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  • Going Paperless

    - by Jesse
    One year ago I came to work for a company where the entire development team is 100% “remote”; we’re spread over 3 time zones and each of us works from home. This seems to be an increasingly popular way for people to work and there are many articles and blog posts out there enumerating the advantages and disadvantages of working this way. I had read a lot about telecommuting before accepting this job and felt as if I had a pretty decent idea of what I was getting into, but I’ve encountered a few things over the past year that I did not expect. Among the most surprising by-products of working from home for me has been a dramatic reduction in the amount of paper that I use on a weekly basis. Hoarding In The Workplace Prior to my current telecommute job I worked in what most would consider pretty traditional office environments. I sat in cubicles furnished with an enormous plastic(ish) modular desks, had a mediocre (at best) PC workstation, and had ready access to a seemingly endless supply of legal pads, pens, staplers and paper clips. The ready access to paper, countless conference room meetings, and abundance of available surface area on my desk and in drawers created a perfect storm for wasting paper. I brought a pad of paper with me to every meeting I ever attended, scrawled some brief notes, and then tore that sheet off to keep next to my keyboard to follow up on any needed action items. Once my immediate need for the notes was fulfilled, that sheet would get shuffled off into a corner of my desk or filed away in a drawer “just in case”. I would guess that for all of the notes that I ever filed away, I might have actually had to dig up and refer to 2% of them (and that’s probably being very generous). That said, on those rare occasions that I did have to dig something up from old notes, it was usually pretty important and I ended up being very glad that I saved them. It was only when I would leave a job or move desks that I would finally gather all those notes together and take them to shredding bin to be disposed of. When I left my last job the amount of paper I had accumulated over my three years there was absurd, and I knew coworkers who had substance-abuse caliber paper wasting addictions that made my bad habit look like nail-biting in comparison. A Product Of My Environment I always hated using all of this paper, but simply couldn’t bring myself to stop. It would look bad if I showed up to an important conference room meeting without a pad of paper. What if someone said something profound! Plus, everyone else always brought paper with them. If you saw someone walking down the hallway with a pad of paper in hand you knew they must be on their way to a conference room meeting. Some people even had fancy looking portfolio notebook sheaths that gave their legal pads all the prestige of a briefcase. No one ever worried about running out of fresh paper because there was an endless supply, and there certainly was no shortage of places to store and file used paper. In short, the traditional office was setup for using tons and tons of paper; it’s baked into the culture there. For that reason, it didn’t take long for me to kick the paper habit once I started working from home. In my home office, desk and drawer space are at a premium. I don’t have the budget (or the tolerance) for huge modular office furniture in my spare bedroom. I also no longer have access to a bottomless pit of office supplies stock piled in cabinets and closets. If I want to use some paper, I have to go out and buy it. Finally (and most importantly), all of the meetings that I have to attend these days are “virtual”. We use instant messaging, VOIP, video conferencing, and e-mail to communicate with each other. All I need to take notes during a meeting is my computer, which I happen to be sitting right in front of all day. I don’t have any hard numbers for this, but my gut feeling is that I actually take a lot more notes now than I ever did when I worked in an office. The big difference is I don’t have to use any paper to do so. This makes it far easier to keep important information safe and organized. The Right Tool For The Job When I first started working from home I tried to find a single application that would fill the gap left by the pen and paper that I always had at my desk when I worked in an office. Well, there are no silver bullets and I’ve evolved my approach over time to try and find the best tool for the job at hand. Here’s a quick summary of how I take notes and keep everything organized. Notepad++ – This is the first application I turn to when I feel like there’s some bit of information that I need to write down and save. I use Launchy, so opening Notepad++ and creating a new file only takes a few keystrokes. If I find that the information I’m trying to get down requires a more sophisticated application I escalate as needed. The Desktop – By default, I save every file or other bit of information to the desktop. Anyone who has ever had to fix their parents computer before knows that this is a dangerous game (any file my mother has ever worked on is saved directly to the desktop and rarely moves anywhere else). I agree that storing things on the desktop isn’t a great long term approach to keeping organized, which is why I treat my desktop a bit like my e-mail inbox. I strive to keep both empty (or as close to empty as I possibly can). If something is on my desktop, it means that it’s something relevant to a task or project that I’m currently working on. About once a week I take things that I’m not longer working on and put them into my ‘Notes’ folder. The ‘Notes’ Folder – As I work on a task, I tend to accumulate multiple files associated with that task. For example, I might have a bit of SQL that I’m working on to gather data for a new report, a quick C# method that I came up with but am not yet ready to commit to source control, a bulleted list of to-do items in a .txt file, etc. If the desktop starts to get too cluttered, I create a new sub-folder in my ‘Notes’ folder. Each sub-folder’s name is the current date followed by a brief description of the task or project. Then all files related to that task or project go into that sub folder. By using the date as the first part of the folder name, these folders are automatically sorted in reverse chronological order. This means that things I worked on recently will generally be near the top of the list. Using the built-in Windows search functionality I now have a pretty quick and easy way to try and find something that I worked on a week ago or six months ago. Dropbox – Dropbox is a free service that lets you store up to 2GB of files “in the cloud” and have those files synced to all of the different computers that you use. My ‘Notes’ folder lives in Dropbox, meaning that it’s contents are constantly backed up and are always available to me regardless of which computer I’m using. They also have a pretty decent iPhone application that lets you browse and view all of the files that you have stored there. The free 2GB edition is probably enough for just storing notes, but I also pay $99/year for the 50GB storage upgrade and keep all of my music, e-books, pictures, and documents in Dropbox. It’s a fantastic service and I highly recommend it. Evernote – I use Evernote mostly to organize information that I access on a fairly regular basis. For example, my Evernote account has a running grocery shopping list, recipes that my wife and I use a lot, and contact information for people I contact infrequently enough that I don’t want to keep them in my phone. I know some people that keep nearly everything in Evernote, but there’s something about it that I find a bit clunky, so I tend to use it sparingly. Google Tasks – One of my biggest paper wasting habits was keeping a running task-list next to my computer at work. Every morning I would sit down, look at my task list, cross off what was done and add new tasks that I thought of during my morning commute. This usually resulted in having to re-copy the task list onto a fresh sheet of paper when I was done. I still keep a running task list at my desk, but I’ve started using Google Tasks instead. This is a dead-simple web-based application for quickly adding, deleting, and organizing tasks in a simple checklist style. You can quickly move tasks up and down on the list (which I use for prioritizing), and even create sub-tasks for breaking down larger tasks into smaller pieces. Balsamiq Mockups – This is a simple and lightweight tool for creating drawings of user interfaces. It’s great for sketching out a new feature, brainstorm the layout of a interface, or even draw up a quick sequence diagram. I’m terrible at drawing, so Balsamiq Mockups not only lets me create sketches that other people can actually understand, but it’s also handy because you can upload a sketch to a common location for other team members to access. I can honestly say that using these tools (and having limited resources at home) have lead me to cut my paper usage down to virtually none. If I ever were to return to a traditional office workplace (hopefully never!) I’d try to employ as many of these applications and techniques as I could to keep paper usage low. I feel far less cluttered and far better organized now.

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  • Ubuntu 11.10 - Everytime i am trying to connect to my box using SSH, its failing not connecting

    - by YumYumYum
    From any other PC doing SSH to my Ubuntu 11.10,is failing. Even the SSH is running: Other PC: retrying over and over $ ping 192.168.0.128 PING 192.168.0.128 (192.168.0.128) 56(84) bytes of data. From 192.168.0.226 icmp_seq=1 Destination Host Unreachable From 192.168.0.226 icmp_seq=2 Destination Host Unreachable From 192.168.0.226 icmp_seq=3 Destination Host Unreachable From 192.168.0.226 icmp_seq=4 Destination Host Unreachable $ sudo service iptables stop Stopping iptables (via systemctl): [ OK ] $ ssh [email protected] ssh: connect to host 192.168.0.128 port 22: No route to host $ ssh [email protected] ssh: connect to host 192.168.0.128 port 22: No route to host $ ssh [email protected] ssh: connect to host 192.168.0.128 port 22: No route to host $ ssh [email protected] ssh: connect to host 192.168.0.128 port 22: No route to host $ ssh [email protected] Connection closed by 192.168.0.128 $ ssh [email protected] [email protected]'s password: Connection closed by UNKNOWN $ ssh [email protected] ssh: connect to host 192.168.0.128 port 22: No route to host $ ssh [email protected] ssh: connect to host 192.168.0.128 port 22: No route to host Follow up: -- checked cable -- using cable tester and other detectors -- no problem found in cable -- used random 10 cables -- adapter is not broken -- checked it using circuit tester by opening the system (card is new so its not network adapter card problem) -- leds are OK showing -- used LiveCD and did same ping test was having same problem -- disabled ipv6 100% to make sure its not the cause -- disabled iptables 100% so its also not the issue -- some more info $ sudo killall dnsmasq -- did not solved the problem -- -- like many other Q/A was suggesting this same --- $ iptables --list Chain INPUT (policy ACCEPT) target prot opt source destination Chain FORWARD (policy ACCEPT) target prot opt source destination Chain OUTPUT (policy ACCEPT) target prot opt source destination $ netstat -nr Kernel IP routing table Destination Gateway Genmask Flags MSS Window irtt Iface 0.0.0.0 192.168.0.1 0.0.0.0 UG 0 0 0 eth0 169.254.0.0 0.0.0.0 255.255.0.0 U 0 0 0 eth0 192.168.0.0 0.0.0.0 255.255.255.0 U 0 0 0 eth0 $ ssh -vvv [email protected] OpenSSH_5.6p1, OpenSSL 1.0.0j-fips 10 May 2012 debug1: Reading configuration data /etc/ssh/ssh_config debug1: Applying options for * debug2: ssh_connect: needpriv 0 debug1: Connecting to 192.168.0.128 [192.168.0.128] port 22. debug1: Connection established. debug3: Not a RSA1 key file /home/sun/.ssh/id_rsa. debug2: key_type_from_name: unknown key type '-----BEGIN' debug3: key_read: missing keytype debug3: key_read: missing whitespace debug3: key_read: missing whitespace debug3: key_read: missing whitespace debug3: key_read: missing whitespace debug3: key_read: missing whitespace debug3: key_read: missing whitespace debug3: key_read: missing whitespace debug3: key_read: missing whitespace debug3: key_read: missing whitespace debug3: key_read: missing whitespace debug3: key_read: missing whitespace debug3: key_read: missing whitespace debug3: key_read: missing whitespace debug3: key_read: missing whitespace debug3: key_read: missing whitespace debug3: key_read: missing whitespace debug3: key_read: missing whitespace debug3: key_read: missing whitespace debug3: key_read: missing whitespace debug3: key_read: missing whitespace debug3: key_read: missing whitespace debug3: key_read: missing whitespace debug3: key_read: missing whitespace debug3: key_read: missing whitespace debug3: key_read: missing whitespace debug2: key_type_from_name: unknown key type '-----END' debug3: key_read: missing keytype debug1: identity file /home/sun/.ssh/id_rsa type 1 debug1: identity file /home/sun/.ssh/id_rsa-cert type -1 debug1: identity file /home/sun/.ssh/id_dsa type -1 debug1: identity file /home/sun/.ssh/id_dsa-cert type -1 debug1: Remote protocol version 2.0, remote software version OpenSSH_5.8p1 Debian-7ubuntu1 debug1: match: OpenSSH_5.8p1 Debian-7ubuntu1 pat OpenSSH* debug1: Enabling compatibility mode for protocol 2.0 debug1: Local version string SSH-2.0-OpenSSH_5.6 debug2: fd 3 setting O_NONBLOCK debug1: SSH2_MSG_KEXINIT sent debug1: SSH2_MSG_KEXINIT received debug2: kex_parse_kexinit: diffie-hellman-group-exchange-sha256,diffie-hellman-group-exchange-sha1,diffie-hellman-group14-sha1,diffie-hellman-group1-sha1 debug2: kex_parse_kexinit: [email protected],[email protected],[email protected],[email protected],ssh-rsa,ssh-dss debug2: kex_parse_kexinit: aes128-ctr,aes192-ctr,aes256-ctr,arcfour256,arcfour128,aes128-cbc,3des-cbc,blowfish-cbc,cast128-cbc,aes192-cbc,aes256-cbc,arcfour,[email protected] debug2: kex_parse_kexinit: aes128-ctr,aes192-ctr,aes256-ctr,arcfour256,arcfour128,aes128-cbc,3des-cbc,blowfish-cbc,cast128-cbc,aes192-cbc,aes256-cbc,arcfour,[email protected] debug2: kex_parse_kexinit: hmac-md5,hmac-sha1,[email protected],hmac-ripemd160,[email protected],hmac-sha1-96,hmac-md5-96 debug2: kex_parse_kexinit: hmac-md5,hmac-sha1,[email protected],hmac-ripemd160,[email protected],hmac-sha1-96,hmac-md5-96 debug2: kex_parse_kexinit: none,[email protected],zlib debug2: kex_parse_kexinit: none,[email protected],zlib debug2: kex_parse_kexinit: debug2: kex_parse_kexinit: debug2: kex_parse_kexinit: first_kex_follows 0 debug2: kex_parse_kexinit: reserved 0 debug2: kex_parse_kexinit: ecdh-sha2-nistp256,ecdh-sha2-nistp384,ecdh-sha2-nistp521,diffie-hellman-group-exchange-sha256,diffie-hellman-group-exchange-sha1,diffie-hellman-group14-sha1,diffie-hellman-group1-sha1 debug2: kex_parse_kexinit: ssh-rsa,ssh-dss,ecdsa-sha2-nistp256 debug2: kex_parse_kexinit: aes128-ctr,aes192-ctr,aes256-ctr,arcfour256,arcfour128,aes128-cbc,3des-cbc,blowfish-cbc,cast128-cbc,aes192-cbc,aes256-cbc,arcfour,[email protected] debug2: kex_parse_kexinit: aes128-ctr,aes192-ctr,aes256-ctr,arcfour256,arcfour128,aes128-cbc,3des-cbc,blowfish-cbc,cast128-cbc,aes192-cbc,aes256-cbc,arcfour,[email protected] debug2: kex_parse_kexinit: hmac-md5,hmac-sha1,[email protected],hmac-ripemd160,[email protected],hmac-sha1-96,hmac-md5-96 debug2: kex_parse_kexinit: hmac-md5,hmac-sha1,[email protected],hmac-ripemd160,[email protected],hmac-sha1-96,hmac-md5-96 debug2: kex_parse_kexinit: none,[email protected] debug2: kex_parse_kexinit: none,[email protected] debug2: kex_parse_kexinit: debug2: kex_parse_kexinit: debug2: kex_parse_kexinit: first_kex_follows 0 debug2: kex_parse_kexinit: reserved 0 debug2: mac_setup: found hmac-md5 debug1: kex: server->client aes128-ctr hmac-md5 none debug2: mac_setup: found hmac-md5 debug1: kex: client->server aes128-ctr hmac-md5 none debug1: SSH2_MSG_KEX_DH_GEX_REQUEST(1024<1024<8192) sent debug1: expecting SSH2_MSG_KEX_DH_GEX_GROUP debug2: dh_gen_key: priv key bits set: 118/256 debug2: bits set: 539/1024 debug1: SSH2_MSG_KEX_DH_GEX_INIT sent debug1: expecting SSH2_MSG_KEX_DH_GEX_REPLY debug3: check_host_in_hostfile: host 192.168.0.128 filename /home/sun/.ssh/known_hosts debug3: check_host_in_hostfile: host 192.168.0.128 filename /home/sun/.ssh/known_hosts debug3: check_host_in_hostfile: match line 139 debug1: Host '192.168.0.128' is known and matches the RSA host key. debug1: Found key in /home/sun/.ssh/known_hosts:139 debug2: bits set: 544/1024 debug1: ssh_rsa_verify: signature correct debug2: kex_derive_keys debug2: set_newkeys: mode 1 debug1: SSH2_MSG_NEWKEYS sent debug1: expecting SSH2_MSG_NEWKEYS debug2: set_newkeys: mode 0 debug1: SSH2_MSG_NEWKEYS received debug1: Roaming not allowed by server debug1: SSH2_MSG_SERVICE_REQUEST sent debug2: service_accept: ssh-userauth debug1: SSH2_MSG_SERVICE_ACCEPT received debug2: key: /home/sun/.ssh/id_rsa (0x213db960) debug2: key: /home/sun/.ssh/id_dsa ((nil)) debug1: Authentications that can continue: publickey,password debug3: start over, passed a different list publickey,password debug3: preferred gssapi-keyex,gssapi-with-mic,publickey,keyboard-interactive,password debug3: authmethod_lookup publickey debug3: remaining preferred: keyboard-interactive,password debug3: authmethod_is_enabled publickey debug1: Next authentication method: publickey debug1: Offering RSA public key: /home/sun/.ssh/id_rsa debug3: send_pubkey_test debug2: we sent a publickey packet, wait for reply debug1: Authentications that can continue: publickey,password debug1: Trying private key: /home/sun/.ssh/id_dsa debug3: no such identity: /home/sun/.ssh/id_dsa debug2: we did not send a packet, disable method debug3: authmethod_lookup password debug3: remaining preferred: ,password debug3: authmethod_is_enabled password debug1: Next authentication method: password [email protected]'s password: debug3: packet_send2: adding 64 (len 60 padlen 4 extra_pad 64) debug2: we sent a password packet, wait for reply debug1: Authentication succeeded (password). Authenticated to 192.168.0.128 ([192.168.0.128]:22). debug1: channel 0: new [client-session] debug3: ssh_session2_open: channel_new: 0 debug2: channel 0: send open debug1: Requesting [email protected] debug1: Entering interactive session. debug2: callback start debug2: client_session2_setup: id 0 debug2: channel 0: request pty-req confirm 1 debug1: Sending environment. debug3: Ignored env ORBIT_SOCKETDIR debug3: Ignored env XDG_SESSION_ID debug3: Ignored env HOSTNAME debug3: Ignored env GIO_LAUNCHED_DESKTOP_FILE_PID debug3: Ignored env IMSETTINGS_INTEGRATE_DESKTOP debug3: Ignored env GPG_AGENT_INFO debug3: Ignored env TERM debug3: Ignored env HARDWARE_PLATFORM debug3: Ignored env SHELL debug3: Ignored env DESKTOP_STARTUP_ID debug3: Ignored env HISTSIZE debug3: Ignored env XDG_SESSION_COOKIE debug3: Ignored env GJS_DEBUG_OUTPUT debug3: Ignored env WINDOWID debug3: Ignored env GNOME_KEYRING_CONTROL debug3: Ignored env QTDIR debug3: Ignored env QTINC debug3: Ignored env GJS_DEBUG_TOPICS debug3: Ignored env IMSETTINGS_MODULE debug3: Ignored env USER debug3: Ignored env LS_COLORS debug3: Ignored env SSH_AUTH_SOCK debug3: Ignored env USERNAME debug3: Ignored env SESSION_MANAGER debug3: Ignored env GIO_LAUNCHED_DESKTOP_FILE debug3: Ignored env PATH debug3: Ignored env MAIL debug3: Ignored env DESKTOP_SESSION debug3: Ignored env QT_IM_MODULE debug3: Ignored env PWD debug1: Sending env XMODIFIERS = @im=none debug2: channel 0: request env confirm 0 debug1: Sending env LANG = en_US.utf8 debug2: channel 0: request env confirm 0 debug3: Ignored env KDE_IS_PRELINKED debug3: Ignored env GDM_LANG debug3: Ignored env KDEDIRS debug3: Ignored env GDMSESSION debug3: Ignored env SSH_ASKPASS debug3: Ignored env HISTCONTROL debug3: Ignored env HOME debug3: Ignored env SHLVL debug3: Ignored env GDL_PATH debug3: Ignored env GNOME_DESKTOP_SESSION_ID debug3: Ignored env LOGNAME debug3: Ignored env QTLIB debug3: Ignored env CVS_RSH debug3: Ignored env DBUS_SESSION_BUS_ADDRESS debug3: Ignored env LESSOPEN debug3: Ignored env WINDOWPATH debug3: Ignored env XDG_RUNTIME_DIR debug3: Ignored env DISPLAY debug3: Ignored env G_BROKEN_FILENAMES debug3: Ignored env COLORTERM debug3: Ignored env XAUTHORITY debug3: Ignored env _ debug2: channel 0: request shell confirm 1 debug2: fd 3 setting TCP_NODELAY debug2: callback done debug2: channel 0: open confirm rwindow 0 rmax 32768 debug2: channel_input_status_confirm: type 99 id 0 debug2: PTY allocation request accepted on channel 0 debug2: channel 0: rcvd adjust 2097152 debug2: channel_input_status_confirm: type 99 id 0 debug2: shell request accepted on channel 0 Welcome to Ubuntu 11.10 (GNU/Linux 3.0.0-12-generic x86_64) * Documentation: https://help.ubuntu.com/ 297 packages can be updated. 92 updates are security updates. New release '12.04 LTS' available. Run 'do-release-upgrade' to upgrade to it. Last login: Fri Jun 8 07:45:15 2012 from 192.168.0.226 sun@SystemAX51:~$ ping 19<--------Lost connection again-------------- Tail follow: -- dmesg is showing a very abnormal logs, like Ubuntu is automatically bringing the eth0 up, where eth0 is getting also auto down. [ 2025.897511] r8169 0000:02:00.0: eth0: link up [ 2029.347649] r8169 0000:02:00.0: eth0: link up [ 2030.775556] r8169 0000:02:00.0: eth0: link up [ 2038.242203] r8169 0000:02:00.0: eth0: link up [ 2057.267801] r8169 0000:02:00.0: eth0: link up [ 2062.871770] r8169 0000:02:00.0: eth0: link up [ 2082.479712] r8169 0000:02:00.0: eth0: link up [ 2285.630797] r8169 0000:02:00.0: eth0: link up [ 2308.417640] r8169 0000:02:00.0: eth0: link up [ 2480.948290] r8169 0000:02:00.0: eth0: link up [ 2824.884798] r8169 0000:02:00.0: eth0: link up [ 3030.022183] r8169 0000:02:00.0: eth0: link up [ 3306.587353] r8169 0000:02:00.0: eth0: link up [ 3523.566881] r8169 0000:02:00.0: eth0: link up [ 3619.839585] r8169 0000:02:00.0: eth0: link up [ 3682.154393] nf_conntrack version 0.5.0 (16384 buckets, 65536 max) [ 3899.866854] r8169 0000:02:00.0: eth0: link up [ 4723.978269] r8169 0000:02:00.0: eth0: link up [ 4807.415682] r8169 0000:02:00.0: eth0: link up [ 5101.865686] r8169 0000:02:00.0: eth0: link up How do i fix it? -- http://ubuntuforums.org/showthread.php?t=1959794 -- apt-get install openipml openhpi-plugin-ipml

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  • ADDS: 1 - Introducing and designing

