Search Results

Search found 5929 results on 238 pages for 'node webkit'.

Page 220/238 | < Previous Page | 216 217 218 219 220 221 222 223 224 225 226 227  | Next Page >

• Why MySQL sat for 2 minutes doing nothing?

  - by Alex R

  This was a one-time thing, not reproducible... But I saved the show innodb status output. Can anybody tell what's going on here? The simple insert took almost 3 minutes to complete. | InnoDB | | ===================================== 110201 15:58:10 INNODB MONITOR OUTPUT ===================================== Per second averages calculated from the last 34 seconds ---------- SEMAPHORES ---------- OS WAIT ARRAY INFO: reservation count 11963, signal count 11766 --Thread 1824 has waited at .\btr\btr0cur.c line 443 for 118.00 seconds the sema phore: S-lock on RW-latch at 09D6453C created in file .\buf\buf0buf.c line 550 a writer (thread id 1824) has reserved it in mode wait exclusive number of readers 1, waiters flag 1 Last time read locked in file .\buf\buf0flu.c line 599 Last time write locked in file .\btr\btr0cur.c line 443 Mutex spin waits 0, rounds 527817, OS waits 7133 RW-shared spins 2532, OS waits 1226; RW-excl spins 1652, OS waits 1118 ------------ TRANSACTIONS ------------ Trx id counter 0 95830 Purge done for trx's n:o < 0 95814 undo n:o < 0 0 History list length 11 LIST OF TRANSACTIONS FOR EACH SESSION: ---TRANSACTION 0 0, not started, OS thread id 3704 MySQL thread id 551, query id 2702112 localhost 127.0.0.1 root show innodb status ---TRANSACTION 0 95829, not started, OS thread id 3132 MySQL thread id 534, query id 2702020 localhost 127.0.0.1 root ---TRANSACTION 0 95828, not started, OS thread id 3152 MySQL thread id 527, query id 2701973 localhost 127.0.0.1 root ---TRANSACTION 0 95827, ACTIVE 118 sec, OS thread id 1824 inserting, thread decl ared inside InnoDB 500 mysql tables in use 1, locked 1 1 lock struct(s), heap size 320, 0 row lock(s) MySQL thread id 526, query id 2701972 localhost 127.0.0.1 root update INSERT INTO log_searchcriteria (userid,search_criteria,date,search_type) VALUES ( NAME_CONST('userid',NULL), NAME_CONST('search_criteria',_latin1' SELECT SQL_C ALC_FOUND_ROWS idx_search.CTCX_LATITUDE, idx_search.CTCX_LONGITUDE, idx_search.b uilding_id, idx_search.LN_LIST_NUMBER, idx_search.LP_LIST_PRICE, idx_search.HSN_ ADRESS_HOUSE_NUMBER, idx_search.STR_ADDRESS_STREET, idx_search.CP_ADDRESS_COMPAS S_POINT, idx_search.UN_UNIT, idx_search.CIT_CITY, idx_search.ZP_ZIP_CODE, idx_se arch.AR_AREA_NAME, idx_search.BR_BEDROOMS, idx_search.BTH_BATHS, idx_search.ST_S TATUS, idx_search.CTCX_STYLE_TYPE, idx_s -------- FILE I/O -------- I/O thread 0 state: wait Windows aio (insert buffer thread) I/O thread 1 state: wait Windows aio (log thread) I/O thread 2 state: wait Windows aio (read thread) I/O thread 3 state: wait Windows aio (write thread) Pending normal aio reads: 0, aio writes: 1, ibuf aio reads: 0, log i/o's: 0, sync i/o's: 0 Pending flushes (fsync) log: 0; buffer pool: 0 151006 OS file reads, 120758 OS file writes, 6844 OS fsyncs 0.00 reads/s, 0 avg bytes/read, 0.00 writes/s, 0.00 fsyncs/s ------------------------------------- INSERT BUFFER AND ADAPTIVE HASH INDEX ------------------------------------- Ibuf: size 1, free list len 5, seg size 7, 24664 inserts, 24664 merged recs, 4612 merges Hash table size 553253, node heap has 629 buffer(s) 0.00 hash searches/s, 0.00 non-hash searches/s --- LOG --- Log sequence number 5 2318193115 Log flushed up to 5 2318193115 Last checkpoint at 5 2318129891 0 pending log writes, 0 pending chkp writes 3036 log i/o's done, 0.00 log i/o's/second ---------------------- BUFFER POOL AND MEMORY ---------------------- Total memory allocated 213459462; in additional pool allocated 1720192 Dictionary memory allocated 240416 Buffer pool size 8192 Free buffers 0 Database pages 7563 Modified db pages 18 Pending reads 0 Pending writes: LRU 0, flush list 18, single page 0 Pages read 150973, created 28788, written 115137 0.00 reads/s, 0.00 creates/s, 0.00 writes/s No buffer pool page gets since the last printout -------------- ROW OPERATIONS -------------- 1 queries inside InnoDB, 0 queries in queue 1 read views open inside InnoDB Main thread id 2992, state: flushing buffer pool pages Number of rows inserted 794294, updated 89203, deleted 13698, read 1453084305 0.00 inserts/s, 0.00 updates/s, 0.00 deletes/s, 0.00 reads/s ---------------------------- END OF INNODB MONITOR OUTPUT ============================ Thanks

  Read the article

• multiple puppet masters set up using inventory

  - by Oli

  I have managed to set up multiple puppet masters with one puppet master acting as a CA and clients are able to get a certificate from this CA server but use their designated puppet master to get their manifests. See this question for more info.. multiple puppet masters. However, there are a couple of things I have had to do to get this working correctly and have an error which I'll get to. First of all, to get inventory working for a puppet-client (PC) connecting to its designated puppet-master (PM), I had to copy the CA certs on PM1 to the PM2 ca directory. I ran this command: scp [email protected]:/var/lib/puppet/ssl/ca/* [email protected]:/var/lib/puppet/ssl/ca/. Once i have done that, I was able to uncomment the SSLCertificateChainFile, SSLCACertificateFile & SSLCARevocationFile section of my rack.conf VH file on the PM2. Once I had done this, inventory started to work. Does this sound an acceptable way to do things? Secondly, in the puppet.conf file, I am setting the designated PM server for that client. Unless there is a better way, this is how it'll work in my production setup. So PC1 will talk to PM1 and PC2 will talk to PM2. This is where I have an error. When PC2 first requests a cert from the CA on PM1, the cert appears and then I sign the cert on the CA on PM1. When I then do a puppet agent --test on PC2 (which has server = PM2 in puppet.conf), I get this error: Warning: Unable to fetch my node definition, but the agent run will continue: Warning: Error 403 on SERVER: Forbidden request: puppet-master2.test.net(10.1.1.161) access to /certificate_revocation_list/ca [find] at :112 However, if I change the PC2 puppet.conf file and specify server = PM1 and the rerun puppet agent --test, i do not get any errors. I can then revert the change in the puppet.conf file back to server = PM2 and everything seems to run normally. Do I have to set up some kind of ProxyPassMatch on PM2 for requests made from clients to /certificate_revocation_list/* and redirect them to PM1? Or how can I fix this error? Cheers, Oli

  Read the article

• PHP at the root directory using Ngnix on Linode and Ubuntu 12.04

  - by Steve Kinney

  I originally set up my Linode to use it with the Sinatra applications using Phusion Passenger that I was developing and I have it working great for that. However, as time goes on, I find myself needing just a wee bit of PHP to do a server-side thing here or there. My basic set up was based off of this Linode recipe (I copied and pasted the parts that I needed—I did not install Redis and Node). If I go to http://scholarsnyc.com/index.php everything works great. If I just go the base URL however, I get a 403 Forbidden error (I have a vanilla HTML page there for now). I've played with file permissions and the same file will work if I call it directly. I've done my homework and nothing I try seems to work. I'm sure there is an obvious error. I'm also sure that there are some rookie mistakes in my Nginx configuration (some of those mistakes are the artifacts of trying different fixes from my research. user www-data www-data; worker_processes 1; events { worker_connections 1024; } upstream php { server 127.0.0.1:9001; } http { passenger_root /usr/local/lib/ruby/gems/1.9.1/gems/passenger-3.0.12; passenger_ruby /usr/local/bin/ruby; include mime.types; default_type application/octet-stream; index index.php index.html index.htm; sendfile on; keepalive_timeout 65; server { server_name localhost scholarsnyc.com www.scholarsnyc.com; root /srv/www/scholarsnyc.com/public; location / { index index.php; } location ~ \.php$ { fastcgi_pass 127.0.0.1:9000; fastcgi_param SCRIPT_FILENAME $document_root$fastcgi_script_name; include fastcgi_params; } } server { server_name data.scholarsnyc.com; root /srv/www/data.scholarsnyc.com/public; passenger_enabled on; } server { server_name tech.scholarsnyc.com; root /srv/www/tech.scholarsnyc.com/public; location / { root /srv/www/tech.scholarsnyc.com/public; index index.php index.html index.htm; } } } Any other optimizations are also appreciated. I literally don't know what to do at this point.

  Read the article

• Can Vagrant point to a directory of Puppet manifests for execution?

