnectionString; using(SqlConnection sqlConn = new SqlConnection(connectionStr)) using (SqlCommand sqlCmd = sqlConn.CreateCommand()) { sqlCmd.CommandText = "spGetRunningTotals"; sqlCmd.CommandType = CommandType.StoredProcedure; sqlCmd.Parameters.Add(highestDayCountParameter); sqlConn.Open(); using (SqlDataReader dr = sqlCmd.ExecuteReader()) { while (dr.Read()) { Sale newSale = new Sale(); newSale.DayCount = dr.GetInt16(0); newSale.Sales = dr.GetDecimal(1); newSale.RunningTotal = dr.GetDecimal(2); sales.Add(newSale); } } } return sales; } } } The static GetSale() method makes a call to the spGetRunningTotals stored procedure and then reads each row from the returned SqlDataReader into an instance of the Sale class, it then returns a List of the Sale objects, as IEnnumerable<Sale>. A reference to System.Configuration needs to be added to the CacheSample.BusinessObjects project so that the connection string can be read from the web.config file. In the CacheSample.UI ASP.NET project, create a single web page called ShowSales.aspx, and make this the default start up page. This page will contain a single button to call the GetSales() method and a label to display the results. The html mark up and the C# code behind are shown below: ShowSales.aspx <%@ Page Language="C#" AutoEventWireup="true" CodeBehind="ShowSales.aspx.cs" Inherits="CacheSample.UI.ShowSales" %> <!DOCTYPE html PUBLIC "-//W3C//DTD XHTML 1.0 Transitional//EN" "http://www.w3.org/TR/xhtml1/DTD/xhtml1-transitional.dtd"> <html xmlns="http://www.w3.org/1999/xhtml"> <head runat="server"> <title>Cache Sample - Show All Sales</title> </head> <body> <form id="form1" runat="server"> <div> <asp:Button ID="btnTest1" runat="server" onclick="btnTest1_Click" Text="Get All Sales" />     <asp:Label ID="lblResults" runat="server"></asp:Label> </div> </form> </body> </html> ShowSales.aspx.cs using System; using System.Collections.Generic; using System.Linq; using System.Web; using System.Web.UI; using System.Web.UI.WebControls; using CacheSample.BusinessObjects; namespace CacheSample.UI { public partial class ShowSales : System.Web.UI.Page { protected void Page_Load(object sender, EventArgs e) { } protected void btnTest1_Click(object sender, EventArgs e) { System.Diagnostics.Stopwatch stopWatch = new System.Diagnostics.Stopwatch(); stopWatch.Start(); var sales = Sale.GetSales(null); var lastSales = sales.Last(); stopWatch.Stop(); lblResults.Text = string.Format( "Count of Sales: {0}, Last DayCount: {1}, Total Sales: {2}. Query took {3} ms", sales.Count(), lastSales.DayCount, lastSales.RunningTotal, stopWatch.ElapsedMilliseconds); } } } Finally we need to add a connection string to the CacheSample SQL Server database, called CacheSample, to the web.config file: <?xmlversion="1.0"?> <configuration> <connectionStrings> <addname="CacheSample" connectionString="data source=.\SQLEXPRESS;Integrated Security=SSPI;Initial Catalog=CacheSample" providerName="System.Data.SqlClient" /> </connectionStrings> <system.web> <compilationdebug="true"targetFramework="4.0" /> </system.web> </configuration> Run the application and click the button a few times to see how long each call to the database takes. On my system, each query takes about 450ms. Next I shall look at a solution to use the ASP.NET caching to cache the data returned by the query, so that subsequent requests to the GetSales() method are much faster. Adding Data Caching Support I am going to create my caching support in a separate project called CacheSample.Caching, so the next step is to add a class library to the solution. We shall be using the application configuration to define the implementation of our caching system, so we need a reference to System.Configuration adding to the project. ICacheProvider<T> Interface The first step in adding caching to our application is to define an interface, called ICacheProvider, in the CacheSample.Caching project, with methods to retrieve any data from the cache or to retrieve the data from the data source if it is not present in the cache. Dependency Injection will then be used to inject an implementation of this interface at runtime, allowing the users of the interface (i.e. the CacheSample.BusinessObjects project) to be completely unaware of how the caching is actually implemented. As data of any type maybe retrieved from the data source, it makes sense to use generics in the interface, with a generic type parameter defining the data type associated with a particular instance of the cache interface implementation. The C# code for the ICacheProvider interface is shown below: using System; using System.Collections.Generic; namespace CacheSample.Caching { public interface ICacheProvider { } public interface ICacheProvider<T> : ICacheProvider { T Fetch(string key, Func<T> retrieveData, DateTime? absoluteExpiry, TimeSpan? relativeExpiry); IEnumerable<T> Fetch(string key, Func<IEnumerable<T>> retrieveData, DateTime? absoluteExpiry, TimeSpan? relativeExpiry); } } The empty non-generic interface will be used as a type in a Dictionary generic collection later to store instances of the ICacheProvider<T> implementation for reuse, I prefer to use a base interface when doing this, as I think the alternative of using object makes for less clear code. The ICacheProvider<T> interface defines two overloaded Fetch methods, the difference between these is that one will return a single instance of the type T and the other will return an IEnumerable<T>, providing support for easy caching of collections of data items. Both methods will take a key parameter, which will uniquely identify the cached data, a delegate of type Func<T> or Func<IEnumerable<T>> which will provide the code to retrieve the data from the store if it is not present in the cache, and absolute or relative expiry policies to define when a cached item should expire. Note that at present there is no support for cache dependencies, but I shall be showing a method of adding this in part two of this article. CacheProviderFactory Class We need a mechanism of creating instances of our ICacheProvider<T> interface, using Dependency Injection to get the implementation of the interface. To do this we shall create a CacheProviderFactory static class in the CacheSample.Caching project. This factory will provide a generic static method called GetCacheProvider<T>(), which shall return instances of ICacheProvider<T>. We can then call this factory method with the relevant data type (for example the Sale class in the CacheSample.BusinessObject project) to get a instance of ICacheProvider for that type (e.g. call CacheProviderFactory.GetCacheProvider<Sale>() to get the ICacheProvider<Sale> implementation). The C# code for the CacheProviderFactory is shown below: using System; using System.Collections.Generic; using CacheSample.Caching.Configuration; namespace CacheSample.Caching { public static class CacheProviderFactory { private static Dictionary<Type, ICacheProvider> cacheProviders = new Dictionary<Type, ICacheProvider>(); private static object syncRoot = new object(); ///<summary> /// Factory method to create or retrieve an implementation of the /// ICacheProvider interface for type <typeparamref name="T"/>. ///</summary> ///<typeparam name="T"> /// The type that this cache provider instance will work with ///</typeparam> ///<returns>An instance of the implementation of ICacheProvider for type ///<typeparamref name="T"/>, as specified by the application /// configuration</returns> public static ICacheProvider<T> GetCacheProvider<T>() { ICacheProvider<T> cacheProvider = null; // Get the Type reference for the type parameter T Type typeOfT = typeof(T); // Lock the access to the cacheProviders dictionary // so multiple threads can work with it lock (syncRoot) { // First check if an instance of the ICacheProvider implementation // already exists in the cacheProviders dictionary for the type T if (cacheProviders.ContainsKey(typeOfT)) cacheProvider = (ICacheProvider<T>)cacheProviders[typeOfT]; else { // There is not already an instance of the ICacheProvider in // cacheProviders for the type T // so we need to create one // Get the Type reference for the application's implementation of // ICacheProvider from the configuration Type cacheProviderType = Type.GetType(CacheProviderConfigurationSection.Current. CacheProviderType); if (cacheProviderType != null) { // Now get a Type reference for the Cache Provider with the // type T generic parameter Type typeOfCacheProviderTypeForT = cacheProviderType.MakeGenericType(new Type[] { typeOfT }); if (typeOfCacheProviderTypeForT != null) { // Create the instance of the Cache Provider and add it to // the cacheProviders dictionary for future use cacheProvider = (ICacheProvider<T>)Activator. CreateInstance(typeOfCacheProviderTypeForT); cacheProviders.Add(typeOfT, cacheProvider); } } } } return cacheProvider; } } } As this code uses Activator.CreateInstance() to create instances of the ICacheProvider<T> implementation, which is a slow process, the factory class maintains a Dictionary of the previously created instances so that a cache provider needs to be created only once for each type. The type of the implementation of ICacheProvider<T> is read from a custom configuration section in the application configuration file, via the CacheProviderConfigurationSection class, which is described below. CacheProviderConfigurationSection Class The implementation of ICacheProvider<T> will be specified in a custom configuration