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  • git post-receive hook throws "command not found" error but seems to run properly and no errors when run manually

    - by Ben
    I have a post-receive hook that runs on a central git repository set up with gitolite to trigger a git pull on a staging server. It seems to work properly, but throws a "command not found" error when it is run. I am trying to track down the source of the error, but have not had any luck. Running the same commands manually does not produce an error. The error changes depending on what was done in the commit that is being pushed to the central repository. For instance, if 'git rm ' was committed and pushed to the central repo the error message will be "remote: hooks/post-receive: line 16: Removed: command not found" and if 'git add ' was committed and pushed to the central repo the error message will be "remote: hooks/post-receive: line 16: Merge: command not found". In either case the 'git pull' run on the staging server works correctly despite the error message. Here is the post-receive script: #!/bin/bash # # This script is triggered by a push to the local git repository. It will # ssh into a remote server and perform a git pull. # # The SSH_USER must be able to log into the remote server with a # passphrase-less SSH key *AND* be able to do a git pull without a passphrase. # # The command to actually perform the pull request on the remost server comes # from the ~/.ssh/authorized_keys file on the REMOTE_HOST and is triggered # by the ssh login. SSH_USER="remoteuser" REMOTE_HOST="staging.server.com" `ssh $SSH_USER@$REMOTE_HOST` # This is line 16 echo "Done!" The command that does the git pull on the staging server is in the ssh user's ~/.ssh/authorized_keys file and is: command="cd /var/www/staging_site; git pull",no-port-forwarding,no-X11-forwarding,no-agent-forwarding, ssh-rsa AAAAB3NzaC1yc2EAAAABIwAA... (the rest of the public key) This is the actual output from removing a file from my local repo, committing it locally, and pushing it to the central git repo: ben@tamarack:~/thejibe/testing/web$ git rm ./testing rm 'testing' ben@tamarack:~/thejibe/testing/web$ git commit -a -m "Remove testing file" [master bb96e13] Remove testing file 1 files changed, 0 insertions(+), 5 deletions(-) delete mode 100644 testing ben@tamarack:~/thejibe/testing/web$ git push Counting objects: 3, done. Delta compression using up to 2 threads. Compressing objects: 100% (2/2), done. Writing objects: 100% (2/2), 221 bytes, done. Total 2 (delta 1), reused 0 (delta 0) remote: From [email protected]:testing remote: aa72ad9..bb96e13 master -> origin/master remote: hooks/post-receive: line 16: Removed: command not found # The error msg remote: Done! To [email protected]:testing aa72ad9..bb96e13 master -> master ben@tamarack:~/thejibe/testing/web$ As you can see the post-receive script gets to the echo "Done!" line and when I look on the staging server the git pull has been successfully run, but there's still that nagging error message. Any suggestions on where to look for the source of the error message would be greatly appreciated. I'm tempted to redirect stderr to /dev/null but would prefer to know what the problem is.

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  • git post-receive hook throws "command not found" error but seems to run properly and no errors when run manually

    - by Ben
    I have a post-receive hook that runs on a central git repository set up with gitolite to trigger a git pull on a staging server. It seems to work properly, but throws a "command not found" error when it is run. I am trying to track down the source of the error, but have not had any luck. Running the same commands manually does not produce an error. The error changes depending on what was done in the commit that is being pushed to the central repository. For instance, if 'git rm ' was committed and pushed to the central repo the error message will be "remote: hooks/post-receive: line 16: Removed: command not found" and if 'git add ' was committed and pushed to the central repo the error message will be "remote: hooks/post-receive: line 16: Merge: command not found". In either case the 'git pull' run on the staging server works correctly despite the error message. Here is the post-receive script: #!/bin/bash # # This script is triggered by a push to the local git repository. It will # ssh into a remote server and perform a git pull. # # The SSH_USER must be able to log into the remote server with a # passphrase-less SSH key *AND* be able to do a git pull without a passphrase. # # The command to actually perform the pull request on the remost server comes # from the ~/.ssh/authorized_keys file on the REMOTE_HOST and is triggered # by the ssh login. SSH_USER="remoteuser" REMOTE_HOST="staging.server.com" `ssh $SSH_USER@$REMOTE_HOST` # This is line 16 echo "Done!" The command that does the git pull on the staging server is in the ssh user's ~/.ssh/authorized_keys file and is: command="cd /var/www/staging_site; git pull",no-port-forwarding,no-X11-forwarding,no-agent-forwarding, ssh-rsa AAAAB3NzaC1yc2EAAAABIwAA... (the rest of the public key) This is the actual output from removing a file from my local repo, committing it locally, and pushing it to the central git repo: ben@tamarack:~/thejibe/testing/web$ git rm ./testing rm 'testing' ben@tamarack:~/thejibe/testing/web$ git commit -a -m "Remove testing file" [master bb96e13] Remove testing file 1 files changed, 0 insertions(+), 5 deletions(-) delete mode 100644 testing ben@tamarack:~/thejibe/testing/web$ git push Counting objects: 3, done. Delta compression using up to 2 threads. Compressing objects: 100% (2/2), done. Writing objects: 100% (2/2), 221 bytes, done. Total 2 (delta 1), reused 0 (delta 0) remote: From [email protected]:testing remote: aa72ad9..bb96e13 master -> origin/master remote: hooks/post-receive: line 16: Removed: command not found # The error msg remote: Done! To [email protected]:testing aa72ad9..bb96e13 master -> master ben@tamarack:~/thejibe/testing/web$ As you can see the post-receive script gets to the echo "Done!" line and when I look on the staging server the git pull has been successfully run, but there's still that nagging error message. Any suggestions on where to look for the source of the error message would be greatly appreciated. I'm tempted to redirect stderr to /dev/null but would prefer to know what the problem is.

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  • Import/rip/convert DVD to Adobe Premiere Pro for Mac

    - by alexyu2010
    For those who want to edit their videos, Adobe Premiere Pro will inevitably a good choice, it is a professional, real time, timeline based video editing software application that supports many video editing cards and plug-ins for accelerated processing, additional file format support and video/audio effects. Although Adobe Premiere Pro is said to be for professionals, is not so complicated that a hobbyist can't excel at using it in an hour or so. General file formats supported by Adobe Premiere Pro Up to now, Adobe Creative Suite has released several versions of Adobe Premiere Pro, including Adobe Premiere 1.0, Adobe Premiere 2.0, Adobe Premiere Pro CS3, Adobe Premiere Pro CS4 and the newly published Adobe Premiere Pro CS5. Although I saw diversity in file formats they support, I did find some common file formats supported by all of them, such as AVI, MOV, MPG. Importing DVD, Adobe Premiere Pro says "NO" It is obvious to all of us that Adobe Premiere Pro will never give DVD a hug, and it isn't rare to see that many people are really confused when they want to import their DVDs to Adobe Premiere Pro for editing. What to do? Yes, you may have noticed that, there is only a way out, that is ripping your DVDs to some formats workable with Adobe Premiere Pro natively, and this is what DVD to Adobe Premiere Pro can do. Importing DVD to Adobe Premiere Pro on Mac DVD to Adobe Premiere Pro converter for Mac is the specially designed application for ripping/converting DVD movies, DVD VOB files or DVD clips to Adobe Premiere Pro compatible AVI, MOV, MPG files with either DVD ripping tool and video converting tool within the versatile DVD to Adobe Premiere Pro converter who is a powerful program for dealing with DVD and videos perfectly. Mac DVD to Adobe Premiere Pro converter can work with a wide variety of files including DVD, VOB, AVI, WMV, MPG, MOV, MP4, DV, FLV, MKV, ASF, SWF, HD video for using with other editing tools like iMovie, FCP etc, play on QuickTime, iTunes, put on portable devices like iPod, iPhone, iPad, iRiver, BlackBerry, Gphone, Mobile Phone or upload to webistes such as YouTube, MySpace. DVD to Adobe Premiere Pro converter for Mac can also help you do some basic editing. You can trim, crop your DVD movie or DVD clip, apply special effect to make it more artistic, merge several DVD clips to a single one or tweak the output parameters for video and audio separately to get a better quality rendering. Besides, to get a good common of the process the preview widnows is also available for you.