  - by SeligkeitIstInGott

  I am using Vagrant to jump start some initial Puppet config and am confused on how to include/run multiple manifests (other than just site.pp) in the puppet execution workflow without making the extra manifests into modules and including them that way. In the puppet manifests directory that I point Vagrant to (see below) I have two manifests that I want executed: site.pp and hierasetup.pp. config.vm.provision "puppet" do |puppet| puppet.manifests_path = "puppet_files/manifests" puppet.module_path = "puppet_files/modules" puppet.manifest_file = "site.pp" puppet.options = "--verbose --debug" end Currently I am having site.pp be the manifest that calls hierasetup.pp. My site.pp looks like this: File { owner => 'root', group => 'root', mode => '0644', } import "hierasetup.pp" include jboss But I get this error about the deprecation of "import": Warning: The use of 'import' is deprecated at /tmp/vagrant-puppet-1/manifests/site.pp:33. See http://links.puppetlabs.com/puppet-import-deprecation (at grammar.ra:610:in `_reduce_190') According to the referenced URL under "Things to try instead" it says "To keep your node definitions in separate files, specify a directory as your main manifest". Further this puppet doc on main manifests says: "Recommended: If you’re using the main manifest heavily instead of relying on an ENC, consider changing the manifest setting to $confdir/manifests. This lets you split up your top-level code into multiple files while avoiding the import keyword. It will also match the behavior of simple environments." It appears that Puppet can reference an entire directory instead of just a specific manifest file, such that I would expect that Vagrant would make a provision for this and allow me to drop the "puppet.manifest_file = "site.pp" line and point to the parent directory instead in which all the *.pp files there will be executed. However removing that line in Vagrant merely generates a complaint about an expected "default.pp" in its stead: puppet provisioner: * The configured Puppet manifest is missing. Please specify a path to an existing manifest: /some/path/puppet_files/manifests/default.pp So: Firstly, do I understand the "new" (non-import) way of calling multiple manifests correctly, in that a directory is to be pointed to in which all the *.pp files inside it will be executed? And secondly, has Vagrant "caught up" with this new change to accommodate the referencing of directories in conjunction with Puppet's deprecation of "import"? Update: Thanks to Shane the issue with #2 (Vagrant's code not being caught up to allow pointing to puppet manifest directories) was reported on Vagrant's GitHub issue tracker site and has since been patched: https://github.com/mitchellh/vagrant/issues/4169

  Read the article

• Innodb Queries Slow

  - by user105196

  I have redHat 5.3 (Tikanga) with Mysql 5.0.86 configued with RIAD 10 HW, I run an application inquiries from Mysql/InnoDB and MyIsam tables, the queries are super fast,but some quires on Innodb tables sometime slow down and took more than 1-3 seconds to run and these queries are simple and optimized, this problem occurred just on innodb tables in different time with random queries. Why is this happening only to Innodb tables? the below is the Innodb status and some Mysql variables: show innodb status\G ************* 1. row ************* Status: 120325 10:54:08 INNODB MONITOR OUTPUT Per second averages calculated from the last 19 seconds SEMAPHORES OS WAIT ARRAY INFO: reservation count 22943, signal count 22947 Mutex spin waits 0, rounds 561745, OS waits 7664 RW-shared spins 24427, OS waits 12201; RW-excl spins 1461, OS waits 1277 TRANSACTIONS Trx id counter 0 119069326 Purge done for trx's n:o < 0 119069326 undo n:o < 0 0 History list length 41 Total number of lock structs in row lock hash table 0 LIST OF TRANSACTIONS FOR EACH SESSION: ---TRANSACTION 0 0, not started, process no 29093, OS thread id 1166043456 MySQL thread id 703985, query id 5807220 localhost root show innodb status FILE I/O I/O thread 0 state: waiting for i/o request (insert buffer thread) I/O thread 1 state: waiting for i/o request (log thread) I/O thread 2 state: waiting for i/o request (read thread) I/O thread 3 state: waiting for i/o request (write thread) Pending normal aio reads: 0, aio writes: 0, ibuf aio reads: 0, log i/o's: 0, sync i/o's: 0 Pending flushes (fsync) log: 0; buffer pool: 0 132777 OS file reads, 689086 OS file writes, 252010 OS fsyncs 0.00 reads/s, 0 avg bytes/read, 0.00 writes/s, 0.00 fsyncs/s INSERT BUFFER AND ADAPTIVE HASH INDEX Ibuf: size 1, free list len 366, seg size 368, 62237 inserts, 62237 merged recs, 52881 merges Hash table size 8850487, used cells 3698960, node heap has 7061 buffer(s) 0.00 hash searches/s, 0.00 non-hash searches/s LOG Log sequence number 15 3415398745 Log flushed up to 15 3415398745 Last checkpoint at 15 3415398745 0 pending log writes, 0 pending chkp writes 218214 log i/o's done, 0.00 log i/o's/second BUFFER POOL AND MEMORY Total memory allocated 4798817080; in additional pool allocated 12342784 Buffer pool size 262144 Free buffers 101603 Database pages 153480 Modified db pages 0 Pending reads 0 Pending writes: LRU 0, flush list 0, single page 0 Pages read 151954, created 1526, written 494505 0.00 reads/s, 0.00 creates/s, 0.00 writes/s No buffer pool page gets since the last printout ROW OPERATIONS 0 queries inside InnoDB, 0 queries in queue 1 read views open inside InnoDB Main thread process no. 29093, id 1162049856, state: waiting for server activity Number of rows inserted 77675, updated 85439, deleted 0, read 14377072495 0.00 inserts/s, 0.00 updates/s, 0.00 deletes/s, 0.00 reads/s END OF INNODB MONITOR OUTPUT 1 row in set, 1 warning (0.02 sec) read_buffer_size = 128M sort_buffer_size = 256M tmp_table_size = 1024M innodb_additional_mem_pool_size = 20M innodb_log_file_size=10M innodb_lock_wait_timeout=100 innodb_buffer_pool_size=4G join_buffer_size = 128M key_buffer_size = 1G can any one help me ?

  Read the article

• Ubuntu getting wrong hostname from DHCP

  - by sam

  When provisioning new Ubuntu Precise (12.04) servers, the hostname they're getting seems to be generated from the DNS search path, not a reverse lookup on the hostname. Take the following configuration BIND is configured with the hostname, and reverse name Normal zone $TTL 600 $ORIGIN srv.local.net. @ IN SOA ns0.local.net. hostmaster.local.net. ( 2014082101 10800 3600 604800 600 ) @ IN NS ns0.local.net. @ IN MX 5 mail.local.net. my-new-server IN A 10.32.2.30 And reverse @ IN SOA ns0.local.net. hostmaster.local.net. ( 2014082101 10800 3600 604800 600 ) @ IN NS ns0.local.net. $ORIGIN 32.10.in-addr.arpa. 30.2 IN PTR my-new-server.srv.local.net. Then DHCPD is configured to hand out static leases based on mac addresses like so subnet 10.32.2.0 netmask 255.255.254.0 { option subnet-mask 255.255.254.0; option routers 10.32.2.1; option domain-name-servers 10.32.2.1; option domain-name "util.of1.local.net of1.local.net srv.local.net"; site-option-space "pxelinux"; option pxelinux.magic f1:00:74:7e; if exists dhcp-parameter-request-list { option dhcp-parameter-request-list = concat(option dhcp-parameter-request-list,d0,d1,d2,d3); } group { option pxelinux.configfile "pxelinux.cfg/pxeboot"; host my-new-server { fixed-address my-new-server.srv.local.net; hardware ethernet aa:aa:aa:bb:bb:bb; } } } So the hostname should be my-new-server.srv.local.net, however when building a Ubuntu 12.04 node, the hostname ends up as my-new-server.util.of1.local.net When building Lucid (10.04) hosts, the hostname will be correct, it's only on Precise/12.04 nodes we have the problem. Doing a normal and reverse lookup on the host and IP returns the correct result Sams-MacBook-Pro:~ sam$ host my-new-server my-new-server.srv.local.net has address 10.32.2.30 Sams-MacBook-Pro:~ sam$ host my-new-server.srv.local.net my-new-server.srv.local.net has address 10.32.2.30 Sams-MacBook-Pro:~ sam$ host 10.32.2.30 30.2.32.10.in-addr.arpa domain name pointer my-new-server.srv.local.net. The contents of the hosts file is incorrect too 127.0.0.1 localhost 127.0.1.1 my-new-server.util.of1.local.net of1.local.net srv.local.net my-new-server So it looks like when it creates the hosts file, it puts the entire contents of the DNS search path into the local address so the FQDN according to the server is the short hostname as defined, then the first domain in the search path. Is there a way to get around this behaviour, or fix this so it gets the hostname correctly? It's picking up the first part of the hostname, then the rest is wrong.

  Read the article

• How do I (robustly) remotely execute tasks on Windows workstations in a domain?

  - by Zac B

  I'm not even sure if "robustly" is a word. Anyway. Context: We have a few hundred Windows 7 workstations on a LAN. We use AD/GPO management pretty heavily, but there are a lot of periodic and/or manual maintenance tasks we need to do that can't be done via GPO/scheduled task. For example, say I want to execute program X (which runs silently, in the background, and doesn't bother the user) on workstation Y, or say I want to execute task A on a workstation group B either on a schedule or on demand. Kicking the users off of their computers to do this (i.e. using RDP) is a no-no, and doesn't work on groups anyway. Question: What's the best way to do this that is robust enough that, after setup, I could give it to beginner support people (read: people who are phobic of the command line, and get confused with GUI interfaces more complicated than Firefox)? I'm a competent programmer, and, if there is a robust set of tools or framework out there for this type of task, I'd consider hacking something together myself if it didn't take too long. If there's some combination of tools or techniques that others use to make remote-workstation-administration doable by beginners, I have yet to find it. For those who care about the "why": I'm midlevel IT, and was told to implement a remote management solution that allows arbitrary/scheduled remote execution, with confirmation that programs actually ran remotely, and the ability to view what they returned. "Why?" I asked, "Can't I just use PsExec and the task scheduler on a dispatcher machine?" "No," I was told, "'Joe' the second-week tech is going to be in charge of this one, and he needs something simple with a GUI." What I've tried: I've played with making a bunch of one-clickable "transfer files to remote computer and run them with PsExec" batch/VB scrips, but those tend to break down and don't easily support running on customizable groups. I've played a little bit with the Windows version of Puppet, but it doesn't support arbitrary-time remote execution (it's ability to group computers into a tree/node structure is really nice though). I've used an older version of Altiris, and, while it does a lot of what I want, it's interface is awful, it's slow, crashes a lot, and is probably too expensive for management. SwiftWater's DMS solution does some of what I want, but it's very underdeveloped, closed-source (not a deal breaker but not ideal), and I get the impression that support and reliability are lacking.

  Read the article

• I want to virtualize my workstation (Tier 1), Looking for Bare Metal Hypervisor for consumer grade components

  - by Chase Florell

  I find myself in this similar bind at least once a year. The bind whereby I'm either upgrading a motherboard, or an OS hard drive. It drives me crazy to have to reinstall Windows, Visual Studio, all my addins, reconfigure my settings etc... every single time. I have a layout and I like and I want to stick with it. My question is... Is there a Bare Metal Hypervisor on the market that will enable me to virtualize my consumer grade workstation? I really want to avoid Host/Client virtualization. Bare Metal is definitely a better way to go for my needs. Is this a good approach, or am I going to suffer some other undesirable side effects by doing this? Clarification My machine has very limited purposes. My primary use is Visual Studio 2010 Professional where I develop ASP.NET MVC Web Applications. The second piece of software that I use (that's system intensive) is Photoshop CS3. Beyond that, my applications are limited to Outlook, Internet Explorer, Firefox, Opera, Chrome, LinqPad, and various other (small) apps. Beyond this, I'm considering working on a node.js project and might run ubuntu on the same hypervisor if possible. System Specs: Gigabyte Motherboard Intel i7 920 12 GB Ram basic 500GB 7200RPM HDD for OS 4 VelociRaptors in Raid 1/0 for build disk Dual GTS250 (512MB) Graphics cards (non SLI) for quad monitors On a side note I also wouldn't be opposed to an alternative suggestion if the limitations are too great. I could install the ESXi (or Zen Server) on my box, and build a separate "thin client" to RDP into the virtual machine. It appears as though RDP supports dual monitors. Edit (Dec 9, 2011) It's been nearly a year since I first asked this question. Since then, there have been a lot of great strides in Hypervisor technology... AND MokaFive is now released for corporate use. I'd love to dig into this question a little more and find out if there is a solid BareMetal Hypervisor for workstations running consumer grade components (IE: not Dell, HP, Lenovo, Etc).