section in the application’s configuration. To handle this create a folder in the CacheSample.Caching project called Configuration, and add a class called CacheProviderConfigurationSection to this folder. This class will extend the System.Configuration.ConfigurationSection class, and will contain a single string property called CacheProviderType. The C# code for this class is shown below: using System; using System.Configuration; namespace CacheSample.Caching.Configuration { internal class CacheProviderConfigurationSection : ConfigurationSection { public static CacheProviderConfigurationSection Current { get { return (CacheProviderConfigurationSection) ConfigurationManager.GetSection("cacheProvider"); } } [ConfigurationProperty("type", IsRequired=true)] public string CacheProviderType { get { return (string)this["type"]; } } } } Adding Data Caching to the Sales Class We now have enough code in place to add caching to the GetSales() method in the CacheSample.BusinessObjects.Sale class, even though we do not yet have an implementation of the ICacheProvider<T> interface. We need to add a reference to the CacheSample.Caching project to CacheSample.BusinessObjects so that we can use the ICacheProvider<T> interface within the GetSales() method. Once the reference is added, we can first create a unique string key based on the method name and the parameter value, so that the same cache key is used for repeated calls to the method with the same parameter values. Then we get an instance of the cache provider for the Sales type, using the CacheProviderFactory, and pass the existing code to retrieve the data from the database as the retrievalMethod delegate in a call to the Cache Provider Fetch() method. The C# code for the modified GetSales() method is shown below: public static IEnumerable<Sale> GetSales(int? highestDayCount) { string cacheKey = string.Format("CacheSample.BusinessObjects.GetSalesWithCache({0})", highestDayCount); return CacheSample.Caching.CacheProviderFactory. GetCacheProvider<Sale>().Fetch(cacheKey, delegate() { List<Sale> sales = new List<Sale>(); SqlParameter highestDayCountParameter = new SqlParameter("@HighestDayCount", SqlDbType.SmallInt); if (highestDayCount.HasValue) highestDayCountParameter.Value = highestDayCount; else highestDayCountParameter.Value = DBNull.Value; string connectionStr = System.Configuration.ConfigurationManager. ConnectionStrings["CacheSample"].ConnectionString; using (SqlConnection sqlConn = new SqlConnection(connectionStr)) using (SqlCommand sqlCmd = sqlConn.CreateCommand()) { sqlCmd.CommandText = "spGetRunningTotals"; sqlCmd.CommandType = CommandType.StoredProcedure; sqlCmd.Parameters.Add(highestDayCountParameter); sqlConn.Open(); using (SqlDataReader dr = sqlCmd.ExecuteReader()) { while (dr.Read()) { Sale newSale = new Sale(); newSale.DayCount = dr.GetInt16(0); newSale.Sales = dr.GetDecimal(1); newSale.RunningTotal = dr.GetDecimal(2); sales.Add(newSale); } } } return sales; }, null, new TimeSpan(0, 10, 0)); } This example passes the code to retrieve the Sales data from the database to the Cache Provider as an anonymous method, however it could also be written as a lambda. The main advantage of using an anonymous function (method or lambda) is that the code inside the anonymous function can access the parameters passed to the GetSales() method. Finally the absolute expiry is set to null, and the relative expiry set to 10 minutes, to indicate that the cache entry should be removed 10 minutes after the last request for the data. As the ICacheProvider<T> has a Fetch() method that returns IEnumerable<T>, we can simply return the results of the Fetch() method to the caller of the GetSales() method. This should be all that is needed for the GetSales() method to now retrieve data from a cache after the first time the data has be retrieved from the database. Implementing a ASP.NET Cache Provider The final step is to actually implement the ICacheProvider<T> interface, and add the implementation details to the web.config file for the dependency injection. The cache provider implementation needs to have access to System.Web. Therefore it could be placed in the CacheSample.UI project, or in its own project that has a reference to System.Web. Implementing the Cache Provider in a separate project is my favoured approach. Create a new project inside the solution called CacheSample.CacheProvider, and add references to System.Web and CacheSample.Caching to this project. Add a class to the project called AspNetCacheProvider. Make the class a generic class by adding the generic parameter <T> and indicate that the class implements ICacheProvider<T>. The C# code for the AspNetCacheProvider class is shown below: using System; using System.Collections.Generic; using System.Linq; using System.Web; using System.Web.Caching; using CacheSample.Caching; namespace CacheSample.CacheProvider { public class AspNetCacheProvider<T> : ICacheProvider<T> { #region ICacheProvider<T> Members public T Fetch(string key, Func<T> retrieveData, DateTime? absoluteExpiry, TimeSpan? relativeExpiry) { return FetchAndCache<T>(key, retrieveData, absoluteExpiry, relativeExpiry); } public IEnumerable<T> Fetch(string key, Func<IEnumerable<T>> retrieveData, DateTime? absoluteExpiry, TimeSpan? relativeExpiry) { return FetchAndCache<IEnumerable<T>>(key, retrieveData, absoluteExpiry, relativeExpiry); } #endregion #region Helper Methods private U FetchAndCache<U>(string key, Func<U> retrieveData, DateTime? absoluteExpiry, TimeSpan? relativeExpiry) { U value; if (!TryGetValue<U>(key, out value)) { value = retrieveData(); if (!absoluteExpiry.HasValue) absoluteExpiry = Cache.NoAbsoluteExpiration; if (!relativeExpiry.HasValue) relativeExpiry = Cache.NoSlidingExpiration; HttpContext.Current.Cache.Insert(key, value, null, absoluteExpiry.Value, relativeExpiry.Value); } return value; } private bool TryGetValue<U>(string key, out U value) { object cachedValue = HttpContext.Current.Cache.Get(key); if (cachedValue == null) { value = default(U); return false; } else { try { value = (U)cachedValue; return true; } catch { value = default(U); return false; } } } #endregion } } The two interface Fetch() methods call a private method called FetchAndCache(). This method first checks for a element in the HttpContext.Current.Cache with the specified cache key, and if so tries to cast this to the specified type (either T or IEnumerable<T>). If the cached element is found, the FetchAndCache() method simply returns it. If it is not found in the cache, the method calls the retrievalMethod delegate to get the data from the data source, and then adds this to the HttpContext.Current.Cache. The final step is to add the AspNetCacheProvider class to the relevant custom configuration section in the CacheSample.UI.Web.Config file. To do this there needs to be a <configSections> element added as the first element in <configuration>. This will match a custom section called <cacheProvider> with the CacheProviderConfigurationSection. Then we add a <cacheProvider> element, with a type property set to the fully qualified assembly name of the AspNetCacheProvider class, as shown below: <?xmlversion="1.0"?> <configuration> <configSections> <sectionname="cacheProvider" type="CacheSample.Base.Configuration.CacheProviderConfigurationSection, CacheSample.Base" /> </configSections> <connectionStrings> <addname="CacheSample" connectionString="data source=.\SQLEXPRESS;Integrated Security=SSPI;Initial Catalog=CacheSample" providerName="System.Data.SqlClient" /> </connectionStrings> <cacheProvidertype="CacheSample.CacheProvider.AspNetCacheProvider`1, CacheSample.CacheProvider, Version=1.0.0.0, Culture=neutral, PublicKeyToken=null"> </cacheProvider> <system.web> <compilationdebug="true"targetFramework="4.0" /> </system.web> </configuration> One point to note is that the fully qualified assembly name of the AspNetCacheProvider class includes the notation `1 after the class name, which indicates that it is a generic class with a single generic type parameter. The CacheSample.UI project needs to have references added to CacheSample.Caching and CacheSample.CacheProvider so that the actual application is aware of the relevant cache provider implementation. Conclusion After implementing this solution, you should have a working cache provider mechanism, that will allow the middle and data access layers to implement caching support when retrieving data, without any knowledge of the actually caching implementation. If the UI is not ASP.NET based, if for example it is Winforms or WPF, the implementation of ICacheProvider<T> would be written around whatever technology is available. It could even be a standalone caching system that takes full responsibility for adding and removing items from a global store. The next part of this article will show how this caching mechanism may be extended to provide support for cache dependencies, such as the System.Web.Caching.SqlCacheDependency. Another possible extension would be to cache the cache provider implementations instead of storing them in a static Dictionary in the CacheProviderFactory. This would prevent a build up of seldom used cache providers in the application memory, as they could be removed from the cache if not used often enough, although in reality there are probably unlikely to be vast numbers of cache provider implementation instances, as most applications do not have a massive number of business object or model types.