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  • kernel panic after LVM setup

    - by Manuel Sopena Ballesteros
    I broke my webserver... My setup is: VMWare ESXi environemt CPanel installed CentOS release 6.5 (Final) 4 CPUs 2G RAM 2x VM disks 100G each LVM system This was my previous storage settings (the server was working fine at this time): # df -h Filesystem Size Used Avail Use% Mounted on /dev/mapper/vg_test01-lv_root 95G 1.4G 88G 2% / tmpfs 939M 0 939M 0% /dev/shm /dev/sdb1 99G 188M 94G 1% /tmp /dev/sda1 485M 54M 407M 12% /boot My web developer asked me to merge /tmp and / disks so this is what I did: Delete /dev/sdb1 partition using fdisk Create a new partition as LVM on /dev/sdb1 using fdisk Create a new physical volume -- pvcreate /dev/sdb1 Extend volume group -- vgextend /dev/sdb1 vg_test01 Extend logical volume -- lvextend -l +100%FREE /dev/vg_test01/lv_root Resize filesystem -- resize2fs /dev/vg_test01/lv_root This is the new configuration: # df -h Filesystem Size Used Avail Use% Mounted on /dev/mapper/vg_test01-lv_root 213G 105G 97G 52% / tmpfs 939M 0 939M 0% /dev/shm /dev/sda1 485M 54M 407M 12% /boot /usr/tmpDSK 4.0G 145M 3.6G 4% /tmp Since I have the new settings my web server is throwing kernel panics quite often (around every 2 days). The message says: INFO: task <taskName>:<pid> blocked for more than 120 seconds. The list of process affected that I can see from the console are: mysqld queueprocd httpd suphp vmtoolsd loop0 auditd The only way I can fix this is reseting (cold reboot) the VM. I don't think it is a hardware issue as sar is not showing any bottleneck: Linux 2.6.32-431.3.1.el6.x86_64 (test01) 08/22/2014 _x86_64_ (4 CPU) 12:00:01 AM CPU %user %nice %system %iowait %steal %idle 12:10:01 AM all 26.86 0.01 0.98 0.57 0.00 71.57 12:20:01 AM all 1.78 0.02 1.03 0.08 0.00 97.09 12:30:01 AM all 26.34 0.02 0.85 0.05 0.00 72.74 12:40:01 AM all 27.12 0.01 1.11 1.22 0.00 70.54 12:50:01 AM all 1.59 0.02 0.94 0.13 0.00 97.32 01:00:01 AM all 26.10 0.01 0.77 0.04 0.00 73.07 01:10:01 AM all 27.51 0.01 1.16 0.14 0.00 71.18 01:20:01 AM all 1.80 0.07 1.06 0.08 0.00 96.99 01:30:01 AM all 26.19 0.01 0.78 0.05 0.00 72.96 01:40:01 AM all 26.62 0.02 0.87 0.05 0.00 72.45 01:50:02 AM all 1.35 0.01 0.87 0.02 0.00 97.75 02:00:01 AM all 26.11 0.02 0.69 0.02 0.00 73.17 02:10:01 AM all 26.73 0.02 0.89 0.14 0.00 72.21 02:20:01 AM all 1.45 0.01 0.92 0.04 0.00 97.58 02:30:01 AM all 26.59 0.01 1.06 0.03 0.00 72.31 02:40:01 AM all 26.27 0.01 0.72 0.05 0.00 72.95 02:50:01 AM all 0.86 0.01 0.50 0.09 0.00 98.53 03:00:01 AM all 25.61 0.02 0.39 0.03 0.00 73.96 03:10:01 AM all 26.30 0.08 0.66 0.14 0.00 72.82 03:20:01 AM all 0.81 0.01 0.51 0.04 0.00 98.63 03:30:02 AM all 26.15 0.02 0.53 0.07 0.00 73.24 03:40:01 AM all 26.06 0.01 0.47 0.04 0.00 73.42 03:50:01 AM all 0.96 0.02 0.51 0.03 0.00 98.48 Average: all 17.69 0.02 0.79 0.14 0.00 81.36 06:58:14 AM LINUX RESTART 07:00:01 AM CPU %user %nice %system %iowait %steal %idle 07:10:01 AM all 1.04 0.02 0.57 0.95 0.00 97.42 07:20:02 AM all 0.66 0.01 0.39 0.06 0.00 98.87 07:30:01 AM all 25.71 0.01 0.45 0.16 0.00 73.67 07:40:01 AM all 25.88 0.01 0.35 0.08 0.00 73.68 07:50:01 AM all 1.13 0.02 0.55 0.11 0.00 98.19 As you can see the server became unresponsive at 03.50 AM and I had to reset the VM at 06.58 AM to bring the website up again. I would appreciate any help/assistance to fix this issue. thank you very much

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  • ffmpeg hangs when creating a video

    - by FearUs
    I am trying to insert an audio channel with a video: first of all I extract the audio from the original video for processing: ffmpeg -i lotr.mp4 lotr.wav I then extract all frames for later processing too: ffmpeg -i lotr.mp4 -f image2 %d.jpg When done processing audio and video streams, I try to create the video ffmpeg -f image2 -r 15 -i %d.jpg new.mp4 then merge with the audio: ffmpeg -i new.mp4 -i lotr.wav -map 0:0 -map 1:0 new_w_audio.mp4 Result: CPU activity = 100%, the process hangs and never returns. PS: I even tried it without modifying the images or the audio (so just trying to unpack then repack the video) but still the same output FFmpeg version SVN-r26400, Copyright (c) 2000-2011 the FFmpeg developers built on Jan 18 2011 04:07:05 with gcc 4.4.2 configuration: --enable-gpl --enable-version3 --enable-libgsm --enable-libvorb is --enable-libtheora --enable-libspeex --enable-libmp3lame --enable-libopenjpeg --enable-libschroedinger --enable-libopencore_amrwb --enable-libopencore_amrnb --enable-libvpx --disable-decoder=libvpx --arch=x86 --enable-runtime-cpudetect - -enable-libxvid --enable-libx264 --enable-librtmp --extra-libs='-lrtmp -lpolarss l -lws2_32 -lwinmm' --target-os=mingw32 --enable-avisynth --enable-w32threads -- cross-prefix=i686-mingw32- --cc='ccache i686-mingw32-gcc' --enable-memalign-hack libavutil 50.36. 0 / 50.36. 0 libavcore 0.16. 1 / 0.16. 1 libavcodec 52.108. 0 / 52.108. 0 libavformat 52.93. 0 / 52.93. 0 libavdevice 52. 2. 3 / 52. 2. 3 libavfilter 1.74. 0 / 1.74. 0 libswscale 0.12. 0 / 0.12. 0 Input #0, mov,mp4,m4a,3gp,3g2,mj2, from 'new.mp4': Metadata: major_brand : isom minor_version : 512 compatible_brands: isomiso2mp41 creation_time : 1970-01-01 00:00:00 encoder : Lavf52.93.0 Duration: 00:00:29.66, start: 0.000000, bitrate: 193 kb/s Stream #0.0(und): Video: mpeg4, yuv420p, 200x134 [PAR 1:1 DAR 100:67], 192 k b/s, 15 fps, 15 tbr, 15 tbn, 15 tbc Metadata: creation_time : 1970-01-01 00:00:00 [wav @ 01fed010] max_analyze_duration reached Input #1, wav, from 'lotr.wav': Duration: 00:00:29.90, bitrate: 176 kb/s Stream #1.0: Audio: pcm_s16le, 11025 Hz, 1 channels, s16, 176 kb/s File 'new_w_audio.mp4' already exists. Overwrite ? [y/N] y [buffer @ 01b03820] w:200 h:134 pixfmt:yuv420p Output #0, mp4, to 'new_w_audio.mp4': Metadata: major_brand : isom minor_version : 512 compatible_brands: isomiso2mp41 creation_time : 1970-01-01 00:00:00 encoder : Lavf52.93.0 Stream #0.0(und): Video: mpeg4, yuv420p, 200x134 [PAR 1:1 DAR 100:67], q=2-3 1, 200 kb/s, 15 tbn, 15 tbc Metadata: creation_time : 1970-01-01 00:00:00 Stream #0.1: Audio: aac, 11025 Hz, 1 channels, s16, 64 kb/s Stream mapping: Stream #0.0 -> #0.0 Stream #1.0 -> #0.1 Press [q] to stop encoding

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  • How to move a ruby on rails application to a new server

    - by ManiacZX
    I have a rails app on an old Ubuntu server I need to move onto a new machine. I haven't worked with ruby on rails so I don't really know anything about the structure of the app. I want to load this onto an Ubuntu 8.04 AMI on Amazon EC2 and am looking for any information regarding the migration process such as: Do I copy over the entire folder defined as the application root in the mongrel config (for ex: /u/apps/myapp/current) or just certain folders? Am I looking for trouble if I go with the latest versions of ruby and the various gems? Any general gotchas to look out for in the process. Current server information: root@webnode001:/# cat /proc/version Linux version 2.6.15-27-server (buildd@terranova) (gcc version 4.0.3 (Ubuntu 4.0.3-1ubuntu5)) #1 SMP Fri Dec 8 18:43:54 UTC 2006 root@webnode001:/# rails -v Rails 1.2.3 root@webnode001:/# mongrel_rails cluster::configure --version Version 1.0.1 root@webnode001:/# gem -v 0.9.0 root@webnode001:/# gem list -l *** LOCAL GEMS *** actionmailer (1.3.3, 1.2.5) Service layer for easy email delivery and testing. actionpack (1.13.3, 1.12.5) Web-flow and rendering framework putting the VC in MVC. actionwebservice (1.2.3, 1.1.6) Web service support for Action Pack. activerecord (1.15.3, 1.15.2, 1.14.4) Implements the ActiveRecord pattern for ORM. activesupport (1.4.2, 1.4.1, 1.3.1) Support and utility classes used by the Rails framework. cgi_multipart_eof_fix (2.1) Fix an exploitable bug in CGI multipart parsing which affects Ruby <= 1.8.5 when multipart boundary attribute contains a non-halting regular expression string. daemons (1.0.7, 1.0.5, 1.0.4, 1.0.2) A toolkit to create and control daemons in different ways eventmachine (0.7.2, 0.7.0) Ruby/EventMachine socket engine library fastercsv (1.2.0, 1.1.0) FasterCSV is CSV, but faster, smaller, and cleaner. fastthread (1.0) Optimized replacement for thread.rb primitives ferret (0.11.4) Ruby indexing library. gem_plugin (0.2.2, 0.2.1) A plugin system based only on rubygems that uses dependencies only mongrel (1.0.1, 0.3.13.4) A small fast HTTP library and server that runs Rails, Camping, Nitro and Iowa apps. mongrel_cluster (0.2.1) Mongrel plugin that provides commands and Capistrano tasks for managing multiple Mongrel processes. mysql (2.7) MySQL/Ruby provides the same functions for Ruby programs that the MySQL C API provides for C programs. piston (1.3.3) Piston is a utility that enables merge tracking of remote repositories. rails (1.2.3, 1.1.6) Web-application framework with template engine, control-flow layer, and ORM. rake (0.7.3, 0.7.1) Ruby based make-like utility. sources (0.0.1) This package provides download sources for remote gem installation swiftiply (0.5.1) A fast clustering proxy for web applications.