  Read the article

• recommendations for efficient offsite remote backup solution of vm's

  - by senorsmile

  I am looking for recommendations for backing up my current 6 vm's(and soon to grow to up to 20). Currently I am running a two node proxmox cluster(which is a debian base using kvm for virtualization with a custom web front end to administer). I have two nearly identical boxes with amd phenom II x4's and asus motherboards. Each has 4 500 GB sata2 hdd's, 1 for the os and other data for the proxmox install, and 3 using mdadm+drbd+lvm to share the 1.5 TB's of storage between the two machines. I mount lvm images to kvm for all of the virtual machines. I currently have the ability to do live transfer from one machine to the other, typically within seconds(it takes about 2 minutes on the largest vm running win2008 with m$ sql server). I am using proxmox's built-in vzdump utility to take snapshots of the vm's and store those on an external harddrive on the network. I then have jungledisk service (using rackspace) to sync the vzdump folder for remote offsite backup. This is all fine and dandy, but it's not very scalable. For one, the backups themselves can take up to a few hours every night. With jungledisk's block level incremental transfers, the sync only transfers a small portion of the data offsite, but that still takes at least a half an hour. The much better solution would of course be something that allows me to instantly take the difference of two time points (say what was written from 6am to 7am), zip it, then send that difference file to the backup server which would instantly transfer to the remote storage on rackspace. I have looked a little into zfs and it's ability to do send/receive. That coupled with a pipe of the data in bzip or something would seem perfect. However, it seems that implementing a nexenta server with zfs would essentially require at least one or two more dedicated storage servers to serve iSCSI block volumes (via zvol's???) to the proxmox servers. I would prefer to keep the setup as minimal as possible (i.e. NOT having separate storage servers) if at all possible. I have also briefly read about zumastor. It looks like it could also do what I want, but it appears to have halted development in 2008. So, zfs, zumastor or other?

  Read the article

• Sycronizing/deploying scripts across several systems

  - by otto

  I have a few time consuming tasks that I like to spread across several computers. These tasks require running an identical ruby or python script (or series of scripts that call each other) on each machine. The machines will a separate config file telling the script what portion of the task to complete. I want to figure out the best way to syncronize the scripts on these machines prior to running them. Up until now, I have been making changes to a copy of the script on a network share and then copying a fresh copy to each machine when I want to run it. But this is cumbersome and leaves a chance for error ( e.g missing a file on the copy or not clicking "copy and replace"). Lets assume the systems are standard windows machines that are not dedicated to this task and I don't need to run these scripts all the time (so I don't want a solution that runs 24/7 and always keeps them up to date, I'd prefer something that pushes/pulls on command). My thoughts on various options: Simple adaptation of my current workflow: Keep the originals on the network drive, but write a batch file that copies over the latest version of the scripts so everything is a one-click operation. Requires action on each system, but that's not the end of the world (since each one usually needs their configuration file changed slightly too). Put everything in a Mercurial/Git reposotory and pull a fresh copy onto each node. Going straight to the repo from each machine would guarantee a current version (and would have the fringe benefit of allowing edits to the script to be made from any machine). Cons would be that it requires VCS to be installed on each machine and there might be some pains dealing with authentication since I wouldn't use a public repo. Open up write access on a shared folder and write a script to use rsync (or similar) to push the changes out to all of the machines at once. This gets a current version on every machine (though you would have to change the script if you want to omit a machine or add a new one). Possible issue would be that each computer has to allow write access. Dropbox is a reasonable suggestion (and could work well) but I dont want to use an external service and I'd prefer not to have to have dropbox running 24/7 on systems that would normally not need it. Is there something simple that I am missing? Some tool designed expressly for doing this kind of thing? Otherwise I am leaning toward just tying all of the systems into Mercurial since, while it requires extra software, it is a little more robust than writing a batch file (e.g. if I split part of a script into a separate module, Mercurial will know what to do whereas I would have to add a line to the batch file).

  Read the article

• How does an NTP host switch among the various modes?

  - by James A. Rosen

  The NTPv3 RFC describes five operating modes: Symmetric Active (1): A host operating in this mode sends periodic messages regardless of the reachability state or stratum of its peer. By operating in this mode the host announces its willingness to synchronize and be synchronized by the peer. Symmetric Passive (2): This type of association is ordinarily created upon arrival of a message from a peer operating in the symmetric active mode and persists only as long as the peer is reachable and operating at a stratum level less than or equal to the host; otherwise, the association is dissolved. However, the association will always persist until at least one message has been sent in reply. By operating in this mode the host announces its willingness to synchronize and be synchronized by the peer. Client (3): A host operating in this mode sends periodic messages regardless of the reachability state or stratum of its peer. By operating in this mode the host, usually a LAN workstation, announces its willingness to be synchronized by, but not to synchronize the peer. Server (4): This type of association is ordinarily created upon arrival of a client request message and exists only in order to reply to that request, after which the association is dissolved. By operating in this mode the host, usually a LAN time server, announces its willingness to synchronize, but not to be synchronized by the peer. Broadcast (5): A host operating in this mode sends periodic messages regardless of the reachability state or stratum of the peers. By operating in this mode the host, usually a LAN time server operating on a high-speed broadcast medium, announces its willingness to synchronize all of the peers, but not to be synchronized by any of them. It seems to me, though, that any host except a leaf node would probably be in several modes. For example, I might have a local area network with three NTP servers, each in Symmetric Active (1) mode with respect to one another. They would also each be clients (3) of one of the many public stratum two time servers. Lastly, they would all server as servers (4) to the many local clients. Is the point that they're only in a given mode for a moment during the synchronization? If so, how does a host know to switch? I'm only looking for enough depth here to discuss the issue in an educated manner, not to write a custom time server.

  Read the article

• Which hardware to VM ratio for Build-Server virtualization?

  - by Martin

  Let's start with saying that I'm a total noob wrt. to server virtualization. That is, I use VMs often during development, but they're simple desktop machine things for me. Now to my problem: We have two (physical) build servers, one master, one slave running Jenkins to do daily tasks and build (Visual C++ Builds) our release packages for our software. As such these machines are critical to our company, because we do lot's releases and without a controlled environment to create them, we can't ship fixes. (And currently there's no proper backup of these machines in place, because they do not hold any data as such - it just would be a major pain to setup them again should they go bust. (But setting up backup that I'd know would work in case of HW failure would even be more pain, so we have skipped that until now.)) Therefore (and for scaling purposes) we would like to go virtual with these machines. Outsourcing to the cloud is not an option, not at all, so we'll have to use on-premises hardware and VM hosts. Each Build-Server (master or slave) is a fully configured (installs, licenses, shares in case of the master, ...) Windows Server box. I would now ideally like to just convert the (two) existing physical nodes to VM images and run them. Later add more VM slave instances as clones of the existing ones. And here begin my questions: Should I go for one VM per one hardware-box or should I go for something where a single hardware runs multiple VMs? That would mean a single point of failure hardware wise and doesn't seem like a good idea ... or?? Since we're doing C++ compilation with Visual Studio, I assume that during a build the hardware (processor cores + disk) will be fully utilized, so going with more than one build-node per hardware doesn't seem to make much sense?? Wrt. to hardware options, does it make any difference which VM software we use (VMWare, MS, Virtualbox, ... ?) (We're using Windows exclusively for our builds.) Regarding budget: We have a normal small company (20 developers) budget for this. ;-) That is, if it's going to cost a few k$ it's going to cost. If it's free - the better. I strongly prefer solutions where there's no multi-k$ maintenance costs per year.

  Read the article

• Tips for maximizing Nginx requests/sec?

  - by linkedlinked

  I'm building an analytics package, and project requirements state that I need to support 1 billion hits per day. Yep, "billion". In other words, no less than 12,000 hits per second sustained, and preferably some room to burst. I know I'll need multiple servers for this, but I'm trying to get maximum performance out of each node before "throwing more hardware at it". Right now, I have the hits-tracking portion completed, and well optimized. I pretty much just save the requests straight into Redis (for later processing with Hadoop). The application is Python/Django with a gunicorn for the gateway. My 2GB Ubuntu 10.04 Rackspace server (not a production machine) can serve about 1200 static files per second (benchmarked using Apache AB against a single static asset). To compare, if I swap out the static file link with my tracking link, I still get about 600 requests per second -- I think this means my tracker is well optimized, because it's only a factor of 2 slower than serving static assets. However, when I benchmark with millions of hits, I notice a few things -- No disk usage -- this is expected, because I've turned off all Nginx logs, and my custom code doesn't do anything but save the request details into Redis. Non-constant memory usage -- Presumably due to Redis' memory managing, my memory usage will gradually climb up and then drop back down, but it's never once been my bottleneck. System load hovers around 2-4, the system is still responsive during even my heaviest benchmarks, and I can still manually view http://mysite.com/tracking/pixel with little visible delay while my (other) server performs 600 requests per second. If I run a short test, say 50,000 hits (takes about 2m), I get a steady, reliable 600 requests per second. If I run a longer test (tried up to 3.5m so far), my r/s degrades to about 250. My questions -- a. Does it look like I'm maxing out this server yet? Is 1,200/s static files nginx performance comparable to what others have experienced? b. Are there common nginx tunings for such high-volume applications? I have worker threads set to 64, and gunicorn worker threads set to 8, but tweaking these values doesn't seem to help or harm me much. c. Are there any linux-level settings that could be limiting my incoming connections? d. What could cause my performance to degrade to 250r/s on long-running tests? Again, the memory is not maxing out during these tests, and HDD use is nil. Thanks in advance, all :)

  Read the article

• How should we serve files in a small bioinformatics cluster?