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Cluster Node Recovery Using Second Node in Solaris Cluster

- by Onur Bingul

Assumptions:Node 0a is the cluster node that has crashed and could not boot anymore.Node 0b is the node in cluster and in production with services active.Both nodes have their boot disk mirrored via SDS/SVM.We have many options to clone the boot disk from node 0b:- make a copy via network using the ufsdump command and pipe to ufsrestore - make a copy inserting the disk locally on node 0b and creating the third mirror with SDS- make a copy inserting the disk locally on node 0b using dd commandIn this procedure we are going to use dd command (from my experience this is the best option).Bare in mind that in the examples provided we work on Sun Fire V240 systems which have SCSI internal disks. In the case of Fibre Channel (FC) internal disks you must pay attention to the unique identifier, or World Wide Name (WWN), associated with each FC disk (in this case take a look at infodoc #40133 in order to recreate the device tree correctly).Procedure:On node 0b the boot disk is c1t0d0 (c1t1d0 mirror) and this is the VTOC:* Partition Tag Flags Sector Count Sector Mount Directory 0 2 00 0 2106432 2106431 1 3 01 2106432 74630784 76737215 2 5 00 0 143349312 143349311 4 7 00 76737216 50340672 127077887 5 4 00 127077888 14683968 141761855 6 0 00 141761856 1058304 142820159 7 0 00 142820160 529152 143349311We will insert the new disk on node 0b and it will be seen as c1t2d0.1) On node 0b we make a copy via dd from disk c1t0d0s2 to disk c1t2d0s2# dd if=/dev/rdsk/c1t0d0s2 of=/dev/rdsk/c1t2d0s2 bs=8192kA copy of a 72GB disk will take approximately about 45 minutes.Note: as an alternative to make identical copy of root over network follow Document ID: 47498Title: Sun[TM] Cluster 3.0: How to Rebuild a node with Veritas Volume Manager2) Perform an fsck on disk c1t2d0 data slices: 1. fsck -o f /dev/rdsk/c1t2d0s0 (root) 2. fsck -o f /dev/rdsk/c1t2d0s4 (/var) 3. fsck -o f /dev/rdsk/c1t2d0s5 (/usr) 4. fsck -o f /dev/rdsk/c1t2d0s6 (/globaldevices)3) Mount the root file system in order to edit following files for changing the node name:# mount /dev/dsk/c1t2d0s0 /mntChange the hostname from 0b to 0a:# cd /mnt/etc# vi hosts # vi hostname.bge0 # vi hostname.bge2 # vi nodename 4) Change the /mnt/etc/vfstab from the actual:/dev/md/dsk/d201 - - swap - no -/dev/md/dsk/d200 /dev/md/rdsk/d200 / ufs 1 no -/dev/md/dsk/d205 /dev/md/rdsk/d205 /usr ufs 1 no logging/dev/md/dsk/d204 /dev/md/rdsk/d204 /var ufs 1 no logging#/dev/md/dsk/d206 /dev/md/rdsk/d206 /globaldevices ufs 2 yes loggingswap - /tmp tmpfs - yes -/dev/md/dsk/d206 /dev/md/rdsk/d206 /global/.devices/node@2 ufs 2 noglobalto this (unencapsulate disk from SDS/SVM):/dev/dsk/c1t0d0s1 - - swap - no -/dev/dsk/c1t0d0s0 /dev/rdsk/c1t0d0s0 / ufs 1 no -/dev/dsk/c1t0d0s5 /dev/rdsk/c1t0d0s5 /usr ufs 1 no logging/dev/dsk/c1t0d0s4 /dev/rdsk/c1t0d0s4 /var ufs 1 no logging#/dev/md/dsk/d206 /dev/md/rdsk/d206 /globaldevices ufs 2 yes loggingswap - /tmp tmpfs - yes -/dev/dsk/c1t0d0s6 /dev/rdsk/c1t0d0s6 /global/.devices/node@1 ufs 2 no globalIt is important that global device partition (slice 6) in the new vfstab will point to the physical partition of the disk (in our case slice 6).Be careful with the name you use for the new disk. In this case we define it as c1t0d0 because we will insert it as target 0 in node 0a.But this could be different based on the configuration you are working on.5) Remove following entry from /mnt/etc/system (part of unencapsulation procedure):rootdev:/pseudo/md@0:0,200,blk6) Correct the link shared -> ../../global/.devices/node@2/dev/md/shared in order to point to the nodeid of node 0a (in our case nodeid 1):# cd /mnt/dev/mdhow it is now.... node 0b has nodeid 2lrwxrwxrwx 1 root root 42 Mar 10 2005 shared ->../../global/.devices/node@2/dev/md/shared# rm shared# ln -s ../../global/.devices/node@1/dev/md/shared sharedhow is going to be... with nodeid 1 for node 0alrwxrwxrwx 1 root root 42 Mar 10 2005 shared ->../../global/.devices/node@1/dev/md/shared7) Change nodeid (in our case from 2 to 1):# cd /mnt/etc/cluster# vi nodeid8) Change the file /mnt/etc/path_to_inst in order to reflect the correct nodeid for node 0a:# cd /mnt/etc# vi path_to_instChange entries from node@2 to node@1 with the vi command ":%s/node@2/node@1/g"9) Write the bootblock to the disk... just in case:# /usr/sbin/installboot /usr/platform/sun4u/lib/fs/ufs/bootblk /dev/rdsk/c1t2d0s0Now the disk is ready to be inserted in node 0a in order to bootup the node.10) Bootup node 0a with command "boot -sx"... this is becasue we need to make some changes in ccr files in order to recreate did environment.11) Modify cluster ccr:# cd /etc/cluster/ccr# rm did_instances# rm did_instances.bak# vi directory - remove the did_instances line.# /usr/cluster/lib/sc/ccradm -i /etc/cluster/ccr/directory # grep ccr_gennum /etc/cluster/ccr/directory ccr_gennum -1 # /usr/cluster/lib/sc/ccradm -i /etc/cluster/ccr/infrastructure # grep ccr_gennum /etc/cluster/ccr/infrastructure ccr_gennum -112) Bring the node 0a down again to the ok prompt and then issue the command "boot -r"Now the node will join the cluster and from scstat and metaset command you can verify functionality. Next step is to encapsulate the boot disk in SDS/SVM and create the mirrors.In our case node 0b has metadevice name starting from d200. For this reason on node 0a we need to create metadevice starting from d100. This is just an example, you can have different names.The important thing to remember is that metadevice boot disks have different names on each node.13) Remove metadevice pointing to the boot and mirror disks (inherit from node 0b):# metaclear -r -f d200# metaclear -r -f d201# metaclear -r -f d204# metaclear -r -f d205# metaclear -r -f d206verify from metastat that no metadevices are set for boot and mirror disks.14) Encapsulate the boot disk:# metainit -f d110 1 1 c1t0d0s0# metainit d100 -m d110# metaroot d10015) Reboot node 0a.16) Create all the metadevice for slices remaining on boot disk# metainit -f d111 1 1 c1t0d0s1# metainit d101 -m d111# metainit -f d114 1 1 c1t0d0s4# metainit d104 -m d114# metainit -f d115 1 1 c1t0d0s5# metainit d105 -m d115# metainit -f d116 1 1 c1t0d0s6# metainit d106 -m d11617) Edit the vfstab in order to specifiy metadevices created:old:/dev/dsk/c1t0d0s1 - - swap - no -/dev/md/dsk/d100 /dev/md/rdsk/d100 / ufs 1 no -/dev/dsk/c1t0d0s5 /dev/rdsk/c1t0d0s5 /usr ufs 1 no logging/dev/dsk/c1t0d0s4 /dev/rdsk/c1t0d0s4 /var ufs 1 no logging#/dev/md/dsk/d206 /dev/md/rdsk/d206 /globaldevices ufs 2 yes loggingswap - /tmp tmpfs - yes -/dev/dsk/c1t0d0s6 /dev/rdsk/c1t0d0s6 /global/.devices/node@1 ufs 2 no globalnew:/dev/md/dsk/d101 - - swap - no -/dev/md/dsk/d100 /dev/md/rdsk/d100 / ufs 1 no -/dev/md/dsk/d105 /dev/md/rdsk/d105 /usr ufs 1 no logging/dev/md/dsk/d104 /dev/md/rdsk/d104 /var ufs 1 no logging#/dev/md/dsk/106 /dev/md/rdsk/d106 /globaldevices ufs 2 yes loggingswap - /tmp tmpfs - yes -/dev/md/dsk/d106 /dev/md/rdsk/d106 /global/.devices/node@1 ufs 2 noglobal18) Reboot node 0a in order to check new SDS/SVM boot configuration.19) Label the mirror disk c1t1d0 with the VTOC of boot disk c1t0d0:# prtvtoc /dev/dsk/c1t0d0s2 > /var/tmp/VTOC_c1t0d0 # fmthard -s /var/tmp/VTOC_c1t0d0 /dev/rdsk/c1t1d0s220) Put DB replica on slice 7 of disk c1t1d0:# metadb -a -c 3 /dev/dsk/c1t1d0s721) Create metadevice for mirror disk c1t1d0 and attach the new mirror side:# metainit d120 1 1 c1t1d0s0# metattach d100 d120# metainit d121 1 1 c1t1d0s1# metattach d101 d121# metainit d124 1 1 c1t1d0s4# metattach d104 d124# metainit d125 1 1 c1t1d0s5# metattach d105 d125# metainit d126 1 1 c1t1d0s6# metattach d106 d126

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iPhone SDK vs. Windows Phone 7 Series SDK Challenge, Part 2: MoveMe

In this series, I will be taking sample applications from the iPhone SDK and implementing them on Windows Phone 7 Series. My goal is to do as much of an apples-to-apples comparison as I can. This series will be written to not only compare and contrast how easy or difficult it is to complete tasks on either platform, how many lines of code, etc., but Id also like it to be a way for iPhone developers to either get started on Windows Phone 7 Series development, or for developers in general to learn the platform. Heres my methodology: Run the iPhone SDK app in the iPhone Simulator to get a feel for what it does and how it works, without looking at the implementation Implement the equivalent functionality on Windows Phone 7 Series using Silverlight. Compare the two implementations based on complexity, functionality, lines of code, number of files, etc. Add some functionality to the Windows Phone 7 Series app that shows off a way to make the scenario more interesting or leverages an aspect of the platform, or uses a better design pattern to implement the functionality. You can download Microsoft Visual Studio 2010 Express for Windows Phone CTP here, and the Expression Blend 4 Beta here. If youre seeing this series for the first time, check out Part 1: Hello World. A note on methodologyin the prior post there was some feedback about lines of code not being a very good metric for this exercise. I dont really disagree, theres a lot more to this than lines of code but I believe that is a relevant metric, even if its not the ultimate one. And theres no perfect answer here. So I am going to continue to report the number of lines of code that I, as a developer would need to write in these apps as a data point, and Ill leave it up to the reader to determine how that fits in with overall complexity, etc. The first example was so basic that I think it was difficult to talk about in real terms. I think that as these apps get more complex, the subjective differences in concept count and will be more important. MoveMe The MoveMe app is the main end-to-end app writing example in the iPhone SDK, called Creating an iPhone Application. This application demonstrates a few concepts, including handling touch input, how to do animations, and how to do some basic transforms. The behavior of the application is pretty simple. User touches the button: The button does a throb type animation where it scales up and then back down briefly. User drags the button: After a touch begins, moving the touch point will drag the button around with the touch. User lets go of the button: The button animates back to its original position, but does a few small bounces as it reaches its original point, which makes the app fun and gives it an extra bit of interactivity. Now, how would I write an app that meets this spec for Windows Phone 7 Series, and how hard would it be? Lets find out! Implementing the UI Okay, lets build the UI for this application. In the HelloWorld example, we did all the UI design in Visual Studio and/or by hand in XAML. In this example, were going to use the Expression Blend 4 Beta. You might be wondering when to use Visual Studio, when to use Blend, and when to do