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  • How to unlock and remove a protected partition from Prestigio USB stick?

    - by mr.b
    Ok, so, I have one of those fancy schmancy devices, which is given to me by a frustrated friend of mine. Device is a Prestigio Leather 8GB, which identifies itself to Linux host as: Bus 001 Device 006: ID 1307:0165 Transcend Information, Inc. 2GB/4GB Flash Drive Kernel messages as USB device is plugged in: kernel: [ 2769.580042] usb 1-9: new high speed USB device using ehci_hcd and address 7 kernel: [ 2769.714782] scsi8 : usb-storage 1-9:1.0 kernel: [ 2770.713937] scsi 8:0:0:0: Direct-Access 8192MB flash drive 1.00 PQ: 0 ANSI: 2 kernel: [ 2770.714535] scsi 8:0:0:1: Direct-Access 8192MB flash drive 1.00 PQ: 0 ANSI: 2 kernel: [ 2770.715734] sd 8:0:0:0: Attached scsi generic sg3 type 0 kernel: [ 2770.716108] sd 8:0:0:1: Attached scsi generic sg4 type 0 kernel: [ 2770.722175] sd 8:0:0:0: [sdc] 962560 512-byte logical blocks: (492 MB/470 MiB) kernel: [ 2770.722657] sd 8:0:0:0: [sdc] Write Protect is on kernel: [ 2770.731078] sd 8:0:0:1: [sdd] 14012416 512-byte logical blocks: (7.17 GB/6.68 GiB) kernel: [ 2770.731215] sdc: kernel: [ 2770.738251] sd 8:0:0:1: [sdd] Write Protect is off kernel: [ 2770.880328] kernel: [ 2770.885876] sd 8:0:0:0: [sdc] Attached SCSI removable disk kernel: [ 2770.887442] sdd: unknown partition table kernel: [ 2771.049605] sd 8:0:0:1: [sdd] Attached SCSI removable disk So, symptoms are typical for U3-like devices: two separate devices inside of a single flash device. Windows sees it also as two identical usb devices, and mounts two separate drives to system, whereas first one presents itself as a CDROM device, holding a write-protected content, and second is a regular flash-disk partition, that "can" be written to. However, it seems like it's broken in some weird way, since it won't let me write anything to it, format it, nothing, but that's not the issue right now. Question: How can I unlock entire USB stick so it appears to system as a single, 8GB device which can be partitioned and used normally, without restrictions? Since it appeared to be an U3 device, I have tried standard utilities: both U3 Uninstaller by u3.com (found on SoftPedia), and opensource u3_tool from sourceforge (on both Windows and Linux). First utility failed to even detect USB stick as U3 device (simply stood idle while I re-plugged stick several times), while second tool failed with some obscure error about SCSI command unable to do something (I might be able to provide exact errors when I switch back to windows). u3_tool -i /dev/sg3 (Display device info) fails with u3_partition_info() failed: Device reported command failed: status 1 ...and every other option fails with same error, minus first part which states which command precisely has failed. So, apparently, this isn't a U3 device. Or, if it is, it doesn't behave like one. I read on a few occasions that this device protection is done by special command sent to device which tells it to lock itself, and so there should be an unlock command, that would set drive straight. Does anyone have any idea about what could I do to this device to fix it? P.S. I also mentioned a problem with being unable to use second "drive", but I'll tackle that problem when (and if) I manage to merge those two devices into one...

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  • Restore single users Exchange 2003 mailbox from backup

    - by Campo
    I take weekly backups of exchange in full. I also take complete weekly backups of the entire server. It is a Server 2003 R2 with AD and Exchange 2003 all on one box. One users inbox has disappeared. She has 19000+ junk items now. It is possible the inbox got mixed into the junk. Regardless it is such a huge mess she is not going to go through all of that.... I want to restore he mailbox from the backup. I followed this MS KB http://support.microsoft.com/kb/823176 I had to use Method 3. I have a VM of Server 2003 R2 with exchange but I am having failures on the restore from NT backup. The backup log just states to check the application log.... Application log points to backup log... Only info Is failed to restore Only thing different is the computer name... The only error I can find is in the Application log. Information Store Database not found All others just say that the backup failed. Any assistance is greatly appreciated. UPDATE I have successfully proven I can restore the DB into a recovery storage group in my VM Unfortunately due to the actual account being on a different store I am unable to do the recovery... Error is The attempt to log on to the Microsoft Exchange Server computer has failed. The MAPI provider failed. Microsoft Exchange Server Information Store ID no: 8004011d-0512-00000000 Two questions QUESTION 1 Should I repeat my steps on the production exchange server in the recover storage group? then merge into her original account? I am just concerned with doing recovery like that on the live server.... QUESTION 2 Is there any way I can extract her .PST from my recovery VM and then import into her outlook? On the Recovery VM: I restored the raw DB from my full backup repaired it with ESEUTIL then mounted in the recovery store. Was thinking I could just repeat and mount in the main store on the VM? Thanks for the suggestions.

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  • Resizing Partitions on Live RHEL/cPanel Server

    - by Timothy R. Butler
    I've resized many partitions over the years on Linux, Windows and Mac OS X -- but always using a GUI. However, the time has come where the preset partition sizes my data center placed on my server aren't the right sizes and I need to resize a production server's disks. I could fiddle with it and probably do OK, but given that it is a production server, I wanted to get some advice about the right way to do this. I do have KVM over IP access, so if it is best to take the server offline and boot off a rescue partition, I can do that. root [/var/lib/mysql]# df -h Filesystem Size Used Avail Use% Mounted on /dev/sda2 9.9G 2.1G 7.3G 23% / tmpfs 7.8G 0 7.8G 0% /dev/shm /dev/sda1 99M 77M 18M 82% /boot /dev/sda8 884G 463G 376G 56% /home /dev/sda3 9.9G 8.0G 1.5G 85% /usr /dev/sda5 9.9G 9.1G 308M 97% /var /usr/tmpDSK 2.0G 38M 1.8G 3% /tmp As you can see /var and /usr are quite close to being full and I've actually had to symlink some logs on /usr to directories in /home to balance things out. What I would like to do is to add 6-10 GB each to /usr and /var, presumably taking the space from /home. As I think about how the disk is arranged, the best thought I've come up with is to reduce /home by 16 GB, say, and move /var to the spot freed up, then allocating /var's space to /usr. However, that would put /var at the far end of the disk, which seems less than idea, given that MySQL has all of its data on that partition. I'd love to take the space out of the closer end of /usr, but I assume that would take a very arduous (and perhaps risky) process of moving all of the data in /usr around. I seem to recall having such a process fail for me on a computer in the past. The other option might be to merge / and /usr since / is underutilized, though I'm not sure if that's a good idea. Do you have any suggestions both on the best reallocation plan and the commands to use to accomplish it? UPDATE: I should add -- here's the partition table. There's one unused partition, which, if memory serves, was the original tmp location before I created a tmp image: Name Flags Part Type FS Type [Label] Size (MB) ------------------------------------------------------------------------------ Unusable 1.05* sda1 Boot Primary Linux ext2 106.96* sda2 Primary Linux ext3 10737.42* sda3 Primary Linux ext3 10737.42* sda5 NC Logical Linux ext3 10738.47* sda6 NC Logical Linux swap / Solaris 2148.54* sda7 NC Logical Linux ext3 1074.80* sda8 NC Logical Linux ext3 964098.53*

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  • R glm standard error estimate differences to SAS PROC GENMOD