    - by marc dekeyser
    Normal 0 false false false EN-GB X-NONE X-NONE /* 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-parent:""; mso-padding-alt:0cm 5.4pt 0cm 5.4pt; mso-para-margin-top:0cm; mso-para-margin-right:0cm; mso-para-margin-bottom:10.0pt; mso-para-margin-left:0cm; line-height:115%; mso-pagination:widow-orphan; font-size:11.0pt; font-family:"Calibri","sans-serif"; mso-ascii-font-family:Calibri; mso-ascii-theme-font:minor-latin; mso-hansi-font-family:Calibri; mso-hansi-theme-font:minor-latin; mso-bidi-font-family:"Times New Roman"; mso-bidi-theme-font:minor-bidi; mso-fareast-language:EN-US;} What is ADDS?  Every Microsoft oriented infrastructure in today's enterprises will depend largely on the active directory version built by Microsoft. It is the foundation stone on which all other products (Exchange, update services, office communicator, the system center family, etc) rely on to get their information. And that is just looking at it from an infrastructure perspective. A well designed and implemented Active Directory implementation makes life for IT personnel and user alike a lot easier. Centralised management and the abilities opened up  by having it in place are ample.  But what is Active Directory Domain Services? We can look at ADDS as a centralised directory containing all objects your infrastructure runs on in one way or another. Since it is a Microsoft product you'll obviously not be seeing linux or mac clients listed in here (exceptions exist) but in general we can say it contains everything your company has in place in one form or another.  The domain name services. The domain naming service (or DNS for short) is a service which translates IP address (the identifiers for each computer in your domain) into readable and easy to understand names. This service is a prequisite for ADDA to work and having wrong record in a DNS server will make any ADDS service fail. Generally speaking a DNS service will be run on the same server as the ADDS service but it is worth wile to remember that this is not necessary. You could, for example, run your DNS services on a linux box (which would need special preparing to host an ADDS integrated DNS zone) and run the ADDS service of another box… Where to start? If the aim is to put in place a first time implementation of ADDS in your enterprise there are plenty of things to consider depending on what you are going to do in the long run. Great care has to be taken when first designing and implementing as having it set up wrong will cause a headache down the line. It is for that reason that I like to start building from the bottom up and start with a generic installation of ADDS (which will still differ for every client) and make it adaptable for future services which can hook in to the existing environment. Adapting existing environments is out of scope for this document (and series) although it is possible to take the pointers and change your existing environment to run in a smoother manor. Take great care when changing things as one small slip of the hand can give you a forest wide failure… Whenever starting with an ADDS deployment I ask the client the following questions:  What are your long term plans and goals?  How flexible do you want it? Are you currently linux heavy and want to keep this or can we go for an all Microsoft design? Those three questions should give some sort of indicator what direction can be taken and if the client has thought about some things themselves :).  The technical side of things  What is next to consider is what kind of infrastructure is already in place. For these series I'll keep it simple and introduce some general concepts without going in to depth on integrating ADDS with other DNS services.  Building from the ground up means we need to consider our layers on which our infrastructure will rely. In my view that goes as follows:  Network (WAN/LAN links and physical sites DNS Namespacing All in one domain or split up in different domains/forests? Security (both for ADDS and physical sites) The network side of things  Looking at how the network is currently set up can potentially teach us a large deal about the client. Do they have multiple physical site? What network speeds exist between these sites, etc… Depending on this information we will design our site links (which controls replication) in future stages. DNS Namespacing Maybe the single most intresting thing to know is what the domain will be named (ADDS will need a DNS domain with the same name) and where this will be hosted. Note that active directory can be set up with a singe name (aka contoso instead of contoso.com) but it is highly recommended to never do this. If you do end up with a domain like that for some reason there will be a lot of services that are going to give you good grief in the future (exchange being one of them). So one of the best practises would be always to use a double name (contoso.com or contoso.lan for example). Internal namespace A single namespace is just what it sounds like. You have a DNS domain which is different internally from what the client has as an external namespace. f.e. contoso.com as an external name (out on the internet) and contoso.lan on the internal network. his setup is has its advantages in that you have more obscurity from the internet in the DNS side of this but it will require additional work to publish services to the web. External namespace Quite like the internal namespace only here you do not differ the internal namespace of the company from what is known on the internet. In this implementation you would host your own DNS servers for the external domain inside the network. Or in other words, any external computer doing a DNS lookup would contact your internal DNS server for the resolution. Generally speaking this set up is a bad idea from the security side of things. Split DNS Whilst using an external namespace design is fairly easy it involves a lot of security risks. Opening up you ADDS DSN servers for lookups exposes your entire network to the internet and should be avoided at any cost. And that is where the "split DNS" design comes in. In this setup up would still have the same namespace internally and externally but you would be using different DNS servers for lookups on the external network who have no records of your internal resources unless you explicitly publish them. All in one or not? In determining your active directory design you can look at the following possibilities:  Single forest, Single domain Single forest, multiple domains Multiple forests, multiple domains I've listed the possibilities for design in increasing order of administrative magnitude. Microsoft recommends trying to use a single forest, single domain in as much situations as possible. It is, however, always possible that you require your services to be seperated from your users in a resource forest with trusts set up between the different forests. To start out I would go with the single forest design to avoid complexity unless there are strict requirements to have multiple forests. Security What kind of security is required on the domain and does this reflect the physical security on the sites? Not every client can afford to have a domain controller in a secluded server room on every site and it is exactly for that reason that Microsoft introduced the RODC (read only domain controller). A RODC is a domain controller that has been limited in functionality, in essence it will only cache the data you explicitly tell it to cache and in the case of a DC compromise (it being stolen) only a limited number of accounts will need to be affected. Th- Th- Th- That’s all folks! Well at least for now! In future editions of this series we’ll be walking through the different task that need to be done and the thought which needs to be put in to it. But for all editions we’ll be going from the concept of running a single forest, single domain with a split DNS setup… See you next time!

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  • Enum driving a Visual State change via the ViewModel

    - by Chris Skardon
    Exciting title eh? So, here’s the problem, I want to use my ViewModel to drive my Visual State, I’ve used the ‘DataStateBehavior’ before, but the trouble with it is that it only works for bool values, and the minute you jump to more than 2 Visual States, you’re kind of screwed. A quick search has shown up a couple of points of interest, first, the DataStateSwitchBehavior, which is part of the Expression Samples (on Codeplex), and also available via Pete Blois’ blog. The second interest is to use a DataTrigger with GoToStateAction (from the Silverlight forums). So, onwards… first let’s create a basic switch Visual State, so, a DataObj with one property: IsAce… public class DataObj : NotifyPropertyChanger { private bool _isAce; public bool IsAce { get { return _isAce; } set { _isAce = value; RaisePropertyChanged("IsAce"); } } } The ‘NotifyPropertyChanger’ is literally a base class with RaisePropertyChanged, implementing INotifyPropertyChanged. OK, so we then create a ViewModel: public class MainPageViewModel : NotifyPropertyChanger { private DataObj _dataObj; public MainPageViewModel() { DataObj = new DataObj {IsAce = true}; ChangeAcenessCommand = new RelayCommand(() => DataObj.IsAce = !DataObj.IsAce); } public ICommand ChangeAcenessCommand { get; private set; } public DataObj DataObj { get { return _dataObj; } set { _dataObj = value; RaisePropertyChanged("DataObj"); } } } Aaaand finally – hook it all up to the XAML, which is a very simple UI: A Rectangle, a TextBlock and a Button. The Button is hooked up to ChangeAcenessCommand, the TextBlock is bound to the ‘DataObj.IsAce’ property and the Rectangle has 2 visual states: IsAce and NotAce. To make the Rectangle change it’s visual state I’ve used a DataStateBehavior inside the Layout Root Grid: <i:Interaction.Behaviors> <ei:DataStateBehavior Binding="{Binding DataObj.IsAce}" Value="true" TrueState="IsAce" FalseState="NotAce"/> </i:Interaction.Behaviors> So now we have the button changing the ‘IsAce’ property and giving us the other visual state: Great! So – the next stage is to get that to work inside a DataTemplate… Which (thankfully) is easy money. All we do is add a ListBox to the View and an ObservableCollection to the ViewModel. Well – ok, a little bit more than that. Once we’ve got the ListBox with it’s ItemsSource property set, it’s time to add the DataTemplate itself. Again, this isn’t exactly taxing, and is purely going to be a Grid with a Textblock and a Rectangle (again, I’m nothing if not consistent). Though, to be a little jazzy I’ve swapped the rectangle to the other side (living the dream). So, all that’s left is to add some States to the template.. (Yes – you can do that), these can be the same names as the others, or indeed, something else, I have chosen to stick with the same names and take the extra confusion hit right on the nose. Once again, I add the DataStateBehavior to the root Grid element: <i:Interaction.Behaviors> <ei:DataStateBehavior Binding="{Binding IsAce}" Value="true" TrueState="IsAce" FalseState="NotAce"/> </i:Interaction.Behaviors> The key difference here is the ‘Binding’ attribute, where I’m now binding to the IsAce property directly, and boom! It’s all gravy!   So far, so good. We can use boolean values to change the visual states, and (crucially) it works in a DataTemplate, bingo! Now. Onwards to the Enum part of this (finally!). Obviously we can’t use the DataStateBehavior, it' only gives us true/false options. So, let’s give the GoToStateAction a go. Now, I warn you, things get a bit complex from here, instead of a bool with 2 values, I’m gonna max it out and bring in an Enum with 3 (count ‘em) 3 values: Red, Amber and Green (those of you with exceptionally sharp minds will be reminded of traffic lights). We’re gonna have a rectangle which also has 3 visual states – cunningly called ‘Red’, ‘Amber’ and ‘Green’. A new class called DataObj2: public class DataObj2 : NotifyPropertyChanger { private Status _statusValue; public DataObj2(Status status) { StatusValue = status; } public Status StatusValue { get { return _statusValue; } set { _statusValue = value; RaisePropertyChanged("StatusValue"); } } } Where ‘Status’ is my enum. Good times are here! Ok, so let’s get to the beefy stuff. So, we’ll start off in the same manner as the last time, we will have a single DataObj2 instance available to the Page and bind to that. Let’s add some Triggers (these are in the LayoutRoot again). <i:Interaction.Triggers> <ei:DataTrigger Binding="{Binding DataObject2.StatusValue}" Value="Amber"> <ei:GoToStateAction StateName="Amber" UseTransitions="False" /> </ei:DataTrigger> <ei:DataTrigger Binding="{Binding DataObject2.StatusValue}" Value="Green"> <ei:GoToStateAction StateName="Green" UseTransitions="False" /> </ei:DataTrigger> <ei:DataTrigger Binding="{Binding DataObject2.StatusValue}" Value="Red"> <ei:GoToStateAction StateName="Red" UseTransitions="False" /> </ei:DataTrigger> </i:Interaction.Triggers> So what we’re saying here is that when the DataObject2.StatusValue is equal to ‘Red’ then we’ll go to the ‘Red’ state. Same deal for Green and Amber (but you knew that already). Hook it all up and start teh project. Hmm. Just grey. Not what I wanted. Ok, let’s add a ‘ChangeStatusCommand’, hook that up to a button and give it a whirl: Right, so the DataTrigger isn’t picking up the data on load. On the plus side, changing the status is making the visual states change. So. We’ll cross the ‘Grey’ hurdle in a bit, what about doing the same in the DataTemplate? <Codey Codey/> Grey again, but if we press the button: (I should mention, pressing the button sets the StatusValue property on the DataObj2 being represented to the next colour). Right. Let’s look at this ‘Grey’ issue. First ‘fix’ (and I use the term ‘fix’ in a very loose way): The Dispatcher Fix This involves using the Dispatcher on the View to call something like ‘RefreshProperties’ on the ViewModel, which will in turn raise all the appropriate ‘PropertyChanged’ events on the data objects being represented. So, here goes, into turdcode-ville – population – me: First, add the ‘RefreshProperties’ method to the DataObj2: internal void RefreshProperties() { RaisePropertyChanged("StatusValue"); } (shudder) Now, add it to the hosting ViewModel: public void RefreshProperties() { DataObject2.RefreshProperties(); if (DataObjects != null && DataObjects.Count > 0) { foreach (DataObj2 dataObject in DataObjects) dataObject.RefreshProperties(); } } (double shudder) and now for the cream on the cake, adding the following line to the code behind of the View: Dispatcher.BeginInvoke(() => ((MoreVisualStatesViewModel)DataContext).RefreshProperties()); So, what does this *ahem* code give us: Awesome, it makes the single bound data object show the colour, but frankly ignores the DataTemplate items. This (by the way) is the same output you get from: Dispatcher.BeginInvoke(() => ((MoreVisualStatesViewModel)DataContext).ChangeStatusCommand.Execute(null)); So… Where does that leave me? What about adding a button to the Page to refresh the properties – maybe it’s a timer thing? Yes, that works. Right, what about using the Loaded event then eh? Loaded += (s, e) => ((MoreVisualStatesViewModel) DataContext).RefreshProperties(); Ahhh No. What about converting the DataTemplate into a UserControl? Anything is worth a shot.. Though – I still suspect I’m going to have to ‘RefreshProperties’ if I want the rectangles to update. Still. No. This DataTemplate DataTrigger binding is becoming a bit of a pain… I can’t add a ‘refresh’ button to the actual code base, it’s not exactly user friendly. I’m going to end this one now, and put some investigating into the use of the DataStateSwitchBehavior (all the ones I’ve found, well, all 2 of them are working in SL3, but not 4…)

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  • Get Started using Build-Deploy-Test Workflow with TFS 2012

    - by Jakob Ehn
    TFS 2012 introduces a new type of Lab environment called Standard Environment. This allows you to setup a full Build Deploy Test (BDT) workflow that will build your application, deploy it to your target machine(s) and then run a set of tests on that server to verify the deployment. In TFS 2010, you had to use System Center Virtual Machine Manager and involve half of your IT department to get going. Now all you need is a server (virtual or physical) where you want to deploy and test your application. You don’t even have to install a test agent on the machine, TFS 2012 will do this for you! Although each step is rather simple, the entire process of setting it up consists of a bunch of steps. So I thought that it could be useful to run through a typical setup.I will also link to some good guidance from MSDN on each topic. High Level Steps Install and configure Visual Studio 2012 Test Controller on Target Server Create Standard Environment Create Test Plan with Test Case Run Test Case Create Coded UI Test from Test Case Associate Coded UI Test with Test Case Create Build Definition using LabDefaultTemplate 1. Install and Configure Visual Studio 2012 Test Controller on Target Server First of all, note that you do not have to have the Test Controller running on the target server. It can be running on another server, as long as the Test Agent can communicate with the test controller and the test controller can communicate with the TFS server. If you have several machines in your environment (web server, database server etc..), the test controller can be installed either on one of those machines or on a dedicated machine. To install the test controller, simply mount the Visual Studio Agents media on the server and browse to the vstf_controller.exe file located in the TestController folder. Run through the installation, you might need to reboot the server since it installs .NET 4.5. When the test controller is installed, the Test Controller configuration tool will launch automatically (if it doesn’t, you can start it from the Start menu). Here you will supply the credentials of the account running the test controller service. Note that this account will be given the necessary permissions in TFS during the configuration. Make sure that you have entered a valid account by pressing the Test link. Also, you have to register the test controller with the TFS collection where your test plan is located (and usually the code base of course) When you press Apply Settings, all the configuration will be done. You might get some warnings at the end, that might or might not cause a problem later. Be sure to read them carefully.   For more information about configuring your test controllers, see Setting Up Test Controllers and Test Agents to Manage Tests with Visual Studio 2. Create Standard Environment Now you need to create a Lab environment in Microsoft Test Manager. Since we are using an existing physical or virtual machine we will create a Standard Environment. Open MTM and go to Lab Center. Click New to create a new environment Enter a name for the environment. Since this environment will only contain one machine, we will use the machine name for the environment (TargetServer in this case) On the next page, click Add to add a machine to the environment. Enter the name of the machine (TargetServer.Domain.Com), and give it the Web Server role. The name must be reachable both from your machine during configuration and from the TFS app tier server. You also need to supply an account that is a local administration on the target server. This is needed in order to automatically install a test agent later on the machine. On the next page, you can add tags to the machine. This is not needed in this scenario so go to the next page. Here you will specify which test controller to use and that you want to run UI tests on this environment. This will in result in a Test Agent being automatically installed and configured on the target server. The name of the machine where you installed the test controller should be available on the drop down list (TargetServer in this sample). If you can’t see it, you might have selected a different TFS project collection. Press Next twice and then Verify to verify all the settings: Press finish. This will now create and prepare the environment, which means that it will remote install a test agent on the machine. As part of this installation, the remote server will be restarted. 3-5. Create Test Plan, Run Test Case, Create Coded UI Test I will not cover step 3-5 here, there are plenty of information on how you create test plans and test cases and automate them using Coded UI Tests. In this example I have a test plan called My Application and it contains among other things a test suite called Automated Tests where I plan to put test cases that should be automated and executed as part of the BDT workflow. For more information about Coded UI Tests, see Verifying Code by Using Coded User Interface Tests   6. Associate Coded UI Test with Test Case OK, so now we want to automate our Coded UI Test and have it run as part of the BDT workflow. You might think that you coded UI test already is automated, but the meaning of the term here is that you link your coded UI Test to an existing Test Case, thereby making the Test Case automated. And the test case should be part of the test suite that we will run during the BDT. Open the solution that contains the coded UI test method. Open the Test Case work item that you want to automate. Go to the Associated Automation tab and click on the “…” button. Select the coded UI test that you corresponds to the test case: Press OK and the save the test case For more information about associating an automated test case with a test case, see How to: Associate an Automated Test with a Test Case 7. Create Build Definition using LabDefaultTemplate Now we are ready to create a build definition that will implement the full BDT workflow. For this purpose we will use the LabDefaultTemplate.11.xaml that comes out of the box in TFS 2012. This build process template lets you take the output of another build and deploy it to each target machine. Since the deployment process will be running on the target server, you will have less problem with permissions and firewalls than if you were to remote deploy your solution. So, before creating a BDT workflow build definition, make sure that you have an existing build definition that produces a release build of your application. Go to the Builds hub in Team Explorer and select New Build Definition Give the build definition a meaningful name, here I called it MyApplication.Deploy Set the trigger to Manual Define a workspace for the build definition. Note that a BDT build doesn’t really need a workspace, since all it does is to launch another build definition and deploy the output of that build. But TFS doesn’t allow you to save a build definition without adding at least one mapping. On Build Defaults, select the build controller. Since this build actually won’t produce any output, you can select the “This build does not copy output files to a drop folder” option. On the process tab, select the LabDefaultTemplate.11.xaml. This is usually located at $/TeamProject/BuildProcessTemplates/LabDefaultTemplate.11.xaml. To configure it, press the … button on the Lab Process Settings property First, select the environment that you created before: Select which build that you want to deploy and test. The “Select an existing build” option is very useful when developing the BDT workflow, because you do not have to run through the target build every time, instead it will basically just run through the deployment and test steps which speeds up the process. Here I have selected to queue a new build of the MyApplication.Test build definition On the deploy tab, you need to specify how the application should be installed on the target server. You can supply a list of deployment scripts with arguments that will be executed on the target server. In this example I execute the generated web deploy command file to deploy the solution. If you for example have databases you can use sqlpackage.exe to deploy the database. If you are producing MSI installers in your build, you can run them using msiexec.exe and so on. A good practice is to create a batch file that contain the entire deployment that you can run both locally and on the target server. Then you would just execute the deployment batch file here in one single step. The workflow defines some variables that are useful when running the deployments. These variables are: $(BuildLocation) The full path to where your build files are located $(InternalComputerName_<VM Name>) The computer name for a virtual machine in a SCVMM environment $(ComputerName_<VM Name>) The fully qualified domain name of the virtual machine As you can see, I specify the path to the myapplication.deploy.cmd file using the $(BuildLocation) variable, which is the drop folder of the MyApplication.Test build. Note: The test agent account must have read permission in this drop location. You can find more information here on Building your Deployment Scripts On the last tab, we specify which tests to run after deployment. Here I select the test plan and the Automated Tests test suite that we saw before: Note that I also selected the automated test settings (called TargetServer in this case) that I have defined for my test plan. In here I define what data that should be collected as part of the test run. For more information about test settings, see Specifying Test Settings for Microsoft Test Manager Tests We are done! Queue your BDT build and wait for it to finish. If the build succeeds, your build summary should look something like this:

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  • Making a Statement: How to retrieve the T-SQL statement that caused an event

    - by extended_events
    If you’ve done any troubleshooting of T-SQL, you know that sooner or later, probably sooner, you’re going to want to take a look at the actual statements you’re dealing with. In extended events we offer an action (See the BOL topic that covers Extended Events Objects for a description of actions) named sql_text that seems like it is just the ticket. Well…not always – sounds like a good reason for a blog post. When is a statement not THE statement? The sql_text action returns the same information that is returned from DBCC INPUTBUFFER, which may or may not be what you want. For example, if you execute a stored procedure, the sql_text action will return something along the lines of “EXEC sp_notwhatiwanted” assuming that is the statement you sent from the client. Often times folks would like something more specific, like the actual statements that are being run from within the stored procedure or batch. Enter the stack Extended events offers another action, this one with the descriptive name of tsql_stack, that includes the sql_handle and offset information about the statements being run when an event occurs. With the sql_handle and offset values you can retrieve the specific statement you seek using the DMV dm_exec_sql_statement. The BOL topic for dm_exec_sql_statement provides an example for how to extract this information, so I’ll cover the gymnastics required to get the sql_handle and offset values out of the tsql_stack data collected by the action. I’m the first to admit that this isn’t pretty, but this is what we have in SQL Server 2008 and 2008 R2. We will be making it easier to get statement level information in the next major release of SQL Server. The sample code For this example I have a stored procedure that includes multiple statements and I have a need to differentiate between those two statements in my tracing. I’m going to track two events: module_end tracks the completion of the stored procedure execution and sp_statement_completed tracks the execution of each statement within a stored procedure. I’m adding the tsql_stack events (since that’s the topic of this post) and the sql_text action for comparison sake. (If you have questions about creating event sessions, check out Pedro’s post Introduction to Extended Events.) USE AdventureWorks2008GO -- Test SPCREATE PROCEDURE sp_multiple_statementsASSELECT 'This is the first statement'SELECT 'this is the second statement'GO -- Create a session to look at the spCREATE EVENT SESSION track_sprocs ON SERVERADD EVENT sqlserver.module_end (ACTION (sqlserver.tsql_stack, sqlserver.sql_text)),ADD EVENT sqlserver.sp_statement_completed (ACTION (sqlserver.tsql_stack, sqlserver.sql_text))ADD TARGET package0.ring_bufferWITH (MAX_DISPATCH_LATENCY = 1 SECONDS)GO -- Start the sessionALTER EVENT SESSION track_sprocs ON SERVERSTATE = STARTGO -- Run the test procedureEXEC sp_multiple_statementsGO -- Stop collection of events but maintain ring bufferALTER EVENT SESSION track_sprocs ON SERVERDROP EVENT sqlserver.module_end,DROP EVENT sqlserver.sp_statement_completedGO Aside: Altering the session to drop the events is a neat little trick that allows me to stop collection of events while keeping in-memory targets such as the ring buffer available for use. If you stop the session the in-memory target data is lost. Now that we’ve collected some events related to running the stored procedure, we need to do some processing of the data. I’m going to do this in multiple steps using temporary tables so you can see what’s going on; kind of like having to “show your work” on a math test. The first step is to just cast the target data into XML so I can work with it. After that you can pull out the interesting columns, for our purposes I’m going to limit the output to just the event name, object name, stack and sql text. You can see that I’ve don a second CAST, this time of the tsql_stack column, so that I can further process this data. -- Store the XML data to a temp tableSELECT CAST( t.target_data AS XML) xml_dataINTO #xml_event_dataFROM sys.dm_xe_sessions s INNER JOIN sys.dm_xe_session_targets t    ON s.address = t.event_session_addressWHERE s.name = 'track_sprocs' SELECT * FROM #xml_event_data -- Parse the column data out of the XML blockSELECT    event_xml.value('(./@name)', 'varchar(100)') as [event_name],    event_xml.value('(./data[@name="object_name"]/value)[1]', 'varchar(255)') as [object_name],    CAST(event_xml.value('(./action[@name="tsql_stack"]/value)[1]','varchar(MAX)') as XML) as [stack_xml],    event_xml.value('(./action[@name="sql_text"]/value)[1]', 'varchar(max)') as [sql_text]INTO #event_dataFROM #xml_event_data    CROSS APPLY xml_data.nodes('//event') n (event_xml) SELECT * FROM #event_data event_name object_name stack_xml sql_text sp_statement_completed NULL <frame level="1" handle="0x03000500D0057C1403B79600669D00000100000000000000" line="4" offsetStart="94" offsetEnd="172" /><frame level="2" handle="0x01000500CF3F0331B05EC084000000000000000000000000" line="1" offsetStart="0" offsetEnd="-1" /> EXEC sp_multiple_statements sp_statement_completed NULL <frame level="1" handle="0x03000500D0057C1403B79600669D00000100000000000000" line="6" offsetStart="174" offsetEnd="-1" /><frame level="2" handle="0x01000500CF3F0331B05EC084000000000000000000000000" line="1" offsetStart="0" offsetEnd="-1" /> EXEC sp_multiple_statements module_end sp_multiple_statements <frame level="1" handle="0x03000500D0057C1403B79600669D00000100000000000000" line="0" offsetStart="0" offsetEnd="0" /><frame level="2" handle="0x01000500CF3F0331B05EC084000000000000000000000000" line="1" offsetStart="0" offsetEnd="-1" /> EXEC sp_multiple_statements After parsing the columns it’s easier to see what is recorded. You can see that I got back two sp_statement_completed events, which makes sense given the test procedure I’m running, and I got back a single module_end for the entire statement. As described, the sql_text isn’t telling me what I really want to know for the first two events so a little extra effort is required. -- Parse the tsql stack information into columnsSELECT    event_name,    object_name,    frame_xml.value('(./@level)', 'int') as [frame_level],    frame_xml.value('(./@handle)', 'varchar(MAX)') as [sql_handle],    frame_xml.value('(./@offsetStart)', 'int') as [offset_start],    frame_xml.value('(./@offsetEnd)', 'int') as [offset_end]INTO #stack_data    FROM #event_data        CROSS APPLY    stack_xml.nodes('//frame') n (frame_xml)    SELECT * from #stack_data event_name object_name frame_level sql_handle offset_start offset_end sp_statement_completed NULL 1 0x03000500D0057C1403B79600669D00000100000000000000 94 172 sp_statement_completed NULL 2 0x01000500CF3F0331B05EC084000000000000000000000000 0 -1 sp_statement_completed NULL 1 0x03000500D0057C1403B79600669D00000100000000000000 174 -1 sp_statement_completed NULL 2 0x01000500CF3F0331B05EC084000000000000000000000000 0 -1 module_end sp_multiple_statements 1 0x03000500D0057C1403B79600669D00000100000000000000 0 0 module_end sp_multiple_statements 2 0x01000500CF3F0331B05EC084000000000000000000000000 0 -1 Parsing out the stack information doubles the fun and I get two rows for each event. If you examine the stack from the previous table, you can see that each stack has two frames and my query is parsing each event into frames, so this is expected. There is nothing magic about the two frames, that’s just how many I get for this example, it could be fewer or more depending on your statements. The key point here is that I now have a sql_handle and the offset values for those handles, so I can use dm_exec_sql_statement to get the actual statement. Just a reminder, this DMV can only return what is in the cache – if you have old data it’s possible your statements have been ejected from the cache. “Old” is a relative term when talking about caches and can be impacted by server load and how often your statement is actually used. As with most things in life, your mileage may vary. SELECT    qs.*,     SUBSTRING(st.text, (qs.offset_start/2)+1,         ((CASE qs.offset_end          WHEN -1 THEN DATALENGTH(st.text)         ELSE qs.offset_end         END - qs.offset_start)/2) + 1) AS statement_textFROM #stack_data AS qsCROSS APPLY sys.dm_exec_sql_text(CONVERT(varbinary(max),sql_handle,1)) AS st event_name object_name frame_level sql_handle offset_start offset_end statement_text sp_statement_completed NULL 1 0x03000500D0057C1403B79600669D00000100000000000000 94 172 SELECT 'This is the first statement' sp_statement_completed NULL 1 0x03000500D0057C1403B79600669D00000100000000000000 174 -1 SELECT 'this is the second statement' module_end sp_multiple_statements 1 0x03000500D0057C1403B79600669D00000100000000000000 0 0 C Now that looks more like what we were after, the statement_text field is showing the actual statement being run when the sp_statement_completed event occurs. You’ll notice that it’s back down to one row per event, what happened to frame 2? The short answer is, “I don’t know.” In SQL Server 2008 nothing is returned from dm_exec_sql_statement for the second frame and I believe this to be a bug; this behavior has changed in the next major release and I see the actual statement run from the client in frame 2. (In other words I see the same statement that is returned by the sql_text action  or DBCC INPUTBUFFER) There is also something odd going on with frame 1 returned from the module_end event; you can see that the offset values are both 0 and only the first letter of the statement is returned. It seems like the offset_end should actually be –1 in this case and I’m not sure why it’s not returning this correctly. This behavior is being investigated and will hopefully be corrected in the next major version. You can workaround this final oddity by ignoring the offsets and just returning the entire cached statement. SELECT    event_name,    sql_handle,    ts.textFROM #stack_data    CROSS APPLY sys.dm_exec_sql_text(CONVERT(varbinary(max),sql_handle,1)) as ts event_name sql_handle text sp_statement_completed 0x0300070025999F11776BAF006F9D00000100000000000000 CREATE PROCEDURE sp_multiple_statements AS SELECT 'This is the first statement' SELECT 'this is the second statement' sp_statement_completed 0x0300070025999F11776BAF006F9D00000100000000000000 CREATE PROCEDURE sp_multiple_statements AS SELECT 'This is the first statement' SELECT 'this is the second statement' module_end 0x0300070025999F11776BAF006F9D00000100000000000000 CREATE PROCEDURE sp_multiple_statements AS SELECT 'This is the first statement' SELECT 'this is the second statement' Obviously this gives more than you want for the sp_statement_completed events, but it’s the right information for module_end. I leave it to you to determine when this information is needed and use the workaround when appropriate. Aside: You might think it’s odd that I’m showing apparent bugs with my samples, but you’re going to see this behavior if you use this method, so you need to know about it.I’m all about transparency. Happy Eventing- Mike Share this post: email it! | bookmark it! | digg it! | reddit! | kick it! | live it!