XAML by hand. Different people will have different takes on this, but heres mine: XAML by hand simple UI that doesnt contain animations, gradients, etc., and or UI that I want to really optimize and craft when I know exactly what I want to do. Visual Studio Basic UI layout, property setting, data binding, etc. Blend Any serious design work needs to be done in Blend, including animations, handling states and transitions, styling and templating, editing resources. As in Part 1, go ahead and fire up Visual Studio 2010 Express for Windows Phone (yes, soon it will take longer to say the name of our products than to start them up!), and create a new Windows Phone Application. As in Part 1, clear out the XAML from the designer. An easy way to do this is to just: Click on the design surface Hit Control+A Hit Delete Theres a little bit left over (the Grid.RowDefinitions element), just go ahead and delete that element so were starting with a clean state of only one outer Grid element. To use Blend, we need to save this project. See, when you create a project with Visual Studio Express, it doesnt commit it to the disk (well, in a place where you can find it, at least) until you actually save the project. This is handy if youre doing some fooling around, because it doesnt clutter your disk with WindowsPhoneApplication23-like directories. But its also kind of dangerous, since when you close VS, if you dont save the projectits all gone. Yes, this has bitten me since I was saving files and didnt remember that, so be careful to save the project/solution via Save All, at least once. So, save and note the location on disk. Start Expression Blend 4 Beta, and chose File > Open Project/Solution, and load your project. You should see just about the same thing you saw over in VS: a blank, black designer surface. Now, thinking about this application, we dont really need a button, even though it looks like one. We never click it. So were just going to create a visual and use that. This is also true in the iPhone example above, where the visual is actually not a button either but a jpg image with a nice gradient and round edges. Well do something simple here that looks pretty good. In Blend, look in the tool pane on the left for the icon that looks like the below (the highlighted one on the left), and hold it down to get the popout menu, and choose Border: Okay, now draw out a box in the middle of the design surface of about 300x100. The Properties Pane to the left should show the properties for this item. First, lets make it more visible by giving it a border brush. Set the BorderBrush to white by clicking BorderBrush and dragging the color selector all the way to the upper right in the palette. Then, down a bit farther, make the BorderThickness 4 all the way around, and the CornerRadius set to 6. In the Layout section, do the following to Width, Height, Horizontal and Vertical Alignment, and Margin (all 4 margin values): Youll see the outline now is in the middle of the design surface. Now lets give it a background color. Above BorderBrush select Background, and click the third tab over: Gradient Brush. Youll see a gradient slider at the bottom, and if you click the markers, you can edit the gradient stops individually (or add more). In this case, you can select something you like, but wheres what I chose: Left stop: #BFACCFE2 (I just picked a spot on the palette and set opacity to 75%, no magic here, feel free to fiddle these or just enter these numbers into the hex area and be done with it) Right stop: #FF3E738F Okay, looks pretty good. Finally set the name of the element in the Name field at the top of the Properties pane to welcome. Now lets add some text. Just hit T and itll select the TextBlock tool automatically: Now draw out some are inside our welcome visual and type Welcome!, then click on the design surface (to exit text entry mode) and hit V to go back into selection mode (or the top item in the tool pane that looks like a mouse pointer). Click on the text again to select it in the tool pane. Just like the border, we want to center this. So set HorizontalAlignment and VerticalAlignment to Center, and clear the Margins: Thats it for the UI. Heres how it looks, on the design surface: Not bad! Okay, now the fun part Adding Animations Using Blend to build animations is a lot of fun, and its easy. In XAML, I can not only declare elements and visuals, but also I can declare animations that will affect those visuals. These are called Storyboards. To recap, well be doing two animations: The throb animation when the element is touched The center animation when the element is released after being dragged. The throb animation is just a scale transform, so well do that first. In the Objects and Timeline Pane (left side, bottom half), click the little + icon to add a new Storyboard called touchStoryboard: The timeline view will appear. In there, click a bit to the right of 0 to create a keyframe at .2 seconds: Now, click on our welcome element (the Border, not the TextBlock in it), and scroll to the bottom of the Properties Pane. Open up Transform, click the third tab ("Scale), and set X and Y to 1.2: This all of this says that, at .2 seconds, I want the X and Y size of this element to scale to 1.2. In fact you can see this happen. Push the Play arrow in the timeline view, and youll see the animation run! Lets make two tweaks. First, we want the animation to automatically reverse so it scales up then back down nicely. Click in the dropdown that says touchStoryboard in Objects and Timeline, then in the Properties pane check Auto Reverse: Now run it again, and youll see it go both ways. Lets even make it nicer by adding an easing function. First, click on the Render Transform item in the Objects tree, then, in the Property Pane, youll see a bunch of easing functions to choose from. Feel free to play with this, then seeing how each runs. I chose Circle In, but some other ones are fun. Try them out! Elastic In is kind of fun, but well stick with Circle In. Thats it for that animation. Now, we also want an animation to move the Border back to its original position when the user ends the touch gesture. This is exactly the same process as above, but just targeting a different transform property. Create a new animation called releaseStoryboard Select a timeline point at 1.2 seconds. Click on the welcome Border element again Scroll to the Transforms panel at the bottom of the Properties Pane Choose the first tab (Translate), which may already be selected Set both X and Y values to 0.0 (we do this just to make the values stick, because the value is already 0 and we need Blend to know we want to save that value) Click on RenderTransform in the Objects tree In the properties pane, choose Bounce Out Set Bounces to 6, and Bounciness to 4 (feel free to play with these as well) Okay, were done. Note, if you want to test this Storyboard, you have to do something a little tricky because the final value is the same as the initial value, so playing it does nothing. If you want to play with it, do the following: Next to the selection dropdown, hit the little "x (Close Storyboard) Go to the Translate Transform value for welcome Set X,Y to 50, 200, respectively (or whatever) Select releaseStoryboard again from the dropdown Hit play, see it run Go into the object tree and select RenderTransform to change the easing function. When youre done, hit the Close Storyboard x again and set the values in Transform/Translate back to 0 Wiring Up the Animations Okay, now go back to Visual Studio. Youll get a prompt due to the modification of MainPage.xaml. Hit Yes. In the designer, click on the welcome Border element. In the Property Browser, hit the Events button, then double click each of ManipulationStarted, ManipulationDelta, ManipulationCompleted. Youll need to flip back to the designer from code, after each double click. Its code time. Here we go. Here, three event handlers have been created for us: welcome_ManipulationStarted: This will execute when a manipulation begins. Think of it as MouseDown. welcome_ManipulationDelta: This executes each time a manipulation changes. Think MouseMove. welcome_ManipulationCompleted: This will execute when the manipulation ends. Think MouseUp. Now, in ManipuliationStarted, we want to kick off the throb animation that we called touchAnimation. Thats easy: 1: private void welcome_ManipulationStarted(object sender, ManipulationStartedEventArgs e) 2: { 3: touchStoryboard.Begin(); 4: } .csharpcode, .csharpcode pre { font-size: small; color: black; font-family: consolas, "Courier New", courier, monospace; background-color: #ffffff; /*white-space: pre;*/ } .csharpcode pre { margin: 0em; } .csharpcode .rem { color: #008000; } .csharpcode .kwrd { color: #0000ff; } .csharpcode .str { color: #006080; } .csharpcode .op { color: #0000c0; } .csharpcode .preproc { color: #cc6633; } .csharpcode .asp { background-color: #ffff00; } .csharpcode .html { color: #800000; } .csharpcode .attr { color: #ff0000; } .csharpcode .alt { background-color: #f4f4f4; width: 100%; margin: 0em; } .csharpcode .lnum { color: #606060; } Likewise, when the manipulation completes, we want to re-center the welcome visual with our bounce animation: 1: private void welcome_ManipulationCompleted(object sender, ManipulationCompletedEventArgs e) 2: { 3: releaseStoryboard.Begin(); 4: } .csharpcode, .csharpcode pre { font-size: small; color: black; font-family: consolas, "Courier New", courier, monospace; background-color: #ffffff; /*white-space: pre;*/ } .csharpcode pre { margin: 0em; } .csharpcode .rem { color: #008000; } .csharpcode .kwrd { color: #0000ff; } .csharpcode .str { color: #006080; } .csharpcode .op { color: #0000c0; } .csharpcode .preproc { color: #cc6633; } .csharpcode .asp { background-color: #ffff00; } .csharpcode .html { color: #800000; } .csharpcode .attr { color: #ff0000; } .csharpcode .alt { background-color: #f4f4f4; width: 100%; margin: 0em; } .csharpcode .lnum { color: #606060; } Note there is actually a way to kick off these animations from Blend directly via something called Triggers, but I think its clearer to show whats going on like this. A Trigger basically allows you to say When this event fires, trigger this Storyboard, so its the exact same logical process as above, but without the code. But how do we get the object to move? Well, for that we really dont want an animation because we want it to respond immediately to user input. We do this by directly modifying the transform to match the offset for the manipulation, and then well let the animation bring it back to zero when the manipulation completes. The manipulation events do a great job of keeping track of all the stuff that you usually had to do yourself when doing drags: where you started from, how far youve moved, etc. So we can easily modify the position as below: 1: private void welcome_ManipulationDelta(object sender, ManipulationDeltaEventArgs e) 2: { 3: CompositeTransform transform = (CompositeTransform)welcome.RenderTransform; 4: 5: transform.TranslateX = e.CumulativeManipulation.Translation.X; 6: transform.TranslateY = e.CumulativeManipulation.Translation.Y; 7: } .csharpcode, .csharpcode pre { font-size: small; color: black; font-family: consolas, "Courier New", courier, monospace; background-color: #ffffff; /*white-space: pre;*/ } .csharpcode pre { margin: 0em; } .csharpcode .rem { color: #008000; } .csharpcode .kwrd { color: #0000ff; } .csharpcode .str { color: #006080; } .csharpcode .op { color: #0000c0; } .csharpcode .preproc { color: #cc6633; } .csharpcode .asp { background-color: #ffff00; } .csharpcode .html { color: #800000; } .csharpcode .attr { color: #ff0000; } .csharpcode .alt { background-color: #f4f4f4; width: 100%; margin: 0em; } .csharpcode .lnum { color: #606060; } Thats it! Go ahead and run the app in the emulator. I suggest running without the debugger, its a little faster (CTRL+F5). If youve got a machine that supports DirectX 10, youll see nice smooth GPU accelerated graphics, which also what it looks like on the phone, running at about 60 frames per second. If your machine does not support DX10 (like the laptop Im writing this on!), it wont be quite a smooth so youll have to take my word for it! Comparing Against the iPhone This is an example where the flexibility and power of XAML meets the tooling of Visual Studio and Blend, and the whole experience really shines. So, for several things that are declarative and 100% toolable with the Windows Phone 7 Series, this example does them with code on the iPhone. In parens is the lines of code that I count to do these operations. PlacardView.m: 19 total LOC Creating the view that hosts the button-like image and the text Drawing the image that is the background of the button Drawing the Welcome text over the image (I think you could technically do this step and/or the prior one using Interface Builder) MoveMeView.m: 63 total LOC Constructing and running the scale (throb) animation (25) Constructing the path describing the animation back to center plus bounce effect (38) Beyond the code count, yy experience with doing this kind of thing in code is that its VERY time intensive. When I was a developer back on Windows Forms, doing GDI+ drawing, we did this stuff a lot, and it took forever! You write some code and even once you get it basically working, you see its not quite right, you go back, tweak the interval, or the math a bit, run it again, etc. You can take a look at the iPhone code here to judge for yourself. Scroll down to animatePlacardViewToCenter toward the bottom. I dont think this code is terribly complicated, but its not what Id call simple and its not at all simple to get right. And then theres a few other lines of code running around for setting up the ViewController and the Views, about 15 lines between MoveMeAppDelegate, PlacardView, and MoveMeView, plus the assorted decls in the h files. Adding those up, I conservatively get something like 100 lines of code (19+63+15+decls) on iPhone that I have to write, by hand, to make this project work. The lines of code that I wrote in the examples above is 5 lines of code on Windows Phone 7 Series. In terms of incremental concept counts beyond the HelloWorld app, heres a shot at that: iPhone: Drawing Images Drawing Text Handling touch events Creating animations Scaling animations Building a path and animating along that Windows Phone 7 Series: Laying out UI in Blend Creating & testing basic animations in Blend Handling touch events Invoking animations from code This was actually the first example I tried converting, even before I did the HelloWorld, and I was pretty surprised. Some of this is luck that this app happens to match up with the Windows Phone 7 Series platform just perfectly. In terms of time, I wrote the above application, from scratch, in about 10 minutes. I dont know how long it would take a very skilled iPhone developer to write MoveMe on that iPhone from scratch, but if I was to write it on Silverlight in the same way (e.g. all via code), I think it would likely take me at least an hour or two to get it all working right, maybe more if I ended up picking the wrong strategy or couldnt get the math right, etc. Making Some Tweaks Silverlight contains a feature called Projections to do a variety of 3D-like effects with a 2D surface. So lets play with that a bit. Go back to Blend and select the welcome Border in the object tree. In its properties, scroll down to the bottom, open Transform, and see Projection at the bottom. Set X,Y,Z to 90. Youll see the element kind of disappear, replaced by a thin blue line. Now Create a new animation called startupStoryboard. Set its key time to .5 seconds in the timeline view Set the projection values above to 0 for X, Y, and Z. Save Go back to Visual Studio, and in the constructor, add the following bold code (lines 7-9 to the constructor: 1: public MainPage() 2: { 3: InitializeComponent(); 4: 5: SupportedOrientations = SupportedPageOrientation.Portrait; 6: 7: this.Loaded += (s, e) => 8: { 9: startupStoryboard.Begin(); 10: }; 11: } .csharpcode, .csharpcode pre { font-size: small; color: black; font-family: consolas, "Courier New", courier, monospace; background-color: #ffffff; /*white-space: pre;*/ } .csharpcode pre { margin: 0em; } .csharpcode .rem { color: #008000; } .csharpcode .kwrd { color: #0000ff; } .csharpcode .str { color: #006080; } .csharpcode .op { color: #0000c0; } .csharpcode .preproc { color: #cc6633; } .csharpcode .asp { background-color: #ffff00; } .csharpcode .html { color: #800000; } .csharpcode .attr { color: #ff0000; } .csharpcode .alt { background-color: #f4f4f4; width: 100%; margin: 0em; } .csharpcode .lnum { color: #606060; } If the code above looks funny, its using something called a lambda in C#, which is an inline anonymous method. Its just a handy shorthand for creating a handler like the manipulation ones above. So with this youll get a nice 3D looking fly in effect when the app starts up. Here it is, in flight: Pretty cool!Did you know that DotNetSlackers also publishes .net articles written by top known .net Authors? We already have over 80 articles in several categories including Silverlight. Take a look: here.

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Informed TDD – Kata “To Roman Numerals”

- by Ralf Westphal

Originally posted on: http://geekswithblogs.net/theArchitectsNapkin/archive/2014/05/28/informed-tdd-ndash-kata-ldquoto-roman-numeralsrdquo.aspxIn a comment on my article on what I call Informed TDD (ITDD) reader gustav asked how this approach would apply to the kata “To Roman Numerals”. And whether ITDD wasn´t a violation of TDD´s principle of leaving out “advanced topics like mocks”. I like to respond with this article to his questions. There´s more to say than fits into a commentary. Mocks and TDD I don´t see in how far TDD is avoiding or opposed to mocks. TDD and mocks are orthogonal. TDD is about pocess, mocks are about structure and costs. Maybe by moving forward in tiny red+green+refactor steps less need arises for mocks. But then… if the functionality you need to implement requires “expensive” resource access you can´t avoid using mocks. Because you don´t want to constantly run all your tests against the real resource. True, in ITDD mocks seem to be in almost inflationary use. That´s not what you usually see in TDD demonstrations. However, there´s a reason for that as I tried to explain. I don´t use mocks as proxies for “expensive” resource. Rather they are stand-ins for functionality not yet implemented. They allow me to get a test green on a high level of abstraction. That way I can move forward in a top-down fashion. But if you think of mocks as “advanced” or if you don´t want to use a tool like JustMock, then you don´t need to use mocks. You just need to stand the sight of red tests for a little longer ;-) Let me show you what I mean by that by doing a kata. ITDD for “To Roman Numerals” gustav asked for the kata “To Roman Numerals”. I won´t explain the requirements again. You can find descriptions and TDD demonstrations all over the internet, like this one from Corey Haines. Now here is, how I would do this kata differently. 1. Analyse A demonstration of TDD should never skip the analysis phase. It should be made explicit. The requirements should be formalized and acceptance test cases should be compiled. “Formalization” in this case to me means describing the API of the required functionality. “[D]esign a program to work with Roman numerals” like written in this “requirement document” is not enough to start software development. Coding should only begin, if the interface between the “system under development” and its context is clear. If this interface is not readily recognizable from the requirements, it has to be developed first. Exploration of interface alternatives might be in order. It might be necessary to show several interface mock-ups to the customer – even if that´s you fellow developer. Designing the interface is a task of it´s own. It should not be mixed with implementing the required functionality behind the interface. Unfortunately, though, this happens quite often in TDD demonstrations. TDD is used to explore the API and implement it at the same time. To me that´s a violation of the Single Responsibility Principle (SRP) which not only should hold for software functional units but also for tasks or activities. In the case of this kata the API fortunately is obvious. Just one function is needed: string ToRoman(int arabic). And it lives in a class ArabicRomanConversions. Now what about acceptance test cases? There are hardly any stated in the kata descriptions. Roman numerals are explained, but no specific test cases from the point of view of a customer. So I just “invent” some acceptance test cases by picking roman numerals from a wikipedia article. They are supposed to be just “typical examples” without special meaning. Given the acceptance test cases I then try to develop an understanding of the problem domain. I´ll spare you that. The domain is trivial and is explain in almost all kata descriptions. How roman numerals are built is not difficult to understand. What´s more difficult, though, might be to find an efficient solution to convert into them automatically. 