    - by Michelle
    I am converting a SAS PROC GENMOD example into R, using glm in R. The SAS code was: proc genmod data=data0 namelen=30; model boxcoxy=boxcoxxy ~ AGEGRP4 + AGEGRP5 + AGEGRP6 + AGEGRP7 + AGEGRP8 + RACE1 + RACE3 + WEEKEND + SEQ/dist=normal; FREQ REPLICATE_VAR; run; My R code is: parmsg2 <- glm(boxcoxxy ~ AGEGRP4 + AGEGRP5 + AGEGRP6 + AGEGRP7 + AGEGRP8 + RACE1 + RACE3 + WEEKEND + SEQ , data=data0, family=gaussian, weights = REPLICATE_VAR) When I use summary(parmsg2) I get the same coefficient estimates as in SAS, but my standard errors are wildly different. The summary output from SAS is: Name df Estimate StdErr LowerWaldCL UpperWaldCL ChiSq ProbChiSq Intercept 1 6.5007436 .00078884 6.4991975 6.5022897 67911982 0 agegrp4 1 .64607262 .00105425 .64400633 .64813891 375556.79 0 agegrp5 1 .4191395 .00089722 .41738099 .42089802 218233.76 0 agegrp6 1 -.22518765 .00083118 -.22681672 -.22355857 73401.113 0 agegrp7 1 -1.7445189 .00087569 -1.7462352 -1.7428026 3968762.2 0 agegrp8 1 -2.2908855 .00109766 -2.2930369 -2.2887342 4355849.4 0 race1 1 -.13454883 .00080672 -.13612997 -.13296769 27817.29 0 race3 1 -.20607036 .00070966 -.20746127 -.20467944 84319.131 0 weekend 1 .0327884 .00044731 .0319117 .03366511 5373.1931 0 seq2 1 -.47509583 .00047337 -.47602363 -.47416804 1007291.3 0 Scale 1 2.9328613 .00015586 2.9325559 2.9331668 -127 The summary output from R is: Coefficients: Estimate Std. Error t value Pr(>|t|) (Intercept) 6.50074 0.10354 62.785 < 2e-16 AGEGRP4 0.64607 0.13838 4.669 3.07e-06 AGEGRP5 0.41914 0.11776 3.559 0.000374 AGEGRP6 -0.22519 0.10910 -2.064 0.039031 AGEGRP7 -1.74452 0.11494 -15.178 < 2e-16 AGEGRP8 -2.29089 0.14407 -15.901 < 2e-16 RACE1 -0.13455 0.10589 -1.271 0.203865 RACE3 -0.20607 0.09315 -2.212 0.026967 WEEKEND 0.03279 0.05871 0.558 0.576535 SEQ -0.47510 0.06213 -7.646 2.25e-14 The importance of the difference in the standard errors is that the SAS coefficients are all statistically significant, but the RACE1 and WEEKEND coefficients in the R output are not. I have found a formula to calculate the Wald confidence intervals in R, but this is pointless given the difference in the standard errors, as I will not get the same results. Apparently SAS uses a ridge-stabilized Newton-Raphson algorithm for its estimates, which are ML. The information I read about the glm function in R is that the results should be equivalent to ML. What can I do to change my estimation procedure in R so that I get the equivalent coefficents and standard error estimates that were produced in SAS? To update, thanks to Spacedman's answer, I used weights because the data are from individuals in a dietary survey, and REPLICATE_VAR is a balanced repeated replication weight, that is an integer (and quite large, in the order of 1000s or 10000s). The website that describes the weight is here. I don't know why the FREQ rather than the WEIGHT command was used in SAS. I will now test by expanding the number of observations using REPLICATE_VAR and rerunning the analysis.

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  • Apache proxy pass in nginx

    - by summerbulb
    I have the following configuration in Apache: RewriteEngine On #APP ProxyPass /abc/ http://remote.com/abc/ ProxyPassReverse /abc/ http://remote.com/abc/ #APP2 ProxyPass /efg/ http://remote.com/efg/ ProxyPassReverse /efg/ http://remote.com/efg/ I am trying to have the same configuration in nginx. After reading some links, this is what I have : server { listen 8081; server_name localhost; proxy_redirect http://localhost:8081/ http://remote.com/; location ^~/abc/ { proxy_set_header X-Forwarded-Host $host; proxy_set_header X-Forwarded-Server $host; proxy_set_header X-Forwarded-For $proxy_add_x_forwarded_for; proxy_pass http://remote.com/abc/; } location ^~/efg/ { proxy_set_header X-Forwarded-Host $host; proxy_set_header X-Forwarded-Server $host; proxy_set_header X-Forwarded-For $proxy_add_x_forwarded_for; proxy_pass http://remote.com/efg/; } } I already have the following configuration: server { listen 8080; server_name localhost; location / { root html; index index.html index.htm; } location ^~/myAPP { alias path/to/app; index main.html; } location ^~/myAPP/images { alias another/path/to/images autoindex on; } } The idea here is to overcome a same-origin-policy problem. The main pages are on localhost:8080 but we need ajax calls to http://remote.com/abc. Both domains are under my control. Using the above configuration, the ajax calls either don't reach the remote server or get cut off because of the cross origin. The above solution worked in Apache and isn't working in nginx, so I am assuming it's a configuration problem. I think there is an implicit question here: should I have two server declarations or should I somehow merge them into one? EDIT: Added some more information EDIT2: I've moved all the proxy_pass configuration into the main server declaration and changed all the ajax calls to go through port 8080. I am now getting a new error: 502 Connection reset by peer. Wireshark shows packets going out to http://remote.com with a bad IP header checksum.

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  • Can't connect to svnserve on localhost - connection actively refused

    - by RMorrisey
    When I try to connect using Tortoise to my SVN server using: svn://localhost/ Tortoise tells me: "Can't connect to host 'localhost'. No connection could be made because the target machine actively refused it." How can I fix this? I am trying to set up a subversion server on my local PC for personal use. I am running Windows Vista, with SlikSVN and TortoiseSVN installed. I previously had everything working correctly, but I found that I couldn't merge(!), apparently due to a version mismatch between the SVN client and server. Anyway... I now have the following setup: I created a repository using svnadmin create; it resides at C:\svnGrove C:\svnGrove\conf\svnserve.conf (# comments omitted): [general] anon-access=read auth-access=write password-db=passwd #authz-db=authz realm=svnGrove C:\svnGrove\conf\passwd: [users] myname=mypass My Subversion Server service is pointed to: C:\Program Files\SlikSvn\bin\svnserve.exe --service -r C:\svnGrove It shows the TCP/IP service as a dependency. I have also tried running svnserve from the command line, with similar results. The below is provided by the 'about' option in TortoiseSVN: TortoiseSVN 1.6.10, Build 19898 - 32 Bit , 2010/07/16 15:46:08 Subversion 1.6.12, apr 1.3.8 apr-utils 1.3.9 neon 0.29.3 OpenSSL 0.9.8o 01 Jun 2010 zlib 1.2.3 The following is from svn --version on the command line (not sure why it says CollabNet, CollabNet was the previous SVN binary that I had set up. The uninstaller failed to remove everything gracefully): svn, version 1.6.12 (SlikSvn/1.6.12) WIN32 compiled Jun 22 2010, 20:45:29 Copyright (C) 2000-2009 CollabNet. Subversion is open source software, see http://subversion.tigris.org/ This product includes software developed by CollabNet (http://www.Collab.Net/). The following repository access (RA) modules are available: * ra_neon : Module for accessing a repository via WebDAV protocol using Neon. - handles 'http' scheme - handles 'https' scheme * ra_svn : Module for accessing a repository using the svn network protocol. - with Cyrus SASL authentication - handles 'svn' scheme * ra_local : Module for accessing a repository on local disk. - handles 'file' scheme * ra_serf : Module for accessing a repository via WebDAV protocol using serf. - handles 'http' scheme - handles 'https' scheme I disabled my Windows Firewall and CA Internet Security, without success in resolving the issue. Edit The old version of svnserve was still set up as a service after the uninstall, pointed to this path: C:\Program Files\Subversion\svn-win32-1.4.6\bin I edited the registry key for the service to point to the new path (shown above). Whether I run svnserve as a service, or using -d, I do not see an entry for that port number in the listing generated by netstat -anp tcp.

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  • The best way to separate admin functionality from a public site?

    - by AndrewO
    I'm working on a site that's grown both in terms of user-base and functionality to the point where it's becoming evident that some of the admin tasks should be separate from the public website. I was wondering what the best way to do this would be. For example, the site has a large social component to it, and a public sales interface. But at the same time, there's back office tasks, bulk upload processing, dashboards (with long running queries), and customer relations tools in the admin section that I would like to not be effected by spikes in public traffic (or effect the public-facing response time). The site is running on a fairly standard Rails/MySQL/Linux stack, but I think this is more of an architecture problem than an implementation one: mainly, how does one keep the data and business logic in sync between these different applications? Some strategies that I'm evaluating: 1) Create a slave database of the public facing database on another machine. Extract out all of the model and library code so that it can be shared between the applications. Create new controllers and views for the admin interfaces. I have limited experience with replication and am not even sure that it's supposed to be used this way (most of the time I've seen it, it's been for scaling out the read capabilities of the same application, rather than having multiple different ones). I'm also worried about the potential for latency issues if the slave is not on the same network. 2) Create new more task/department-specific applications and use a message oriented middleware to integrate them. I read Enterprise Integration Patterns awhile back and they seemed to advocate this for distributed systems. (Alternatively, in some cases the basic Rails-style RESTful API functionality might suffice.) But, I have nightmares about data synchronization issues and the massive re-architecting that this would entail. 3) Some mixture of the two. For example, the only public information necessary for some of the back office tasks is a read-only completion time or status. Would it make sense to have that on a completely separate system and send the data to public? Meanwhile, the user/group admin functionality would be run on a separate system sharing the database? The downside is, this seems to keep many of the concerns I have with the first two, especially the re-architecting. I'm sure the answers are going to be highly dependent on a site's specific needs, but I'd love to hear success (or failure) stories.