  - by cespinoza

  We have a small cluster of six ubuntu servers. We run bioinformatics analyses on these clusters. Each analysis takes about 24 hours to complete, each core i7 server can handle 2 at a time, takes as input about 5GB data and outputs about 10-25GB of data. We run dozens of these a week. The software is a hodgepodge of custom perl scripts and 3rd party sequence alignment software written in C/C++. Currently, files are served from two of the compute nodes (yes, we're using compute nodes as file servers)-- each node has 5 1TB sata drives mounted separately (no raid) and is pooled via glusterfs 2.0.1. They each have as 3 bonded intel ethernet pci gigabit ethernet cards, attached to a d-link DGS-1224T switch ($300 24 port consumer-level). We are not currently using jumbo frames (not sure why, actually). The two file-serving compute nodes are then mirrored via glusterfs. Each of the four other nodes mounts the files via glusterfs. The files are all large (4gb+), and are stored as bare files (no database/etc) if that matters. As you can imagine, this is a bit of a mess that grew organically without forethought and we want to improve it now that we're running out of space. Our analyses are I/O intensive and it is a bottle neck-- we're only getting 140mB/sec between the two fileservers, maybe 50mb/sec from the clients (which only have single NICs). We have a flexible budget which I can probably get up $5k or so. How should we spend our budget? We need at least 10TB of storage fast enough to serve all nodes. How fast/big does the cpu/memory of such a file server have to be? Should we use NFS, ATA over Ethernet, iSCSI, Glusterfs, or something else? Should we buy two or more servers and create some sort of storage cluster, or is 1 server enough for such a small number of nodes? Should we invest in faster NICs (say, PCI-express cards with multiple connectors)? The switch? Should we use raid, if so, hardware or software? and which raid (5, 6, 10, etc)? Any ideas appreciated. We're biologists, not IT gurus.

  Read the article

• ERROR: Linux route add command failed: external program exited with error status: 4

  - by JohnMerlino

  A remote machine running fedora uses openvpn, and multiple developers were successfully able to connect to it via their client openvpn. However, I am running Ubuntu 12.04 and I am having trouble connecting to the server via vpn. I copied ca.crt, home.key, and home.crt from the server to my local machine to /etc/openvpn folder. My client.conf file looks like this: ############################################## # Sample client-side OpenVPN 2.0 config file # # for connecting to multi-client server. # # # # This configuration can be used by multiple # # clients, however each client should have # # its own cert and key files. # # # # On Windows, you might want to rename this # # file so it has a .ovpn extension # ############################################## # Specify that we are a client and that we # will be pulling certain config file directives # from the server. client # Use the same setting as you are using on # the server. # On most systems, the VPN will not function # unless you partially or fully disable # the firewall for the TUN/TAP interface. ;dev tap dev tun # Windows needs the TAP-Win32 adapter name # from the Network Connections panel # if you have more than one. On XP SP2, # you may need to disable the firewall # for the TAP adapter. ;dev-node MyTap # Are we connecting to a TCP or # UDP server? Use the same setting as # on the server. ;proto tcp proto udp # The hostname/IP and port of the server. # You can have multiple remote entries # to load balance between the servers. remote xx.xxx.xx.130 1194 ;remote my-server-2 1194 # Choose a random host from the remote # list for load-balancing. Otherwise # try hosts in the order specified. ;remote-random # Keep trying indefinitely to resolve the # host name of the OpenVPN server. Very useful # on machines which are not permanently connected # to the internet such as laptops. resolv-retry infinite # Most clients don't need to bind to # a specific local port number. nobind # Downgrade privileges after initialization (non-Windows only) ;user nobody ;group nogroup # Try to preserve some state across restarts. persist-key persist-tun # If you are connecting through an # HTTP proxy to reach the actual OpenVPN # server, put the proxy server/IP and # port number here. See the man page # if your proxy server requires # authentication. ;http-proxy-retry # retry on connection failures ;http-proxy [proxy server] [proxy port #] # Wireless networks often produce a lot # of duplicate packets. Set this flag # to silence duplicate packet warnings. ;mute-replay-warnings # SSL/TLS parms. # See the server config file for more # description. It's best to use # a separate .crt/.key file pair # for each client. A single ca # file can be used for all clients. ca ca.crt cert home.crt key home.key # Verify server certificate by checking # that the certicate has the nsCertType # field set to "server". This is an # important precaution to protect against # a potential attack discussed here: # http://openvpn.net/howto.html#mitm # # To use this feature, you will need to generate # your server certificates with the nsCertType # field set to "server". The build-key-server # script in the easy-rsa folder will do this. ns-cert-type server # If a tls-auth key is used on the server # then every client must also have the key. ;tls-auth ta.key 1 # Select a cryptographic cipher. # If the cipher option is used on the server # then you must also specify it here. ;cipher x # Enable compression on the VPN link. # Don't enable this unless it is also # enabled in the server config file. comp-lzo # Set log file verbosity. verb 3 # Silence repeating messages ;mute 20 But when I start server and look in /var/log/syslog, I notice the following error: May 27 22:13:51 myuser ovpn-client[5626]: /sbin/route add -net 10.27.12.1 netmask 255.255.255.252 gw 10.27.12.37 May 27 22:13:51 myuser ovpn-client[5626]: ERROR: Linux route add command failed: external program exited with error status: 4 May 27 22:13:51 myuser ovpn-client[5626]: /sbin/route add -net 172.27.12.0 netmask 255.255.255.0 gw 10.27.12.37 May 27 22:13:51 myuser ovpn-client[5626]: /sbin/route add -net 10.27.12.1 netmask 255.255.255.255 gw 10.27.12.37 And I am unable to connect to the server via openvpn: $ ssh [email protected] ssh: connect to host xxx.xx.xx.130 port 22: No route to host What may I be doing wrong?

  Read the article

• Commercial Drupal Modules & Themes

  - by Ravish

  A discussion at Drupal.org forums prompted me to give my input about commercial ecosystem around Open Source Content Management Systems. WordPress and Joomla have been growing rapidly since past few years. But, growth rate of Drupal seems to be almost flat. Despite being the most powerful CMS around, Drupal is still not being adopted by masses. Many people will argue that Drupal is not targeted towards masses, but developers. I agree, Drupal is more of a development platform than a consumer CMS. Drupal is ‘many things to many people’, and I can build almost any type of website with it. Drupal is being used for building blogs, corporate websites, Intranet portals, social networking and even a project management system. Looking at the wide array of Drupal implementations, it deserves to be the most widely adopted CMS. I believe there are few challenges that Drupal community needs to overcome. To understand these challenges, I surveyed some webmasters who use Joomla or WordPress but not Drupal. I asked them why they don’t want to use Drupal, following are the responses I got from them: Drupal is too complicated, takes time to learn. Drupal is great, but its admin panel is overwhelming. I couldn’t find any nice themes for Drupal. There is no WYSIWYG editor in Drupal. Most Drupal modules do not work out of the box. There aren’t enough modules like Ubercart which provides any out of the box functionality. I tried modules like CCK, Views and Panels. After wasting several hours struggling with them, I decided to give up on Drupal. I don’t use Drupal because of pushbutton and Garland theme. I had hard time trying to customize Garland and it messed up the whole layout. There are no premium modules and themes for Drupal. Joomla has tons of awesome themes and modules. I don’t want a million hacks like CCK, Views, Tokens, Pathauto, ImageCache and CTools just to run a simple website. Most of the complaints from users are related to the learning and development curve involved with Drupal, and the lack of ecosystem. While most of the problems will be gone in Drupal 7, ecosystem is something that needs to be built by the Drupal community. Drupal distributions are a great step forward. There are few awesome Drupal distributions available like Open Publish, Open Atrium and Drupal Commons. I predict, there will be a wave of many powerful Drupal distributions after Drupal 7 release. Many of them will be user-friendly and commercial supported. Following is my post at Drupal.org forums: Quote from: http://drupal.org/node/863776#comment-3313836 Brian Gardner (StudioPress) and Woo Themes launched premium WordPress themes in 2007, the developer community did not accept it at first. Moreover, they were not even GPL licensed. There was an outcry in WordPress community against them. Following that, most premium theme providers switched to GPL licensing. Despite controversies, users voted for premium theme and plugins by buying them. Inspired by their success, hundreds of other developers started to sell premium themes and plugins. It is now the acceptable and in fact most popular business model among WordPress community. Matt Mullenweg once told me, they would not support premium themes. If he supported, developers would no more give out free GPL themes & plugins. He pointed me towards Joomla, there were hardly any nice free themes & modules available. Now two years forward, premium products are not just accepted but embraced by the WordPress community – http://wordpress.org/extend/themes/commercial/ The quality and number of themes & modules has increased, even the free ones. This also helped to boost the adoption and ecosystem of WordPress. Today, state of Drupal is like WordPress was in 2007. There are hardly any out of the box solutions available for Drupal. Ubercart, Open Publish and Open Atrium are the only ones I can think of. Many of the popular Drupal modules are patches and hole-fillers. Thankfully, these hole-filler modules are going to be in Drupal 7 core. Drupal 7 and distributions will spawn a new array of solutions built upon Drupal. Soon, we will have more like Ubercarts and Open Atriums. If commercial solutions can help fuel this ecosystem and growth, Drupal community will accept them eventually. This debate will not stop your customers from buying your product. If your product is awesome, they will vote for you by buying your product.

  Read the article

• How to Troubleshoot TFS Build Server Failure?