2. Solve The usual TDD demonstration skips a solution finding phase. Like the interface exploration it´s mixed in with the implementation. But I don´t think this is how it should be done. I even think this is not how it really works for the people demonstrating TDD. They´re simplifying their true software development process because they want to show a streamlined TDD process. I doubt this is helping anybody. Before you code you better have a plan what to code. This does not mean you have to do “Big Design Up-Front”. It just means: Have a clear picture of the logical solution in your head before you start to build a physical solution (code). Evidently such a solution can only be as good as your understanding of the problem. If that´s limited your solution will be limited, too. Fortunately, in the case of this kata your understanding does not need to be limited. Thus the logical solution does not need to be limited or preliminary or tentative. That does not mean you need to know every line of code in advance. It just means you know the rough structure of your implementation beforehand. Because it should mirror the process described by the logical or conceptual solution. Here´s my solution approach: The arabic “encoding” of numbers represents them as an ordered set of powers of 10. Each digit is a factor to multiply a power of ten with. The “encoding” 123 is the short form for a set like this: {1*10^2, 2*10^1, 3*10^0}. And the number is the sum of the set members. The roman “encoding” is different. There is no base (like 10 for arabic numbers), there are just digits of different value, and they have to be written in descending order. The “encoding” XVI is short for [10, 5, 1]. And the number is still the sum of the members of this list. The roman “encoding” thus is simpler than the arabic. Each “digit” can be taken at face value. No multiplication with a base required. But what about IV which looks like a contradiction to the above rule? It is not – if you accept roman “digits” not to be limited to be single characters only. Usually I, V, X, L, C, D, M are viewed as “digits”, and IV, IX etc. are viewed as nuisances preventing a simple solution. All looks different, though, once IV, IX etc. are taken as “digits”. Then MCMLIV is just a sum: M+CM+L+IV which is 1000+900+50+4. Whereas before it would have been understood as M-C+M+L-I+V – which is more difficult because here some “digits” get subtracted. Here´s the list of roman “digits” with their values: {1, I}, {4, IV}, {5, V}, {9, IX}, {10, X}, {40, XL}, {50, L}, {90, XC}, {100, C}, {400, CD}, {500, D}, {900, CM}, {1000, M} Since I take IV, IX etc. as “digits” translating an arabic number becomes trivial. I just need to find the values of the roman “digits” making up the number, e.g. 1954 is made up of 1000, 900, 50, and 4. I call those “digits” factors. If I move from the highest factor (M=1000) to the lowest (I=1) then translation is a two phase process: Find all the factors Translate the factors found Compile the roman representation Translation is just a look-up. Finding, though, needs some calculation: Find the highest remaining factor fitting in the value Remember and subtract it from the value Repeat with remaining value and remaining factors Please note: This is just an algorithm. It´s not code, even though it might be close. Being so close to code in my solution approach is due to the triviality of the problem. In more realistic examples the conceptual solution would be on a higher level of abstraction. With this solution in hand I finally can do what TDD advocates: find and prioritize test cases. As I can see from the small process description above, there are two aspects to test: Test the translation Test the compilation Test finding the factors Testing the translation primarily means to check if the map of factors and digits is comprehensive. That´s simple, even though it might be tedious. Testing the compilation is trivial. Testing factor finding, though, is a tad more complicated. I can think of several steps: First check, if an arabic number equal to a factor is processed correctly (e.g. 1000=M). Then check if an arabic number consisting of two consecutive factors (e.g. 1900=[M,CM]) is processed correctly. Then check, if a number consisting of the same factor twice is processed correctly (e.g. 2000=[M,M]). Finally check, if an arabic number consisting of non-consecutive factors (e.g. 1400=[M,CD]) is processed correctly. I feel I can start an implementation now. If something becomes more complicated than expected I can slow down and repeat this process. 3. Implement First I write a test for the acceptance test cases. It´s red because there´s no implementation even of the API. That´s in conformance with “TDD lore”, I´d say: Next I implement the API: The acceptance test now is formally correct, but still red of course. This will not change even now that I zoom in. Because my goal is not to most quickly satisfy these tests, but to implement my solution in a stepwise manner. That I do by “faking” it: I just “assume” three functions to represent the transformation process of my solution: My hypothesis is that those three functions in conjunction produce correct results on the API-level. I just have to implement them correctly. That´s what I´m trying now – one by one. I start with a simple “detail function”: Translate(). And I start with all the test cases in the obvious equivalence partition: As you can see I dare to test a private method. Yes. That´s a white box test. But as you´ll see it won´t make my tests brittle. It serves a purpose right here and now: it lets me focus on getting one aspect of my solution right. Here´s the implementation to satisfy the test: It´s as simple as possible. Right how TDD wants me to do it: KISS. Now for the second equivalence partition: translating multiple factors. (It´a pattern: if you need to do something repeatedly separate the tests for doing it once and doing it multiple times.) In this partition I just need a single test case, I guess. Stepping up from a single translation to multiple translations is no rocket science: Usually I would have implemented the final code right away. Splitting it in two steps is just for “educational purposes” here. How small your implementation steps are is a matter of your programming competency. Some “see” the final code right away before their mental eye – others need to work their way towards it. Having two tests I find more important. Now for the next low hanging fruit: compilation. It´s even simpler than translation. A single test is enough, I guess. And normally I would not even have bothered to write that one, because the implementation is so simple. I don´t need to test .NET framework functionality. But again: if it serves the educational purpose… Finally the most complicated part of the solution: finding the factors. There are several equivalence partitions. But still I decide to write just a single test, since the structure of the test data is the same for all partitions: Again, I´m faking the implementation first: I focus on just the first test case. No looping yet. Faking lets me stay on a high level of abstraction. I can write down the implementation of the solution without bothering myself with details of how to actually accomplish the feat. That´s left for a drill down with a test of the fake function: There are two main equivalence partitions, I guess: either the first factor is appropriate or some next. The implementation seems easy. Both test cases are green. (Of course this only works on the premise that there´s always a matching factor. Which is the case since the smallest factor is 1.) And the first of the equivalence partitions on the higher level also is satisfied: Great, I can move on. Now for more than a single factor: Interestingly not just one test becomes green now, but all of them. Great! You might say, then I must have done not the simplest thing possible. And I would reply: I don´t care. I did the most obvious thing. But I also find this loop very simple. Even simpler than a recursion of which I had thought briefly during the problem solving phase. And by the way: Also the acceptance tests went green: Mission accomplished. At least functionality wise. Now I´ve to tidy up things a bit. TDD calls for refactoring. Not uch refactoring is needed, because I wrote the code in top-down fashion. I faked it until I made it. I endured red tests on higher levels while lower levels weren´t perfected yet. But this way I saved myself from refactoring tediousness. At the end, though, some refactoring is required. But maybe in a different way than you would expect. That´s why I rather call it “cleanup”. First I remove duplication. There are two places where factors are defined: in Translate() and in Find_factors(). So I factor the map out into a class constant. Which leads to a small conversion in Find_factors(): And now for the big cleanup: I remove all tests of private methods. They are scaffolding tests to me. They only have temporary value. They are brittle. Only acceptance tests need to remain. However, I carry over the single “digit” tests from Translate() to the acceptance test. I find them valuable to keep, since the other acceptance tests only exercise a subset of all roman “digits”. This then is my final test class: And this is the final production code: Test coverage as reported by NCrunch is 100%: Reflexion Is this the smallest possible code base for this kata? Sure not. You´ll find more concise solutions on the internet. But LOC are of relatively little concern – as long as I can understand the code quickly. So called “elegant” code, however, often is not easy to understand. The same goes for KISS code – especially if left unrefactored, as it is often the case. That´s why I progressed from requirements to final code the way I did. I first understood and solved the problem on a conceptual level. Then I implemented it top down according to my design. I also could have implemented it bottom-up, since I knew some bottom of the solution. That´s the leaves of the functional decomposition tree. Where things became fuzzy, since the design did not cover any more details as with Find_factors(), I repeated the process in the small, so to speak: fake some top level, endure red high level tests, while first solving a simpler problem. Using scaffolding tests (to be thrown away at the end) brought two advantages: Encapsulation of the implementation details was not compromised. Naturally private methods could stay private. I did not need to make them internal or public just to be able to test them. I was able to write focused tests for small aspects of the solution. No need to test everything through the solution root, the API. The bottom line thus for me is: Informed TDD produces cleaner code in a systematic way. It conforms to core principles of programming: Single Responsibility Principle and/or Separation of Concerns. Distinct roles in development – being a researcher, being an engineer, being a craftsman – are represented as different phases. First find what, what there is. Then devise a solution. Then code the solution, manifest the solution in code. Writing tests first is a good practice. But it should not be taken dogmatic. And above all it should not be overloaded with purposes. And finally: moving from top to bottom through a design produces refactored code right away. Clean code thus almost is inevitable – and not left to a refactoring step at the end which is skipped often for different reasons. PS: Yes, I have done this kata several times. But that has only an impact on the time needed for phases 1 and 2. I won´t skip them because of that. And there are no shortcuts during implementation because of that.