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  • Rebasing a branch which is public

    - by Dror
    I'm failing to understand how to use git-rebase, and I consider the following example. Let's start a repository in ~/tmp/repo: $ git init Then add a file foo $ echo "hello world" > foo which is then added and committed: $ git add foo $ git commit -m "Added foo" Next, I started a remote repository. In ~/tmp/bare.git I ran $ git init --bare In order to link repo to bare.git I ran $ git remote add origin ../bare.git/ $ git push --set-upstream origin master Next, lets branch, add a file and set an upstream for the new branch b1: $ git checkout -b b1 $ echo "bar" > foo2 $ git add foo2 $ git commit -m "add foo2 in b1" $ git push --set-upstream origin b1 Now it is time to switch back to master and change something there: $ echo "change foo" > foo $ git commit -a -m "changed foo in master" $ git push At this point in master the file foo contain changed foo, while in b1 it is still hello world. Finally, I want to sync b1 with the progress made in master. $ git checkout b1 $ git fetch origin $ git rebase origin/master At this point git st returns: # On branch b1 # Your branch and 'origin/b1' have diverged, # and have 2 and 1 different commit each, respectively. # (use "git pull" to merge the remote branch into yours) # nothing to commit, working directory clean At this point the content of foo in the branch b1 is change foo as well. So what does this warning mean? I expected I should do a git push, git suggests to do git pull... According to this answer, this is more or less it, and in his comment @FrerichRaabe explicitly say that I don't need to do a pull. What's going on here? What is the danger, how should one proceed? How should the history be kept consistent? What is the interplay between the case described above and the following citation: Do not rebase commits that you have pushed to a public repository. taken from pro git book. I guess it is somehow related, and if not I would love to know why. What's the relation between the above scenario and the procedure I described in this post.

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  • freebsd-update from 8.3-RELEASE to 9.0-RELEASE: How to deal with dozens of diffs?

    - by Stefan Lasiewski
    I am upgrading a FreeBSD 8.3-RELEASE system to FreeBSD 9.0-RELEASE using freebsd-update. This is my first time performing a major version upgrade in FreeBSD. At one point in the process, freebsd-update performs a diff on files which are different then what is expected for the 9.0-RELEASE. It compares the current version on the system with the new changes added from 9.0-RELEASE. There are dozens of files in the list. Thus, I am presented with dozens and dozens of diffs which open in a vi window and look like this: The following file could not be merged automatically: /etc/ntp.conf Press Enter to edit this file in vi and resolve the conflicts manually... ### vi window opens <<<<<<< current version driftfile /etc/ntp/drift ======= # # $FreeBSD: release/9.0.0/etc/ntp.conf 195652 2009-07-13 05:51:33Z dwmalone $ # # Default NTP servers for the FreeBSD operating system. # # Don't forget to enable ntpd in /etc/rc.conf with: # ntpd_enable="YES" # # The driftfile is by default /var/db/ntpd.drift, check # /etc/defaults/rc.conf on how to change the location. # >>>>>>> 9.0-RELEASE restrict default notrust nomodify ignore And so on. This requires that I manually edit each file and remove the strings like <<<<<<< current version >>>>>>> 9.0-RELEASE and =======. As I discovered afterwards, if I don't remove these strings, they end up in the file afterwards. There are dozens of files which differ between 8.3 and 9.0, and I have a dozen local modifications myself. It appears that freebsd-update is using a diff, sdiff or mergemaster function of some sort, but I can't tell what it is doing exactly. Processing these files is tedious. Is there a way that I can just say "Accept new version" or "keep old version" or "Your merge is correct"? There has got to be an easier way to deal with these files. I must be missing something. This isn't a huge problem for one machine, but eventually I'll be doing this dozens of times and I want to find an easier way.

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  • Why do I sometimes get 'sh: $'\302\211 ... ': command not found' in xterm/sh?

    - by amn
    Sometimes when I simply type a valid command like 'find ...', or anything really, I get back the following, which is completely unexpected and confusing (... is command name I type): sh: $'\302\211...': command not found There is some corruption going on I think. I don't use color in my prompt, I am using the Bash shell in POSIX mode as sh (chsh to /bin/sh and so on - $SHELL is sh). What is going on and why does this keep happening? Anything I can debug? I think this is more of an xterm issue than sh, or at least a combination of the two. Files, for context: My /etc/profile, as distributed with Arch Linux x86-64: # /etc/profile #Set our umask umask 022 # Set our default path PATH="/usr/local/sbin:/usr/local/bin:/usr/bin" export PATH # Load profiles from /etc/profile.d if test -d /etc/profile.d/; then for profile in /etc/profile.d/*.sh; do test -r "$profile" && . "$profile" done unset profile fi # Source global bash config if test "$PS1" && test "$BASH" && test -r /etc/bash.bashrc; then . /etc/bash.bashrc fi # Termcap is outdated, old, and crusty, kill it. unset TERMCAP # Man is much better than us at figuring this out unset MANPATH My /etc/shrc, which I created as a way to have sh parse some file on startup, when non-login shell. This is achieved using ENV variable set in /etc/environment with the line ENV=/etc/shrc: PS1='\u@\H \w \$ ' alias ls='ls -F --color' alias grep='grep -i --color' [ -f ~/.shrc ] && . ~/.shrc My ~/.profile, I am launching X when logging in through first virtual tty: [[ -z $DISPLAY && $XDG_VTNR -eq 1 ]] && exec xinit -- -dpi 111 My ~/.xinitc, as you can see I am using the system as a Virtual Box guest: xrdb -merge ~/.Xresources VBoxClient-all awesome & exec xterm And finally, my ~/.Xresources, no fancy stuff here I guess: *faceName: Inconsolata *faceSize: 10 xterm*VT100*translations: #override <Btn1Up>: select-end(PRIMARY, CLIPBOARD, CUT_BUFFER0) xterm*colorBDMode: true xterm*colorBD: #ff8000 xterm*cursorColor: S_red Since ~/.profile references among other things /etc/bash.bashrc, here is its content: # # /etc/bash.bashrc # # If not running interactively, don't do anything [[ $- != *i* ]] && return PS1='[\u@\h \W]\$ ' PS2='> ' PS3='> ' PS4='+ ' case ${TERM} in xterm*|rxvt*|Eterm|aterm|kterm|gnome*) PROMPT_COMMAND=${PROMPT_COMMAND:+$PROMPT_COMMAND; }'printf "\033]0;%s@%s:%s\007" "${USER}" "${HOSTNAME%%.*}" "${PWD/#$HOME/~}"' ;; screen) PROMPT_COMMAND=${PROMPT_COMMAND:+$PROMPT_COMMAND; }'printf "\033_%s@%s:%s\033\\" "${USER}" "${HOSTNAME%%.*}" "${PWD/#$HOME/~}"' ;; esac [ -r /usr/share/bash-completion/bash_completion ] && . /usr/share/bash-completion/bash_completion I have no idea what that case statement does, by the way, it does look a bit suspicious though, but then again, who am I to know.

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  • Difference between Cloud and Virtualization

    - by Akash Kava
    Ops: This does not belong to ServerFault because it focuses on Programing Architecture. I have following questions regarding differences between Cloud and Virtualization.. How Cloud is different then Virtualization? Currently I tried to find out pricing of Rackspace, Amazone and all similar cloud providers, I found that our current 6 dedicated servers came cheaper then their pricing. So how one can claim cloud is cheaper? Is it cheaper only in comparison of normal hosting? We re organized our infrastructure in virtual environment to reduce or configuration overhead at time of failure, we did not have to rewrite any peice of code that is already written for earlier setup. So moving to virtualization does not require any re programming. But cloud is absoltely different and it will require entire reprogramming right? Is it really worth to recode when our current IT costs are 3-4 times lower then cloud hosting including raid backups and all sort of clustering for high availability? New programming architecture means new overheads of training staff, new methods of testing and new deployment schemes, does it justify over "on demand resource usage" words of cloud? We are having current development architecture with simple Server side ASP.NET WebServices with no local context and on client side Flex/Silverlight which offers pretty good REST architecture and its highly scalable. How does cloud differs from REST model of deployment? On storage, SQL Server or MySQL offers pretty good replication and high availibility then what is advantage in cloud? Data guarantee, one of our vendor hosting some other customer's app on cloud (one of most used), lost Entire Hard Disk (the virtual) and entire module in first 6 months. Second provider said its your duty to take backup, fine I agree, but no provider gives SLA for data guarantee, they give 99% uptime. However in most business apps, uptime is less important then data integrity. In our 10 years of dedicated hosting experience we had only one hard disk crash. This makes me little skeptical to go for cloud and loosing control over data. And I feel its just a big marketing buzz to sell virtulization in different form. Size of data, currently all providers charge very heavy for large data, if you are hosting only below 100GB cloud can be good alternative, but I think virtual servers and dedicated servers above 100GB to few TBs are still cheaper. Why would want to pay so high on cloud when there is no data guarentee as well as it doesnt say anything about redundancy. (I wish SO had something for spell check for Internet Explorer, sorry for wrong spellings in my post)

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  • 'Timeout Expired' error against local SQL Express on only 2 LINQ Methods...