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  • Solaris: What comes next?

    - by alanc
    As you probably know by now, a few months ago, we released Solaris 11 after years of development. That of course means we now need to figure out what comes next - if Solaris 11 is “The First Cloud OS”, then what do we need to make future releases of Solaris be, to be modern and competitive when they're released? So we've been having planning and brainstorming meetings, and I've captured some notes here from just one of those we held a couple weeks ago with a number of the Silicon Valley based engineers. Now before someone sees an idea here and calls their product rep wanting to know what's up, please be warned what follows are rough ideas, and as I'll discuss later, none of them have any committment, schedule, working code, or even plan for integration in any possible future product at this time. (Please don't make me force you to read the full Oracle future product disclaimer here, you should know it by heart already from the front of every Oracle product slide deck.) To start with, we did some background research, looking at ideas from other Oracle groups, and competitive OS'es. We examined what was hot in the technology arena and where the interesting startups were heading. We then looked at Solaris to see where we could apply those ideas. Making Network Admins into Socially Networking Admins We all know an admin who has grumbled about being the only one stuck late at work to fix a problem on the server, or having to work the weekend alone to do scheduled maintenance. But admins are humans (at least most are), and crave companionship and community with their fellow humans. And even when they're alone in the server room, they're never far from a network connection, allowing access to the wide world of wonders on the Internet. Our solution here is not building a new social network - there's enough of those already, and Oracle even has its own Oracle Mix social network already. What we proposed is integrating Solaris features to help engage our system admins with these social networks, building community and bringing them recognition in the workplace, using achievement recognition systems as found in many popular gaming platforms. For instance, if you had a Facebook account, and a group of admin friends there, you could register it with our Social Network Utility For Facebook, and then your friends might see: Alan earned the achievement Critically Patched (April 2012) for patching all his servers. Matt is only at 50% - encourage him to complete this achievement today! To avoid any undue risk of advertising who has unpatched servers that are easier targets for hackers to break into, this information would be tightly protected via Facebook's world-renowned privacy settings to avoid it falling into the wrong hands. A related form of gamification we considered was replacing simple certfications with role-playing-game-style Experience Levels. Instead of just knowing an admin passed a test establishing a given level of competency, these would provide recruiters with a more detailed level of how much real-world experience an admin has. Achievements such as the one above would feed into it, but larger numbers of experience points would be gained by tougher or more critical tasks - such as recovering a down system, or migrating a service to a new platform. (As long as it was an Oracle platform of course - migrating to an HP or IBM platform would cause the admin to lose points with us.) Unfortunately, we couldn't figure out a good way to prevent (if you will) “gaming” the system. For instance, a disgruntled admin might decide to start ignoring warnings from FMA that a part is beginning to fail or skip preventative maintenance, in the hopes that they'd cause a catastrophic failure to earn more points for bolstering their resume as they look for a job elsewhere, and not worrying about the effect on your business of a mission critical server going down. More Z's for ZFS Our suggested new feature for ZFS was inspired by the worlds most successful Z-startup of all time: Zynga. Using the Social Network Utility For Facebook described above, we'd tie it in with ZFS monitoring to help you out when you find yourself in a jam needing more disk space than you have, and can't wait a month to get a purchase order through channels to buy more. Instead with the click of a button you could post to your group: Alan can't find any space in his server farm! Can you help? Friends could loan you some space on their connected servers for a few weeks, knowing that you'd return the favor when needed. ZFS would create a new filesystem for your use on their system, and securely share it with your system using Kerberized NFS. If none of your friends have space, then you could buy temporary use space in small increments at affordable rates right there in Facebook, using your Facebook credits, and then file an expense report later, after the urgent need has passed. Universal Single Sign On One thing all the engineers agreed on was that we still had far too many "Single" sign ons to deal with in our daily work. On the web, every web site used to have its own password database, forcing us to hope we could remember what login name was still available on each site when we signed up, and which unique password we came up with to avoid having to disclose our other passwords to a new site. In recent years, the web services world has finally been reducing the number of logins we have to manage, with many services allowing you to login using your identity from Google, Twitter or Facebook. So we proposed following their lead, introducing PAM modules for web services - no more would you have to type in whatever login name IT assigned and try to remember the password you chose the last time password aging forced you to change it - you'd simply choose which web service you wanted to authenticate against, and would login to your Solaris account upon reciept of a cookie from their identity service. Pinning notes to the cloud We also all noted that we all have our own pile of notes we keep in our daily work - in text files in our home directory, in notebooks we carry around, on white boards in offices and common areas, on sticky notes on our monitors, or on scraps of paper pinned to our bulletin boards. The contents of the notes vary, some are things just for us, some are useful for our groups, some we would share with the world. For instance, when our group moved to a new building a couple years ago, we had a white board in the hallway listing all the NIS & DNS servers, subnets, and other network configuration information we needed to set up our Solaris machines after the move. Similarly, as Solaris 11 was finishing and we were all learning the new network configuration commands, we shared notes in wikis and e-mails with our fellow engineers. Users may also remember one of the popular features of Sun's old BigAdmin site was a section for sharing scripts and tips such as these. Meanwhile, the online "pin board" at Pinterest is taking the web by storm. So we thought, why not mash those up to solve this problem? We proposed a new BigAddPin site where users could “pin” notes, command snippets, configuration information, and so on. For instance, once they had worked out the ideal Automated Installation manifest for their app server, they could pin it up to share with the rest of their group, or choose to make it public as an example for the world. Localized data, such as our group's notes on the servers for our subnet, could be shared only to users connecting from that subnet. And notes that they didn't want others to see at all could be marked private, such as the list of phone numbers to call for late night pizza delivery to the machine room, the birthdays and anniversaries they can never remember but would be sleeping on the couch if they forgot, or the list of automatically generated completely random, impossible to remember root passwords to all their servers. For greater integration with Solaris, we'd put support right into the command shells — redirect output to a pinned note, set your path to include pinned notes as scripts you can run, or bring up your recent shell history and pin a set of commands to save for the next time you need to remember how to do that operation. Location service for Solaris servers A longer term plan would involve convincing the hardware design groups to put GPS locators with wireless transmitters in future server designs. This would help both admins and service personnel trying to find servers in todays massive data centers, and could feed into location presence apps to help show potential customers that while they may not see many Solaris machines on the desktop any more, they are all around. For instance, while walking down Wall Street it might show “There are over 2000 Solaris computers in this block.” [Note: this proposal was made before the recent media coverage of a location service aggregrator app with less noble intentions, and in hindsight, we failed to consider what happens when such data similarly falls into the wrong hands. We certainly wouldn't want our app to be misinterpreted as “There are over $20 million dollars of SPARC servers in this building, waiting for you to steal them.” so it's probably best it was rejected.] Harnessing the power of the GPU for Security Most modern OS'es make use of the widespread availability of high powered GPU hardware in today's computers, with desktop environments requiring 3-D graphics acceleration, whether in Ubuntu Unity, GNOME Shell on Fedora, or Aero Glass on Windows, but we haven't yet made Solaris fully take advantage of this, beyond our basic offering of Compiz on the desktop. Meanwhile, more businesses are interested in increasing security by using biometric authentication, but must also comply with laws in many countries preventing discrimination against employees with physical limations such as missing eyes or fingers, not to mention the lost productivity when employees can't login due to tinted contacts throwing off a retina scan or a paper cut changing their fingerprint appearance until it heals. Fortunately, the two groups considering these problems put their heads together and found a common solution, using 3D technology to enable authentication using the one body part all users are guaranteed to have - pam_phrenology.so, a new PAM module that uses an array USB attached web cams (or just one if the user is willing to spin their chair during login) to take pictures of the users head from all angles, create a 3D model and compare it to the one in the authentication database. While Mythbusters has shown how easy it can be to fool common fingerprint scanners, we have not yet seen any evidence that people can impersonate the shape of another user's cranium, no matter how long they spend beating their head against the wall to reshape it. This could possibly be extended to group users, using modern versions of some of the older phrenological studies, such as giving all users with long grey beards access to the System Architect role, or automatically placing users with pointy spikes in their hair into an easy use mode. Unfortunately, there are still some unsolved technical challenges we haven't figured out how to overcome. Currently, a visit to the hair salon causes your existing authentication to expire, and some users have found that shaving their heads is the only way to avoid bad hair days becoming bad login days. Reaction to these ideas After gathering all our notes on these ideas from the engineering brainstorming meeting, we took them in to present to our management. Unfortunately, most of their reaction cannot be printed here, and they chose not to accept any of these ideas as they were, but they did have some feedback for us to consider as they sent us back to the drawing board. They strongly suggested our ideas would be better presented if we weren't trying to decipher ink blotches that had been smeared by the condensation when we put our pint glasses on the napkins we were taking notes on, and to that end let us know they would not be approving any more engineering offsites in Irish themed pubs on the Friday of a Saint Patrick's Day weekend. (Hopefully they mean that situation specifically and aren't going to deny the funding for travel to this year's X.Org Developer's Conference just because it happens to be in Bavaria and ending on the Friday of the weekend Oktoberfest starts.) They recommended our research techniques could be improved over just sitting around reading blogs and checking our Facebook, Twitter, and Pinterest accounts, such as considering input from alternate viewpoints on topics such as gamification. They also mentioned that Oracle hadn't fully adopted some of Sun's common practices and we might have to try harder to get those to be accepted now that we are one unified company. So as I said at the beginning, don't pester your sales rep just yet for any of these, since they didn't get approved, but if you have better ideas, pass them on and maybe they'll get into our next batch of planning.

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  • Table Variables: an empirical approach.

    - by Phil Factor
    It isn’t entirely a pleasant experience to publish an article only to have it described on Twitter as ‘Horrible’, and to have it criticized on the MVP forum. When this happened to me in the aftermath of publishing my article on Temporary tables recently, I was taken aback, because these critics were experts whose views I respect. What was my crime? It was, I think, to suggest that, despite the obvious quirks, it was best to use Table Variables as a first choice, and to use local Temporary Tables if you hit problems due to these quirks, or if you were doing complex joins using a large number of rows. What are these quirks? Well, table variables have advantages if they are used sensibly, but this requires some awareness by the developer about the potential hazards and how to avoid them. You can be hit by a badly-performing join involving a table variable. Table Variables are a compromise, and this compromise doesn’t always work out well. Explicit indexes aren’t allowed on Table Variables, so one cannot use covering indexes or non-unique indexes. The query optimizer has to make assumptions about the data rather than using column distribution statistics when a table variable is involved in a join, because there aren’t any column-based distribution statistics on a table variable. It assumes a reasonably even distribution of data, and is likely to have little idea of the number of rows in the table variables that are involved in queries. However complex the heuristics that are used might be in determining the best way of executing a SQL query, and they most certainly are, the Query Optimizer is likely to fail occasionally with table variables, under certain circumstances, and produce a Query Execution Plan that is frightful. The experienced developer or DBA will be on the lookout for this sort of problem. In this blog, I’ll be expanding on some of the tests I used when writing my article to illustrate the quirks, and include a subsequent example supplied by Kevin Boles. A simplified example. We’ll start out by illustrating a simple example that shows some of these characteristics. We’ll create two tables filled with random numbers and then see how many matches we get between the two tables. We’ll forget indexes altogether for this example, and use heaps. We’ll try the same Join with two table variables, two table variables with OPTION (RECOMPILE) in the JOIN clause, and with two temporary tables. It is all a bit jerky because of the granularity of the timing that isn’t actually happening at the millisecond level (I used DATETIME). However, you’ll see that the table variable is outperforming the local temporary table up to 10,000 rows. Actually, even without a use of the OPTION (RECOMPILE) hint, it is doing well. What happens when your table size increases? The table variable is, from around 30,000 rows, locked into a very bad execution plan unless you use OPTION (RECOMPILE) to provide the Query Analyser with a decent estimation of the size of the table. However, if it has the OPTION (RECOMPILE), then it is smokin’. Well, up to 120,000 rows, at least. It is performing better than a Temporary table, and in a good linear fashion. What about mixed table joins, where you are joining a temporary table to a table variable? You’d probably expect that the query analyzer would throw up its hands and produce a bad execution plan as if it were a table variable. After all, it knows nothing about the statistics in one of the tables so how could it do any better? Well, it behaves as if it were doing a recompile. And an explicit recompile adds no value at all. (we just go up to 45000 rows since we know the bigger picture now)   Now, if you were new to this, you might be tempted to start drawing conclusions. Beware! We’re dealing with a very complex beast: the Query Optimizer. It can come up with surprises What if we change the query very slightly to insert the results into a Table Variable? We change nothing else and just measure the execution time of the statement as before. Suddenly, the table variable isn’t looking so much better, even taking into account the time involved in doing the table insert. OK, if you haven’t used OPTION (RECOMPILE) then you’re toast. Otherwise, there isn’t much in it between the Table variable and the temporary table. The table variable is faster up to 8000 rows and then not much in it up to 100,000 rows. Past the 8000 row mark, we’ve lost the advantage of the table variable’s speed. Any general rule you may be formulating has just gone for a walk. What we can conclude from this experiment is that if you join two table variables, and can’t use constraints, you’re going to need that Option (RECOMPILE) hint. Count Dracula and the Horror Join. These tables of integers provide a rather unreal example, so let’s try a rather different example, and get stuck into some implicit indexing, by using constraints. What unusual words are contained in the book ‘Dracula’ by Bram Stoker? Here we get a table of all the common words in the English language (60,387 of them) and put them in a table. We put them in a Table Variable with the word as a primary key, a Table Variable Heap and a Table Variable with a primary key. We then take all the distinct words used in the book ‘Dracula’ (7,558 of them). We then create a table variable and insert into it all those uncommon words that are in ‘Dracula’. i.e. all the words in Dracula that aren’t matched in the list of common words. To do this we use a left outer join, where the right-hand value is null. The results show a huge variation, between the sublime and the gorblimey. If both tables contain a Primary Key on the columns we join on, and both are Table Variables, it took 33 Ms. If one table contains a Primary Key, and the other is a heap, and both are Table Variables, it took 46 Ms. If both Table Variables use a unique constraint, then the query takes 36 Ms. If neither table contains a Primary Key and both are Table Variables, it took 116383 Ms. Yes, nearly two minutes!! If both tables contain a Primary Key, one is a Table Variables and the other is a temporary table, it took 113 Ms. If one table contains a Primary Key, and both are Temporary Tables, it took 56 Ms.If both tables are temporary tables and both have primary keys, it took 46 Ms. Here we see table variables which are joined on their primary key again enjoying a  slight performance advantage over temporary tables. Where both tables are table variables and both are heaps, the query suddenly takes nearly two minutes! So what if you have two heaps and you use option Recompile? If you take the rogue query and add the hint, then suddenly, the query drops its time down to 76 Ms. If you add unique indexes, then you've done even better, down to half that time. Here are the text execution plans.So where have we got to? Without drilling down into the minutiae of the execution plans we can begin to create a hypothesis. If you are using table variables, and your tables are relatively small, they are faster than temporary tables, but as the number of rows increases you need to do one of two things: either you need to have a primary key on the column you are using to join on, or else you need to use option (RECOMPILE) If you try to execute a query that is a join, and both tables are table variable heaps, you are asking for trouble, well- slow queries, unless you give the table hint once the number of rows has risen past a point (30,000 in our first example, but this varies considerably according to context). Kevin’s Skew In describing the table-size, I used the term ‘relatively small’. Kevin Boles produced an interesting case where a single-row table variable produces a very poor execution plan when joined to a very, very skewed table. In the original, pasted into my article as a comment, a column consisted of 100000 rows in which the key column was one number (1) . To this was added eight rows with sequential numbers up to 9. When this was joined to a single-tow Table Variable with a key of 2 it produced a bad plan. This problem is unlikely to occur in real usage, and the Query Optimiser team probably never set up a test for it. Actually, the skew can be slightly less extreme than Kevin made it. The following test showed that once the table had 54 sequential rows in the table, then it adopted exactly the same execution plan as for the temporary table and then all was well. Undeniably, real data does occasionally cause problems to the performance of joins in Table Variables due to the extreme skew of the distribution. We've all experienced Perfectly Poisonous Table Variables in real live data. As in Kevin’s example, indexes merely make matters worse, and the OPTION (RECOMPILE) trick does nothing to help. In this case, there is no option but to use a temporary table. However, one has to note that once the slight de-skew had taken place, then the plans were identical across a huge range. Conclusions Where you need to hold intermediate results as part of a process, Table Variables offer a good alternative to temporary tables when used wisely. They can perform faster than a temporary table when the number of rows is not great. For some processing with huge tables, they can perform well when only a clustered index is required, and when the nature of the processing makes an index seek very effective. Table Variables are scoped to the batch or procedure and are unlikely to hang about in the TempDB when they are no longer required. They require no explicit cleanup. Where the number of rows in the table is moderate, you can even use them in joins as ‘Heaps’, unindexed. Beware, however, since, as the number of rows increase, joins on Table Variable heaps can easily become saddled by very poor execution plans, and this must be cured either by adding constraints (UNIQUE or PRIMARY KEY) or by adding the OPTION (RECOMPILE) hint if this is impossible. Occasionally, the way that the data is distributed prevents the efficient use of Table Variables, and this will require using a temporary table instead. Tables Variables require some awareness by the developer about the potential hazards and how to avoid them. If you are not prepared to do any performance monitoring of your code or fine-tuning, and just want to pummel out stuff that ‘just runs’ without considering namby-pamby stuff such as indexes, then stick to Temporary tables. If you are likely to slosh about large numbers of rows in temporary tables without considering the niceties of processing just what is required and no more, then temporary tables provide a safer and less fragile means-to-an-end for you.