  - by Tarun Arora

  Ever found your self in this helpless situation where you think you have tried every possible suggestion on the internet to bring the build server back but it just won’t work. Well some times before hunting around for a solution it is important to understand what the problem is, if the error messages in the build logs don’t seem to help you can always enable tracing on the build server to get more information on what could possibly be the root cause of failure. In this blog post today I’ll be showing you how to enable tracing on, - TFS 2010/11 Server - Build Server - Client Enable Tracing on Team Foundation Server 2010/2011 On the Team Foundation Server navigate to C:\Program Files\Microsoft Team Foundation Server 2010\Application Tier\Web Services, right click web.config and from the context menu select edit.          Search for the <appSettings> node in the config file and set the value of the key ‘traceWriter’ to true.          In the <System.diagnostics> tag set the value of switches from 0 to 4 to set the trace level to maximum to write diagnostics level trace information.          Restart the TFS Application pool to force this change to take effect. The application pool restart will impact any one using the TFS server at present. Note - It is recommended that you do not make any changes to the TFS production application server, this can have serious consequences and can even jeopardize the installation of your server.          Download the Debug view tool from http://technet.microsoft.com/en-us/sysinternals/bb896647.aspx and set it to capture “Global Events”. Perform any actions in the Team Explorer on the client machine, you should be able to see a series of trace data in the debug view tool now.         Enable Tracing on Build Controller/Agents Log on to the Build Controller/Agent and Navigate to the directory C:\Program Files\Microsoft Team Foundation Server 2010\Tools         Look for the configuration file ‘TFSBuildServiceHost.exe.config’ if it is not already there create a new text file and rename it to ‘TFSBuildServiceHost.exe.config’         To Enable tracing uncomment the <system.diagnostics> and paste the snippet below if it is not already there. <configuration> <system.diagnostics> <switches> <add name="BuildServiceTraceLevel" value="4"/> </switches> <trace autoflush="true" indentsize="4"> <listeners> <add name="myListener" type="Microsoft.TeamFoundation.TeamFoundationTextWriterTraceListener, Microsoft.TeamFoundation.Common, Version=10.0.0.0, Culture=neutral, PublicKeyToken=b03f5f7f11d50a3a" initializeData="c:\logs\TFSBuildServiceHost.exe.log" /> <remove name="Default" /> </listeners> </trace> </system.diagnostics> </configuration> The highlighted path above is where the Log file will be created. If the folder is not already there then create the folder, also, make sure that the account running the build service has access to write to this folder.         Restart the build Controller/Agent service from the administration console (or net stop tfsbuildservicehost & net start tfsbuildservicehost) in order for the new setting to be picked up.         Enable TFS Tracing on the Client Machine On the client machine, shut down Visual Studio, navigate to C:\Program Files\Microsoft Visual Studio 10.0\Common 7\IDE          Search for devenv.exe.config, make a backup copy of the config file and right click the file and from the context menu select edit. If its not already there create this file.          Edit devenv.exe.config by adding the below code snippet before the last </configuration> tag <system.diagnostics> <switches> <add name="TeamFoundationSoapProxy" value="4" /> <add name="VersionControl" value="4" /> </switches> <trace autoflush="true" indentsize="3"> <listeners> <add name="myListener" type="Microsoft.TeamFoundation.TeamFoundationTextWriterTraceListener,Microsoft.TeamFoundation.Common, Version=10.0.0.0, Culture=neutral, PublicKeyToken=b03f5f7f11d50a3a" initializeData="c:\tf.log" /> <add name="perfListener" type="Microsoft.TeamFoundation.Client.PerfTraceListener, Microsoft.TeamFoundation.Client, Version=10.0.0.0, Culture=neutral, PublicKeyToken=b03f5f7f11d50a3a"/> </listeners> </trace> </system.diagnostics> The highlighted path above is where the Log file will be created. If the folder is not already there then create the folder. Start Visual Studio and after a bit of activity you should be able to see the new log file being created on the folder specified in the config file. Other Resources Below are some Key resource you might like to review. I would highly recommend the documentation, walkthroughs and videos available on MSDN.   Thank you for taking the time out and reading this blog post. If you enjoyed the post, remember to subscribe to http://feeds.feedburner.com/TarunArora. Have you come across an interesting one to one with the build server, please share your experience here. Questions/Feedback/Suggestions, etc please leave a comment. Thank You! Share this post : CodeProject

  Read the article

• Lessons from a SAN Failure

  - by Bill Graziano

  At 1:10AM Sunday morning the main SAN at one of my clients suffered a “partial” failure.  Partial means that the SAN was still online and functioning but the LUNs attached to our two main SQL Servers “failed”.  Failed means that SQL Server wouldn’t start and the MDF and LDF files mostly showed a zero file size.  But they were online and responding and most other LUNs were available.  I’m not sure how SANs know to fail at 1AM on a Saturday night but they seem to.  From a personal standpoint this worked out poorly: I was out with friends and after more than a few drinks.  From a work standpoint this was about the best time to fail you could imagine.  Everything was running well before Monday morning.  But it was a long, long Sunday.  I started tipsy, got tired and ended up hung over later in the day. Note to self: Try not to go out drinking right before the SAN fails. This caught us at an interesting time.  We’re in the process of migrating to an entirely new set of servers so some things were partially moved.  This made it difficult to follow our procedures as cleanly as we’d like.  The benefit was that we had much better documentation of everything on the server.  I would encourage everyone to really think through the process of implementing your DR plan and document as much as possible.  Following a checklist is much easier than trying to remember at night under pressure in a hurry after a few drinks. I had a series of estimates on how long things would take.  They were accurate for any single server failure.  They weren’t accurate for a SAN failure that took two servers down.  This wasn’t bad but we should have communicated better. Don’t forget how many things are outside the database.  Logins, linked servers, DTS packages (yikes!), jobs, service broker, DTC (especially DTC), database triggers and any objects in the master database are all things you need backed up.  We’d done a decent job on this and didn’t find significant problems here.  That said this still took a lot of time.  There were many annoyances as a result of this.  Small settings like a login’s default database had a big impact on whether an application could run.  This is probably the single biggest area of concern when looking to recreate a server.  I’d encourage everyone to go through every single node of SSMS and look for user created objects or settings outside the database. Script out your logins with the proper SID and already encrypted passwords and keep it updated.  This makes life so much easier.  I used an approach based on KB246133 that worked well.  I’ll get my scripts posted over the next few days. The disaster can cause your DR process to fail in unexpected ways.  We have a job that scripts out all logins and role memberships and writes it to a file.  This runs on the DR server and pulls from the production server.  Upon opening the file I found that the contents were a “server not found” error.  Fortunately we had other copies and didn’t need to try and restore the master database.  This now runs on the production server and pushes the script to the DR site.  Soon we’ll get it pushed to our version control software. One of the biggest challenges is keeping your DR resources up to date.  Any server change (new linked server, new SQL Server Agent job, etc.) means that your DR plan (and scripts) is out of date.  It helps to automate the generation of these resources if possible. Take time now to test your database restore process.  We test ours quarterly.  If you have a large database I’d also encourage you to invest in a compressed backup solution.  Restoring backups was the single larger consumer of time during our recovery. And yes, there’s a database mirroring solution planned in our new architecture. I didn’t have much involvement in things outside SQL Server but this caused many, many things to change in our environment.  Many applications today aren’t just executables or web sites.  They are a combination of those plus network infrastructure, reports, network ports, IP addresses, DTS and SSIS packages, batch systems and many other things.  These all needed a little bit of attention to make sure they were functioning properly. Profiler turned out to be a handy tool.  I started a trace for failed logins and kept that running.  That let me fix a number of problems before people were able to report them.  I also ran traces to capture exceptions.  This helped identify problems with linked servers. Overall the thing that gave me the most problem was linked servers.  In order for a linked server to function properly you need to be pointed to the right server, have the proper login information, have the network routes available and have MSDTC configured properly.  We have a lot of linked servers and this created many failure points.  Some of the older linked servers used IP addresses and not DNS names.  This meant we had to go in and touch all those linked servers when the servers moved.