    - by Refracted Paladin
    I am going to sum up my problem first and then offer massive details and what I have already tried. Summary: I have an internal winform app that uses Linq 2 Sql to connect to a local SQL Express database. Each user has there own DB and the DB stay in sync through Merge Replication with a Central DB. All DB's are SQL 2005(sp2or3). We have been using this app for over 5 months now but recently our users are getting a Timeout expired. The timeout period elapsed prior to completion of the operation or the server is not responding. Detailed: The strange part is they get that in two differnt locations(2 differnt LINQ Methods) and only the first time they fire in a given time period(~5mins). One LINQ method is pulling all records that match a FK ID and then Manipulating them to form a Heirarchy View for a TreeView. The second is pulling all records that match a FK ID and dumping them into a DataGridView. The only things I can find in common with the 2 are that the first IS an IEnumerable and the second converts itself from IQueryable - IEnumerable - DataTable... I looked at the query's in Profiler and they 'seemed' normal. They are not very complicated querys. They are only pulling back 10 - 90 records, from one table. Any thoughts, suggestions, hints whatever would be greatly appreciated. I am at my wit's end on this.... public IList<CaseNoteTreeItem> GetTreeViewDataAsList(int personID) { var myContext = MatrixDataContext.Create(); var caseNotesTree = from cn in myContext.tblCaseNotes where cn.PersonID == personID orderby cn.ContactDate descending, cn.InsertDate descending select new CaseNoteTreeItem { CaseNoteID = cn.CaseNoteID, NoteContactDate = Convert.ToDateTime(cn.ContactDate). ToShortDateString(), ParentNoteID = cn.ParentNote, InsertUser = cn.InsertUser, ContactDetailsPreview = cn.ContactDetails.Substring(0, 75) }; return caseNotesTree.ToList<CaseNoteTreeItem>(); } AND THIS ONE public static DataTable GetAllCNotes(int personID) { using (var context = MatrixDataContext.Create()) { var caseNotes = from cn in context.tblCaseNotes where cn.PersonID == personID orderby cn.ContactDate select new { cn.ContactDate, cn.ContactDetails, cn.TimeSpentUnits, cn.IsCaseLog, cn.IsPreEnrollment, cn.PresentAtContact, cn.InsertDate, cn.InsertUser, cn.CaseNoteID, cn.ParentNote }; return caseNotes.ToList().CopyLinqToDataTable(); } }

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  • PHP, PEAR, and oci8 configuration

    - by zack_falcon
    I'll make this quick. I installed Oracle 11g (with appropriate database, users, etc), Apache 2.4.6, and PHP 5.5.4 on a Fedora 19 system. I wanted to connect PHP to Oracle. What I really wanted to do was to download MDB2_Driver_oci8, which I thought would be easy, but before I can do such a thing, PHP needs to have that plug-in enabled, so here's what I did: Tried to install oci8 via the following: pecl install oci8 When that didn't exactly work the first few times, I figured out I, for some reason, needed "Development tools" - via yum groupinstall "Development Tools" Then I figured out later that PHP actually doesn't do oci8 - it's PHP Devel. So, I had to install that too, via yum install php-devel. And then, I finally got to install oci8. It asked for the Oracle Directory, and that was that. But it said the following: Configuration option 'php_ini' is not set to php.ini location You should add 'extensions=oci8.so' to php.ini First, I did a locate oci8.so - found it in /usr/lib64/php/modules/ Second, I added what it told me to, to the php.ini file. Third, I checked the usual php_info() test page - no mention of OCI8. Uh-oh. Fourth, running both php -i and php -m listed oci8 as one of the modules. Weird. In desperation, I went ahead and downloaded the MDB2_Driver_oci8. Maybe that will fix things. Nope. When I loaded my PHP Webpage, it returned the following: Error message: extension oci8 is not compiled into PHP As well as: MDB2 error: not found Strange. And then I decided to check the error logs: PHP Startup - unable to load dynamic library '/usr/lib64/php/modules/oci8.so' - libclntsh.so.11.1: cannot open shared object file: No such file or directory in Unknown on line 0 And now I'm stuck. I tried going into the php.ini, and found that the extension_dir was commented out. I put it back in, which only seemed to break stuff. Things of note: I followed this (link) guide on how to configure PHP and install oci8. ./configure --with-oci8 doesn't work. Fedora says no such directory. As both the webpage files and the actual server reside on the same PC, I did not install the Oracle Client files. The extension_dir is commented out by default in the php.ini. This is just one of my problems in a long line of problems concerning the replication of an already existing and working, but dying, setup. It seems whenever I want to solve a problem, I have to do X first. And by doing X, I uncover another problem, which I have to solve by doing Y, which has its own problems, etc, etc. Any help would be much appreciated. Thanks.

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  • Add Your Own Domain to Your WordPress.com Blog

    - by Matthew Guay
    Now that you’ve got a nice blog on WordPress.com, why not get your own domain to brand your site?  Here’s how you can easily register a new domain or move your existing domain to your WordPress site. By default, your free WordPress address is yourblog’sname.wordpress.com.  But whether this is a personal or a company blog, it can be nice to have your own domain to really brand your site and make it your own.  Or, if you already have another website and want to use WordPress as a blog for it, you could even add blog.yoursite.com or any other subdomain. Adding a domain to your WordPress.com is a paid upgrade; registering and mapping a new domain to your account costs $14.97 a year, while mapping a domain you already own to your WordPress blog costs $9.97 a year. Getting Started Login to your blog’s dashboard, click the arrow beside Upgrades in the sidebar, and select Domains. Enter the domain or subdomain you want to add to your site in the text box, and click Add domain to blog.   If you entered a new domain you want to register, WordPress will make sure the domain is available and then present you a registration form to register the domain.  Enter your information, and then click Register Domain.   Or, if you enter a domain that’s already registered, you will see the following prompt. If this domain is a domain you own, you can map it to WordPress.com.  Login to your domain registrar account and switch your nameserver to: NS1.WORDPRESS.COM NS2.WORDPRESS.COM NS3.WORDPRESS.COM Your DNS settings page for your domain may be different, depending on your registrar.  Here’s how our domain settings looked. Alternately, if you’re wanting to map a subdomain, such as blog.yoursite.com to your WordPress blog, create the following CNAME record on your domain register.  You may have to contact your domain registrar’s support to do this.  Substitute your subdomain, domain, and blog name when creating the record. subdomain.yourdomain.com. IN CNAME yourblog.wordpress.com. Once your settings are correct, click Try Again in your WordPress dashboard.  The DNS settings may take a while to update, but once WordPress can tell your DNS settings point to it, you will see the following confirmation screen.  Click Map Domain to add this domain to your WordPress blog. Now you’re ready to pay for your domain mapping or registration.  Depending on your purchase, the information and price shown may be different.  Here we’re mapping a domain we already have registered, so it costs $9.97.  Select your method of payment, enter your payment information or signin with your Paypal account, and continue as usual. Once your purchase is finished, you’ll be returned to the Domains page on WordPress.  Try going to your new domain, and make sure it opens your blog.  If it works, then click the bullet beside the new domain, and click Update Primary Domain.  Now, when people visit your WordPress site, they’ll see your new domain in the address bar.  You can still access your blog from your old yourname.wordpress.com address, but it will redirect to you new domain. Conclusion Having a personalized domain is a great way to make your blog more professional, while still taking advantage of the ease of use that WordPress.com offers.  And, if you have your own domain, you can easily move to your site traffic to a different hosting provider in the future if you need to.  The process is slightly complicated, but for $15/year we found this one of the best upgrades you could do to your WordPress.com blog. If you want to see an example of a site created with Wordpress, check out Matthew’s tech site techinch.com. And, if you’re just getting started with WordPress, check out our series on how to Start your WordPress.com blog, Personalize it, and Easily Post Content to it from anywhere. Similar Articles Productive Geek Tips Add Social Bookmarking (Digg This!) Links to your Wordpress BlogHow-To Geek SoftwareHow To Start Your Own Professional Blog with WordPressDisable Logon to Windows Computers When Not Connected to a DomainMake a Backup Copy of your Production Wordpress Blog on Ubuntu TouchFreeze Alternative in AutoHotkey The Icy Undertow Desktop Windows Home Server – Backup to LAN The Clear & Clean Desktop Use This Bookmarklet to Easily Get Albums Use AutoHotkey to Assign a Hotkey to a Specific Window Latest Software Reviews Tinyhacker Random Tips Xobni Plus for Outlook All My Movies 5.9 CloudBerry Online Backup 1.5 for Windows Home Server Snagit 10 Use ILovePDF To Split and Merge PDF Files TimeToMeet is a Simple Online Meeting Planning Tool Easily Create More Bookmark Toolbars in Firefox Filevo is a Cool File Hosting & Sharing Site Get a free copy of WinUtilities Pro 2010 World Cup Schedule

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  • BNF – how to read syntax?