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  • Everytime i am trying to connect to my box using SSH, its failing not connecting

    - by YumYumYum
    From any other PC doing SSH to my Ubuntu 11.10,is failing. My network setup: Telenet ISP (Belgium) Fiber cable < RJ45 cable straight to Ubuntu PC Even the SSH is running: Other PC: retrying over and over $ ping 192.168.0.128 PING 192.168.0.128 (192.168.0.128) 56(84) bytes of data. From 192.168.0.226 icmp_seq=1 Destination Host Unreachable From 192.168.0.226 icmp_seq=2 Destination Host Unreachable From 192.168.0.226 icmp_seq=3 Destination Host Unreachable From 192.168.0.226 icmp_seq=4 Destination Host Unreachable $ sudo service iptables stop Stopping iptables (via systemctl): [ OK ] $ ssh [email protected] ssh: connect to host 192.168.0.128 port 22: No route to host $ ssh [email protected] ssh: connect to host 192.168.0.128 port 22: No route to host $ ssh [email protected] ssh: connect to host 192.168.0.128 port 22: No route to host $ ssh [email protected] ssh: connect to host 192.168.0.128 port 22: No route to host $ ssh [email protected] Connection closed by 192.168.0.128 $ ssh [email protected] [email protected]'s password: Connection closed by UNKNOWN $ ssh [email protected] ssh: connect to host 192.168.0.128 port 22: No route to host $ ssh [email protected] ssh: connect to host 192.168.0.128 port 22: No route to host Follow up: -- checked cable -- using cable tester and other detectors -- no problem found in cable -- used random 10 cables -- adapter is not broken -- checked it using circuit tester by opening the system (card is new so its not network adapter card problem) -- leds are OK showing -- used LiveCD and did same ping test was having same problem -- disabled ipv6 100% to make sure its not the cause -- disabled iptables 100% so its also not the issue -- some more info $ nmap 192.168.0.128 Starting Nmap 5.50 ( http://nmap.org ) at 2012-06-08 19:11 CEST Nmap scan report for 192.168.0.128 Host is up (0.00045s latency). All 1000 scanned ports on 192.168.0.128 are closed (842) or filtered (158) Nmap done: 1 IP address (1 host up) scanned in 6.86 seconds ubuntu@ubuntu:~$ netstat -aunt | head Active Internet connections (servers and established) Proto Recv-Q Send-Q Local Address Foreign Address State tcp 0 0 127.0.0.1:631 0.0.0.0:* LISTEN tcp 0 1 192.168.0.128:58616 74.125.132.99:80 FIN_WAIT1 tcp 0 0 192.168.0.128:56749 199.7.57.72:80 ESTABLISHED tcp 0 1 192.168.0.128:58614 74.125.132.99:80 FIN_WAIT1 tcp 0 0 192.168.0.128:49916 173.194.65.113:443 ESTABLISHED tcp 0 1 192.168.0.128:45699 64.34.119.101:80 SYN_SENT tcp 0 0 192.168.0.128:48404 64.34.119.12:80 ESTABLISHED tcp 0 0 192.168.0.128:54161 67.201.31.70:80 TIME_WAIT $ sudo killall dnsmasq -- did not solved the problem -- -- like many other Q/A was suggesting this same --- $ iptables --list Chain INPUT (policy ACCEPT) target prot opt source destination Chain FORWARD (policy ACCEPT) target prot opt source destination Chain OUTPUT (policy ACCEPT) target prot opt source destination $ netstat -nr Kernel IP routing table Destination Gateway Genmask Flags MSS Window irtt Iface 0.0.0.0 192.168.0.1 0.0.0.0 UG 0 0 0 eth0 169.254.0.0 0.0.0.0 255.255.0.0 U 0 0 0 eth0 192.168.0.0 0.0.0.0 255.255.255.0 U 0 0 0 eth0 $ ssh -vvv [email protected] OpenSSH_5.6p1, OpenSSL 1.0.0j-fips 10 May 2012 debug1: Reading configuration data /etc/ssh/ssh_config debug1: Applying options for * debug2: ssh_connect: needpriv 0 debug1: Connecting to 192.168.0.128 [192.168.0.128] port 22. debug1: Connection established. debug3: Not a RSA1 key file /home/sun/.ssh/id_rsa. debug2: key_type_from_name: unknown key type '-----BEGIN' debug3: key_read: missing keytype debug3: key_read: missing whitespace debug3: key_read: missing whitespace debug3: key_read: missing whitespace debug3: key_read: missing whitespace debug3: key_read: missing whitespace debug3: key_read: missing whitespace debug3: key_read: missing whitespace debug3: key_read: missing whitespace debug3: key_read: missing whitespace debug3: key_read: missing whitespace debug3: key_read: missing whitespace debug3: key_read: missing whitespace debug3: key_read: missing whitespace debug3: key_read: missing whitespace debug3: key_read: missing whitespace debug3: key_read: missing whitespace debug3: key_read: missing whitespace debug3: key_read: missing whitespace debug3: key_read: missing whitespace debug3: key_read: missing whitespace debug3: key_read: missing whitespace debug3: key_read: missing whitespace debug3: key_read: missing whitespace debug3: key_read: missing whitespace debug3: key_read: missing whitespace debug2: key_type_from_name: unknown key type '-----END' debug3: key_read: missing keytype debug1: identity file /home/sun/.ssh/id_rsa type 1 debug1: identity file /home/sun/.ssh/id_rsa-cert type -1 debug1: identity file /home/sun/.ssh/id_dsa type -1 debug1: identity file /home/sun/.ssh/id_dsa-cert type -1 debug1: Remote protocol version 2.0, remote software version OpenSSH_5.8p1 Debian-7ubuntu1 debug1: match: OpenSSH_5.8p1 Debian-7ubuntu1 pat OpenSSH* debug1: Enabling compatibility mode for protocol 2.0 debug1: Local version string SSH-2.0-OpenSSH_5.6 debug2: fd 3 setting O_NONBLOCK debug1: SSH2_MSG_KEXINIT sent debug1: SSH2_MSG_KEXINIT received debug2: kex_parse_kexinit: diffie-hellman-group-exchange-sha256,diffie-hellman-group-exchange-sha1,diffie-hellman-group14-sha1,diffie-hellman-group1-sha1 debug2: kex_parse_kexinit: [email protected],[email protected],[email protected],[email protected],ssh-rsa,ssh-dss debug2: kex_parse_kexinit: aes128-ctr,aes192-ctr,aes256-ctr,arcfour256,arcfour128,aes128-cbc,3des-cbc,blowfish-cbc,cast128-cbc,aes192-cbc,aes256-cbc,arcfour,[email protected] debug2: kex_parse_kexinit: aes128-ctr,aes192-ctr,aes256-ctr,arcfour256,arcfour128,aes128-cbc,3des-cbc,blowfish-cbc,cast128-cbc,aes192-cbc,aes256-cbc,arcfour,[email protected] debug2: kex_parse_kexinit: hmac-md5,hmac-sha1,[email protected],hmac-ripemd160,[email protected],hmac-sha1-96,hmac-md5-96 debug2: kex_parse_kexinit: hmac-md5,hmac-sha1,[email protected],hmac-ripemd160,[email protected],hmac-sha1-96,hmac-md5-96 debug2: kex_parse_kexinit: none,[email protected],zlib debug2: kex_parse_kexinit: none,[email protected],zlib debug2: kex_parse_kexinit: debug2: kex_parse_kexinit: debug2: kex_parse_kexinit: first_kex_follows 0 debug2: kex_parse_kexinit: reserved 0 debug2: kex_parse_kexinit: ecdh-sha2-nistp256,ecdh-sha2-nistp384,ecdh-sha2-nistp521,diffie-hellman-group-exchange-sha256,diffie-hellman-group-exchange-sha1,diffie-hellman-group14-sha1,diffie-hellman-group1-sha1 debug2: kex_parse_kexinit: ssh-rsa,ssh-dss,ecdsa-sha2-nistp256 debug2: kex_parse_kexinit: aes128-ctr,aes192-ctr,aes256-ctr,arcfour256,arcfour128,aes128-cbc,3des-cbc,blowfish-cbc,cast128-cbc,aes192-cbc,aes256-cbc,arcfour,[email protected] debug2: kex_parse_kexinit: aes128-ctr,aes192-ctr,aes256-ctr,arcfour256,arcfour128,aes128-cbc,3des-cbc,blowfish-cbc,cast128-cbc,aes192-cbc,aes256-cbc,arcfour,[email protected] debug2: kex_parse_kexinit: hmac-md5,hmac-sha1,[email protected],hmac-ripemd160,[email protected],hmac-sha1-96,hmac-md5-96 debug2: kex_parse_kexinit: hmac-md5,hmac-sha1,[email protected],hmac-ripemd160,[email protected],hmac-sha1-96,hmac-md5-96 debug2: kex_parse_kexinit: none,[email protected] debug2: kex_parse_kexinit: none,[email protected] debug2: kex_parse_kexinit: debug2: kex_parse_kexinit: debug2: kex_parse_kexinit: first_kex_follows 0 debug2: kex_parse_kexinit: reserved 0 debug2: mac_setup: found hmac-md5 debug1: kex: server->client aes128-ctr hmac-md5 none debug2: mac_setup: found hmac-md5 debug1: kex: client->server aes128-ctr hmac-md5 none debug1: SSH2_MSG_KEX_DH_GEX_REQUEST(1024<1024<8192) sent debug1: expecting SSH2_MSG_KEX_DH_GEX_GROUP debug2: dh_gen_key: priv key bits set: 118/256 debug2: bits set: 539/1024 debug1: SSH2_MSG_KEX_DH_GEX_INIT sent debug1: expecting SSH2_MSG_KEX_DH_GEX_REPLY debug3: check_host_in_hostfile: host 192.168.0.128 filename /home/sun/.ssh/known_hosts debug3: check_host_in_hostfile: host 192.168.0.128 filename /home/sun/.ssh/known_hosts debug3: check_host_in_hostfile: match line 139 debug1: Host '192.168.0.128' is known and matches the RSA host key. debug1: Found key in /home/sun/.ssh/known_hosts:139 debug2: bits set: 544/1024 debug1: ssh_rsa_verify: signature correct debug2: kex_derive_keys debug2: set_newkeys: mode 1 debug1: SSH2_MSG_NEWKEYS sent debug1: expecting SSH2_MSG_NEWKEYS debug2: set_newkeys: mode 0 debug1: SSH2_MSG_NEWKEYS received debug1: Roaming not allowed by server debug1: SSH2_MSG_SERVICE_REQUEST sent debug2: service_accept: ssh-userauth debug1: SSH2_MSG_SERVICE_ACCEPT received debug2: key: /home/sun/.ssh/id_rsa (0x213db960) debug2: key: /home/sun/.ssh/id_dsa ((nil)) debug1: Authentications that can continue: publickey,password debug3: start over, passed a different list publickey,password debug3: preferred gssapi-keyex,gssapi-with-mic,publickey,keyboard-interactive,password debug3: authmethod_lookup publickey debug3: remaining preferred: keyboard-interactive,password debug3: authmethod_is_enabled publickey debug1: Next authentication method: publickey debug1: Offering RSA public key: /home/sun/.ssh/id_rsa debug3: send_pubkey_test debug2: we sent a publickey packet, wait for reply debug1: Authentications that can continue: publickey,password debug1: Trying private key: /home/sun/.ssh/id_dsa debug3: no such identity: /home/sun/.ssh/id_dsa debug2: we did not send a packet, disable method debug3: authmethod_lookup password debug3: remaining preferred: ,password debug3: authmethod_is_enabled password debug1: Next authentication method: password [email protected]'s password: debug3: packet_send2: adding 64 (len 60 padlen 4 extra_pad 64) debug2: we sent a password packet, wait for reply debug1: Authentication succeeded (password). Authenticated to 192.168.0.128 ([192.168.0.128]:22). debug1: channel 0: new [client-session] debug3: ssh_session2_open: channel_new: 0 debug2: channel 0: send open debug1: Requesting [email protected] debug1: Entering interactive session. debug2: callback start debug2: client_session2_setup: id 0 debug2: channel 0: request pty-req confirm 1 debug1: Sending environment. debug3: Ignored env ORBIT_SOCKETDIR debug3: Ignored env XDG_SESSION_ID debug3: Ignored env HOSTNAME debug3: Ignored env GIO_LAUNCHED_DESKTOP_FILE_PID debug3: Ignored env IMSETTINGS_INTEGRATE_DESKTOP debug3: Ignored env GPG_AGENT_INFO debug3: Ignored env TERM debug3: Ignored env HARDWARE_PLATFORM debug3: Ignored env SHELL debug3: Ignored env DESKTOP_STARTUP_ID debug3: Ignored env HISTSIZE debug3: Ignored env XDG_SESSION_COOKIE debug3: Ignored env GJS_DEBUG_OUTPUT debug3: Ignored env WINDOWID debug3: Ignored env GNOME_KEYRING_CONTROL debug3: Ignored env QTDIR debug3: Ignored env QTINC debug3: Ignored env GJS_DEBUG_TOPICS debug3: Ignored env IMSETTINGS_MODULE debug3: Ignored env USER debug3: Ignored env LS_COLORS debug3: Ignored env SSH_AUTH_SOCK debug3: Ignored env USERNAME debug3: Ignored env SESSION_MANAGER debug3: Ignored env GIO_LAUNCHED_DESKTOP_FILE debug3: Ignored env PATH debug3: Ignored env MAIL debug3: Ignored env DESKTOP_SESSION debug3: Ignored env QT_IM_MODULE debug3: Ignored env PWD debug1: Sending env XMODIFIERS = @im=none debug2: channel 0: request env confirm 0 debug1: Sending env LANG = en_US.utf8 debug2: channel 0: request env confirm 0 debug3: Ignored env KDE_IS_PRELINKED debug3: Ignored env GDM_LANG debug3: Ignored env KDEDIRS debug3: Ignored env GDMSESSION debug3: Ignored env SSH_ASKPASS debug3: Ignored env HISTCONTROL debug3: Ignored env HOME debug3: Ignored env SHLVL debug3: Ignored env GDL_PATH debug3: Ignored env GNOME_DESKTOP_SESSION_ID debug3: Ignored env LOGNAME debug3: Ignored env QTLIB debug3: Ignored env CVS_RSH debug3: Ignored env DBUS_SESSION_BUS_ADDRESS debug3: Ignored env LESSOPEN debug3: Ignored env WINDOWPATH debug3: Ignored env XDG_RUNTIME_DIR debug3: Ignored env DISPLAY debug3: Ignored env G_BROKEN_FILENAMES debug3: Ignored env COLORTERM debug3: Ignored env XAUTHORITY debug3: Ignored env _ debug2: channel 0: request shell confirm 1 debug2: fd 3 setting TCP_NODELAY debug2: callback done debug2: channel 0: open confirm rwindow 0 rmax 32768 debug2: channel_input_status_confirm: type 99 id 0 debug2: PTY allocation request accepted on channel 0 debug2: channel 0: rcvd adjust 2097152 debug2: channel_input_status_confirm: type 99 id 0 debug2: shell request accepted on channel 0 Welcome to Ubuntu 11.10 (GNU/Linux 3.0.0-12-generic x86_64) * Documentation: https://help.ubuntu.com/ 297 packages can be updated. 92 updates are security updates. New release '12.04 LTS' available. Run 'do-release-upgrade' to upgrade to it. Last login: Fri Jun 8 07:45:15 2012 from 192.168.0.226 sun@SystemAX51:~$ ping 19<--------Lost connection again-------------- Tail follow: -- dmesg is showing a very abnormal logs, like Ubuntu is automatically bringing the eth0 up, where eth0 is getting also auto down. [ 2025.897511] r8169 0000:02:00.0: eth0: link up [ 2029.347649] r8169 0000:02:00.0: eth0: link up [ 2030.775556] r8169 0000:02:00.0: eth0: link up [ 2038.242203] r8169 0000:02:00.0: eth0: link up [ 2057.267801] r8169 0000:02:00.0: eth0: link up [ 2062.871770] r8169 0000:02:00.0: eth0: link up [ 2082.479712] r8169 0000:02:00.0: eth0: link up [ 2285.630797] r8169 0000:02:00.0: eth0: link up [ 2308.417640] r8169 0000:02:00.0: eth0: link up [ 2480.948290] r8169 0000:02:00.0: eth0: link up [ 2824.884798] r8169 0000:02:00.0: eth0: link up [ 3030.022183] r8169 0000:02:00.0: eth0: link up [ 3306.587353] r8169 0000:02:00.0: eth0: link up [ 3523.566881] r8169 0000:02:00.0: eth0: link up [ 3619.839585] r8169 0000:02:00.0: eth0: link up [ 3682.154393] nf_conntrack version 0.5.0 (16384 buckets, 65536 max) [ 3899.866854] r8169 0000:02:00.0: eth0: link up [ 4723.978269] r8169 0000:02:00.0: eth0: link up [ 4807.415682] r8169 0000:02:00.0: eth0: link up [ 5101.865686] r8169 0000:02:00.0: eth0: link up How do i fix it? -- http://ubuntuforums.org/showthread.php?t=1959794 $ apt-get install openipml openhpi-plugin-ipml $ openipmish > help redisp_cmd on|off > redisp_cmd on redisp set Final follow up: Step 1: BUG for network card driver r8169 Step 2: get the latest build version http://www.realtek.com/downloads/downloadsView.aspx?Langid=1&PNid=4&PFid=4&Level=5&Conn=4&DownTypeID=3&GetDown=false&Downloads=true#RTL8110SC(L) Step 3: build / make $ cd /var/tmp/driver $ tar xvfj r8169.tar.bz2 $ make clean modules && make install $ rmmod r8169 $ depmod $ cp src/r8169.ko /lib/modules/3.xxxx/kernel/drivers/net/r8169.ko $ modprobe r8169 $ update-initramfs -u $ init 6 Voila!!

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  • How do I prevent missing network from slowing down boot-up?

    - by Ravi S Ghosh
    I have been having rather slow boot on Ubuntu 12.04. Lately, I tried to figure out the reason and it seems to be the network connection which does not get connected and requires multiple attempts. Here is part of dmesg [ 2.174349] EXT4-fs (sda2): INFO: recovery required on readonly filesystem [ 2.174352] EXT4-fs (sda2): write access will be enabled during recovery [ 2.308172] firewire_core: created device fw0: GUID 384fc00005198d58, S400 [ 2.333457] usb 7-1.2: new low-speed USB device number 3 using uhci_hcd [ 2.465896] EXT4-fs (sda2): recovery complete [ 2.466406] EXT4-fs (sda2): mounted filesystem with ordered data mode. Opts: (null) [ 2.589440] usb 7-1.3: new low-speed USB device number 4 using uhci_hcd **[ 18.292029] ADDRCONF(NETDEV_UP): eth0: link is not ready** [ 18.458958] udevd[377]: starting version 175 [ 18.639482] Adding 4200960k swap on /dev/sda5. Priority:-1 extents:1 across:4200960k [ 19.314127] wmi: Mapper loaded [ 19.426602] r592 0000:09:01.2: PCI INT B -> GSI 18 (level, low) -> IRQ 18 [ 19.426739] r592: driver successfully loaded [ 19.460105] input: Dell WMI hotkeys as /devices/virtual/input/input5 [ 19.493629] lp: driver loaded but no devices found [ 19.497012] cfg80211: Calling CRDA to update world regulatory domain [ 19.535523] ACPI Warning: _BQC returned an invalid level (20110623/video-480) [ 19.539457] acpi device:03: registered as cooling_device2 [ 19.539520] input: Video Bus as /devices/LNXSYSTM:00/device:00/PNP0A08:00/device:01/LNXVIDEO:00/input/input6 [ 19.539568] ACPI: Video Device [M86] (multi-head: yes rom: no post: no) [ 19.578060] Linux video capture interface: v2.00 [ 19.667708] dcdbas dcdbas: Dell Systems Management Base Driver (version 5.6.0-3.2) [ 19.763171] r852 0000:09:01.3: PCI INT B -> GSI 18 (level, low) -> IRQ 18 [ 19.763258] r852: driver loaded successfully [ 19.854769] input: Microsoft Comfort Curve Keyboard 2000 as /devices/pci0000:00/0000:00:1d.1/usb7/7-1/7-1.2/7-1.2:1.0/input/input7 [ 19.854864] generic-usb 0003:045E:00DD.0001: input,hidraw0: USB HID v1.11 Keyboard [Microsoft Comfort Curve Keyboard 2000] on usb-0000:00:1d.1-1.2/input0 [ 19.878605] input: Microsoft Comfort Curve Keyboard 2000 as /devices/pci0000:00/0000:00:1d.1/usb7/7-1/7-1.2/7-1.2:1.1/input/input8 [ 19.878698] generic-usb 0003:045E:00DD.0002: input,hidraw1: USB HID v1.11 Device [Microsoft Comfort Curve Keyboard 2000] on usb-0000:00:1d.1-1.2/input1 [ 19.902779] input: DELL DELL USB Laser Mouse as /devices/pci0000:00/0000:00:1d.1/usb7/7-1/7-1.3/7-1.3:1.0/input/input9 [ 19.925034] generic-usb 0003:046D:C063.0003: input,hidraw2: USB HID v1.10 Mouse [DELL DELL USB Laser Mouse] on usb-0000:00:1d.1-1.3/input0 [ 19.925057] usbcore: registered new interface driver usbhid [ 19.925059] usbhid: USB HID core driver [ 19.942362] uvcvideo: Found UVC 1.00 device Laptop_Integrated_Webcam_2M (0c45:63ea) [ 19.947004] input: Laptop_Integrated_Webcam_2M as /devices/pci0000:00/0000:00:1a.7/usb1/1-6/1-6:1.0/input/input10 [ 19.947075] usbcore: registered new interface driver uvcvideo [ 19.947077] USB Video Class driver (1.1.1) [ 20.145232] Intel(R) Wireless WiFi Link AGN driver for Linux, in-tree: [ 20.145235] Copyright(c) 2003-2011 Intel Corporation [ 20.145327] iwlwifi 0000:04:00.0: PCI INT A -> GSI 17 (level, low) -> IRQ 17 [ 20.145357] iwlwifi 0000:04:00.0: setting latency timer to 64 [ 20.145402] iwlwifi 0000:04:00.0: pci_resource_len = 0x00002000 [ 20.145404] iwlwifi 0000:04:00.0: pci_resource_base = ffffc90000674000 [ 20.145407] iwlwifi 0000:04:00.0: HW Revision ID = 0x0 [ 20.145531] iwlwifi 0000:04:00.0: irq 46 for MSI/MSI-X [ 20.145613] iwlwifi 0000:04:00.0: Detected Intel(R) WiFi Link 5100 AGN, REV=0x54 [ 20.145720] iwlwifi 0000:04:00.0: L1 Enabled; Disabling L0S [ 20.167535] iwlwifi 0000:04:00.0: device EEPROM VER=0x11f, CALIB=0x4 [ 20.167538] iwlwifi 0000:04:00.0: Device SKU: 0Xf0 [ 20.167567] iwlwifi 0000:04:00.0: Tunable channels: 13 802.11bg, 24 802.11a channels [ 20.172779] fglrx: module license 'Proprietary. (C) 2002 - ATI Technologies, Starnberg, GERMANY' taints kernel. [ 20.172783] Disabling lock debugging due to kernel taint [ 20.250115] [fglrx] Maximum main memory to use for locked dma buffers: 3759 MBytes. [ 20.250567] [fglrx] vendor: 1002 device: 9553 count: 1 [ 20.251256] [fglrx] ioport: bar 1, base 0x2000, size: 0x100 [ 20.251271] pci 0000:01:00.0: PCI INT A -> GSI 16 (level, low) -> IRQ 16 [ 20.251277] pci 0000:01:00.0: setting latency timer to 64 [ 20.251559] [fglrx] Kernel PAT support is enabled [ 20.251578] [fglrx] module loaded - fglrx 8.96.4 [Mar 12 2012] with 1 minors [ 20.310385] iwlwifi 0000:04:00.0: loaded firmware version 8.83.5.1 build 33692 [ 20.310598] Registered led device: phy0-led [ 20.310628] cfg80211: Ignoring regulatory request Set by core since the driver uses its own custom regulatory domain [ 20.372306] ieee80211 phy0: Selected rate control algorithm 'iwl-agn-rs' [ 20.411015] psmouse serio1: synaptics: Touchpad model: 1, fw: 7.2, id: 0x1c0b1, caps: 0xd04733/0xa40000/0xa0000 [ 20.454232] input: SynPS/2 Synaptics TouchPad as /devices/platform/i8042/serio1/input/input11 [ 20.545636] cfg80211: Ignoring regulatory request Set by core since the driver uses its own custom regulatory domain [ 20.545640] cfg80211: World regulatory domain updated: [ 20.545642] cfg80211: (start_freq - end_freq @ bandwidth), (max_antenna_gain, max_eirp) [ 20.545644] cfg80211: (2402000 KHz - 2472000 KHz @ 40000 KHz), (300 mBi, 2000 mBm) [ 20.545647] cfg80211: (2457000 KHz - 2482000 KHz @ 20000 KHz), (300 mBi, 2000 mBm) [ 20.545649] cfg80211: (2474000 KHz - 2494000 KHz @ 20000 KHz), (300 mBi, 2000 mBm) [ 20.545652] cfg80211: (5170000 KHz - 5250000 KHz @ 40000 KHz), (300 mBi, 2000 mBm) [ 20.545654] cfg80211: (5735000 KHz - 5835000 KHz @ 40000 KHz), (300 mBi, 2000 mBm) [ 20.609484] type=1400 audit(1340502633.160:2): apparmor="STATUS" operation="profile_load" name="/sbin/dhclient" pid=693 comm="apparmor_parser" [ 20.609494] type=1400 audit(1340502633.160:3): apparmor="STATUS" operation="profile_replace" name="/sbin/dhclient" pid=642 comm="apparmor_parser" [ 20.609843] type=1400 audit(1340502633.160:4): apparmor="STATUS" operation="profile_load" name="/usr/lib/NetworkManager/nm-dhcp-client.action" pid=693 comm="apparmor_parser" [ 20.609852] type=1400 audit(1340502633.160:5): apparmor="STATUS" operation="profile_replace" name="/usr/lib/NetworkManager/nm-dhcp-client.action" pid=642 comm="apparmor_parser" [ 20.610047] type=1400 audit(1340502633.160:6): apparmor="STATUS" operation="profile_load" name="/usr/lib/connman/scripts/dhclient-script" pid=693 comm="apparmor_parser" [ 20.610060] type=1400 audit(1340502633.160:7): apparmor="STATUS" operation="profile_replace" name="/usr/lib/connman/scripts/dhclient-script" pid=642 comm="apparmor_parser" [ 20.610476] type=1400 audit(1340502633.160:8): apparmor="STATUS" operation="profile_replace" name="/sbin/dhclient" pid=814 comm="apparmor_parser" [ 20.610829] type=1400 audit(1340502633.160:9): apparmor="STATUS" operation="profile_replace" name="/usr/lib/NetworkManager/nm-dhcp-client.action" pid=814 comm="apparmor_parser" [ 20.611035] type=1400 audit(1340502633.160:10): apparmor="STATUS" operation="profile_replace" name="/usr/lib/connman/scripts/dhclient-script" pid=814 comm="apparmor_parser" [ 20.661912] snd_hda_intel 0000:00:1b.0: PCI INT A -> GSI 22 (level, low) -> IRQ 22 [ 20.661982] snd_hda_intel 0000:00:1b.0: irq 47 for MSI/MSI-X [ 20.662013] snd_hda_intel 0000:00:1b.0: setting latency timer to 64 [ 20.770289] input: HDA Intel Mic as /devices/pci0000:00/0000:00:1b.0/sound/card0/input12 [ 20.770689] snd_hda_intel 0000:01:00.1: PCI INT B -> GSI 17 (level, low) -> IRQ 17 [ 20.770786] snd_hda_intel 0000:01:00.1: irq 48 for MSI/MSI-X [ 20.770815] snd_hda_intel 0000:01:00.1: setting latency timer to 64 [ 20.994040] HDMI status: Codec=0 Pin=3 Presence_Detect=0 ELD_Valid=0 [ 20.994189] input: HDA ATI HDMI HDMI/DP,pcm=3 as /devices/pci0000:00/0000:00:01.0/0000:01:00.1/sound/card1/input13 [ 21.554799] vesafb: mode is 1024x768x32, linelength=4096, pages=0 [ 21.554802] vesafb: scrolling: redraw [ 21.554804] vesafb: Truecolor: size=0:8:8:8, shift=0:16:8:0 [ 21.557342] vesafb: framebuffer at 0xd0000000, mapped to 0xffffc90011800000, using 3072k, total 3072k [ 21.557498] Console: switching to colour frame buffer device 128x48 [ 21.557516] fb0: VESA VGA frame buffer device [ 21.987338] EXT4-fs (sda2): re-mounted. Opts: errors=remount-ro [ 22.184693] EXT4-fs (sda6): mounted filesystem with ordered data mode. Opts: (null) [ 27.362440] iwlwifi 0000:04:00.0: RF_KILL bit toggled to disable radio. [ 27.436988] init: failsafe main process (986) killed by TERM signal [ 27.970112] ppdev: user-space parallel port driver [ 28.198917] Bluetooth: Core ver 2.16 [ 28.198935] NET: Registered protocol family 31 [ 28.198937] Bluetooth: HCI device and connection manager initialized [ 28.198940] Bluetooth: HCI socket layer initialized [ 28.198941] Bluetooth: L2CAP socket layer initialized [ 28.198947] Bluetooth: SCO socket layer initialized [ 28.226135] Bluetooth: RFCOMM TTY layer initialized [ 28.226141] Bluetooth: RFCOMM socket layer initialized [ 28.226143] Bluetooth: RFCOMM ver 1.11 [ 28.445620] Bluetooth: BNEP (Ethernet Emulation) ver 1.3 [ 28.445623] Bluetooth: BNEP filters: protocol multicast [ 28.524578] type=1400 audit(1340502641.076:11): apparmor="STATUS" operation="profile_load" name="/usr/lib/cups/backend/cups-pdf" pid=1052 comm="apparmor_parser" [ 28.525018] type=1400 audit(1340502641.076:12): apparmor="STATUS" operation="profile_load" name="/usr/sbin/cupsd" pid=1052 comm="apparmor_parser" [ 28.629957] type=1400 audit(1340502641.180:13): apparmor="STATUS" operation="profile_replace" name="/sbin/dhclient" pid=1105 comm="apparmor_parser" [ 28.630325] type=1400 audit(1340502641.180:14): apparmor="STATUS" operation="profile_replace" name="/usr/lib/NetworkManager/nm-dhcp-client.action" pid=1105 comm="apparmor_parser" [ 28.630535] type=1400 audit(1340502641.180:15): apparmor="STATUS" operation="profile_replace" name="/usr/lib/connman/scripts/dhclient-script" pid=1105 comm="apparmor_parser" [ 28.645266] type=1400 audit(1340502641.196:16): apparmor="STATUS" operation="profile_load" name="/usr/lib/lightdm/lightdm/lightdm-guest-session-wrapper" pid=1104 comm="apparmor_parser" **[ 28.751922] ADDRCONF(NETDEV_UP): wlan0: link is not ready** [ 28.753653] tg3 0000:08:00.0: irq 49 for MSI/MSI-X **[ 28.856127] ADDRCONF(NETDEV_UP): eth0: link is not ready [ 28.857034] ADDRCONF(NETDEV_UP): eth0: link is not ready** [ 28.871080] type=1400 audit(1340502641.420:17): apparmor="STATUS" operation="profile_load" name="/usr/lib/telepathy/mission-control-5" pid=1108 comm="apparmor_parser" [ 28.871519] type=1400 audit(1340502641.420:18): apparmor="STATUS" operation="profile_load" name="/usr/lib/telepathy/telepathy-*" pid=1108 comm="apparmor_parser" [ 28.874905] type=1400 audit(1340502641.424:19): apparmor="STATUS" operation="profile_replace" name="/usr/lib/cups/backend/cups-pdf" pid=1113 comm="apparmor_parser" [ 28.875354] type=1400 audit(1340502641.424:20): apparmor="STATUS" operation="profile_replace" name="/usr/sbin/cupsd" pid=1113 comm="apparmor_parser" [ 30.477976] tg3 0000:08:00.0: eth0: Link is up at 100 Mbps, full duplex [ 30.477979] tg3 0000:08:00.0: eth0: Flow control is on for TX and on for RX **[ 30.478390] ADDRCONF(NETDEV_CHANGE): eth0: link becomes ready** [ 31.110269] fglrx_pci 0000:01:00.0: irq 50 for MSI/MSI-X [ 31.110859] [fglrx] Firegl kernel thread PID: 1327 [ 31.111021] [fglrx] Firegl kernel thread PID: 1329 [ 31.111408] [fglrx] Firegl kernel thread PID: 1330 [ 31.111543] [fglrx] IRQ 50 Enabled [ 31.712938] [fglrx] Gart USWC size:1224 M. [ 31.712941] [fglrx] Gart cacheable size:486 M. [ 31.712945] [fglrx] Reserved FB block: Shared offset:0, size:1000000 [ 31.712948] [fglrx] Reserved FB block: Unshared offset:fc2b000, size:3d5000 [ 31.712950] [fglrx] Reserved FB block: Unshared offset:1fffb000, size:5000 [ 41.312020] eth0: no IPv6 routers present As you can see I get multiple instances of [ 28.856127] ADDRCONF(NETDEV_UP): eth0: link is not ready and then finally it becomes read and I get the message [ 30.478390] ADDRCONF(NETDEV_CHANGE): eth0: link becomes ready. I searched askubuntun, ubuntuforum, and the web but couldn't find a solution. Any help would be very much appreciated. Here is the bootchart