  Read the article

• Android exception i don't understand after loading webpage in a webview

  - by DixieFlatline

  I have a webview that loads a webpage. I also have a reload button. Sometimes it works but sometimes it crashes when i hit reload and i get this exceptions: 05-14 10:08:33.958: ERROR/WindowManager(918): Activity com.poslji.gor.Uvod has leaked window com.android.internal.policy.impl.PhoneWindow$DecorView@435da698 that was originally added here 05-14 10:08:33.958: ERROR/WindowManager(918): android.view.WindowLeaked: Activity com.poslji.gor.Uvod has leaked window com.android.internal.policy.impl.PhoneWindow$DecorView@435da698 that was originally added here 05-14 10:08:33.958: ERROR/WindowManager(918): at android.view.ViewRoot.(ViewRoot.java:217) 05-14 10:08:33.958: ERROR/WindowManager(918): at android.view.WindowManagerImpl.addView(WindowManagerImpl.java:148) 05-14 10:08:33.958: ERROR/WindowManager(918): at android.view.WindowManagerImpl.addView(WindowManagerImpl.java:91) 05-14 10:08:33.958: ERROR/WindowManager(918): at android.view.Window$LocalWindowManager.addView(Window.java:392) 05-14 10:08:33.958: ERROR/WindowManager(918): at android.app.Dialog.show(Dialog.java:231) 05-14 10:08:33.958: ERROR/WindowManager(918): at android.app.ProgressDialog.show(ProgressDialog.java:107) 05-14 10:08:33.958: ERROR/WindowManager(918): at android.app.ProgressDialog.show(ProgressDialog.java:90) 05-14 10:08:33.958: ERROR/WindowManager(918): at com.poslji.gor.Odgovori$2.onClick(Odgovori.java:120) 05-14 10:08:33.958: ERROR/WindowManager(918): at android.view.View.performClick(View.java:2179) 05-14 10:08:33.958: ERROR/WindowManager(918): at android.view.View.onTouchEvent(View.java:3828) 05-14 10:08:33.958: ERROR/WindowManager(918): at android.widget.TextView.onTouchEvent(TextView.java:6307) 05-14 10:08:33.958: ERROR/WindowManager(918): at android.view.View.dispatchTouchEvent(View.java:3368) 05-14 10:08:33.958: ERROR/WindowManager(918): at android.view.ViewGroup.dispatchTouchEvent(ViewGroup.java:903) 05-14 10:08:33.958: ERROR/WindowManager(918): at android.view.ViewGroup.dispatchTouchEvent(ViewGroup.java:903) 05-14 10:08:33.958: ERROR/WindowManager(918): at android.view.ViewGroup.dispatchTouchEvent(ViewGroup.java:903) 05-14 10:08:33.958: ERROR/WindowManager(918): at com.android.internal.policy.impl.PhoneWindow$DecorView.superDispatchTouchEvent(PhoneWindow.java:1752) 05-14 10:08:33.958: ERROR/WindowManager(918): at com.android.internal.policy.impl.PhoneWindow.superDispatchTouchEvent(PhoneWindow.java:1206) 05-14 10:08:33.958: ERROR/WindowManager(918): at android.app.Activity.dispatchTouchEvent(Activity.java:1997) 05-14 10:08:33.958: ERROR/WindowManager(918): at com.android.internal.policy.impl.PhoneWindow$DecorView.dispatchTouchEvent(PhoneWindow.java:1736) 05-14 10:08:33.958: ERROR/WindowManager(918): at android.view.ViewGroup.dispatchTouchEvent(ViewGroup.java:903) 05-14 10:08:33.958: ERROR/WindowManager(918): at android.view.ViewGroup.dispatchTouchEvent(ViewGroup.java:903) 05-14 10:08:33.958: ERROR/WindowManager(918): at android.view.ViewGroup.dispatchTouchEvent(ViewGroup.java:903) 05-14 10:08:33.958: ERROR/WindowManager(918): at android.view.ViewGroup.dispatchTouchEvent(ViewGroup.java:903) 05-14 10:08:33.958: ERROR/WindowManager(918): at android.view.ViewGroup.dispatchTouchEvent(ViewGroup.java:903) 05-14 10:08:33.958: ERROR/WindowManager(918): at com.android.internal.policy.impl.PhoneWindow$DecorView.superDispatchTouchEvent(PhoneWindow.java:1752) 05-14 10:08:33.958: ERROR/WindowManager(918): at com.android.internal.policy.impl.PhoneWindow.superDispatchTouchEvent(PhoneWindow.java:1206) 05-14 10:08:33.958: ERROR/WindowManager(918): at android.app.Activity.dispatchTouchEvent(Activity.java:1997) 05-14 10:08:33.958: ERROR/WindowManager(918): at com.android.internal.policy.impl.PhoneWindow$DecorView.dispatchTouchEvent(PhoneWindow.java:1736) 05-14 10:08:33.958: ERROR/WindowManager(918): at android.view.ViewRoot.handleMessage(ViewRoot.java:1761) 05-14 10:08:33.958: ERROR/WindowManager(918): at android.os.Handler.dispatchMessage(Handler.java:99) 05-14 10:08:33.958: ERROR/WindowManager(918): at android.os.Looper.loop(Looper.java:123) 05-14 10:08:33.958: ERROR/WindowManager(918): at android.app.ActivityThread.main(ActivityThread.java:3948) 05-14 10:08:33.958: ERROR/WindowManager(918): at java.lang.reflect.Method.invokeNative(Native Method) 05-14 10:08:33.958: ERROR/WindowManager(918): at java.lang.reflect.Method.invoke(Method.java:521) 05-14 10:08:33.958: ERROR/WindowManager(918): at com.android.internal.os.ZygoteInit$MethodAndArgsCaller.run(ZygoteInit.java:782) 05-14 10:08:33.958: ERROR/WindowManager(918): at com.android.internal.os.ZygoteInit.main(ZygoteInit.java:540) 05-14 10:08:33.958: ERROR/WindowManager(918): at dalvik.system.NativeStart.main(Native Method) 05-14 10:08:36.768: ERROR/AndroidRuntime(918): Uncaught handler: thread main exiting due to uncaught exception 05-14 10:08:36.778: ERROR/AndroidRuntime(918): java.lang.IllegalArgumentException: View not attached to window manager 05-14 10:08:36.778: ERROR/AndroidRuntime(918): at android.view.WindowManagerImpl.findViewLocked(WindowManagerImpl.java:356) 05-14 10:08:36.778: ERROR/AndroidRuntime(918): at android.view.WindowManagerImpl.removeView(WindowManagerImpl.java:201) 05-14 10:08:36.778: ERROR/AndroidRuntime(918): at android.view.Window$LocalWindowManager.removeView(Window.java:400) 05-14 10:08:36.778: ERROR/AndroidRuntime(918): at android.app.Dialog.dismissDialog(Dialog.java:268) 05-14 10:08:36.778: ERROR/AndroidRuntime(918): at android.app.Dialog.access$000(Dialog.java:69) 05-14 10:08:36.778: ERROR/AndroidRuntime(918): at android.app.Dialog$1.run(Dialog.java:103) 05-14 10:08:36.778: ERROR/AndroidRuntime(918): at android.app.Dialog.dismiss(Dialog.java:252) 05-14 10:08:36.778: ERROR/AndroidRuntime(918): at com.poslji.gor.Odgovori$HelloWebViewClient.onPageFinished(Odgovori.java:180) 05-14 10:08:36.778: ERROR/AndroidRuntime(918): at android.webkit.CallbackProxy.handleMessage(CallbackProxy.java:225) 05-14 10:08:36.778: ERROR/AndroidRuntime(918): at android.os.Handler.dispatchMessage(Handler.java:99) 05-14 10:08:36.778: ERROR/AndroidRuntime(918): at android.os.Looper.loop(Looper.java:123) 05-14 10:08:36.778: ERROR/AndroidRuntime(918): at android.app.ActivityThread.main(ActivityThread.java:3948) 05-14 10:08:36.778: ERROR/AndroidRuntime(918): at java.lang.reflect.Method.invokeNative(Native Method) 05-14 10:08:36.778: ERROR/AndroidRuntime(918): at java.lang.reflect.Method.invoke(Method.java:521) 05-14 10:08:36.778: ERROR/AndroidRuntime(918): at com.android.internal.os.ZygoteInit$MethodAndArgsCaller.run(ZygoteInit.java:782) 05-14 10:08:36.778: ERROR/AndroidRuntime(918): at com.android.internal.os.ZygoteInit.main(ZygoteInit.java:540) 05-14 10:08:36.778: ERROR/AndroidRuntime(918): at dalvik.system.NativeStart.main(Native Method) What is going wrong here?

  Read the article

• How to create a new Team Project Collection in TFS2010:

  - by jehan

  TFS 2010 has introduced the notion of Team Project Collection (TPC).  I have already discussed about TPC in my earlier post, you can check it out here. In this post, I will demonstrate how to create a new Team Project Collection in TFS2010. First, you have to open the TFS Administration Console (Start à All Programs à Microsoft Team Foundation Server 2010 à Team Foundation Server Administration Console), expand the Application Tier node in TFS Administration Console and click on Team Project Collection. Here you will see the TPC’s which are already exist, I am having only one TPC named New Collection and I’m going to create a new TPC called Demo Collection. To create a new Team Project Collection, you need to click on Create Collection; it will open the Create New Team Project Collection window.     Under the Name tab, you have to enter the name of Collection which you want to give for your new TPC (I naming it as Demo Collection). You can also provide some description about your TPC in Description tab which is optional and click next. Here, you need to enter the name of SQL Server Instance where you want your new TPC data to reside. You have the option either to choose the creating a Database for this TPC or use the already existing empty database and then click next.   In next screen, you have to choose SharePoint configuration. Here you have the options to either configure SharePoint Site for TPC at default collections or you can specify the your existing SharePoint site and  you can also choose not  to configure the SharePoint for this collection, if you choose last option then you cannot configure the Share Point sites for the all the Team Projects under this Project Collection. You also have the flexibility to create a Share Point site for this TPC later on, then if you need you have to configure SharePoint site for the existing team projects manually.   In next screen, you will have the Reports configuration. Here you have the options to either configure the Reports for TPC at default path or you can specify the path for at existing Reports folder, you can also choose not to configure the Reports for this collection, if you choose last option then you cannot create  the Reports  for the all the Team Projects under this Project Collection. Here also you can enable reporting for this TPC later on. The next screen is related to Lab Management Configuration, Lab Management is the new feature in TFS2010 which enables the users to create and manage virtual test environments where you can deploy and test your application. There are no options available here as I don’t have the Lab Management configured for my Team Foundation Server. The next screen is Review Configuration window, which will show up all the configuration settings you have specified, so that you can review the configurations before creating the Team Project Collection. If you want to make any changes to the configurations then you can go back to the previous windows and can make the changes. After Reviewing the configuration settings, you can click on verify button. Which will verify that if you’re Team Project Collection is ready to be created or not, it will show up the errors and warning (if any) which can make your Team Project Collection fail. You can then choose to create the Team Project Collection if the verify option doesn’t throw any warnings and errors. If the verify option throws any errors, then it is strongly suggested that you have to first rectify the issues then only go for TPC creation especially in case of warnings as it is a common practice to overlook the warnings.   If you choose the create TPC option, then it will start the process of creating a Team Project Collection  and once its completed you can check the status of configuration different components  during Team Project Collection. You can see in below screen that all the components are configured successfully.   In next screen, you can find the location of log file created for this Team Project Creation, this log file is really important in case of Team Project creation failure because it will help you to find  the root cause for the failure. Now, you can see that the New Team Projection (Demo Collection) which was created is now available in Team Foundation Collection tab and its status is Online.   You can now try to connect to this Team Project Collection from Team Explorer. Choose the newly created Team Project Collection and click on connect.     This Team Project Collection is empty because no Team Projects are created yet. Now, you can create the new Team Projects and start working.

  Read the article

• .NET 3.5 Installation Problems in Windows 8

  - by Rick Strahl

  Windows 8 installs with .NET 4.5. A default installation of Windows 8 doesn't seem to include .NET 3.0 or 3.5, although .NET 2.0 does seem to be available by default (presumably because Windows has app dependencies on that). I ran into some pretty nasty compatibility issues regarding .NET 3.5 which I'll describe in this post. I'll preface this by saying that depending on how you install Windows 8 you may not run into these issues. In fact, it's probably a special case, but one that might be common with developer folks reading my blog. Specifically it's the install order that screwed things up for me -  installing Visual Studio before explicitly installing .NET 3.5 from Windows Features - in particular. If you install Visual Studio 2010 I highly recommend you install .NET 3.5 from Windows features BEFORE you install Visual Studio 2010 and save yourself the trouble I went through. So when I installed Windows 8, and then looked at the Windows Features to install after the fact in the Windows Feature dialog, I thought - .NET 3.5 - who needs it. I'd be happy to not have to install .NET 3.5, but unfortunately I found out quite a while after initial installation that one of my applications/tools (DevExpress's awesome CodeRush) depends on it and won't install without it. Enabling .NET 3.5 in Windows 8 If you want to run .NET 3.5 on Windows 8, don't download an installer - those installers don't work on Windows 8, and you don't need to do this because you can use the Windows Features dialog to enable .NET 3.5: And that *should* do the trick. If you do this before you install other apps that require .NET 3.5 and install a non-SP1 one version of it, you are going to have no problems. Unfortunately for me, even after I've installed the above, when I run the CodeRush installer I still get this lovely dialog: Now I double checked to see if .NET 3.5 is installed - it is, both for 32 bit and 64 bit. I went as far as creating a small .NET Console app and running it to verify that it actually runs. And it does… So naturally I thought the CodeRush installer is a little whacky. After some back and forth Alex Skorkin on Twitter pointed me in the right direction: He asked me to look in the registry for exact info on which version of .NET 3.5 is installed here: HKEY_LOCAL_MACHINE\SOFTWARE\Microsoft\NET Framework Setup\NDP where I found that .NET 3.5 SP1 was installed. This is the 64 bit key which looks all correct. However, when I looked under the 32 bit node I found: HKEY_LOCAL_MACHINE\SOFTWARE\Wow6432Node\Microsoft\NET Framework Setup\NDP\v3.5 Notice that the service pack number is set to 0, rather than 1 (which it was for the 64 bit install), which is what the installer requires. So to summarize: the 64 bit version is installed with SP1, the 32 bit version is not. Uhm, Ok… thanks for that! Easy to fix, you say - just install SP1. Nope, not so easy because the standalone installer doesn't work on Windows 8. I can't get either .NET 3.5 installer or the SP 1 installer to even launch. They simply start and hang (or exit immediately) without messages. I also tried to get Windows to update .NET 3.5 by checking for Windows Updates, which should pick up on the dated version of .NET 3.5 and pull down SP1, but that's also no go. Check for Updates doesn't bring down any updates for me yet. I'm sure at some random point in the future Windows will deem it necessary to update .NET 3.5 to SP1, but at this point it's not letting me coerce it to do it explicitly. How did this happen I'm not sure exactly whether this is the cause and effect, but I suspect the story goes like this: Installed Windows 8 without support for .NET 3.5 Installed Visual Studio 2010 which installs .NET 3.5 (no SP) I now had .NET 3.5 installed but without SP1. I then: Tried to install CodeRush - Error: .NET 3.5 SP1 required Enabled .NET 3.5 in Windows Features I figured enabling the .NET 3.5 Windows Features would do the trick. But still no go. Now I suspect Visual Studio installed the 32 bit version of .NET 3.5 on my machine and Windows Features detected the previous install and didn't reinstall it. This left the 32 bit install at least with no SP1 installed. How to Fix it My final solution was to completely uninstall .NET 3.5 *and* to reboot: Go to Windows Features Uncheck the .NET Framework 3.5 Restart Windows Go to Windows Features Check .NET Framework 3.5 and voila, I now have a proper installation of .NET 3.5. I tried this before but without the reboot step in between which did not work. Make sure you reboot between uninstalling and reinstalling .NET 3.5! More Problems The above fixed me right up, but in looking for a solution it seems that a lot of people are also having problems with .NET 3.5 installing properly from the Windows Features dialog. The problem there is that the feature wasn't properly loading from the installer disks or not downloading the proper components for updates. It turns out you can explicitly install Windows features using the DISM tool in Windows.dism.exe /online /enable-feature /featurename:NetFX3 /Source:f:\sources\sxs You can try this without the /Source flag first - which uses the hidden Windows installer files if you kept those. Otherwise insert the DVD or ISO and point at the path \sources\sxs path where the installer lives. This also gives you a little more information if something does go wrong.© Rick Strahl, West Wind Technologies, 2005-2012Posted in Windows  .NET   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); })();