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How to shoot yourself in the foot (DO NOT Read in the office)

- by TATWORTH

Originally posted on: http://geekswithblogs.net/TATWORTH/archive/2013/06/21/how-to-shoot-yourself-in-the-foot-do-not-read.aspxLet me make it absolutely clear - the following is:merely collated by your Geek from http://www.codeproject.com/Lounge.aspx?msg=3917012#xx3917012xxvery, very very funny so you read it in the presence of others at your own riskso here is the list - you have been warned!C You shoot yourself in the foot. C++ You accidently create a dozen instances of yourself and shoot them all in the foot. Providing emergency medical assistance is impossible since you can't tell which are bitwise copies and which are just pointing at others and saying "That's me, over there." FORTRAN You shoot yourself in each toe, iteratively, until you run out of toes, then you read in the next foot and repeat. If you run out of bullets, you continue anyway because you have no exception-handling facility. Modula-2 After realizing that you can't actually accomplish anything in this language, you shoot yourself in the head. COBOL USEing a COLT 45 HANDGUN, AIM gun at LEG.FOOT, THEN place ARM.HAND.FINGER on HANDGUN.TRIGGER and SQUEEZE. THEN return HANDGUN to HOLSTER. CHECK whether shoelace needs to be retied. Lisp You shoot yourself in the appendage which holds the gun with which you shoot yourself in the appendage which holds the gun with which you shoot yourself in the appendage which holds... BASIC Shoot yourself in the foot with a water pistol. On big systems, continue until entire lower body is waterlogged. Forth Foot yourself in the shoot. APL You shoot yourself in the foot; then spend all day figuring out how to do it in fewer characters. Pascal The compiler won't let you shoot yourself in the foot. Snobol If you succeed, shoot yourself in the left foot. If you fail, shoot yourself in the right foot. HyperTalk Put the first bullet of the gun into foot left of leg of you. Answer the result. Prolog You tell your program you want to be shot in the foot. The program figures out how to do it, but the syntax doesn't allow it to explain. 370 JCL You send your foot down to MIS with a 4000-page document explaining how you want it to be shot. Three years later, your foot comes back deep-fried. FORTRAN-77 You shoot yourself in each toe, iteratively, until you run out of toes, then you read in the next foot and repeat. If you run out of bullets, you continue anyway because you still can't do exception-processing. Modula-2 (alternative) You perform a shooting on what might be currently a foot with what might be currently a bullet shot by what might currently be a gun. BASIC (compiled) You shoot yourself in the foot with a BB using a SCUD missile launcher. Visual Basic You'll really only appear to have shot yourself in the foot, but you'll have so much fun doing it that you won't care. Forth (alternative) BULLET DUP3 * GUN LOAD FOOT AIM TRIGGER PULL BANG! EMIT DEAD IF DROP ROT THEN (This takes about five bytes of memory, executes in two to ten clock cycles on any processor and can be used to replace any existing function of the language as well as in any future words). (Welcome to bottom up programming - where you, too, can perform compiler pre-processing instead of writing code) APL (alternative) You hear a gunshot and there's a hole in your foot, but you don't remember enough linear algebra to understand what happened. or @#&^$%&%^ foot Pascal (alternative) Same as Modula-2 except that the bullet is not the right type for the gun and your hand is blown off. Snobol (alternative) You grab your foot with your hand, then rewrite your hand to be a bullet. The act of shooting the original foot then changes your hand/bullet into yet another foot (a left foot). Prolog (alternative) You attempt to shoot yourself in the foot, but the bullet, failing to find its mark, backtracks to the gun, which then explodes in your face. COMAL You attempt to shoot yourself in the foot with a water pistol, but the bore is clogged, and the pressure build-up blows apart both the pistol and your hand. or draw_pistol aim_at_foot(left) pull_trigger hop(swearing) Scheme As Lisp, but none of the other appendages are aware of this happening. Algol You shoot yourself in the foot with a musket. The musket is aesthetically fascinating and the wound baffles the adolescent medic in the emergency room. Ada If you are dumb enough to actually use this language, the United States Department of Defense will kidnap you, stand you up in front of a firing squad and tell the soldiers, "Shoot at the feet." or The Department of Defense shoots you in the foot after offering you a blindfold and a last cigarette. or After correctly packaging your foot, you attempt to concurrently load the gun, pull the trigger, scream and shoot yourself in the foot. When you try, however, you discover that your foot is of the wrong type. or After correctly packing your foot, you attempt to concurrently load the gun, pull the trigger, scream, and confidently aim at your foot knowing it is safe. However the cordite in the round does an Unchecked Conversion, fires and shoots you in the foot anyway. Eiffel You create a GUN object, two FOOT objects and a BULLET object. The GUN passes both the FOOT objects a reference to the BULLET. The FOOT objects increment their hole counts and forget about the BULLET. A little demon then drives a garbage truck over your feet and grabs the bullet (both of it) on the way. Smalltalk You spend so much time playing with the graphics and windowing system that your boss shoots you in the foot, takes away your workstation and makes you develop in COBOL on a character terminal. or You send the message shoot to gun, with selectors bullet and myFoot. A window pops up saying Gunpowder doesNotUnderstand: spark. After several fruitless hours spent browsing the methods for Trigger, FiringPin and IdealGas, you take the easy way out and create ShotFoot, a subclass of Foot with an additional instance variable bulletHole. Object Oriented Pascal You perform a shooting on what might currently be a foot with what might currently be a bullet fired from what might currently be a gun. PL/I You consume all available system resources, including all the offline bullets. The Data Processing & Payroll Department doubles its size, triples its budget, acquires four new mainframes and drops the original one on your foot. Postscript foot bullets 6 locate loadgun aim gun shoot showpage or It takes the bullet ten minutes to travel from the gun to your foot, by which time you're long since gone out to lunch. The text comes out great, though. PERL You stab yourself in the foot repeatedly with an incredibly large and very heavy Swiss Army knife. or You pick up the gun and begin to load it. The gun and your foot begin to grow to huge proportions and the world around you slows down, until the gun fires. It makes a tiny hole, which you don't feel. Assembly Language You crash the OS and overwrite the root disk. The system administrator arrives and shoots you in the foot. After a moment of contemplation, the administrator shoots himself in the foot and then hops around the room rabidly shooting at everyone in sight. or You try to shoot yourself in the foot only to discover you must first reinvent the gun, the bullet, and your foot.or The bullet travels to your foot instantly, but it took you three weeks to load the round and aim the gun. BCPL You shoot yourself somewhere in the leg -- you can't get any finer resolution than that. Concurrent Euclid You shoot yourself in somebody else's foot. Motif You spend days writing a UIL description of your foot, the trajectory, the bullet and the intricate scrollwork on the ivory handles of the gun. When you finally get around to pulling the trigger, the gun jams. Powerbuilder While attempting to load the gun you discover that the LoadGun system function is buggy; as a work around you tape the bullet to the outside of the gun and unsuccessfully attempt to fire it with a nail. In frustration you club your foot with the butt of the gun and explain to your client that this approximates the functionality of shooting yourself in the foot and that the next version of Powerbuilder will fix it. Standard ML By the time you get your code to typecheck, you're using a shoot to foot yourself in the gun. MUMPS You shoot 583149 AK-47 teflon-tipped, hollow-point, armour-piercing bullets into even-numbered toes on odd-numbered feet of everyone in the building -- with one line of code. Three weeks later you shoot yourself in the head rather than try to modify that line. Java You locate the Gun class, but discover that the Bullet class is abstract, so you extend it and write the missing part of the implementation. Then you implement the ShootAble interface for your foot, and recompile the Foot class. The interface lets the bullet call the doDamage method on the Foot, so the Foot can damage itself in the most effective way. Now you run the program, and call the doShoot method on the instance of the Gun class. First the Gun creates an instance of Bullet, which calls the doFire method on the Gun. The Gun calls the hit(Bullet) method on the Foot, and the instance of Bullet is passed to the Foot. But this causes an IllegalHitByBullet exception to be thrown, and you die. Unix You shoot yourself in the foot or % ls foot.c foot.h foot.o toe.c toe.o % rm * .o rm: .o: No such file or directory % ls % 370 JCL (alternative) You shoot yourself in the head just thinking about it. DOS JCL You first find the building you're in in the phone book, then find your office number in the corporate phone book. Then you have to write this down, then describe, in cubits, your exact location, in relation to the door (right hand side thereof). Then you need to write down the location of the gun (loading it is a proprietary utility), then you load it, and the COBOL program, and run them, and, with luck, it may be run tonight. VMS $ MOUNT/DENSITY=.45/LABEL=BULLET/MESSAGE="BYE" BULLET::BULLET$GUN SYS$BULLET $ SET GUN/LOAD/SAFETY=OFF/SIGHT=NONE/HAND=LEFT/CHAMBER=1/ACTION=AUTOMATIC/ LOG/ALL/FULL SYS$GUN_3$DUA3:[000000]GUN.GNU $ SHOOT/LOG/AUTO SYS$GUN SYS$SYSTEM:[FOOT]FOOT.FOOT %DCL-W-ACTIMAGE, error activating image GUN -CLI-E-IMGNAME, image file $3$DUA240:[GUN]GUN.EXE;1 -IMGACT-F-NOTNATIVE, image is not an OpenVMS Alpha AXP image or %SYS-F-FTSHT, foot shot (fifty lines of traceback omitted) sh,csh, etc You can't remember the syntax for anything, so you spend five hours reading manual pages, then your foot falls asleep. You shoot the computer and switch to C. Apple System 7 Double click the gun icon and a window giving a selection for guns, target areas, plus balloon help with medical remedies, and assorted sound effects. Click "shoot" button and a small bomb appears with note "Error of Type 1 has occurred." Windows 3.1 Double click the gun icon and wait. Eventually a window opens giving a selection for guns, target areas, plus balloon help with medical remedies, and assorted sound effects. Click "shoot" button and a small box appears with note "Unable to open Shoot.dll, check that path is correct." Windows 95 Your gun is not compatible with this OS and you must buy an upgrade and install it before you can continue. Then you will be informed that you don't have enough memory. CP/M I remember when shooting yourself in the foot with a BB gun was a big deal. DOS You finally found the gun, but can't locate the file with the foot for the life of you. MSDOS You shoot yourself in the foot, but can unshoot yourself with add-on software. Access You try to point