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  • Java JRE 1.7.0_60 Certified with Oracle E-Business Suite

    - by Steven Chan (Oracle Development)
    Java Runtime Environment 7u60 (a.k.a. JRE 7u60-b19) and later updates on the JRE 7 codeline are now certified with Oracle E-Business Suite Release 11i and 12.0, 12.1, and 12.2 for Windows-based desktop clients. Effects of new support dates on Java upgrades for EBS environments Support dates for the E-Business Suite and Java have changed.  Please review the sections below for more details: What does this mean for Oracle E-Business Suite users? Will EBS users be forced to upgrade to JRE 7 for Windows desktop clients? Will EBS users be forced to upgrade to JDK 7 for EBS application tier servers? All JRE 6 and 7 releases are certified with EBS upon release Our standard policy is that all E-Business Suite customers can apply all JRE updates to end-user desktops from JRE 1.6.0_03 and later updates on the 1.6 codeline, and from JRE 7u10 and later updates on the JRE 7 codeline.  We test all new JRE 1.6 and JRE 7 releases in parallel with the JRE development process, so all new JRE 1.6 and 7 releases are considered certified with the E-Business Suite on the same day that they're released by our Java team.  You do not need to wait for a certification announcement before applying new JRE 1.6 or JRE 7 releases to your EBS users' desktops. What's new in JRE 1.7.0_60? JDK 7u60 contains IANA time zone data version 2014b. For more information, refer to Timezone Data Versions in the JRE Software. It is strongly recommended that all customers upgrade to this release.  Details about update in this release are listed in the release notes. 32-bit and 64-bit versions certified This certification includes both the 32-bit and 64-bit JRE versions for various Windows operating systems. See the respective Recommended Browser documentation for your EBS release for details. Where are the official patch requirements documented? All patches required for ensuring full compatibility of the E-Business Suite with JRE 7 are documented in these Notes: For EBS 11i: Deploying Sun JRE (Native Plug-in) for Windows Clients in Oracle E-Business Suite Release 11i (Note 290807.1) Upgrading Developer 6i with Oracle E-Business Suite 11i (Note 125767.1) For EBS 12.0, 12.1, 12.2 Deploying Sun JRE (Native Plug-in) for Windows Clients in Oracle E-Business Suite Release 12 (Note 393931.1) Upgrading OracleAS 10g Forms and Reports in Oracle E-Business Suite Release 12 (Note 437878.1) EBS + Discoverer 11g Users JRE 1.7.0_60 is certified for Discoverer 11g in E-Business Suite environments with the following minimum requirements: Discoverer (11g) 11.1.1.6 plus Patch 13877486 and later  Reference: How To Find Oracle BI Discoverer 10g and 11g Certification Information (Document 233047.1) Worried about the 'mismanaged session cookie' issue? No need to worry -- it's fixed.  To recap: JRE releases 1.6.0_18 through 1.6.0_22 had issues with mismanaging session cookies that affected some users in some circumstances. The fix for those issues was first included in JRE 1.6.0_23. These fixes will carry forward and continue to be fixed in all future JRE releases on the JRE 6 and 7 codelines.  In other words, if you wish to avoid the mismanaged session cookie issue, you should apply any release after JRE 1.6.0_22 on the JRE 6 codeline, and JRE 7u10 and later JRE 7 codeline updates. Implications of Java 6 End of Public Updates for EBS Users The Support Roadmap for Oracle Java is published here: Oracle Java SE Support Roadmap The latest updates to that page (as of Sept. 19, 2012) state (emphasis added): Java SE 6 End of Public Updates Notice After February 2013, Oracle will no longer post updates of Java SE 6 to its public download sites. Existing Java SE 6 downloads already posted as of February 2013 will remain accessible in the Java Archive on Oracle Technology Network. Developers and end-users are encouraged to update to more recent Java SE versions that remain available for public download. For enterprise customers, who need continued access to critical bug fixes and security fixes as well as general maintenance for Java SE 6 or older versions, long term support is available through Oracle Java SE Support . What does this mean for Oracle E-Business Suite users? EBS users fall under the category of "enterprise users" above.  Java is an integral part of the Oracle E-Business Suite technology stack, so EBS users will continue to receive Java SE 6 updates from February 2013 to the end of Java SE 6 Extended Support in June 2017. In other words, nothing changes for EBS users after February 2013.  EBS users will continue to receive critical bug fixes and security fixes as well as general maintenance for Java SE 6 until the end of Java SE 6 Extended Support in June 2017. How can EBS customers obtain Java 6 updates after the public end-of-life? EBS customers can download Java 6 patches from My Oracle Support.  For a complete list of all Java SE patch numbers, see: All Java SE Downloads on MOS (Note 1439822.1) Both JDK and JRE packages are contained in a single combined download after 6u45.  Download the "JDK" package for both the desktop client JRE and the server-side JDK package.  Will EBS users be forced to upgrade to JRE 7 for Windows desktop clients? This upgrade is highly recommended but remains optional while Java 6 is covered by Extended Support. Updates will be delivered via My Oracle Support, where you can continue to receive critical bug fixes and security fixes as well as general maintenance for JRE 6 desktop clients.  Java 6 is covered by Extended Support until June 2017.  All E-Business Suite customers must upgrade to JRE 7 by June 2017. Coexistence of JRE 6 and JRE 7 on Windows desktops The upgrade to JRE 7 is highly recommended for EBS users, but some users may need to run both JRE 6 and 7 on their Windows desktops for reasons unrelated to the E-Business Suite. Most EBS configurations with IE and Firefox use non-static versioning by default. JRE 7 will be invoked instead of JRE 6 if both are installed on a Windows desktop. For more details, see "Appendix B: Static vs. Non-static Versioning and Set Up Options" in Notes 290807.1 and 393931.1. Applying Updates to JRE 6 and JRE 7 to Windows desktops Auto-update will keep JRE 7 up-to-date for Windows users with JRE 7 installed. Auto-update will only keep JRE 7 up-to-date for Windows users with both JRE 6 and 7 installed.  JRE 6 users are strongly encouraged to apply the latest Critical Patch Updates as soon as possible after each release. The Jave SE CPUs will be available via My Oracle Support.  EBS users can find more information about JRE 6 and 7 updates here: Information Center: Installation & Configuration for Oracle Java SE (Note 1412103.2) The dates for future Java SE CPUs can be found on the Critical Patch Updates, Security Alerts and Third Party Bulletin.  An RSS feed is available on that site for those who would like to be kept up-to-date. What do Mac users need? Mac users running Mac OS X 10.9 can run JRE 7 plug-ins.  See this article: EBS Release 12 Certified with Mac OS X 10.9 with Safari 7 and JRE 7 Will EBS users be forced to upgrade to JDK 7 for EBS application tier servers? JRE is used for desktop clients.  JDK is used for application tier servers JDK upgrades for E-Business Suite application tier servers are highly recommended but currently remain optional while Java 6 is covered by Extended Support. Updates will be delivered via My Oracle Support, where you can continue to receive critical bug fixes and security fixes as well as general maintenance for JDK 6 for application tier servers.  Java SE 6 is covered by Extended Support until June 2017.  All EBS customers with application tier servers on Windows, Solaris, and Linux must upgrade to JDK 7 by June 2017. EBS customers running their application tier servers on other operating systems should check with their respective vendors for the support dates for those platforms. JDK 7 is certified with E-Business Suite 12.  See: Java (JDK) 7 Certified for E-Business Suite 12.0 and 12.1 Servers Java (JDK) 7 Certified with E-Business Suite 12.2 Servers References Recommended Browsers for Oracle Applications 11i (Metalink Note 285218.1) Upgrading Sun JRE (Native Plug-in) with Oracle Applications 11i for Windows Clients (Metalink Note 290807.1) Recommended Browsers for Oracle Applications 12 (MetaLink Note 389422.1) Upgrading JRE Plugin with Oracle Applications R12 (MetaLink Note 393931.1) Related Articles Mismanaged Session Cookie Issue Fixed for EBS in JRE 1.6.0_23 Roundup: Oracle JInitiator 1.3 Desupported for EBS Customers in July 2009

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  • CodePlex Daily Summary for Saturday, November 02, 2013

    CodePlex Daily Summary for Saturday, November 02, 2013Popular ReleasesWsus Package Publisher: Release v1.3.1311.02: Add three new Actions in Custom Updates : Work with Files (Copy, Delete, Rename), Work with Folders (Add, Delete, Rename) and Work with Registry Keys (Add, Delete, Rename). Fix a bug, where after resigning an update, the display is not refresh. Modify the way WPP sort rows in 'Updates Detail Viewer' and 'COmputer List Viewer' so that dates are correctly sorted. Add a Tab in the settings form to set Proxy settings when WPP needs to go on Internet. Fix a bug where 'Manage Catalogs Subsc...uComponents: uComponents v6.0.0: This release of uComponents will compile against and support the new API in Umbraco v6.1.0. What's new in uComponents v6.0.0? New DataTypesImage Point XML DropDownList XPath Templatable List New features / Resolved issuesThe following workitems have been implemented and/or resolved: 14781 14805 14808 14818 14854 14827 14868 14859 14790 14853 14790 DataType Grid 14788 14810 14873 14833 14864 14855 / 14860 14816 14823 Drag & Drop support for rows Su...SmartStore.NET - Free ASP.NET MVC Ecommerce Shopping Cart Solution: SmartStore.NET 1.2.1: New FeaturesAdded option Limit to current basket subtotal to HadSpentAmount discount rule Items in product lists can be labelled as NEW for a configurable period of time Product templates can optionally display a discount sign when discounts were applied Added the ability to set multiple favicons depending on stores and/or themes Plugin management: multiple plugins can now be (un)installed in one go Added a field for the HTML body id to store entity (Developer) New property 'Extra...CodeGen Code Generator: CodeGen 4.3.2: Changes in this release include: Removed old tag tokens from several example templates. Fixed a bug which was causing the default author and company names not to be picked up from the registry under .NET. Added several additional tag loop expressions: <IF FIRST_TAG>, <IF LAST_TAG>, <IF MULTIPLE_TAGS> and<IF SINGLE_TAG>. Upgraded to Synergy/DE 10.1.1b, Visual Studio 2013 and Installshield Limited Edition 2013.Community Forums NNTP bridge: Community Forums NNTP Bridge V54 (LiveConnect): This is the first release which can be used with the new LiveConnect authentication. Fixes the problem that the authentication will not work after 1 hour. Also a logfile will now be stored in "%AppData%\Community\CommunityForumsNNTPServer". If you have any problems please feel free to sent me the file "LogFile.txt".Aricie - Friendlier Url Provider: Aricie - Friendlier Url Provider Version 2.5.3: This is mainly a maintenance release to stabilize the new Url Group paradigm. As usual, don't forget to install the Aricie - Shared extension first Highlights Fixed: UI bugs Min Requirements: .Net 3.5+ DotNetNuke 4.8.1+ Aricie - Shared 1.7.7+Aricie Shared: Aricie.Shared Version 1.7.7: This is mainly a maintenance version. Fixes in Property Editor: list import/export Min Requirements: DotNetNuke 4.8.1+ .Net 3.5+WPF Extended DataGrid: WPF Extended DataGrid 2.0.0.9 binaries: Fixed issue with ICollectionView containg null values (AutoFilter issue)Community TFS Build Extensions: October 2013: The October 2013 release contains Scripts - a new addition to our delivery. These are a growing library of PowerShell scripts to use with VS2013. See our documentation for more on scripting. VS2010 Activities(target .NET 4.0) VS2012 Activities (target .NET 4.5) VS2013 Activities (target .NET 4.5.1) Community TFS Build Manager VS2012 Community TFS Build Manager VS2013 The Community TFS Build Managers for VS2010, 2012 and 2013 can also be found in the Visual Studio Gallery where upda...SuperSocket, an extensible socket server framework: SuperSocket 1.6 stable: Changes included in this release: Process level isolation SuperSocket ServerManager (include server and client) Connect to client from server side initiatively Client certificate validation New configuration attributes "textEncoding", "defaultCulture", and "storeLocation" (certificate node) Many bug fixes http://docs.supersocket.net/v1-6/en-US/New-Features-and-Breaking-ChangesBarbaTunnel: BarbaTunnel 8.1: Check Version History for more information about this release.Mugen MVVM Toolkit: Mugen MVVM Toolkit 2.1: v 2.1 Added the 'Should' class instead of the 'Validate' class. The 'Validate' class is now obsolete. Added 'Toolkit.Annotations' to support the Mugen MVVM Toolkit ReSharper plugin. Updated JetBrains annotations within the project. Added the 'GlobalSettings.DefaultActivationPolicy' property to represent the default activation policy. Removed the 'GetSettings' method from the 'ViewModelBase' class. Instead of it, the 'GlobalSettings.DefaultViewModelSettings' property is used. Updated...NAudio: NAudio 1.7: full release notes available at http://mark-dot-net.blogspot.co.uk/2013/10/naudio-17-release-notes.htmlDirectX Tool Kit: October 2013: October 28, 2013 Updated for Visual Studio 2013 and Windows 8.1 SDK RTM Added DGSLEffect, DGSLEffectFactory, VertexPositionNormalTangentColorTexture, and VertexPositionNormalTangentColorTextureSkinning Model loading and effect factories support loading skinned models MakeSpriteFont now has a smooth vs. sharp antialiasing option: /sharp Model loading from CMOs now handles UV transforms for texture coordinates A number of small fixes for EffectFactory Minor code and project cleanup ...ExtJS based ASP.NET Controls: FineUI v4.0beta1: +2013-10-28 v4.0 beta1 +?????Collapsed???????????????。 -????:window/group_panel.aspx??,???????,???????,?????????。 +??????SelectedNodeIDArray???????????????。 -????:tree/checkbox/tree_checkall.aspx??,?????,?????,????????????。 -??TimerPicker???????(????、????ing)。 -??????????????????????(???)。 -?????????????,??type=text/css(??~`)。 -MsgTarget???MessageTarget,???None。 -FormOffsetRight?????20px??5px。 -?Web.config?PageManager??FormLabelAlign???。 -ToolbarPosition??Left/Right。 -??Web.conf...CODE Framework: 4.0.31028.0: See change notes in the documentation section for details on what's new. Note: If you download the class reference help file with, you have to right-click the file, pick "Properties", and then unblock the file, as many browsers flag the file as blocked during download (for security reasons) and thus hides all content.VidCoder: 1.5.10 Beta: Broke out all the encoder-specific passthrough options into their own dropdown. This should make what they do a bit more clear and clean up the codec list a bit. Updated HandBrake core to SVN 5855.Indent Guides for Visual Studio: Indent Guides v14: ImportantThis release has a separate download for Visual Studio 2010. The first link is for VS 2012 and later. Version History Changed in v14 Improved performance when scrolling and editing Fixed potential crash when Resharper is installed Fixed highlight of guides split around pragmas in C++/C# Restored VS 2010 support as a separate download Changed in v13 Added page width guide lines Added guide highlighting options Fixed guides appearing over collapsed blocks Fixed guides not...ASP.net MVC Awesome - jQuery Ajax Helpers: 3.5.3 (mvc5): version 3.5.3 - support for mvc5 version 3.5.2 - fix for setting single value to multivalue controls - datepicker min max date offset fix - html encoding for keys fix - enable Column.ClientFormatFunc to be a function call that will return a function version 3.5.1 ========================== - fixed html attributes rendering - fixed loading animation rendering - css improvements version 3.5 ========================== - autosize for all popups ( can be turned off by calling in js...Media Companion: Media Companion MC3.585b: IMDB plot scraping Fixed. New* Movie - Rename Folder using Movie Set, option to move ignored articles to end of Movie Set, only for folder renaming. Fixed* Media Companion - Fixed if using profiles, config files would blown up in size due to some settings duplicating. * Ignore Article of An was cutting of last character of movie title. * If Rescraping title, sort title changed depending on 'Move article to end of Sort Title' setting. * Movie - If changing Poster source order, list would beco...New ProjectsCarTrade.dk: program that will help make it easier for cardealers to do trade cars amongst themselvesCMSPORTAL: Nothing CruxOMatic Contributions: Crux-O-Matic-Contrib is the contribution project where the Crux-O-Matic community can contribute Crux-O-Matic components and extenders.Dynamics CRM 2013 Easy Solution Importer: Dynamics CRM 2013 Easy Solution ImporterGlobal Excel Automation PowerShell Library: The goal of the library is to automate common infrastructure tasks.HappyBin AutoUpdater: HappyBin is an auto-updater for .NET apps. It is designed as an api, and can be used as a boot-strap or a passive updater. Every app deserves a HappyBin!HashTag WCF Membership and Role Provider: Membership provider for using WCF to connect to a hosted membership endpont. Also includes comprehensive tests for custom membership and role providers.Luwx-Mobile-Store: Xây d?ng trang web bán di?n tho?i trên MVC4Pokémon Bank Online: Pokémon Bank for PBO/POQuickMatch Game: CLI multiplayer game on a single console.seizyUtils: SQL Mapping For Windows Embedded Compact SharePoint 2013 Global Metadata Navigation: Deploy your SharePoint 2013 Managed Metadata navigation term set to an unlimited number of site collections using JQuery, SCOM, CSS, and a custom master page.Simple ASP.NET MVC 3: a simple blog - Guest: view posts and add comments. - Admin: view, edit, delete posts and delete comments. User must be log in.SPRepository: SPRepository aims to be a generic Repository Design Pattern implementation for accessing SharePoint objects (but is not inherently SP).WebSocket Security: WebSocket SecutrityWildfire Ttraining Server: Server for application that tracks Wildfire around the world??? ???? 2013: ??????? ? ?????? ????? ???????????? 2013

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  • Applications: The mathematics of movement, Part 1