  Read the article

• Debugging Tips for Skinning

  - by Christian David Straub

  Another guest post by Jeanne Waldman.If you are developing a skin for your Fusion Application in JDeveloper you should know these tips.   1. Firebug is your friend 2. Uncompress the css style classes 3. CHECK_FILE_MODIFICATION so that you see your skinning changes right away 4. View the generated CSS File   1. Firebug is your friend Install Firebug (http://getfirebug.com/layout) into Firefox and use it to view your rendered jspx page in the browser. You can select the HTML dom nodes on your page and you can see the css styles applied to each dom node.   2. Uncompress the css style classes By default the styleclasses that are rendered are compressed. You may see style classes like class="x10" and class="x2e". But in your skin css file you have styleclasses like: af|inputText::content or af|panelBox::header   It is easier for you to develop a skin and debug a skin with Firebug if you see the uncompressed styleclasses. To do this, a. open web.xml b. add   <context-param>     <param-name>org.apache.myfaces.trinidad.DISABLE_CONTENT_COMPRESSION</param-name>     <param-value>true</param-value>   </context-param> c. save d. restart the server and re-run your page.   3. CHECK_FILE_MODIFICATION so that you see your skinning changes right away   For performance sake the ADF Faces framework does not check if you skin .css file has changed on every render. But this is exactly what you want to happen when you are developing or debugging a skin. You want your changes to get noticed right away, without restarting the server.   To do this, a. open web.xml b. add   <context-param>     <description>If this parameter is true, there will be an automatic check of the modification date of your JSPs, and saved state will be discarded when JSP's change. It will also automatically check if your skinning css files have changed without you having to restart the server. This makes development easier, but adds overhead. For this reason this parameter should be set to false when your application is deployed.</description>     <param-name>org.apache.myfaces.trinidad.CHECK_FILE_MODIFICATION</param-name>     <param-value>false</param-value>   </context-param> c. save d. restart the server and re-run your page. e. from then on, you can change your skin's .css file, save it and refresh your page and you should see the changes right away   4. View the generated CSS File   There are different ways to view the generated CSS File which is your skin's css file merged in with all the skins it extends and processed and generated to the filesystem and linked to your generated html page. One way is to view it with Firebug. The problem with this approach is you might see something that is a little different than the actual css file because Firebug may do some extra processing. I like to view the generated css file by: Right click on your page in the browser 

  Read the article

• Algorithmia Source Code released on CodePlex

  - by FransBouma

  Following the release of our BCL Extensions Library on CodePlex, we have now released the source-code of Algorithmia on CodePlex! Algorithmia is an algorithm and data-structures library for .NET 3.5 or higher and is one of the pillars LLBLGen Pro v3's designer is built on. The library contains many data-structures and algorithms, and the source-code is well documented and commented, often with links to official descriptions and papers of the algorithms and data-structures implemented. The source-code is shared using Mercurial on CodePlex and is licensed under the friendly BSD2 license. User documentation is not available at the moment but will be added soon. One of the main design goals of Algorithmia was to create a library which contains implementations of well-known algorithms which weren't already implemented in .NET itself. This way, more developers out there can enjoy the results of many years of what the field of Computer Science research has delivered. Some algorithms and datastructures are known in .NET but are re-implemented because the implementation in .NET isn't efficient for many situations or lacks features. An example is the linked list in .NET: it doesn't have an O(1) concat operation, as every node refers to the containing LinkedList object it's stored in. This is bad for algorithms which rely on O(1) concat operations, like the Fibonacci heap implementation in Algorithmia. Algorithmia therefore contains a linked list with an O(1) concat feature. The following functionality is available in Algorithmia: Command, Command management. This system is usable to build a fully undo/redo aware system by building your object graph using command-aware classes. The Command pattern is implemented using a system which allows transparent undo-redo and command grouping so you can use it to make a class undo/redo aware and set properties, use its contents without using commands at all. The Commands namespace is the namespace to start. Classes you'd want to look at are CommandifiedMember, CommandifiedList and KeyedCommandifiedList. See the CommandQueueTests in the test project for examples. Graphs, Graph algorithms. Algorithmia contains a sophisticated graph class hierarchy and algorithms implemented onto them: non-directed and directed graphs, as well as a subgraph view class, which can be used to create a view onto an existing graph class which can be self-maintaining. Algorithms include transitive closure, topological sorting and others. A feature rich depth-first search (DFS) crawler is available so DFS based algorithms can be implemented quickly. All graph classes are undo/redo aware, as they can be set to be 'commandified'. When a graph is 'commandified' it will do its housekeeping through commands, which makes it fully undo-redo aware, so you can remove, add and manipulate the graph and undo/redo the activity automatically without any extra code. If you define the properties of the class you set as the vertex type using CommandifiedMember, you can manipulate the properties of vertices and the graph contents with full undo/redo functionality without any extra code. Heaps. Heaps are data-structures which have the largest or smallest item stored in them always as the 'root'. Extracting the root from the heap makes the heap determine the next in line to be the 'maximum' or 'minimum' (max-heap vs. min-heap, all heaps in Algorithmia can do both). Algorithmia contains various heaps, among them an implementation of the Fibonacci heap, one of the most efficient heap datastructures known today, especially when you want to merge different instances into one. Priority queues. Priority queues are specializations of heaps. Algorithmia contains a couple of them. Sorting. What's an algorithm library without sort algorithms? Algorithmia implements a couple of sort algorithms which sort the data in-place. This aspect is important in situations where you want to sort the elements in a buffer/list/ICollection in-place, so all data stays in the data-structure it already is stored in. PropertyBag. It re-implements Tony Allowatt's original idea in .NET 3.5 specific syntax, which is to have a generic property bag and to be able to build an object in code at runtime which can be bound to a property grid for editing. This is handy for when you have data / settings stored in XML or other format, and want to create an editable form of it without creating many editors. IEditableObject/IDataErrorInfo implementations. It contains default implementations for IEditableObject and IDataErrorInfo (EditableObjectDataContainer for IEditableObject and ErrorContainer for IDataErrorInfo), which make it very easy to implement these interfaces (just a few lines of code) without having to worry about bookkeeping during databinding. They work seamlessly with CommandifiedMember as well, so your undo/redo aware code can use them out of the box. EventThrottler. It contains an event throttler, which can be used to filter out duplicate events in an event stream coming into an observer from an event. This can greatly enhance performance in your UI without needing to do anything other than hooking it up so it's placed between the event source and your real handler. If your UI is flooded with events from data-structures observed by your UI or a middle tier, you can use this class to filter out duplicates to avoid redundant updates to UI elements or to avoid having observers choke on many redundant events. Small, handy stuff. A MultiValueDictionary, which can store multiple unique values per key, instead of one with the default Dictionary, and is also merge-aware so you can merge two into one. A Pair class, to quickly group two elements together. Multiple interfaces for helping with building a de-coupled, observer based system, and some utility extension methods for the defined data-structures. We regularly update the library with new code. If you have ideas for new algorithms or want to share your contribution, feel free to discuss it on the project's Discussions page or send us a pull request. Enjoy!