    - by Piotr Rodak
    A few days ago I read post of Jen McCown (blog) about her idea of blogging about random articles from Books Online. I think this is a great idea, even if Jen says that it’s not exciting or sexy. I noticed that many of the questions that appear on forums and other media arise from pure fact that people asking questions didn’t bother to read and understand the manual – Books Online. Jen came up with a brilliant, concise acronym that describes very well the category of posts about Books Online – RTFM365. I take liberty of tagging this post with the same acronym. I often come across questions of type – ‘Hey, i am trying to create a table, but I am getting an error’. The error often says that the syntax is invalid. 1 CREATE TABLE dbo.Employees 2 (guid uniqueidentifier CONSTRAINT DEFAULT Guid_Default NEWSEQUENTIALID() ROWGUIDCOL, 3 Employee_Name varchar(60) 4 CONSTRAINT Guid_PK PRIMARY KEY (guid) ); 5 The answer is usually(1), ‘Ok, let me check it out.. Ah yes – you have to put name of the DEFAULT constraint before the type of constraint: 1 CREATE TABLE dbo.Employees 2 (guid uniqueidentifier CONSTRAINT Guid_Default DEFAULT NEWSEQUENTIALID() ROWGUIDCOL, 3 Employee_Name varchar(60) 4 CONSTRAINT Guid_PK PRIMARY KEY (guid) ); Why many people stumble on syntax errors? Is the syntax poorly documented? No, the issue is, that correct syntax of the CREATE TABLE statement is documented very well in Books Online and is.. intimidating. Many people can be taken aback by the rather complex block of code that describes all intricacies of the statement. However, I don’t know better way of defining syntax of the statement or command. The notation that is used to describe syntax in Books Online is a form of Backus-Naur notatiion, called BNF for short sometimes. This is a notation that was invented around 50 years ago, and some say that even earlier, around 400 BC – would you believe? Originally it was used to define syntax of, rather ancient now, ALGOL programming language (in 1950’s, not in ancient India). If you look closer at the definition of the BNF, it turns out that the principles of this syntax are pretty simple. Here are a few bullet points: italic_text is a placeholder for your identifier <italic_text_in_angle_brackets> is a definition which is described further. [everything in square brackets] is optional {everything in curly brackets} is obligatory everything | separated | by | operator is an alternative ::= “assigns” definition to an identifier Yes, it looks like these six simple points give you the key to understand even the most complicated syntax definitions in Books Online. Books Online contain an article about syntax conventions – have you ever read it? Let’s have a look at fragment of the CREATE TABLE statement: 1 CREATE TABLE 2 [ database_name . [ schema_name ] . | schema_name . ] table_name 3 ( { <column_definition> | <computed_column_definition> 4 | <column_set_definition> } 5 [ <table_constraint> ] [ ,...n ] ) 6 [ ON { partition_scheme_name ( partition_column_name ) | filegroup 7 | "default" } ] 8 [ { TEXTIMAGE_ON { filegroup | "default" } ] 9 [ FILESTREAM_ON { partition_scheme_name | filegroup 10 | "default" } ] 11 [ WITH ( <table_option> [ ,...n ] ) ] 12 [ ; ] Let’s look at line 2 of the above snippet: This line uses rules 3 and 5 from the list. So you know that you can create table which has specified one of the following. just name – table will be created in default user schema schema name and table name – table will be created in specified schema database name, schema name and table name – table will be created in specified database, in specified schema database name, .., table name – table will be created in specified database, in default schema of the user. Note that this single line of the notation describes each of the naming schemes in deterministic way. The ‘optionality’ of the schema_name element is nested within database_name.. section. You can use either database_name and optional schema name, or just schema name – this is specified by the pipe character ‘|’. The error that user gets with execution of the first script fragment in this post is as follows: Msg 156, Level 15, State 1, Line 2 Incorrect syntax near the keyword 'DEFAULT'. Ok, let’s have a look how to find out the correct syntax. Line number 3 of the BNF fragment above contains reference to <column_definition>. Since column_definition is in angle brackets, we know that this is a reference to notion described further in the code. And indeed, the very next fragment of BNF contains syntax of the column definition. 1 <column_definition> ::= 2 column_name <data_type> 3 [ FILESTREAM ] 4 [ COLLATE collation_name ] 5 [ NULL | NOT NULL ] 6 [ 7 [ CONSTRAINT constraint_name ] DEFAULT constant_expression ] 8 | [ IDENTITY [ ( seed ,increment ) ] [ NOT FOR REPLICATION ] 9 ] 10 [ ROWGUIDCOL ] [ <column_constraint> [ ...n ] ] 11 [ SPARSE ] Look at line 7 in the above fragment. It says, that the column can have a DEFAULT constraint which, if you want to name it, has to be prepended with [CONSTRAINT constraint_name] sequence. The name of the constraint is optional, but I strongly recommend you to make the effort of coming up with some meaningful name yourself. So the correct syntax of the CREATE TABLE statement from the beginning of the article is like this: 1 CREATE TABLE dbo.Employees 2 (guid uniqueidentifier CONSTRAINT Guid_Default DEFAULT NEWSEQUENTIALID() ROWGUIDCOL, 3 Employee_Name varchar(60) 4 CONSTRAINT Guid_PK PRIMARY KEY (guid) ); That is practically everything you should know about BNF. I encourage you to study the syntax definitions for various statements and commands in Books Online, you can find really interesting things hidden there. Technorati Tags: SQL Server,t-sql,BNF,syntax   (1) No, my answer usually is a question – ‘What error message? What does it say?’. You’d be surprised to know how many people think I can go through time and space and look at their screen at the moment they received the error.

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  • Long pause when accessing DFS namespace

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

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  • Your thoughts on Best Practices for Scientific Computing?

    - by John Smith
    A recent paper by Wilson et al (2014) pointed out 24 Best Practices for scientific programming. It's worth to have a look. I would like to hear opinions about these points from experienced programmers in scientific data analysis. Do you think these advices are helpful and practical? Or are they good only in an ideal world? Wilson G, Aruliah DA, Brown CT, Chue Hong NP, Davis M, Guy RT, Haddock SHD, Huff KD, Mitchell IM, Plumbley MD, Waugh B, White EP, Wilson P (2014) Best Practices for Scientific Computing. PLoS Biol 12:e1001745. http://www.plosbiology.org/article/info%3Adoi%2F10.1371%2Fjournal.pbio.1001745 Box 1. Summary of Best Practices Write programs for people, not computers. (a) A program should not require its readers to hold more than a handful of facts in memory at once. (b) Make names consistent, distinctive, and meaningful. (c) Make code style and formatting consistent. Let the computer do the work. (a) Make the computer repeat tasks. (b) Save recent commands in a file for re-use. (c) Use a build tool to automate workflows. Make incremental changes. (a) Work in small steps with frequent feedback and course correction. (b) Use a version control system. (c) Put everything that has been created manually in version control. Don’t repeat yourself (or others). (a) Every piece of data must have a single authoritative representation in the system. (b) Modularize code rather than copying and pasting. (c) Re-use code instead of rewriting it. Plan for mistakes. (a) Add assertions to programs to check their operation. (b) Use an off-the-shelf unit testing library. (c) Turn bugs into test cases. (d) Use a symbolic debugger. Optimize software only after it works correctly. (a) Use a profiler to identify bottlenecks. (b) Write code in the highest-level language possible. Document design and purpose, not mechanics. (a) Document interfaces and reasons, not implementations. (b) Refactor code in preference to explaining how it works. (c) Embed the documentation for a piece of software in that software. Collaborate. (a) Use pre-merge code reviews. (b) Use pair programming when bringing someone new up to speed and when tackling particularly tricky problems. (c) Use an issue tracking tool. I'm relatively new to serious programming for scientific data analysis. When I tried to write code for pilot analyses of some of my data last year, I encountered tremendous amount of bugs both in my code and data. Bugs and errors had been around me all the time, but this time it was somewhat overwhelming. I managed to crunch the numbers at last, but I thought I couldn't put up with this mess any longer. Some actions must be taken. Without a sophisticated guide like the article above, I started to adopt "defensive style" of programming since then. A book titled "The Art of Readable Code" helped me a lot. I deployed meticulous input validations or assertions for every function, renamed a lot of variables and functions for better readability, and extracted many subroutines as reusable functions. Recently, I introduced Git and SourceTree for version control. At the moment, because my co-workers are much more reluctant about these issues, the collaboration practices (8a,b,c) have not been introduced. Actually, as the authors admitted, because all of these practices take some amount of time and effort to introduce, it may be generally hard to persuade your reluctant collaborators to comply them. I think I'm asking your opinions because I still suffer from many bugs despite all my effort on many of these practices. Bug fix may be, or should be, faster than before, but I couldn't really measure the improvement. Moreover, much of my time has been invested on defence, meaning that I haven't actually done much data analysis (offence) these days. Where is the point I should stop at in terms of productivity? I've already deployed: 1a,b,c, 2a, 3a,b,c, 4b,c, 5a,d, 6a,b, 7a,7b I'm about to have a go at: 5b,c Not yet: 2b,c, 4a, 7c, 8a,b,c (I could not really see the advantage of using GNU make (2c) for my purpose. Could anyone tell me how it helps my work with MATLAB?)

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  • IIS7 web farm - local or shared content?