    - by TechTwaddle
    Before you continue reading this post, a suggestion; if you haven’t read “Programming Windows Phone 7 Series” by Charles Petzold, go read it. Now. If you find 150+ pages a little too long, at least go through Chapter 5, Principles of Movement, especially the section “A Brief Review of Vectors”. This post is largely inspired from this chapter. At this point I assume you know what vectors are, how they are represented using the pair (x, y), what a unit vector is, and given a vector how you would normalize the vector to get a unit vector. Our task in this post is simple, a marble is drawn at a point on the screen, the user clicks at a random point on the device, say (destX, destY), and our program makes the marble move towards that point and stop when it is reached. The tricky part of this task is the word “towards”, it adds a direction to our problem. Making a marble bounce around the screen is simple, all you have to do is keep incrementing the X and Y co-ordinates by a certain amount and handle the boundary conditions. Here, however, we need to find out exactly how to increment the X and Y values, so that the marble appears to move towards the point where the user clicked. And this is where vectors can be so helpful. The code I’ll show you here is not ideal, we’ll be working with C# on Windows Mobile 6.x, so there is no built-in vector class that I can use, though I could have written one and done all the math inside the class. I think it is trivial to the actual problem that we are trying to solve and can be done pretty easily once you know what’s going on behind the scenes. In other words, this is an excuse for me being lazy. The first approach, uses the function Atan2() to solve the “towards” part of the problem. Atan2() takes a point (x, y) as input, Atan2(y, x), note that y goes first, and then it returns an angle in radians. What angle you ask. Imagine a line from the origin (0, 0), to the point (x, y). The angle which Atan2 returns is the angle the positive X-axis makes with that line, measured clockwise. The figure below makes it clear, wiki has good details about Atan2(), give it a read. The pair (x, y) also denotes a vector. A vector whose magnitude is the length of that line, which is Sqrt(x*x + y*y), and a direction ?, as measured from positive X axis clockwise. If you’ve read that chapter from Charles Petzold’s book, this much should be clear. Now Sine and Cosine of the angle ? are special. Cosine(?) divides x by the vectors length (adjacent by hypotenuse), thus giving us a unit vector along the X direction. And Sine(?) divides y by the vectors length (opposite by hypotenuse), thus giving us a unit vector along the Y direction. Therefore the vector represented by the pair (cos(?), sin(?)), is the unit vector (or normalization) of the vector (x, y). This unit vector has a length of 1 (remember sin2(?) + cos2(?) = 1 ?), and a direction which is the same as vector (x, y). Now if I multiply this unit vector by some amount, then I will always get a point which is a certain distance away from the origin, but, more importantly, the point will always be on that line. For example, if I multiply the unit vector with the length of the line, I get the point (x, y). Thus, all we have to do to move the marble towards our destination point, is to multiply the unit vector by a certain amount each time and draw the marble, and the marble will magically move towards the click point. Now time for some code. The application, uses a timer based frame draw method to draw the marble on the screen. The timer is disabled initially and whenever the user clicks on the screen, the timer is enabled. The callback function for the timer follows the standard Update and Draw cycle. private double totLenToTravelSqrd = 0; private double startPosX = 0, startPosY = 0; private double destX = 0, destY = 0; private void Form1_MouseUp(object sender, MouseEventArgs e) {     destX = e.X;     destY = e.Y;     double x = marble1.x - destX;     double y = marble1.y - destY;     //calculate the total length to be travelled     totLenToTravelSqrd = x * x + y * y;     //store the start position of the marble     startPosX = marble1.x;     startPosY = marble1.y;     timer1.Enabled = true; } private void timer1_Tick(object sender, EventArgs e) {     UpdatePosition();     DrawMarble(); } Form1_MouseUp() method is called when ever the user touches and releases the screen. In this function we save the click point in destX and destY, this is the destination point for the marble and we also enable the timer. We store a few more values which we will use in the UpdatePosition() method to detect when the marble has reached the destination and stop the timer. So we store the start position of the marble and the square of the total length to be travelled. I’ll leave out the term ‘sqrd’ when speaking of lengths from now on. The time out interval of the timer is set to 40ms, thus giving us a frame rate of about ~25fps. In the timer callback, we update the marble position and draw the marble. We know what DrawMarble() does, so here, we’ll only look at how UpdatePosition() is implemented; private void UpdatePosition() {     //the vector (x, y)     double x = destX - marble1.x;     double y = destY - marble1.y;     double incrX=0, incrY=0;     double distanceSqrd=0;     double speed = 6;     //distance between destination and current position, before updating marble position     distanceSqrd = x * x + y * y;     double angle = Math.Atan2(y, x);     //Cos and Sin give us the unit vector, 6 is the value we use to magnify the unit vector along the same direction     incrX = speed * Math.Cos(angle);     incrY = speed * Math.Sin(angle);     marble1.x += incrX;     marble1.y += incrY;     //check for bounds     if ((int)marble1.x < MinX + marbleWidth / 2)     {         marble1.x = MinX + marbleWidth / 2;     }     else if ((int)marble1.x > (MaxX - marbleWidth / 2))     {         marble1.x = MaxX - marbleWidth / 2;     }     if ((int)marble1.y < MinY + marbleHeight / 2)     {         marble1.y = MinY + marbleHeight / 2;     }     else if ((int)marble1.y > (MaxY - marbleHeight / 2))     {         marble1.y = MaxY - marbleHeight / 2;     }     //distance between destination and current point, after updating marble position     x = destX - marble1.x;     y = destY - marble1.y;     double newDistanceSqrd = x * x + y * y;     //length from start point to current marble position     x = startPosX - (marble1.x);     y = startPosY - (marble1.y);     double lenTraveledSqrd = x * x + y * y;     //check for end conditions     if ((int)lenTraveledSqrd >= (int)totLenToTravelSqrd)     {         System.Console.WriteLine("Stopping because destination reached");         timer1.Enabled = false;     }     else if (Math.Abs((int)distanceSqrd - (int)newDistanceSqrd) < 4)     {         System.Console.WriteLine("Stopping because no change in Old and New position");         timer1.Enabled = false;     } } Ok, so in this function, first we subtract the current marble position from the destination point to give us a vector. The first three lines of the function construct this vector (x, y). The vector (x, y) has the same length as the line from (marble1.x, marble1.y) to (destX, destY) and is in the direction pointing from (marble1.x, marble1.y) to (destX, destY). Note that marble1.x and marble1.y denote the center point of the marble. Then we use Atan2() to get the angle which this vector makes with the positive X axis and use Cosine() and Sine() of that angle to get the unit vector along that same direction. We multiply this unit vector with 6, to get the values which the position of the marble should be incremented by. This variable, speed, can be experimented with and determines how fast the marble moves towards the destination. After this, we check for bounds to make sure that the marble stays within the screen limits and finally we check for the end condition and stop the timer. The end condition has two parts to it. The first case is the normal case, where the user clicks well inside the screen. Here, we stop when the total length travelled by the marble is greater than or equal to the total length to be travelled. Simple enough. The second case is when the user clicks on the very corners of the screen. Like I said before, the values marble1.x and marble1.y denote the center point of the marble. When the user clicks on the corner, the marble moves towards the point, and after some time tries to go outside of the screen, this is when the bounds checking comes into play and corrects the marble position so that the marble stays inside the screen. In this case the marble will never travel a distance of totLenToTravelSqrd, because of the correction is its position. So here we detect the end condition when there is not much change in marbles position. I use the value 4 in the second condition above. After experimenting with a few values, 4 seemed to work okay. There is a small thing missing in the code above. In the normal case, case 1, when the update method runs for the last time, marble position over shoots the destination point. This happens because the position is incremented in steps (which are not small enough), so in this case too, we should have corrected the marble position, so that the center point of the marble sits exactly on top of the destination point. I’ll add this later and update the post. This has been a pretty long post already, so I’ll leave you with a video of how this program looks while running. Notice in the video that the marble moves like a bot, without any grace what so ever. And that is because the speed of the marble is fixed at 6. In the next post we will see how to make the marble move a little more elegantly. And also, if Atan2(), Sine() and Cosine() are a little too much to digest, we’ll see how to achieve the same effect without using them, in the next to next post maybe. Ciao!

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  • The Stub Proto: Not Just For Stub Objects Anymore

    - by user9154181
    One of the great pleasures of programming is to invent something for a narrow purpose, and then to realize that it is a general solution to a broader problem. In hindsight, these things seem perfectly natural and obvious. The stub proto area used to build the core Solaris consolidation has turned out to be one of those things. As discussed in an earlier article, the stub proto area was invented as part of the effort to use stub objects to build the core ON consolidation. Its purpose was merely as a place to hold stub objects. However, we keep finding other uses for it. It turns out that the stub proto should be more properly thought of as an auxiliary place to put things that we would like to put into the proto to help us build the product, but which we do not wish to package or deliver to the end user. Stub objects are one example, but private lint libraries, header files, archives, and relocatable objects, are all examples of things that might profitably go into the stub proto. Without a stub proto, these items were handled in a variety of ad hoc ways: If one part of the workspace needed private header files, libraries, or other such items, it might modify its Makefile to reach up and over to the place in the workspace where those things live and use them from there. There are several problems with this: Each component invents its own approach, meaning that programmers maintaining the system have to invest extra effort to understand what things mean. In the past, this has created makefile ghettos in which only the person who wrote the makefiles feels confident to modify them, while everyone else ignores them. This causes many difficulties and benefits no one. These interdependencies are not obvious to the make, utility, and can lead to races. They are not obvious to the human reader, who may therefore not realize that they exist, and break them. Our policy in ON is not to deliver files into the proto unless those files are intended to be packaged and delivered to the end user. However, sometimes non-shipping files were copied into the proto anyway, causing a different set of problems: It requires a long list of exceptions to silence our normal unused proto item error checking. In the past, we have accidentally shipped files that we did not intend to deliver to the end user. Mixing cruft with valuable items makes it hard to discern which is which. The stub proto area offers a convenient and robust solution. Files needed to build the workspace that are not delivered to the end user can instead be installed into the stub proto. No special exceptions or custom make rules are needed, and the intent is always clear. We are already accessing some private lint libraries and compilation symlinks in this manner. Ultimately, I'd like to see all of the files in the proto that have a packaging exception delivered to the stub proto instead, and for the elimination of all existing special case makefile rules. This would include shared objects, header files, and lint libraries. I don't expect this to happen overnight — it will be a long term case by case project, but the overall trend is clear. The Stub Proto, -z assert_deflib, And The End Of Accidental System Object Linking We recently used the stub proto to solve an annoying build issue that goes back to the earliest days of Solaris: How to ensure that we're linking to the OS bits we're building instead of to those from the running system. The Solaris product is made up of objects and files from a number of different consolidations, each of which is built separately from the others from an independent code base called a gate. The core Solaris OS consolidation is ON, which stands for "Operating System and Networking". You will frequently also see ON called the OSnet. There are consolidations for X11 graphics, the desktop environment, open source utilities, compilers and development tools, and many others. The collection of consolidations that make up Solaris is known as the "Wad Of Stuff", usually referred to simply as the WOS. None of these consolidations is self contained. Even the core ON consolidation has some dependencies on libraries that come from other consolidations. The build server used to build the OSnet must be running a relatively recent version of Solaris, which means that its objects will be very similar to the new ones being built. However, it is necessarily true that the build system objects will always be a little behind, and that incompatible differences may exist. The objects built by the OSnet link to other objects. Some of these dependencies come from the OSnet, while others come from other consolidations. The objects from other consolidations are provided by the standard library directories on the build system (/lib, /usr/lib). The objects from the OSnet itself are supposed to come from the proto areas in the workspace, and not from the build server. In order to achieve this, we make use of the -L command line option to the link-editor. The link-editor finds dependencies by looking in the directories specified by the caller using the -L command line option. If the desired dependency is not found in one of these locations, ld will then fall back to looking at the default locations (/lib, /usr/lib). In order to use OSnet objects from the workspace instead of the system, while still accessing non-OSnet objects from the system, our Makefiles set -L link-editor options that point at the workspace proto areas. In general, this works well and dependencies are found in the right places. However, there have always been failures: Building objects in the wrong order might mean that an OSnet dependency hasn't been built before an object that needs it. If so, the dependency will not be seen in the proto, and the link-editor will silently fall back to the one on the build server. Errors in the makefiles can wipe out the -L options that our top level makefiles establish to cause ld to look at the workspace proto first. In this case, all objects will be found on the build server. These failures were rarely if ever caught. As I mentioned earlier, the objects on the build server are generally quite close to the objects built in the workspace. If they offer compatible linking interfaces, then the objects that link to them will behave properly, and no issue will ever be seen. However, if they do not offer compatible linking interfaces, the failure modes can be puzzling and hard to pin down. Either way, there won't be a compile-time warning or error. The advent of the stub proto eliminated the first type of failure. With stub objects, there is no dependency ordering, and the necessary stub object dependency will always be in place for any OSnet object that needs it. However, makefile errors do still occur, and so, the second form of error was still possible. While working on the stub object project, we realized that the stub proto was also the key to solving the second form of failure caused by makefile errors: Due to the way we set the -L options to point at our workspace proto areas, any valid object from the OSnet should be found via a path specified by -L, and not from the default locations (/lib, /usr/lib). Any OSnet object found via the default locations means that we've linked to the build server, which is an error we'd like to catch. Non-OSnet objects don't exist in the proto areas, and so are found via the default paths. However, if we were to create a symlink in the stub proto pointing at each non-OSnet dependency that we require, then the non-OSnet objects would also be found via the paths specified by -L, and not from the link-editor defaults. Given the above, we should not find any dependency objects from the link-editor defaults. Any dependency found via the link-editor defaults means that we have a Makefile error, and that we are linking to the build server inappropriately. All we need to make use of this fact is a linker option to produce a warning when it happens. Although warnings are nice, we in the OSnet have a zero tolerance policy for build noise. The -z fatal-warnings option that was recently introduced with -z guidance can be used to turn the warnings into fatal build errors, forcing the programmer to fix them. This was too easy to resist. I integrated 7021198 ld option to warn when link accesses a library via default path PSARC/2011/068 ld -z assert-deflib option into snv_161 (February 2011), shortly after the stub proto was introduced into ON. This putback introduced the -z assert-deflib option to the link-editor: -z assert-deflib=[libname] Enables warning messages for libraries specified with the -l command line option that are found by examining the default search paths provided by the link-editor. If a libname value is provided, the default library warning feature is enabled, and the specified library is added to a list of libraries for which no warnings will be issued. Multiple -z assert-deflib options can be specified in order to specify multiple libraries for which warnings should not be issued. The libname value should be the name of the library file, as found by the link-editor, without any path components. For example, the following enables default library warnings, and excludes the standard C library. ld ... -z assert-deflib=libc.so ... -z assert-deflib is a specialized option, primarily of interest in build environments where multiple objects with the same name exist and tight control over the library used is required. If is not intended for general use. Note that the definition of -z assert-deflib allows for exceptions to be specified as arguments to the option. In general, the idea of using a symlink from the stub proto is superior because it does not clutter up the link command with a long list of objects. When building the OSnet, we usually use the plain from of -z deflib, and make symlinks for the non-OSnet dependencies. The exception to this are dependencies supplied by the compiler itself, which are usually found at whatever arbitrary location the compiler happens to be installed at. To handle these special cases, the command line version works better. Following the integration of the link-editor change, I made use of -z assert-deflib in OSnet builds with 7021896 Prevent OSnet from accidentally linking to build system which integrated into snv_162 (March 2011). Turning on -z assert-deflib exposed between 10 and 20 existing errors in our Makefiles, which were all fixed in the same putback. The errors we found in our Makefiles underscore how difficult they can be prevent without an automatic system in place to catch them. Conclusions The stub proto is proving to be a generally useful construct for ON builds that goes beyond serving as a place to hold stub objects. Although invented to hold stub objects, it has already allowed us to simplify a number of previously difficult situations in our makefiles and builds. I expect that we'll find uses for it beyond those described here as we go forward.

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  • Key ATG architecture principles

    - by Glen Borkowski
    Overview The purpose of this article is to describe some of the important foundational concepts of ATG.  This is not intended to cover all areas of the ATG platform, just the most important subset - the ones that allow ATG to be extremely flexible, configurable, high performance, etc.  For more information on these topics, please see the online product manuals. Modules The first concept is called the 'ATG Module'.  Simply put, you can think of modules as the building blocks for ATG applications.  The ATG development team builds the out of the box product using modules (these are the 'out of the box' modules).  Then, when a customer is implementing their site, they build their own modules that sit 'on top' of the out of the box ATG modules.  Modules can be very simple - containing minimal definition, and perhaps a small amount of configuration.  Alternatively, a module can be rather complex - containing custom logic, database schema definitions, configuration, one or more web applications, etc.  Modules generally will have dependencies on other modules (the modules beneath it).  For example, the Commerce Reference Store module (CRS) requires the DCS (out of the box commerce) module. Modules have a ton of value because they provide a way to decouple a customers implementation from the out of the box ATG modules.  This allows for a much easier job when it comes time to upgrade the ATG platform.  Modules are also a very useful way to group functionality into a single package which can be leveraged across multiple ATG applications. One very important thing to understand about modules, or more accurately, ATG as a whole, is that when you start ATG, you tell it what module(s) you want to start.  One of the first things ATG does is to look through all the modules you specified, and for each one, determine a list of modules that are also required to start (based on each modules dependencies).  Once this final, ordered list is determined, ATG continues to boot up.  One of the outputs from the ordered list of modules is that each module can contain it's own classes and configuration.  During boot, the ordered list of modules drives the unified classpath and configpath.  This is what determines which classes override others, and which configuration overrides other configuration.  Think of it as a layered approach. The structure of a module is well defined.  It simply looks like a folder in a filesystem that has certain other folders and files within it.  Here is a list of items that can appear in a module: MyModule: META-INF - this is required, along with a file called MANIFEST.MF which describes certain properties of the module.  One important property is what other modules this module depends on. config - this is typically present in most modules.  It defines a tree structure (folders containing properties files, XML, etc) that maps to ATG components (these are described below). lib - this contains the classes (typically in jarred format) for any code defined in this module j2ee - this is where any web-apps would be stored. src - in case you want to include the source code for this module, it's standard practice to put it here sql - if your module requires any additions to the database schema, you should place that schema here Here's a screenshots of a module: Modules can also contain sub-modules.  A dot-notation is used when referring to these sub-modules (i.e. MyModule.Versioned, where Versioned is a sub-module of MyModule). Finally, it is important to completely understand how modules work if you are going to be able to leverage them effectively.  There are many different ways to design modules you want to create, some approaches are better than others, especially if you plan to share functionality between multiple different ATG applications. Components A component in ATG can be thought of as a single item that performs a certain set of related tasks.  An example could be a ProductViews component - used to store information about what products the current customer has viewed.  Components have properties (also called attributes).  The ProductViews component could have properties like lastProductViewed (stores the ID of the last product viewed) or productViewList (stores the ID's of products viewed in order of their being viewed).  The previous examples of component properties would typically also offer get and set methods used to retrieve and store the property values.  Components typically will also offer other types of useful methods aside from get and set.  In the ProductViewed component, we might want to offer a hasViewed method which will tell you if the customer has viewed a certain product or not. Components are organized in a tree like hierarchy called 'nucleus'.  Nucleus is used to locate and instantiate ATG Components.  So, when you create a new ATG component, it will be able to be found 'within' nucleus.  Nucleus allows ATG components to reference one another - this is how components are strung together to perform meaningful work.  It's also a mechanism to prevent redundant configuration - define it once and refer to it from everywhere. Here is a screenshot of a component in nucleus:  Components can be extremely simple (i.e. a single property with a get method), or can be rather complex offering many properties and methods.  To be an ATG component, a few things are required: a class - you can reference an existing out of the box class or you could write your own a properties file - this is used to define your component the above items must be located 'within' nucleus by placing them in the correct spot in your module's config folder Within the properties file, you will need to point to the class you want to use: $class=com.mycompany.myclass You may also want to define the scope of the class (request, session, or global): $scope=session In summary, ATG Components live in nucleus, generally have links to other components, and provide some meaningful type of work.  You can configure components as well as extend their functionality by writing code. Repositories Repositories (a.k.a. Data Anywhere Architecture) is the mechanism that ATG uses to access data primarily stored in relational databases, but also LDAP or other backend systems.  ATG applications are required to be very high performance, and data access is critical in that if not handled properly, it could create a bottleneck.  ATG's repository functionality has been around for a long time - it's proven to be extremely scalable.  Developers new to ATG need to understand how repositories work as this is a critical aspect of the ATG architecture.   Repositories essentially map relational tables to objects in ATG, as well as handle caching.  ATG defines many repositories out of the box (i.e. user profile, catalog, orders, etc), and this is comprised of both the underlying database schema along with the associated repository definition files (XML).  It is fully expected that implementations will extend / change the out of the box repository definitions, so there is a prescribed approach to doing this.  The first thing to be sure of is to encapsulate your repository definition additions / changes within your own module (as described above).  The other important best practice is to never modify the out of the box schema - in other words, don't add columns to existing ATG tables, just create your own new tables.  These will help ensure you can easily upgrade your application at a later date. xml-combination As mentioned earlier, when you start ATG, the order of the modules will determine the final configpath.  Files within this configpath are 'layered' such that modules on top can override configuration of modules below it.  This is the same concept for repository definition files.  If you want to add a few properties to the out of the box user profile, you simply need to create an XML file containing only your additions, and place it in the correct location in your module.  At boot time, your definition will be combined (hence the term xml-combination) with the lower, out of the box modules, with the result being a user profile that contains everything (out of the box, plus your additions).  Aside from just adding properties, there are also ways to remove and change properties. types of properties Aside from the normal 'database backed' properties, there are a few other interesting types: transient properties - these are properties that are in memory, but not backed by any database column.  These are useful for temporary storage. java-backed properties - by nature, these are transient, but in addition, when you access this property (by called the get method) instead of looking up a piece of data, it performs some logic and returns the results.  'Age' is a good example - if you're storing a birth date on the profile, but your business rules are defined in terms of someones age, you could create a simple java-backed property to look at the birth date and compare it to the current date, and return the persons age. derived properties - this is what allows for inheritance within the repository structure.  You could define a property at the category level, and have the product inherit it's value as well as override it.  This is useful for setting defaults, with the ability to override. caching There are a number of different caching modes which are useful at different times depending on the nature of the data being cached.  For example, the simple cache mode is useful for things like user profiles.  This is because the user profile will typically only be used on a single instance of ATG at one time.  Simple cache mode is also useful for read-only types of data such as the product catalog.  Locked cache mode is useful when you need to ensure that only one ATG instance writes to a particular item at a time - an example would be a customers order.  There are many options in terms of configuring caching which are outside the scope of this article - please refer to the product manuals for more details. Other important concepts - out of scope for this article There are a whole host of concepts that are very important pieces to the ATG platform, but are out of scope for this article.  Here's a brief description of some of them: formhandlers - these are ATG components that handle form submissions by users. pipelines - these are configurable chains of logic that are used for things like handling a request (request pipeline) or checking out an order. special kinds of repositories (versioned, files, secure, ...) - there are a couple different types of repositories that are used in various situations.  See the manuals for more information. web development - JSP/ DSP tag library - ATG provides a traditional approach to developing web applications by providing a tag library called the DSP library.  This library is used throughout your JSP pages to interact with all the ATG components. messaging - a message sub-system used as another way for components to interact. personalization - ability for business users to define a personalized user experience for customers.  See the other blog posts related to personalization.