the gun at your foot, but it shoots holes in all your Borland distribution diskettes instead. Paradox Not only can you shoot yourself in the foot, your users can too. dBase You squeeze the trigger, but the bullet moves so slowly that by the time your foot feels the pain, you've forgotten why you shot yourself anyway. or You buy a gun. Bullets are only available from another company and are promised to work so you buy them. Then you find out that the next version of the gun is the one scheduled to actually shoot bullets. DBase IV, V1.0 You pull the trigger, but it turns out that the gun was a poorly designed hand grenade and the whole building blows up. SQL You cut your foot off, send it out to a service bureau and when it returns, it has a hole in it but will no longer fit the attachment at the end of your leg. or Insert into Foot Select Bullet >From Gun.Hand Where Chamber = 'LOADED' And Trigger = 'PULLED' Clipper You grab a bullet, get ready to insert it in the gun so that you can shoot yourself in the foot and discover that the gun that the bullets fits has not yet been built, but should be arriving in the mail _REAL_SOON_NOW_. Oracle The menus for coding foot_shooting have not been implemented yet and you can't do foot shooting in SQL. English You put your foot in your mouth, then bite it off. (For those who don't know, English is a McDonnell Douglas/PICK query language which allegedly requires 110% of system resources to run happily.) Revelation [an implementation of the PICK Operating System] You'll be able to shoot yourself in the foot just as soon as you figure out what all these bullets are for. FlagShip Starting at the top of your head, you aim the gun at yourself repeatedly until, half an hour later, the gun is finally pointing at your foot and you pull the trigger. A new foot with a hole in it appears but you can't work out how to get rid of the old one and your gun doesn't work anymore. FidoNet You put your foot in your mouth, then echo it internationally. PicoSpan [a UNIX-based computer conferencing system] You can't shoot yourself in the foot because you're not a host. or (host variation) Whenever you shoot yourself in the foot, someone opens a topic in policy about it. Internet You put your foot in your mouth, shoot it, then spam the bullet so that everybody gets shot in the foot. troff rmtroff -ms -Hdrwp | lpr -Pwp2 & .*place bullet in footer .B .NR FT +3i .in 4 .bu Shoot! .br .sp .in -4 .br .bp NR HD -2i .* Genetic Algorithms You create 10,000 strings describing the best way to shoot yourself in the foot. By the time the program produces the optimal solution, humans have evolved wings and the problem is moot. CSP (Communicating Sequential Processes) You only fail to shoot everything that isn't your foot. MS-SQL Server MS-SQL Server’s gun comes pre-loaded with an unlimited supply of Teflon coated bullets, and it only has two discernible features: the muzzle and the trigger. If that wasn't enough, MS-SQL Server also puts the gun in your hand, applies local anesthetic to the skin of your forefinger and stitches it to the gun's trigger. Meanwhile, another process has set up a spinal block to numb your lower body. It will then proceeded to surgically remove your foot, cryogenically freeze it for preservation, and attach it to the muzzle of the gun so that no matter where you aim, you will shoot your foot. In order to avoid shooting yourself in the foot, you need to unstitch your trigger finger, remove your foot from the muzzle of the gun, and have it surgically reattached. Then you probably want to get some crutches and go out to buy a book on SQL Server Performance Tuning. Sybase Sybase's gun requires assembly, and you need to go out and purchase your own clip and bullets to load the gun. Assembly is complicated by the fact that Sybase has hidden the gun behind a big stack of reference manuals, but it hasn't told you where that stack is. While you were off finding the gun, assembling it, buying bullets, etc., Sybase was also busy surgically removing your foot and cryogenically freezing it for preservation. Instead of attaching it to the muzzle of the gun, though, it packed your foot on dry ice and sent it UPS-Ground to an unnamed hookah bar somewhere in the middle east. In order to shoot your foot, you must modify your gun with a GPS system for targeting and hire some guy named "Indy" to find the hookah bar and wire the coordinates back to you. By this time, you've probably become so daunted at the tasks stand between you and shooting your foot that you hire a guy who's read all the books on Sybase to help you shoot your foot. If you're lucky, he'll be smart enough both to find your foot and to stop you from shooting it. Magic software You spend 1 week looking up the correct syntax for GUN. When you find it, you realise that GUN will not let you shoot in your own foot. It will allow you to shoot almost anything but your foot. You then decide to build your own gun. You can't use the standard barrel since this will only allow for standard bullets, which will not fire if the barrel is pointed at your foot. After four weeks, you have created your own custom gun. It blows up in your hand without warning, because you failed to initialise the safety catch and it doesn't know whether the initial state is "0", 0, NULL, "ZERO", 0.0, 0,0, "0.0", or "0,00". You fix the problem with your remaining hand by nesting 12 safety catches, and then decide to build the gun without safety catch. You then shoot the management and retire to a happy life where you code in languages that will allow you to shoot your foot in under 10 days.FirefoxLets you shoot yourself in as many feet as you'd like, while using multiple great addons! IEA moving target in terms of standard ammunition size and doesn't always work properly with non-Microsoft ammunition, so sometimes you shoot something other than your foot. However, it's the corporate world's standard foot-shooting apparatus. Hackers seem to enjoy rigging websites up to trigger cascading foot-shooting failures. Windows 98 About the same as Windows 95 in terms of overall bullet capacity and triggering mechanisms. Includes updated DirectShot API. A new version was released later on to support USB guns, Windows 98 SE.WPF:You get your baseball glove and a ball and you head out to your backyard, where you throw balls to your pitchback. Then your unkempt-haired-cargo-shorts-and-sandals-with-white-socks-wearing neighbor uses XAML to sculpt your arm into a gun, the ball into a bullet and the pitchback into your foot. By now, however, only the neighbor can get it to work and he's only around from 6:30 PM - 3:30 AM. LOGO: You very carefully lay out the trajectory of the bullet. Then you start the gun, which fires very slowly. You walk precisely to the point where the bullet will travel and wait, but just before it gets to you, your class time is up and one of the other kids has already used the system to hack into Sony's PS3 network. Flash: Someone has designed a beautiful-looking gun that anyone can shoot their feet with for free. It weighs six hundred pounds. All kinds of people are shooting themselves in the feet, and sending the link to everyone else so that they can too. That is, except for the criminals, who are all stealing iOS devices that the gun won't work with.APL: Its (mostly) all greek to me. Lisp: Place ((gun in ((hand sight (foot then shoot))))) (Lots of Insipid Stupid Parentheses)Apple OS/X and iOS Once a year, Steve Jobs returns from sick leave to tell millions of unwavering fans how they will be able to shoot themselves in the foot differently this year. They retweet and blog about it ad nauseam, and wait in line to be the first to experience "shoot different".Windows ME Usually fails, even at shooting you in the foot. Yo dawg, I heard you like shooting yourself in the foot. So I put a gun in your gun, so you can shoot yourself in the foot while you shoot yourself in the foot. (Okay, I'm not especially proud of this joke.) Windows 2000 Now you really do have to log in, before you are allowed to shoot yourself in the foot.Windows XPYou thought you learned your lesson: Don't use Windows ME. Then, along came this new creature, built on top of Windows NT! So you spend the next couple days installing antivirus software, patches and service packs, just so you can get that driver to install, and then proceed to shoot yourself in the foot. Windows Vista Newer! Glossier! Shootier! Windows 7 The bullets come out a lot smoother. Active Directory Each bullet now has an attached Bullet Identifier, and can be uniquely identified. Policies can be applied to dictate fragmentation, and the gun will occasionally have a confusing delay after the trigger has been pulled. PythonYou try to use import foot; foot.shoot() only to realize that's only available in 3.0, to which you can't yet upgrade from 2.7 because of all those extension libs lacking support. Solaris Shoots best when used on SPARC hardware, but still runs the trigger GUI under Java. After weeks of learning the appropriate STOP command to prevent the trigger from automatically being pressed on boot, you think you've got it under control. Then the one time you ever use dtrace, it hits a bug that fires the gun. MySQL The feature that allows you to shoot yourself in the foot has been in development for about 6 years, and they are adding it into the next version, which is coming out REAL SOON NOW, promise! But you can always check it out of source control and try it yourself (just not in any environment where data integrity is important because it will probably explode.) PostgreSQLAllows you to have a smug look on your face while you shoot yourself in the foot, because those MySQL guys STILL don't have that feature. NoSQL Barrel? Who needs a barrel? Just put the bullet on your foot, and strike it with a hammer. See? It's so much simpler and more efficient that way. You can even strike multiple bullets in one swing if you swing with a good enough arc, because hammers are easy to use. Getting them to synchronize is a little difficult, though.Eclipse There are about a dozen different packages for shooting yourself in the foot, with weird interdependencies on outdated components. Once you finally navigate the morass and get one installed, you then have something to look at while you shoot yourself in the foot with that package: You can watch the screen redraw.Outlook Makes it really easy to let everyone know you shot yourself in the foot!Shooting yourself in the foot using delegates.You really need to shoot yourself in the foot but you hate firearms (you don't want any dependency on the specifics of shooting) so you delegate it to somebody else. You don't care how it is done as long is shooting your foot. You can do it asynchronously in case you know you may faint so you are called back/slapped in the face by your shooter/friend (or background worker) when everything is done.C#You prepare the gun and the bullet, carefully modeling all of the physics of a bullet traveling through a foot. Just before you're about to pull the trigger, you stumble on System.Windows.BodyParts.Foot.ShootAt(System.Windows.Firearms.IGun gun) in the extended framework, realize you just wasted the entire afternoon, and shoot yourself in the head.PHP<?phprequire("foot_safety_check.php");?><!DOCTYPE HTML><html><head> <title>Shooting me in the foot</title></head> <body> <leg> <footer><?php echo (dungSift($_SERVER['HTTP_USER_AGENT'], "ie"))?("Your foot is safe, but you might want to wear a hard hat!"):("<div class=\"shot\">BANG!</div>"); ?></footer></leg> </body> </html>

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