  Read the article

• AngularJs ng-cloak Problems on large Pages

  - by Rick Strahl

  I’ve been working on a rather complex and large Angular page. Unlike a typical AngularJs SPA style ‘application’ this particular page is just that: a single page with a large amount of data on it that has to be visible all at once. The problem is that when this large page loads it flickers and displays template markup briefly before kicking into its actual content rendering. This is is what the Angular ng-cloak is supposed to address, but in this case I had no luck getting it to work properly. This application is a shop floor app where workers need to see all related information in one big screen view, so some of the benefits of Angular’s routing and view swapping features couldn’t be applied. Instead, we decided to have one very big view but lots of ng-controllers and directives to break out the logic for code separation. For code separation this works great – there are a number of small controllers that deal with their own individual and isolated application concerns. For HTML separation we used partial ASP.NET MVC Razor Views which made breaking out the HTML into manageable pieces super easy and made migration of this page from a previous server side Razor page much easier. We were also able to leverage most of our server side localization without a lot of  changes as a bonus. But as a result of this choice the initial HTML document that loads is rather large – even without any data loaded into it, resulting in a fairly large DOM tree that Angular must manage. Large Page and Angular Startup The problem on this particular page is that there’s quite a bit of markup – 35k’s worth of markup without any data loaded, in fact. It’s a large HTML page with a complex DOM tree. There are quite a lot of Angular {{ }} markup expressions in the document. Angular provides the ng-cloak directive to try and hide the element it cloaks so that you don’t see the flash of these markup expressions when the page initially loads before Angular has a chance to render the data into the markup expressions.<div id="mainContainer" class="mainContainer boxshadow" ng-app="app" ng-cloak> Note the ng-cloak attribute on this element, which here is an outer wrapper element of the most of this large page’s content. ng-cloak is supposed to prevent displaying the content below it, until Angular has taken control and is ready to render the data into the templates. Alas, with this large page the end result unfortunately is a brief flicker of un-rendered markup which looks like this: It’s brief, but plenty ugly – right?  And depending on the speed of the machine this flash gets more noticeable with slow machines that take longer to process the initial HTML DOM. ng-cloak Styles ng-cloak works by temporarily hiding the marked up element and it does this by essentially applying a style that does this:[ng\:cloak], [ng-cloak], [data-ng-cloak], [x-ng-cloak], .ng-cloak, .x-ng-cloak { display: none !important; } This style is inlined as part of AngularJs itself. If you looking at the angular.js source file you’ll find this at the very end of the file:!angular.$$csp() && angular.element(document) .find('head') .prepend('<style type="text/css">@charset "UTF-8";[ng\\:cloak],[ng-cloak],' + '[data-ng-cloak],[x-ng-cloak],.ng-cloak,.x-ng-cloak,' + '.ng-hide{display:none !important;}ng\\:form{display:block;}' '.ng-animate-block-transitions{transition:0s all!important;-webkit-transition:0s all!important;}' + '</style>'); This is is meant to initially hide any elements that contain the ng-cloak attribute or one of the other Angular directive permutation markup. Unfortunately on this particular web page ng-cloak had no effect – I still see the flicker. Why doesn’t ng-cloak work? The problem is of course – timing. The problem is that Angular actually needs to get control of the page before it ever starts doing anything like process even the ng-cloak attribute (or style etc). Because this page is rather large (about 35k of non-data HTML) it takes a while for the DOM to actually plow through the HTML. With the Angular <script> tag defined at the bottom of the page after the HTML DOM content there’s a slight delay which causes the flicker. For smaller pages the initial DOM load/parse cycle is so fast that the markup never shows, but with larger content pages it may show and become an annoying problem. Workarounds There a number of simple ways around this issue and some of them are hinted on in the Angular documentation. Load Angular Sooner One obvious thing that would help with this is to load Angular at the top of the page  BEFORE the DOM loads and that would give it much earlier control. The old ng-cloak documentation actually recommended putting the Angular.js script into the header of the page (apparently this was recently removed), but generally it’s not a good practice to load scripts in the header for page load performance. This is especially true if you load other libraries like jQuery which should be loaded prior to loading Angular so it can use jQuery rather than its own jqLite subset. This is not something I normally would like to do and also something that I’d likely forget in the future and end up right back here :-). Use ng-include for Child Content Angular supports nesting of child templates via the ng-include directive which essentially delay loads HTML content. This helps by removing a lot of the template content out of the main page and so getting control to Angular a lot sooner in order to hide the markup template content. In the application in question, I realize that in hindsight it might have been smarter to break this page out with client side ng-include directives instead of MVC Razor partial views we used to break up the page sections. Razor partial views give that nice separation as well, but in the end Razor puts humpty dumpty (ie. the HTML) back together into a whole single and rather large HTML document. Razor provides the logical separation, but still results in a large physical result document. But Razor also ended up being helpful to have a few security related blocks handled via server side template logic that simply excludes certain parts of the UI the user is not allowed to see – something that you can’t really do with client side exclusion like ng-hide/ng-show – client side content is always there whereas on the server side you can simply not send it to the client. Another reason I’m not a huge fan of ng-include is that it adds another HTTP hit to a request as templates are loaded from the server dynamically as needed. Given that this page was already heavy with resources adding another 10 separate ng-include directives wouldn’t be beneficial :-) ng-include is a valid option if you start from scratch and partition your logic. Of course if you don’t have complex pages, having completely separate views that are swapped in as they are accessed are even better, but we didn’t have this option due to the information having to be on screen all at once. Avoid using {{ }}  Expressions The biggest issue that ng-cloak attempts to address isn’t so much displaying the original content – it’s displaying empty {{ }} markup expression tags that get embedded into content. It gives you the dreaded “now you see it, now you don’t” effect where you sometimes see three separate rendering states: Markup junk, empty views, then views filled with data. If we can remove {{ }} expressions from the page you remove most of the perceived double draw effect as you would effectively start with a blank form and go straight to a filled form. To do this you can forego {{ }}  expressions and replace them with ng-bind directives on DOM elements. For example you can turn:<div class="list-item-name listViewOrderNo"> <a href='#'>{{lineItem.MpsOrderNo}}</a> </div>into:<div class="list-item-name listViewOrderNo"> <a href="#" ng-bind="lineItem.MpsOrderNo"></a> </div> to get identical results but because the {{ }}  expression has been removed there’s no double draw effect for this element. Again, not a great solution. The {{ }} syntax sure reads cleaner and is more fluent to type IMHO. In some cases you may also not have an outer element to attach ng-bind to which then requires you to artificially inject DOM elements into the page. This is especially painful if you have several consecutive values like {{Firstname}} {{Lastname}} for example. It’s an option though especially if you think of this issue up front and you don’t have a ton of expressions to deal with. Add the ng-cloak Styles manually You can also explicitly define the .css styles that Angular injects via code manually in your application’s style sheet. By doing so the styles become immediately available and so are applied right when the page loads – no flicker. I use the minimal:[ng-cloak] { display: none !important; } which works for:<div id="mainContainer" class="mainContainer dialog boxshadow" ng-app="app" ng-cloak> If you use one of the other combinations add the other CSS selectors as well or use the full style shown earlier. Angular will still load its version of the ng-cloak styling but it overrides those settings later, but this will do the trick of hiding the content before that CSS is injected into the page. Adding the CSS in your own style sheet works well, and is IMHO by far the best option. The nuclear option: Hiding the Content manually Using the explicit CSS is the best choice, so the following shouldn’t ever be necessary. But I’ll mention it here as it gives some insight how you can hide/show content manually on load for other frameworks or in your own markup based templates. Before I figured out that I could explicitly embed the CSS style into the page, I had tried to figure out why ng-cloak wasn’t doing its job. After wasting an hour getting nowhere I finally decided to just manually hide and show the container. The idea is simple – initially hide the container, then show it once Angular has done its initial processing and removal of the template markup from the page. You can manually hide the content and make it visible after Angular has gotten control. To do this I used:<div id="mainContainer" class="mainContainer boxshadow" ng-app="app" style="display:none"> Notice the display: none style that explicitly hides the element initially on the page. Then once Angular has run its initialization and effectively processed the template markup on the page you can show the content. For Angular this ‘ready’ event is the app.run() function:app.run( function ($rootScope, $location, cellService) { $("#mainContainer").show(); … }); This effectively removes the display:none style and the content displays. By the time app.run() fires the DOM is ready to displayed with filled data or at least empty data – Angular has gotten control. Edge Case Clearly this is an edge case. In general the initial HTML pages tend to be reasonably sized and the load time for the HTML and Angular are fast enough that there’s no flicker between the rendering times. This only becomes an issue as the initial pages get rather large. Regardless – if you have an Angular application it’s probably a good idea to add the CSS style into your application’s CSS (or a common shared one) just to make sure that content is always hidden. You never know how slow of a browser somebody might be running and while your super fast dev machine might not show any flicker, grandma’s old XP box very well might…© Rick Strahl, West Wind Technologies, 2005-2014Posted in Angular  JavaScript  CSS  HTML   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); })();

  Read the article

• Partition Wise Joins

  - by jean-pierre.dijcks

  Some say they are the holy grail of parallel computing and PWJ is the basis for a shared nothing system and the only join method that is available on a shared nothing system (yes this is oversimplified!). The magic in Oracle is of course that is one of many ways to join data. And yes, this is the old flexibility vs. simplicity discussion all over, so I won't go there... the point is that what you must do in a shared nothing system, you can do in Oracle with the same speed and methods. The Theory A partition wise join is a join between (for simplicity) two tables that are partitioned on the same column with the same partitioning scheme. In shared nothing this is effectively hard partitioning locating data on a specific node / storage combo. In Oracle is is logical partitioning. If you now join the two tables on that partitioned column you can break up the join in smaller joins exactly along the partitions in the data. Since they are partitioned (grouped) into the same buckets, all values required to do the join live in the equivalent bucket on either sides. No need to talk to anyone else, no need to redistribute data to anyone else... in short, the optimal join method for parallel processing of two large data sets. PWJ's in Oracle Since we do not hard partition the data across nodes in Oracle we use the Partitioning option to the database to create the buckets, then set the Degree of Parallelism (or run Auto DOP - see here) and get our PWJs. The main questions always asked are: How many partitions should I create? What should my DOP be? In a shared nothing system the answer is of course, as many partitions as there are nodes which will be your DOP. In Oracle we do want you to look at the workload and concurrency, and once you know that to understand the following rules of thumb. Within Oracle we have more ways of joining of data, so it is important to understand some of the PWJ ideas and what it means if you have an uneven distribution across processes. Assume we have a simple scenario where we partition the data on a hash key resulting in 4 hash partitions (H1 -H4). We have 2 parallel processes that have been tasked with reading these partitions (P1 - P2). The work is evenly divided assuming the partitions are the same size and we can scan this in time t1 as shown below. Now assume that we have changed the system and have a 5th partition but still have our 2 workers P1 and P2. The time it takes is actually 50% more assuming the 5th partition has the same size as the original H1 - H4 partitions. In other words to scan these 5 partitions, the time t2 it takes is not 1/5th more expensive, it is a lot more expensive and some other join plans may now start to look exciting to the optimizer. Just to post the disclaimer, it is not as simple as I state it here, but you get the idea on how much more expensive this plan may now look... Based on this little example there are a few rules of thumb to follow to get the partition wise joins. First, choose a DOP that is a factor of two (2). So always choose something like 2, 4, 8, 16, 32 and so on... Second, choose a number of partitions that is larger or equal to 2* DOP. Third, make sure the number of partitions is divisible through 2 without orphans. This is also known as an even number... Fourth, choose a stable partition count strategy, which is typically hash, which can be a sub partitioning strategy rather than the main strategy (range - hash is a popular one). Fifth, make sure you do this on the join key between the two large tables you want to join (and this should be the obvious one...). Translating this into an example: DOP = 8 (determined based on concurrency or by using Auto DOP with a cap due to concurrency) says that the number of partitions >= 16. Number of hash (sub) partitions = 32, which gives each process four partitions to work on. This number is somewhat arbitrary and depends on your data and system. In this case my main reasoning is that if you get more room on the box you can easily move the DOP for the query to 16 without repartitioning... and of course it makes for no leftovers on the table... And yes, we recommend up-to-date statistics. And before you start complaining, do read this post on a cool way to do stats in 11.

  Read the article

< Previous Page | 216 217 218 219 220 221 222 223 224 225 226 227  | Next Page >