    - by rbeier
    We're setting up an IIS7 web farm with two servers. Should each server have its own local copy of the content, or should they pull content directly from a UNC share? What are the pros and cons of each approach? We currently have a single live server WEB1, with content stored locally on a separate partition. A job periodically syncs WEB1 to a standby server WEB2, using robocopy for content and msdeploy for config. If WEB1 goes down, Nagios notifies us, and we manually run a script to move the IP addresses to WEB2's network interface. Both servers are actually VMs running on separate VMWare ESX 4 hosts. The servers are domain-joined. We have around 50-60 live sites on WEB1 - mostly ASP.NET, with a few that are just static HTML. Most are low-traffic "microsites". A few have moderate traffic, but none are massive. We'd like to change this so both WEB1 and WEB2 are actively serving content. This is mainly for reliability - if WEB1 goes down, we don't want to have to manually intervene to fail things over. Spreading the load is also nice, but the load is not high enough right now for us to need this. We're planning to configure our firewall to balance traffic across the two servers. It will detect when a server goes down and will send all the traffic to the remaining live server. We're planning to use sticky sessions for now... eventually we may move to SQL Server session state and stateless load balancing. But we need a way for the servers to share content. We were originally planning to move all the content to a UNC share. Our storage provider says they can set up a highly available SMB share for us. So if we go the UNC route, the storage shouldn't be a single point of failure. But we're wondering about the downsides to this approach: We'll need to change the physical paths for each site and virtual directory. There are also some projects that have absolute paths in their web.config files - we'll have to update those as well. We'll need to create a domain user for the web servers to access the share, and grant that user appropriate permissions. I haven't looked into this yet - I'm not sure if the application pool identity needs to be changed to this user, or if there's another way to tell IIS to use this account when connecting to the share. Sites will no longer be able to access their content if there's ever an Active Directory problem. In general, it just seems a lot more complicated, with more moving parts that could break. Our storage provider would create a volume for us on their redundant SAN. If I understand correctly, this SAN volume would be mounted on a VM running in their redundant VMWare environment; this VM would then expose the SMB share to our web servers. On the other hand, a benefit of the shared content approach is that we'd only need to deploy code to one place, and there would never be a temporary inconsistency between multiple copies of the content. This thread is pretty interesting, though some of these people are working at a much larger scale. I've just been discussing content so far, but we also need to think about configuration. I don't know if we can just use DFS replication for the applicationHost.config and other files, or if it's best to use the shared configuration feature with the config on a UNC share. What do you think? Thanks for your help, Richard

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  • How Oracle Data Integration Customers Differentiate Their Business in Competitive Markets

    - by Irem Radzik
    Normal 0 false false false EN-US X-NONE X-NONE MicrosoftInternetExplorer4 With data being a central force in driving innovation and competing effectively, data integration has become a key IT approach to remove silos and ensure working with consistent and trusted data. Especially with the release of 12c version, Oracle Data Integrator and Oracle GoldenGate offer easy-to-use and high-performance solutions that help companies with their critical data initiatives, including big data analytics, moving to cloud architectures, modernizing and connecting transactional systems and more. In a recent press release we announced the great momentum and analyst recognition Oracle Data Integration products have achieved in the data integration and replication market. In this press release we described some of the key new features of Oracle Data Integrator 12c and Oracle GoldenGate 12c. In addition, a few from our 4500+ customers explained how Oracle’s data integration platform helped them achieve their business goals. In this blog post I would like to go over what these customers shared about their experience. Land O’Lakes is one of America’s premier member-owned cooperatives, and offers an extensive line of agricultural supplies, as well as production and business services. Rich Bellefeuille, manager, ETL & data warehouse for Land O’Lakes told us how GoldenGate helped them modernize their critical ERP system without impacting service and how they are moving to new projects with Oracle Data Integrator 12c: “With Oracle GoldenGate 11g, we've been able to migrate our enterprise-wide implementation of Oracle’s JD Edwards EnterpriseOne, ERP system, to a new database and application server platform with minimal downtime to our business. Using Oracle GoldenGate 11g we reduced database migration time from nearly 30 hours to less than 30 minutes. Given our quick success, we are considering expansion of our Oracle GoldenGate 12c footprint. We are also in the midst of deploying a solution leveraging Oracle Data Integrator 12c to manage our pricing data to handle orders more effectively and provide a better relationship with our clients. We feel we are gaining higher productivity and flexibility with Oracle's data integration products." ICON, a global provider of outsourced development services to the pharmaceutical, biotechnology and medical device industries, highlighted the competitive advantage that a solid data integration foundation brings. Diarmaid O’Reilly, enterprise data warehouse manager, ICON plc said “Oracle Data Integrator enables us to align clinical trials intelligence with the information needs of our sponsors. It helps differentiate ICON’s services in an increasingly competitive drug-development industry."  You can find more info on ICON's implementation here. A popular use case for Oracle GoldenGate’s real-time data integration is offloading operational reporting from critical transaction processing systems. SolarWorld, one of the world’s largest solar-technology producers and the largest U.S. solar panel manufacturer, implemented Oracle GoldenGate for real-time data integration of manufacturing data for fast analysis. Russ Toyama, U.S. senior database administrator for SolarWorld told us real-time data helps their operations and GoldenGate’s solution supports high performance of their manufacturing systems: “We use Oracle GoldenGate for real-time data integration into our decision support system, which performs real-time analysis for manufacturing operations to continuously improve product quality, yield and efficiency. With reliable and low-impact data movement capabilities, Oracle GoldenGate also helps ensure that our critical manufacturing systems are stable and operate with high performance."  You can watch the full interview with SolarWorld's Russ Toyama here. Normal 0 false false false EN-US X-NONE X-NONE MicrosoftInternetExplorer4 /* Style Definitions */ table.MsoNormalTable {mso-style-name:"Table Normal"; mso-tstyle-rowband-size:0; mso-tstyle-colband-size:0; mso-style-noshow:yes; mso-style-priority:99; mso-style-qformat:yes; mso-style-parent:""; mso-padding-alt:0in 5.4pt 0in 5.4pt; mso-para-margin:0in; mso-para-margin-bottom:.0001pt; mso-pagination:widow-orphan; font-size:11.0pt; font-family:"Calibri","sans-serif"; mso-ascii-font-family:Calibri; mso-ascii-theme-font:minor-latin; mso-fareast-font-family:"Times New Roman"; mso-fareast-theme-font:minor-fareast; mso-hansi-font-family:Calibri; mso-hansi-theme-font:minor-latin; mso-bidi-font-family:"Times New Roman"; mso-bidi-theme-font:minor-bidi;} Starwood Hotels and Resorts is one of the many customers that found out how well Oracle Data Integration products work with Oracle Exadata. Gordon Light, senior director of information technology for StarWood Hotels, says they had notable performance gain in loading Oracle Exadata reporting environment: “We leverage Oracle GoldenGate to replicate data from our central reservations systems and other OLTP databases – significantly decreasing the overall ETL duration. Moving forward, we plan to use Oracle GoldenGate to help the company achieve near-real-time reporting.”You can listen about Starwood Hotels' implementation here. Many companies combine the power of Oracle GoldenGate with Oracle Data Integrator to have a single, integrated data integration platform for variety of use cases across the enterprise. Ufone is another good example of that. The leading mobile communications service provider of Pakistan has improved customer service using timely customer data in its data warehouse. Atif Aslam, head of management information systems for Ufone says: “Oracle Data Integrator and Oracle GoldenGate help us integrate information from various systems and provide up-to-date and real-time CRM data updates hourly, rather than daily. The applications have simplified data warehouse operations and allowed business users to make faster and better informed decisions to protect revenue in the fast-moving Pakistani telecommunications market.” You can read more about Ufone's use case here. In our Oracle Data Integration 12c launch webcast back in November we also heard from BT’s CTO Surren Parthab about their use of GoldenGate for moving to private cloud architecture. Surren also shared his perspectives on Oracle Data Integrator 12c and Oracle GoldenGate 12c releases. You can watch the video here. These are only a few examples of leading companies that have made data integration and real-time data access a key part of their data governance and IT modernization initiatives. They have seen real improvements in how their businesses operate and differentiate in today’s competitive markets. You can read about other customer examples in our Ebook: The Path to the Future and access resources including white papers, data sheets, podcasts and more via our Oracle Data Integration resource kit. /* Style Definitions */ table.MsoNormalTable {mso-style-name:"Table Normal"; mso-tstyle-rowband-size:0; mso-tstyle-colband-size:0; mso-style-noshow:yes; mso-style-priority:99; mso-style-qformat:yes; mso-style-parent:""; mso-padding-alt:0in 5.4pt 0in 5.4pt; mso-para-margin:0in; mso-para-margin-bottom:.0001pt; mso-pagination:widow-orphan; font-size:11.0pt; font-family:"Calibri","sans-serif"; mso-ascii-font-family:Calibri; mso-ascii-theme-font:minor-latin; mso-fareast-font-family:"Times New Roman"; mso-fareast-theme-font:minor-fareast; mso-hansi-font-family:Calibri; mso-hansi-theme-font:minor-latin; mso-bidi-font-family:"Times New Roman"; mso-bidi-theme-font:minor-bidi;}

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