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  • Slow boot on Ubuntu 12.04, probable cause the network connection

    - by Ravi S Ghosh
    I have been having rather slow boot on Ubuntu 12.04. Lately, I tried to figure out the reason and it seems to be the network connection which does not get connected and requires multiple attempts. Here is part of dmesg [ 2.174349] EXT4-fs (sda2): INFO: recovery required on readonly filesystem [ 2.174352] EXT4-fs (sda2): write access will be enabled during recovery [ 2.308172] firewire_core: created device fw0: GUID 384fc00005198d58, S400 [ 2.333457] usb 7-1.2: new low-speed USB device number 3 using uhci_hcd [ 2.465896] EXT4-fs (sda2): recovery complete [ 2.466406] EXT4-fs (sda2): mounted filesystem with ordered data mode. Opts: (null) [ 2.589440] usb 7-1.3: new low-speed USB device number 4 using uhci_hcd **[ 18.292029] ADDRCONF(NETDEV_UP): eth0: link is not ready** [ 18.458958] udevd[377]: starting version 175 [ 18.639482] Adding 4200960k swap on /dev/sda5. Priority:-1 extents:1 across:4200960k [ 19.314127] wmi: Mapper loaded [ 19.426602] r592 0000:09:01.2: PCI INT B -> GSI 18 (level, low) -> IRQ 18 [ 19.426739] r592: driver successfully loaded [ 19.460105] input: Dell WMI hotkeys as /devices/virtual/input/input5 [ 19.493629] lp: driver loaded but no devices found [ 19.497012] cfg80211: Calling CRDA to update world regulatory domain [ 19.535523] ACPI Warning: _BQC returned an invalid level (20110623/video-480) [ 19.539457] acpi device:03: registered as cooling_device2 [ 19.539520] input: Video Bus as /devices/LNXSYSTM:00/device:00/PNP0A08:00/device:01/LNXVIDEO:00/input/input6 [ 19.539568] ACPI: Video Device [M86] (multi-head: yes rom: no post: no) [ 19.578060] Linux video capture interface: v2.00 [ 19.667708] dcdbas dcdbas: Dell Systems Management Base Driver (version 5.6.0-3.2) [ 19.763171] r852 0000:09:01.3: PCI INT B -> GSI 18 (level, low) -> IRQ 18 [ 19.763258] r852: driver loaded successfully [ 19.854769] input: Microsoft Comfort Curve Keyboard 2000 as /devices/pci0000:00/0000:00:1d.1/usb7/7-1/7-1.2/7-1.2:1.0/input/input7 [ 19.854864] generic-usb 0003:045E:00DD.0001: input,hidraw0: USB HID v1.11 Keyboard [Microsoft Comfort Curve Keyboard 2000] on usb-0000:00:1d.1-1.2/input0 [ 19.878605] input: Microsoft Comfort Curve Keyboard 2000 as /devices/pci0000:00/0000:00:1d.1/usb7/7-1/7-1.2/7-1.2:1.1/input/input8 [ 19.878698] generic-usb 0003:045E:00DD.0002: input,hidraw1: USB HID v1.11 Device [Microsoft Comfort Curve Keyboard 2000] on usb-0000:00:1d.1-1.2/input1 [ 19.902779] input: DELL DELL USB Laser Mouse as /devices/pci0000:00/0000:00:1d.1/usb7/7-1/7-1.3/7-1.3:1.0/input/input9 [ 19.925034] generic-usb 0003:046D:C063.0003: input,hidraw2: USB HID v1.10 Mouse [DELL DELL USB Laser Mouse] on usb-0000:00:1d.1-1.3/input0 [ 19.925057] usbcore: registered new interface driver usbhid [ 19.925059] usbhid: USB HID core driver [ 19.942362] uvcvideo: Found UVC 1.00 device Laptop_Integrated_Webcam_2M (0c45:63ea) [ 19.947004] input: Laptop_Integrated_Webcam_2M as /devices/pci0000:00/0000:00:1a.7/usb1/1-6/1-6:1.0/input/input10 [ 19.947075] usbcore: registered new interface driver uvcvideo [ 19.947077] USB Video Class driver (1.1.1) [ 20.145232] Intel(R) Wireless WiFi Link AGN driver for Linux, in-tree: [ 20.145235] Copyright(c) 2003-2011 Intel Corporation [ 20.145327] iwlwifi 0000:04:00.0: PCI INT A -> GSI 17 (level, low) -> IRQ 17 [ 20.145357] iwlwifi 0000:04:00.0: setting latency timer to 64 [ 20.145402] iwlwifi 0000:04:00.0: pci_resource_len = 0x00002000 [ 20.145404] iwlwifi 0000:04:00.0: pci_resource_base = ffffc90000674000 [ 20.145407] iwlwifi 0000:04:00.0: HW Revision ID = 0x0 [ 20.145531] iwlwifi 0000:04:00.0: irq 46 for MSI/MSI-X [ 20.145613] iwlwifi 0000:04:00.0: Detected Intel(R) WiFi Link 5100 AGN, REV=0x54 [ 20.145720] iwlwifi 0000:04:00.0: L1 Enabled; Disabling L0S [ 20.167535] iwlwifi 0000:04:00.0: device EEPROM VER=0x11f, CALIB=0x4 [ 20.167538] iwlwifi 0000:04:00.0: Device SKU: 0Xf0 [ 20.167567] iwlwifi 0000:04:00.0: Tunable channels: 13 802.11bg, 24 802.11a channels [ 20.172779] fglrx: module license 'Proprietary. (C) 2002 - ATI Technologies, Starnberg, GERMANY' taints kernel. [ 20.172783] Disabling lock debugging due to kernel taint [ 20.250115] [fglrx] Maximum main memory to use for locked dma buffers: 3759 MBytes. [ 20.250567] [fglrx] vendor: 1002 device: 9553 count: 1 [ 20.251256] [fglrx] ioport: bar 1, base 0x2000, size: 0x100 [ 20.251271] pci 0000:01:00.0: PCI INT A -> GSI 16 (level, low) -> IRQ 16 [ 20.251277] pci 0000:01:00.0: setting latency timer to 64 [ 20.251559] [fglrx] Kernel PAT support is enabled [ 20.251578] [fglrx] module loaded - fglrx 8.96.4 [Mar 12 2012] with 1 minors [ 20.310385] iwlwifi 0000:04:00.0: loaded firmware version 8.83.5.1 build 33692 [ 20.310598] Registered led device: phy0-led [ 20.310628] cfg80211: Ignoring regulatory request Set by core since the driver uses its own custom regulatory domain [ 20.372306] ieee80211 phy0: Selected rate control algorithm 'iwl-agn-rs' [ 20.411015] psmouse serio1: synaptics: Touchpad model: 1, fw: 7.2, id: 0x1c0b1, caps: 0xd04733/0xa40000/0xa0000 [ 20.454232] input: SynPS/2 Synaptics TouchPad as /devices/platform/i8042/serio1/input/input11 [ 20.545636] cfg80211: Ignoring regulatory request Set by core since the driver uses its own custom regulatory domain [ 20.545640] cfg80211: World regulatory domain updated: [ 20.545642] cfg80211: (start_freq - end_freq @ bandwidth), (max_antenna_gain, max_eirp) [ 20.545644] cfg80211: (2402000 KHz - 2472000 KHz @ 40000 KHz), (300 mBi, 2000 mBm) [ 20.545647] cfg80211: (2457000 KHz - 2482000 KHz @ 20000 KHz), (300 mBi, 2000 mBm) [ 20.545649] cfg80211: (2474000 KHz - 2494000 KHz @ 20000 KHz), (300 mBi, 2000 mBm) [ 20.545652] cfg80211: (5170000 KHz - 5250000 KHz @ 40000 KHz), (300 mBi, 2000 mBm) [ 20.545654] cfg80211: (5735000 KHz - 5835000 KHz @ 40000 KHz), (300 mBi, 2000 mBm) [ 20.609484] type=1400 audit(1340502633.160:2): apparmor="STATUS" operation="profile_load" name="/sbin/dhclient" pid=693 comm="apparmor_parser" [ 20.609494] type=1400 audit(1340502633.160:3): apparmor="STATUS" operation="profile_replace" name="/sbin/dhclient" pid=642 comm="apparmor_parser" [ 20.609843] type=1400 audit(1340502633.160:4): apparmor="STATUS" operation="profile_load" name="/usr/lib/NetworkManager/nm-dhcp-client.action" pid=693 comm="apparmor_parser" [ 20.609852] type=1400 audit(1340502633.160:5): apparmor="STATUS" operation="profile_replace" name="/usr/lib/NetworkManager/nm-dhcp-client.action" pid=642 comm="apparmor_parser" [ 20.610047] type=1400 audit(1340502633.160:6): apparmor="STATUS" operation="profile_load" name="/usr/lib/connman/scripts/dhclient-script" pid=693 comm="apparmor_parser" [ 20.610060] type=1400 audit(1340502633.160:7): apparmor="STATUS" operation="profile_replace" name="/usr/lib/connman/scripts/dhclient-script" pid=642 comm="apparmor_parser" [ 20.610476] type=1400 audit(1340502633.160:8): apparmor="STATUS" operation="profile_replace" name="/sbin/dhclient" pid=814 comm="apparmor_parser" [ 20.610829] type=1400 audit(1340502633.160:9): apparmor="STATUS" operation="profile_replace" name="/usr/lib/NetworkManager/nm-dhcp-client.action" pid=814 comm="apparmor_parser" [ 20.611035] type=1400 audit(1340502633.160:10): apparmor="STATUS" operation="profile_replace" name="/usr/lib/connman/scripts/dhclient-script" pid=814 comm="apparmor_parser" [ 20.661912] snd_hda_intel 0000:00:1b.0: PCI INT A -> GSI 22 (level, low) -> IRQ 22 [ 20.661982] snd_hda_intel 0000:00:1b.0: irq 47 for MSI/MSI-X [ 20.662013] snd_hda_intel 0000:00:1b.0: setting latency timer to 64 [ 20.770289] input: HDA Intel Mic as /devices/pci0000:00/0000:00:1b.0/sound/card0/input12 [ 20.770689] snd_hda_intel 0000:01:00.1: PCI INT B -> GSI 17 (level, low) -> IRQ 17 [ 20.770786] snd_hda_intel 0000:01:00.1: irq 48 for MSI/MSI-X [ 20.770815] snd_hda_intel 0000:01:00.1: setting latency timer to 64 [ 20.994040] HDMI status: Codec=0 Pin=3 Presence_Detect=0 ELD_Valid=0 [ 20.994189] input: HDA ATI HDMI HDMI/DP,pcm=3 as /devices/pci0000:00/0000:00:01.0/0000:01:00.1/sound/card1/input13 [ 21.554799] vesafb: mode is 1024x768x32, linelength=4096, pages=0 [ 21.554802] vesafb: scrolling: redraw [ 21.554804] vesafb: Truecolor: size=0:8:8:8, shift=0:16:8:0 [ 21.557342] vesafb: framebuffer at 0xd0000000, mapped to 0xffffc90011800000, using 3072k, total 3072k [ 21.557498] Console: switching to colour frame buffer device 128x48 [ 21.557516] fb0: VESA VGA frame buffer device [ 21.987338] EXT4-fs (sda2): re-mounted. Opts: errors=remount-ro [ 22.184693] EXT4-fs (sda6): mounted filesystem with ordered data mode. Opts: (null) [ 27.362440] iwlwifi 0000:04:00.0: RF_KILL bit toggled to disable radio. [ 27.436988] init: failsafe main process (986) killed by TERM signal [ 27.970112] ppdev: user-space parallel port driver [ 28.198917] Bluetooth: Core ver 2.16 [ 28.198935] NET: Registered protocol family 31 [ 28.198937] Bluetooth: HCI device and connection manager initialized [ 28.198940] Bluetooth: HCI socket layer initialized [ 28.198941] Bluetooth: L2CAP socket layer initialized [ 28.198947] Bluetooth: SCO socket layer initialized [ 28.226135] Bluetooth: RFCOMM TTY layer initialized [ 28.226141] Bluetooth: RFCOMM socket layer initialized [ 28.226143] Bluetooth: RFCOMM ver 1.11 [ 28.445620] Bluetooth: BNEP (Ethernet Emulation) ver 1.3 [ 28.445623] Bluetooth: BNEP filters: protocol multicast [ 28.524578] type=1400 audit(1340502641.076:11): apparmor="STATUS" operation="profile_load" name="/usr/lib/cups/backend/cups-pdf" pid=1052 comm="apparmor_parser" [ 28.525018] type=1400 audit(1340502641.076:12): apparmor="STATUS" operation="profile_load" name="/usr/sbin/cupsd" pid=1052 comm="apparmor_parser" [ 28.629957] type=1400 audit(1340502641.180:13): apparmor="STATUS" operation="profile_replace" name="/sbin/dhclient" pid=1105 comm="apparmor_parser" [ 28.630325] type=1400 audit(1340502641.180:14): apparmor="STATUS" operation="profile_replace" name="/usr/lib/NetworkManager/nm-dhcp-client.action" pid=1105 comm="apparmor_parser" [ 28.630535] type=1400 audit(1340502641.180:15): apparmor="STATUS" operation="profile_replace" name="/usr/lib/connman/scripts/dhclient-script" pid=1105 comm="apparmor_parser" [ 28.645266] type=1400 audit(1340502641.196:16): apparmor="STATUS" operation="profile_load" name="/usr/lib/lightdm/lightdm/lightdm-guest-session-wrapper" pid=1104 comm="apparmor_parser" **[ 28.751922] ADDRCONF(NETDEV_UP): wlan0: link is not ready** [ 28.753653] tg3 0000:08:00.0: irq 49 for MSI/MSI-X **[ 28.856127] ADDRCONF(NETDEV_UP): eth0: link is not ready [ 28.857034] ADDRCONF(NETDEV_UP): eth0: link is not ready** [ 28.871080] type=1400 audit(1340502641.420:17): apparmor="STATUS" operation="profile_load" name="/usr/lib/telepathy/mission-control-5" pid=1108 comm="apparmor_parser" [ 28.871519] type=1400 audit(1340502641.420:18): apparmor="STATUS" operation="profile_load" name="/usr/lib/telepathy/telepathy-*" pid=1108 comm="apparmor_parser" [ 28.874905] type=1400 audit(1340502641.424:19): apparmor="STATUS" operation="profile_replace" name="/usr/lib/cups/backend/cups-pdf" pid=1113 comm="apparmor_parser" [ 28.875354] type=1400 audit(1340502641.424:20): apparmor="STATUS" operation="profile_replace" name="/usr/sbin/cupsd" pid=1113 comm="apparmor_parser" [ 30.477976] tg3 0000:08:00.0: eth0: Link is up at 100 Mbps, full duplex [ 30.477979] tg3 0000:08:00.0: eth0: Flow control is on for TX and on for RX **[ 30.478390] ADDRCONF(NETDEV_CHANGE): eth0: link becomes ready** [ 31.110269] fglrx_pci 0000:01:00.0: irq 50 for MSI/MSI-X [ 31.110859] [fglrx] Firegl kernel thread PID: 1327 [ 31.111021] [fglrx] Firegl kernel thread PID: 1329 [ 31.111408] [fglrx] Firegl kernel thread PID: 1330 [ 31.111543] [fglrx] IRQ 50 Enabled [ 31.712938] [fglrx] Gart USWC size:1224 M. [ 31.712941] [fglrx] Gart cacheable size:486 M. [ 31.712945] [fglrx] Reserved FB block: Shared offset:0, size:1000000 [ 31.712948] [fglrx] Reserved FB block: Unshared offset:fc2b000, size:3d5000 [ 31.712950] [fglrx] Reserved FB block: Unshared offset:1fffb000, size:5000 [ 41.312020] eth0: no IPv6 routers present As you can see I get multiple instances of [ 28.856127] ADDRCONF(NETDEV_UP): eth0: link is not ready and then finally it becomes read and I get the message [ 30.478390] ADDRCONF(NETDEV_CHANGE): eth0: link becomes ready. I searched askubuntun, ubuntuforum, and the web but couldn't find a solution. Any help would be very much appreciated. Here is the bootchart

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  • Working with Timelines with LINQ to Twitter

    - by Joe Mayo
    When first working with the Twitter API, I thought that using SinceID would be an effective way to page through timelines. In practice it doesn’t work well for various reasons. To explain why, Twitter published an excellent document that is a must-read for anyone working with timelines: Twitter Documentation: Working with Timelines This post shows how to implement the recommended strategies in that document by using LINQ to Twitter. You should read the document in it’s entirety before moving on because my explanation will start at the bottom and work back up to the top in relation to the Twitter document. What follows is an explanation of SinceID, MaxID, and how they come together to help you efficiently work with Twitter timelines. The Role of SinceID Specifying SinceID says to Twitter, “Don’t return tweets earlier than this”. What you want to do is store this value after every timeline query set so that it can be reused on the next set of queries.  The next section will explain what I mean by query set, but a quick explanation is that it’s a loop that gets all new tweets. The SinceID is a backstop to avoid retrieving tweets that you already have. Here’s some initialization code that includes a variable named sinceID that will be used to populate the SinceID property in subsequent queries: // last tweet processed on previous query set ulong sinceID = 210024053698867204; ulong maxID; const int Count = 10; var statusList = new List<status>(); Here, I’ve hard-coded the sinceID variable, but this is where you would initialize sinceID from whatever storage you choose (i.e. a database). The first time you ever run this code, you won’t have a value from a previous query set. Initially setting it to 0 might sound like a good idea, but what if you’re querying a timeline with lots of tweets? Because of the number of tweets and rate limits, your query set might take a very long time to run. A caveat might be that Twitter won’t return an entire timeline back to Tweet #0, but rather only go back a certain period of time, the limits of which are documented for individual Twitter timeline API resources. So, to initialize SinceID at too low of a number can result in a lot of initial tweets, yet there is a limit to how far you can go back. What you’re trying to accomplish in your application should guide you in how to initially set SinceID. I have more to say about SinceID later in this post. The other variables initialized above include the declaration for MaxID, Count, and statusList. The statusList variable is a holder for all the timeline tweets collected during this query set. You can set Count to any value you want as the largest number of tweets to retrieve, as defined by individual Twitter timeline API resources. To effectively page results, you’ll use the maxID variable to set the MaxID property in queries, which I’ll discuss next. Initializing MaxID On your first query of a query set, MaxID will be whatever the most recent tweet is that you get back. Further, you don’t know what MaxID is until after the initial query. The technique used in this post is to do an initial query and then use the results to figure out what the next MaxID will be.  Here’s the code for the initial query: var userStatusResponse = (from tweet in twitterCtx.Status where tweet.Type == StatusType.User && tweet.ScreenName == "JoeMayo" && tweet.SinceID == sinceID && tweet.Count == Count select tweet) .ToList(); statusList.AddRange(userStatusResponse); // first tweet processed on current query maxID = userStatusResponse.Min( status => ulong.Parse(status.StatusID)) - 1; The query above sets both SinceID and Count properties. As explained earlier, Count is the largest number of tweets to return, but the number can be less. A couple reasons why the number of tweets that are returned could be less than Count include the fact that the user, specified by ScreenName, might not have tweeted Count times yet or might not have tweeted at least Count times within the maximum number of tweets that can be returned by the Twitter timeline API resource. Another reason could be because there aren’t Count tweets between now and the tweet ID specified by sinceID. Setting SinceID constrains the results to only those tweets that occurred after the specified Tweet ID, assigned via the sinceID variable in the query above. The statusList is an accumulator of all tweets receive during this query set. To simplify the code, I left out some logic to check whether there were no tweets returned. If  the query above doesn’t return any tweets, you’ll receive an exception when trying to perform operations on an empty list. Yeah, I cheated again. Besides querying initial tweets, what’s important about this code is the final line that sets maxID. It retrieves the lowest numbered status ID in the results. Since the lowest numbered status ID is for a tweet we already have, the code decrements the result by one to keep from asking for that tweet again. Remember, SinceID is not inclusive, but MaxID is. The maxID variable is now set to the highest possible tweet ID that can be returned in the next query. The next section explains how to use MaxID to help get the remaining tweets in the query set. Retrieving Remaining Tweets Earlier in this post, I defined a term that I called a query set. Essentially, this is a group of requests to Twitter that you perform to get all new tweets. A single query might not be enough to get all new tweets, so you’ll have to start at the top of the list that Twitter returns and keep making requests until you have all new tweets. The previous section showed the first query of the query set. The code below is a loop that completes the query set: do { // now add sinceID and maxID userStatusResponse = (from tweet in twitterCtx.Status where tweet.Type == StatusType.User && tweet.ScreenName == "JoeMayo" && tweet.Count == Count && tweet.SinceID == sinceID && tweet.MaxID == maxID select tweet) .ToList(); if (userStatusResponse.Count > 0) { // first tweet processed on current query maxID = userStatusResponse.Min( status => ulong.Parse(status.StatusID)) - 1; statusList.AddRange(userStatusResponse); } } while (userStatusResponse.Count != 0 && statusList.Count < 30); Here we have another query, but this time it includes the MaxID property. The SinceID property prevents reading tweets that we’ve already read and Count specifies the largest number of tweets to return. Earlier, I mentioned how it was important to check how many tweets were returned because failing to do so will result in an exception when subsequent code runs on an empty list. The code above protects against this problem by only working with the results if Twitter actually returns tweets. Reasons why there wouldn’t be results include: if the first query got all the new tweets there wouldn’t be more to get and there might not have been any new tweets between the SinceID and MaxID settings of the most recent query. The code for loading the returned tweets into statusList and getting the maxID are the same as previously explained. The important point here is that MaxID is being reset, not SinceID. As explained in the Twitter documentation, paging occurs from the newest tweets to oldest, so setting MaxID lets us move from the most recent tweets down to the oldest as specified by SinceID. The two loop conditions cause the loop to continue as long as tweets are being read or a max number of tweets have been read.  Logically, you want to stop reading when you’ve read all the tweets and that’s indicated by the fact that the most recent query did not return results. I put the check to stop after 30 tweets are reached to keep the demo from running too long – in the console the response scrolls past available buffer and I wanted you to be able to see the complete output. Yet, there’s another point to be made about constraining the number of items you return at one time. The Twitter API has rate limits and making too many queries per minute will result in an error from twitter that LINQ to Twitter raises as an exception. To use the API properly, you’ll have to ensure you don’t exceed this threshold. Looking at the statusList.Count as done above is rather primitive, but you can implement your own logic to properly manage your rate limit. Yeah, I cheated again. Summary Now you know how to use LINQ to Twitter to work with Twitter timelines. After reading this post, you have a better idea of the role of SinceID - the oldest tweet already received. You also know that MaxID is the largest tweet ID to retrieve in a query. Together, these settings allow you to page through results via one or more queries. You also understand what factors affect the number of tweets returned and considerations for potential error handling logic. The full example of the code for this post is included in the downloadable source code for LINQ to Twitter.   @JoeMayo

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  • Introducing Oracle Multitenant

    - by OracleMultitenant
    0 0 1 1142 6510 Oracle Corporation 54 15 7637 14.0 Normal 0 false false false EN-US JA X-NONE /* 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-parent:""; mso-padding-alt:0in 5.4pt 0in 5.4pt; mso-para-margin:0in; mso-para-margin-bottom:.0001pt; mso-pagination:widow-orphan; font-size:10.0pt; font-family:"Times New Roman"; mso-fareast-language:JA;} The First Database Designed for the Cloud Today Oracle announced the general availability (GA) of Oracle Database 12c, the first database designed for the Cloud. Oracle Multitenant, new with Oracle Database 12c, is a key component of this – a new architecture for consolidating databases and simplifying operations in the Cloud. With this, the inaugural post in the Multitenant blog, my goal is to start the conversation about Oracle Multitenant. We are very proud of this new architecture, which we view as a major advance for Oracle. Customers, partners and analysts who have had previews are very excited about its capabilities and its flexibility. This high level review of Oracle Multitenant will touch on our design considerations and how we re-architected our database for the cloud. I’ll briefly describe our new multitenant architecture and explain it’s key benefits. Finally I’ll mention some of the major use cases we see for Oracle Multitenant. Industry Trends We always start by talking to our customers about the pressures and challenges they’re facing and what trends they’re seeing in the industry. Some things don’t change. They face the same pressures and the same requirements as ever: Pressure to do more with less; be faster, leaner, cheaper, and deliver services 24/7. Big companies have achieved scale. Now they want to realize economies of scale. As ever, DBAs are faced with the challenges of patching and upgrading large numbers of databases, and provisioning new ones.  Requirements are familiar: Performance, scalability, reliability and high availability are non-negotiable. They need ever more security in this threatening climate. There’s no time to stop and retool with new applications. What’s new are the trends. These are the techniques to use to respond to these pressures within the constraints of the requirements. With the advent of cloud computing and availability of massively powerful servers – even engineered systems such as Exadata – our customers want to consolidate many applications into fewer larger servers. There’s a move to standardized services – even self-service. Consolidation Consolidation is not new; companies have tried various different approaches to consolidation of databases in the cloud. One approach is to partition a powerful server between several virtual machines, one per application. A downside of this is that you have the resource and management overheads of OS and RDBMS per VM – that is, per application. Another is that you have replaced physical sprawl with virtual sprawl and virtual sprawl is still expensive to manage. In the dedicated database model, we have a single physical server supporting multiple databases, one per application. So there’s a shared OS overhead, but RDBMS process and memory overhead are replicated per application. Let's think about our traditional Oracle Database architecture. Every time we create a database, be it a production database, a development or a test database, what do we do? We create a set of files, we allocate a bunch of memory for managing the data, and we kick off a series of background processes. This is replicated for every one of the databases that we create. As more and more databases are fired up, these replicated overheads quickly consume the available server resources and this limits the number of applications we can run on any given server. In Oracle Database 11g and earlier the highest degree of consolidation could be achieved by what we call schema consolidation. In this model we have one big server with one big database. Individual applications are installed in separate schemas or table-owners. Database overheads are shared between all applications, which affords maximum consolidation. The shortcomings are that application changes are often required. There is no tenant isolation. One bad apple can spoil the whole batch. New Architecture & Benefits In Oracle Database 12c, we have a new multitenant architecture, featuring pluggable databases. This delivers all the resource utilization advantages of schema consolidation with none of the downsides. There are two parts to the term “pluggable database”: "pluggable", which is new, and "database", which is familiar.  Before we get to the exciting new stuff let’s discuss what hasn’t changed. A pluggable database is a fully functional Oracle database. It’s not watered down in any way. From the perspective of an application or an end user it hasn’t changed at all. This is very important because it means that no application changes are required to adopt this new architecture. There are many thousands of applications built on Oracle databases and they are all ready to run on Oracle Multitenant. So we have these self-contained pluggable databases (PDBs), and as their name suggests, they are plugged into a multitenant container database (CDB). The CDB behaves as a single database from the operations point of view. Very much as we had with the schema consolidation model, we only have a single set of Oracle background processes and a single, shared database memory requirement. This gives us very high consolidation density, which affords maximum reduction in capital expenses (CapEx). By performing management operations at the CDB level – “managing many as one” – we can achieve great reductions in operating expenses (OpEx) as well, but we retain granular control where appropriate. Furthermore, the “pluggability” capability gives us portability and this adds a tremendous amount of agility. We can simply unplug a PDB from one CDB and plug it into another CDB, for example to move it from one SLA tier to another. I'll explore all these new capabilities in much more detail in a future posting.  Use Cases We can identify a number of use cases for Oracle Multitenant. Here are a few of the major ones. 0 0 1 113 650 Oracle Corporation 5 1 762 14.0 Normal 0 false false false EN-US JA X-NONE /* 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-parent:""; mso-padding-alt:0in 5.4pt 0in 5.4pt; mso-para-margin:0in; mso-para-margin-bottom:.0001pt; mso-pagination:widow-orphan; font-size:10.0pt; font-family:"Times New Roman"; mso-fareast-language:JA;} Development / Testing where individual engineers need rapid provisioning and recycling of private copies of a few "master test databases" Consolidation of disparate applications using fewer, more powerful servers Software as a Service deploying separate copies of identical applications to individual tenants Database as a Service typically self-service provisioning of databases on the private cloud Application Distribution from ISV / Installation by Customer Eliminating many typical installation steps (create schema, import seed data, import application code PL/SQL…) - just plug in a PDB! High volume data distribution literally via disk drives in envelopes distributed by truck! - distribution of things like GIS or MDM master databases …various others! Benefits Previous approaches to consolidation have involved a trade-off between reductions in Capital Expenses (CapEx) and Operating Expenses (OpEx), and they’ve usually come at the expense of agility. With Oracle Multitenant you can have your cake and eat it: Minimize CapEx More Applications per server Minimize OpEx Manage many as one Standardized procedures and services Rapid provisioning Maximize Agility Cloning for development and testing Portability through pluggability Scalability with RAC Ease of Adoption Applications run unchanged It’s a pure deployment choice. Neither the database backend nor the application needs to be changed. In future postings I’ll explore various aspects in more detail. However, if you feel compelled to devour everything you can about Oracle Multitenant this very minute, have no fear. Visit the Multitenant page on OTN and explore the various resources we have available there. Among these, Oracle Distinguished Product Manager Bryn Llewellyn has written an excellent, thorough, and exhaustively detailed White Paper about Oracle Multitenant, which is available here.  Follow me  I tweet @OraclePDB #OracleMultitenant

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  • StreamInsight 2.1, meet LINQ

    - by Roman Schindlauer
    Someone recently called LINQ “magic” in my hearing. I leapt to LINQ’s defense immediately. Turns out some people don’t realize “magic” is can be a pejorative term. I thought LINQ needed demystification. Here’s your best demystification resource: http://blogs.msdn.com/b/mattwar/archive/2008/11/18/linq-links.aspx. I won’t repeat much of what Matt Warren says in his excellent series, but will talk about some core ideas and how they affect the 2.1 release of StreamInsight. Let’s tell the story of a LINQ query. Compile time It begins with some code: IQueryable<Product> products = ...; var query = from p in products             where p.Name == "Widget"             select p.ProductID; foreach (int id in query) {     ... When the code is compiled, the C# compiler (among other things) de-sugars the query expression (see C# spec section 7.16): ... var query = products.Where(p => p.Name == "Widget").Select(p => p.ProductID); ... Overload resolution subsequently binds the Queryable.Where<Product> and Queryable.Select<Product, int> extension methods (see C# spec sections 7.5 and 7.6.5). After overload resolution, the compiler knows something interesting about the anonymous functions (lambda syntax) in the de-sugared code: they must be converted to expression trees, i.e.,“an object structure that represents the structure of the anonymous function itself” (see C# spec section 6.5). The conversion is equivalent to the following rewrite: ... var prm1 = Expression.Parameter(typeof(Product), "p"); var prm2 = Expression.Parameter(typeof(Product), "p"); var query = Queryable.Select<Product, int>(     Queryable.Where<Product>(         products,         Expression.Lambda<Func<Product, bool>>(Expression.Property(prm1, "Name"), prm1)),         Expression.Lambda<Func<Product, int>>(Expression.Property(prm2, "ProductID"), prm2)); ... If the “products” expression had type IEnumerable<Product>, the compiler would have chosen the Enumerable.Where and Enumerable.Select extension methods instead, in which case the anonymous functions would have been converted to delegates. At this point, we’ve reduced the LINQ query to familiar code that will compile in C# 2.0. (Note that I’m using C# snippets to illustrate transformations that occur in the compiler, not to suggest a viable compiler design!) Runtime When the above program is executed, the Queryable.Where method is invoked. It takes two arguments. The first is an IQueryable<> instance that exposes an Expression property and a Provider property. The second is an expression tree. The Queryable.Where method implementation looks something like this: public static IQueryable<T> Where<T>(this IQueryable<T> source, Expression<Func<T, bool>> predicate) {     return source.Provider.CreateQuery<T>(     Expression.Call(this method, source.Expression, Expression.Quote(predicate))); } Notice that the method is really just composing a new expression tree that calls itself with arguments derived from the source and predicate arguments. Also notice that the query object returned from the method is associated with the same provider as the source query. By invoking operator methods, we’re constructing an expression tree that describes a query. Interestingly, the compiler and operator methods are colluding to construct a query expression tree. The important takeaway is that expression trees are built in one of two ways: (1) by the compiler when it sees an anonymous function that needs to be converted to an expression tree, and; (2) by a query operator method that constructs a new queryable object with an expression tree rooted in a call to the operator method (self-referential). Next we hit the foreach block. At this point, the power of LINQ queries becomes apparent. The provider is able to determine how the query expression tree is evaluated! The code that began our story was intentionally vague about the definition of the “products” collection. Maybe it is a queryable in-memory collection of products: var products = new[]     { new Product { Name = "Widget", ProductID = 1 } }.AsQueryable(); The in-memory LINQ provider works by rewriting Queryable method calls to Enumerable method calls in the query expression tree. It then compiles the expression tree and evaluates it. It should be mentioned that the provider does not blindly rewrite all Queryable calls. It only rewrites a call when its arguments have been rewritten in a way that introduces a type mismatch, e.g. the first argument to Queryable.Where<Product> being rewritten as an expression of type IEnumerable<Product> from IQueryable<Product>. The type mismatch is triggered initially by a “leaf” expression like the one associated with the AsQueryable query: when the provider recognizes one of its own leaf expressions, it replaces the expression with the original IEnumerable<> constant expression. I like to think of this rewrite process as “type irritation” because the rewritten leaf expression is like a foreign body that triggers an immune response (further rewrites) in the tree. The technique ensures that only those portions of the expression tree constructed by a particular provider are rewritten by that provider: no type irritation, no rewrite. Let’s consider the behavior of an alternative LINQ provider. If “products” is a collection created by a LINQ to SQL provider: var products = new NorthwindDataContext().Products; the provider rewrites the expression tree as a SQL query that is then evaluated by your favorite RDBMS. The predicate may ultimately be evaluated using an index! In this example, the expression associated with the Products property is the “leaf” expression. StreamInsight 2.1 For the in-memory LINQ to Objects provider, a leaf is an in-memory collection. For LINQ to SQL, a leaf is a table or view. When defining a “process” in StreamInsight 2.1, what is a leaf? To StreamInsight a leaf is logic: an adapter, a sequence, or even a query targeting an entirely different LINQ provider! How do we represent the logic? Remember that a standing query may outlive the client that provisioned it. A reference to a sequence object in the client application is therefore not terribly useful. But if we instead represent the code constructing the sequence as an expression, we can host the sequence in the server: using (var server = Server.Connect(...)) {     var app = server.Applications["my application"];     var source = app.DefineObservable(() => Observable.Range(0, 10, Scheduler.NewThread));     var query = from i in source where i % 2 == 0 select i; } Example 1: defining a source and composing a query Let’s look in more detail at what’s happening in example 1. We first connect to the remote server and retrieve an existing app. Next, we define a simple Reactive sequence using the Observable.Range method. Notice that the call to the Range method is in the body of an anonymous function. This is important because it means the source sequence definition is in the form of an expression, rather than simply an opaque reference to an IObservable<int> object. The variation in Example 2 fails. Although it looks similar, the sequence is now a reference to an in-memory observable collection: var local = Observable.Range(0, 10, Scheduler.NewThread); var source = app.DefineObservable(() => local); // can’t serialize ‘local’! Example 2: error referencing unserializable local object The Define* methods support definitions of operator tree leaves that target the StreamInsight server. These methods all have the same basic structure. The definition argument is a lambda expression taking between 0 and 16 arguments and returning a source or sink. The method returns a proxy for the source or sink that can then be used for the usual style of LINQ query composition. The “define” methods exploit the compile-time C# feature that converts anonymous functions into translatable expression trees! Query composition exploits the runtime pattern that allows expression trees to be constructed by operators taking queryable and expression (Expression<>) arguments. The practical upshot: once you’ve Defined a source, you can compose LINQ queries in the familiar way using query expressions and operator combinators. Notably, queries can be composed using pull-sequences (LINQ to Objects IQueryable<> inputs), push sequences (Reactive IQbservable<> inputs), and temporal sequences (StreamInsight IQStreamable<> inputs). You can even construct processes that span these three domains using “bridge” method overloads (ToEnumerable, ToObservable and To*Streamable). Finally, the targeted rewrite via type irritation pattern is used to ensure that StreamInsight computations can leverage other LINQ providers as well. Consider the following example (this example depends on Interactive Extensions): var source = app.DefineEnumerable((int id) =>     EnumerableEx.Using(() =>         new NorthwindDataContext(), context =>             from p in context.Products             where p.ProductID == id             select p.ProductName)); Within the definition, StreamInsight has no reason to suspect that it ‘owns’ the Queryable.Where and Queryable.Select calls, and it can therefore defer to LINQ to SQL! Let’s use this source in the context of a StreamInsight process: var sink = app.DefineObserver(() => Observer.Create<string>(Console.WriteLine)); var query = from name in source(1).ToObservable()             where name == "Widget"             select name; using (query.Bind(sink).Run("process")) {     ... } When we run the binding, the source portion which filters on product ID and projects the product name is evaluated by SQL Server. Outside of the definition, responsibility for evaluation shifts to the StreamInsight server where we create a bridge to the Reactive Framework (using ToObservable) and evaluate an additional predicate. It’s incredibly easy to define computations that span multiple domains using these new features in StreamInsight 2.1! Regards, The StreamInsight Team

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  • Unable to use factory girl with Cucumber and rails 3 (bundler problem)

    - by jbpros
    Hi there, I'm trying to run cucumber features with factory girl factories on a fresh Rails 3 application. Here is my Gemfile: source "http://gemcutter.org" gem "rails", "3.0.0.beta" gem "pg" gem "factory_girl", :git => "git://github.com/thoughtbot/factory_girl.git", :branch => "rails3" gem "rspec-rails", ">= 2.0.0.beta.4" gem "capybara" gem "database_cleaner" gem "cucumber-rails", :require => false Then the bundle install commande just runs smoothly: $ bundle install /usr/lib/ruby/gems/1.8/gems/bundler-0.9.3/lib/bundler/installer.rb:81:Warning: Gem::Dependency#version_requirements is deprecated and will be removed on or after August 2010. Use #requirement Updating git://github.com/thoughtbot/factory_girl.git Fetching source index from http://gemcutter.org Resolving dependencies Installing abstract (1.0.0) from system gems Installing actionmailer (3.0.0.beta) from system gems Installing actionpack (3.0.0.beta) from system gems Installing activemodel (3.0.0.beta) from system gems Installing activerecord (3.0.0.beta) from system gems Installing activeresource (3.0.0.beta) from system gems Installing activesupport (3.0.0.beta) from system gems Installing arel (0.2.1) from system gems Installing builder (2.1.2) from system gems Installing bundler (0.9.13) from system gems Installing capybara (0.3.6) from system gems Installing cucumber (0.6.3) from system gems Installing cucumber-rails (0.3.0) from system gems Installing culerity (0.2.9) from system gems Installing database_cleaner (0.5.0) from system gems Installing diff-lcs (1.1.2) from system gems Installing erubis (2.6.5) from system gems Installing factory_girl (1.2.3) from git://github.com/thoughtbot/factory_girl.git (at rails3) Installing ffi (0.6.3) from system gems Installing i18n (0.3.6) from system gems Installing json_pure (1.2.3) from system gems Installing mail (2.1.3) from system gems Installing memcache-client (1.7.8) from system gems Installing mime-types (1.16) from system gems Installing nokogiri (1.4.1) from system gems Installing pg (0.9.0) from system gems Installing polyglot (0.3.0) from system gems Installing rack (1.1.0) from system gems Installing rack-mount (0.4.7) from system gems Installing rack-test (0.5.3) from system gems Installing rails (3.0.0.beta) from system gems Installing railties (3.0.0.beta) from system gems Installing rake (0.8.7) from system gems Installing rspec (2.0.0.beta.4) from system gems Installing rspec-core (2.0.0.beta.4) from system gems Installing rspec-expectations (2.0.0.beta.4) from system gems Installing rspec-mocks (2.0.0.beta.4) from system gems Installing rspec-rails (2.0.0.beta.4) from system gems Installing selenium-webdriver (0.0.17) from system gems Installing term-ansicolor (1.0.5) from system gems Installing text-format (1.0.0) from system gems Installing text-hyphen (1.0.0) from system gems Installing thor (0.13.4) from system gems Installing treetop (1.4.4) from system gems Installing tzinfo (0.3.17) from system gems Installing webrat (0.7.0) from system gems Your bundle is complete! When I run cucumber, here is the error I get: $ rake cucumber (in /home/jbpros/projects/deorbitburn) /usr/lib/ruby/gems/1.8/gems/bundler-0.9.3/lib/bundler/resolver.rb:97:Warning: Gem::Dependency#version_requirements is deprecated and will be removed on or after August 2010. Use #requirement NOTICE: CREATE TABLE will create implicit sequence "posts_id_seq" for serial column "posts.id" NOTICE: CREATE TABLE / PRIMARY KEY will create implicit index "posts_pkey" for table "posts" /usr/bin/ruby1.8 -I "/usr/lib/ruby/gems/1.8/gems/cucumber-0.6.3/lib:lib" "/usr/lib/ruby/gems/1.8/gems/cucumber-0.6.3/bin/cucumber" --profile default Using the default profile... git://github.com/thoughtbot/factory_girl.git (at rails3) is not checked out. Please run `bundle install` (Bundler::PathError) /home/jbpros/.bundle/gems/bundler-0.9.13/lib/bundler/source.rb:282:in `load_spec_files' /home/jbpros/.bundle/gems/bundler-0.9.13/lib/bundler/source.rb:190:in `local_specs' /home/jbpros/.bundle/gems/bundler-0.9.13/lib/bundler/environment.rb:36:in `runtime_gems' /home/jbpros/.bundle/gems/bundler-0.9.13/lib/bundler/environment.rb:35:in `each' /home/jbpros/.bundle/gems/bundler-0.9.13/lib/bundler/environment.rb:35:in `runtime_gems' /home/jbpros/.bundle/gems/bundler-0.9.13/lib/bundler/index.rb:5:in `build' /home/jbpros/.bundle/gems/bundler-0.9.13/lib/bundler/environment.rb:34:in `runtime_gems' /home/jbpros/.bundle/gems/bundler-0.9.13/lib/bundler/environment.rb:14:in `index' /home/jbpros/.bundle/gems/bundler-0.9.13/lib/bundler/index.rb:5:in `build' /home/jbpros/.bundle/gems/bundler-0.9.13/lib/bundler/environment.rb:13:in `index' /home/jbpros/.bundle/gems/bundler-0.9.13/lib/bundler/environment.rb:55:in `resolve_locally' /home/jbpros/.bundle/gems/bundler-0.9.13/lib/bundler/environment.rb:28:in `specs' /home/jbpros/.bundle/gems/bundler-0.9.13/lib/bundler/environment.rb:65:in `specs_for' /home/jbpros/.bundle/gems/bundler-0.9.13/lib/bundler/environment.rb:23:in `requested_specs' /home/jbpros/.bundle/gems/bundler-0.9.13/lib/bundler/runtime.rb:18:in `setup' /home/jbpros/.bundle/gems/bundler-0.9.13/lib/bundler.rb:68:in `setup' /home/jbpros/projects/deorbitburn/config/boot.rb:7 /usr/local/lib/site_ruby/1.8/rubygems/custom_require.rb:31:in `gem_original_require' /usr/local/lib/site_ruby/1.8/rubygems/custom_require.rb:31:in `polyglot_original_require' /usr/lib/ruby/gems/1.8/gems/polyglot-0.3.0/lib/polyglot.rb:65:in `require' /home/jbpros/projects/deorbitburn/config/application.rb:1 /usr/local/lib/site_ruby/1.8/rubygems/custom_require.rb:31:in `gem_original_require' /usr/local/lib/site_ruby/1.8/rubygems/custom_require.rb:31:in `polyglot_original_require' /usr/lib/ruby/gems/1.8/gems/polyglot-0.3.0/lib/polyglot.rb:65:in `require' /home/jbpros/projects/deorbitburn/config/environment.rb:2 /usr/local/lib/site_ruby/1.8/rubygems/custom_require.rb:31:in `gem_original_require' /usr/local/lib/site_ruby/1.8/rubygems/custom_require.rb:31:in `polyglot_original_require' /usr/lib/ruby/gems/1.8/gems/polyglot-0.3.0/lib/polyglot.rb:65:in `require' /home/jbpros/projects/deorbitburn/features/support/env.rb:8 /usr/local/lib/site_ruby/1.8/rubygems/custom_require.rb:31:in `gem_original_require' /usr/local/lib/site_ruby/1.8/rubygems/custom_require.rb:31:in `polyglot_original_require' /usr/lib/ruby/gems/1.8/gems/polyglot-0.3.0/lib/polyglot.rb:65:in `require' /usr/lib/ruby/gems/1.8/gems/cucumber-0.6.3/bin/../lib/cucumber/rb_support/rb_language.rb:124:in `load_code_file' /usr/lib/ruby/gems/1.8/gems/cucumber-0.6.3/bin/../lib/cucumber/step_mother.rb:85:in `load_code_file' /usr/lib/ruby/gems/1.8/gems/cucumber-0.6.3/bin/../lib/cucumber/step_mother.rb:77:in `load_code_files' /usr/lib/ruby/gems/1.8/gems/cucumber-0.6.3/bin/../lib/cucumber/step_mother.rb:76:in `each' /usr/lib/ruby/gems/1.8/gems/cucumber-0.6.3/bin/../lib/cucumber/step_mother.rb:76:in `load_code_files' /usr/lib/ruby/gems/1.8/gems/cucumber-0.6.3/bin/../lib/cucumber/cli/main.rb:48:in `execute!' /usr/lib/ruby/gems/1.8/gems/cucumber-0.6.3/bin/../lib/cucumber/cli/main.rb:20:in `execute' /usr/lib/ruby/gems/1.8/gems/cucumber-0.6.3/bin/cucumber:8 rake aborted! Command failed with status (1): [/usr/bin/ruby1.8 -I "/usr/lib/ruby/gems/1....] (See full trace by running task with --trace) Do I have to do something special for bundler to check out factory girl's repository on github?

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  • Debugging cucumber/gem dependencies

    - by mobmad
    How do you debug and fix gem errors like below? Although the below case is very specific, I'm also looking for solution to related problems like "gem already activated [...]", and resources to gem management/debugging. mycomputer:projectfolder username$ cucumber features Using the default profile... WARNING: No DRb server is running. Running features locally: /Users/username/.gem/ruby/1.8/gems/rails-2.3.5/lib/rails/gem_dependency.rb:119:Warning: Gem::Dependency#version_requirements is deprecated and will be removed on or after August 2010. Use #requirement can't activate , already activated ruby-hmac-0.4.0 (Gem::Exception) /Users/username/.gem/ruby/1.8/gems/rails-2.3.5/lib/rails/gem_dependency.rb:101:in `specification' /Users/username/.gem/ruby/1.8/gems/rails-2.3.5/lib/rails/plugin/locator.rb:81:in `plugins' /Library/Ruby/Site/1.8/rubygems/custom_require.rb:31:in `inject' /Users/username/.gem/ruby/1.8/gems/rails-2.3.5/lib/rails/plugin/locator.rb:81:in `each' /Users/username/.gem/ruby/1.8/gems/rails-2.3.5/lib/rails/plugin/locator.rb:81:in `inject' /Users/username/.gem/ruby/1.8/gems/rails-2.3.5/lib/rails/plugin/locator.rb:81:in `plugins' /Users/username/.gem/ruby/1.8/gems/rails-2.3.5/lib/rails/plugin/loader.rb:109:in `locate_plugins' /Users/username/.gem/ruby/1.8/gems/rails-2.3.5/lib/rails/plugin/loader.rb:108:in `map' /Users/username/.gem/ruby/1.8/gems/rails-2.3.5/lib/rails/plugin/loader.rb:108:in `locate_plugins' /Users/username/.gem/ruby/1.8/gems/rails-2.3.5/lib/rails/plugin/loader.rb:32:in `all_plugins' /Users/username/.gem/ruby/1.8/gems/rails-2.3.5/lib/rails/plugin/loader.rb:22:in `plugins' /Users/username/.gem/ruby/1.8/gems/rails-2.3.5/lib/rails/plugin/loader.rb:53:in `add_plugin_load_paths' /Users/username/.gem/ruby/1.8/gems/rails-2.3.5/lib/initializer.rb:294:in `add_plugin_load_paths' /Users/username/.gem/ruby/1.8/gems/rails-2.3.5/lib/initializer.rb:136:in `process' /Users/username/.gem/ruby/1.8/gems/rails-2.3.5/lib/initializer.rb:113:in `send' /Users/username/.gem/ruby/1.8/gems/rails-2.3.5/lib/initializer.rb:113:in `run' /Users/username/Documents/projectfolder.0/sites/projectfolder/config/environment.rb:9 /Library/Ruby/Site/1.8/rubygems/custom_require.rb:31:in `gem_original_require' /Library/Ruby/Site/1.8/rubygems/custom_require.rb:31:in `polyglot_original_require' /Library/Ruby/Gems/1.8/gems/polyglot-0.2.9/lib/polyglot.rb:70:in `require' ./features/support/env.rb:12 /Library/Ruby/Gems/1.8/gems/spork-0.7.5/lib/spork.rb:23:in `prefork' ./features/support/env.rb:9 /Library/Ruby/Site/1.8/rubygems/custom_require.rb:31:in `gem_original_require' /Library/Ruby/Site/1.8/rubygems/custom_require.rb:31:in `polyglot_original_require' /Library/Ruby/Gems/1.8/gems/polyglot-0.2.9/lib/polyglot.rb:70:in `require' /Library/Ruby/Gems/1.8/gems/cucumber-0.4.4/bin/../lib/cucumber/rb_support/rb_language.rb:124:in `load_code_file' /Library/Ruby/Gems/1.8/gems/cucumber-0.4.4/bin/../lib/cucumber/step_mother.rb:84:in `load_code_file' /Library/Ruby/Gems/1.8/gems/cucumber-0.4.4/bin/../lib/cucumber/step_mother.rb:76:in `load_code_files' /Library/Ruby/Gems/1.8/gems/cucumber-0.4.4/bin/../lib/cucumber/step_mother.rb:75:in `each' /Library/Ruby/Gems/1.8/gems/cucumber-0.4.4/bin/../lib/cucumber/step_mother.rb:75:in `load_code_files' /Library/Ruby/Gems/1.8/gems/cucumber-0.4.4/bin/../lib/cucumber/cli/main.rb:47:in `execute!' /Library/Ruby/Gems/1.8/gems/cucumber-0.4.4/bin/../lib/cucumber/cli/main.rb:24:in `execute' /Library/Ruby/Gems/1.8/gems/cucumber-0.4.4/bin/cucumber:8 /usr/bin/cucumber:19:in `load' /usr/bin/cucumber:19 And this is the output from gem list actionmailer (2.3.5, 2.2.2, 1.3.6) actionpack (2.3.5, 2.2.2, 1.13.6) actionwebservice (1.2.6) activerecord (2.3.5, 2.2.2, 1.15.6) activeresource (2.3.5, 2.2.2) activesupport (2.3.5, 2.2.2, 1.4.4) acts_as_ferret (0.4.4, 0.4.3) adamwiggins-rest-client (1.0.4) aslakhellesoy-webrat (0.4.4.1) aslakjo-comatose (2.0.5.12) authlogic (2.1.3) authlogic-oid (1.0.4) builder (2.1.2) capistrano (2.5.17, 2.5.2) cgi_multipart_eof_fix (2.5.0) configuration (1.1.0) cucumber (0.4.4) cucumber-rails (0.3.0) daemons (1.0.10) database_cleaner (0.5.0) diff-lcs (1.1.2) dnssd (1.3.1, 0.6.0) fakeweb (1.2.8) fastthread (1.0.7, 1.0.1) fcgi (0.8.8, 0.8.7) ferret (0.11.6) gem_plugin (0.2.3) gemcutter (0.4.1) heroku (1.8.0) highline (1.5.2, 1.5.0) hoe (2.5.0) hpricot (0.8.2, 0.6.164) json (1.2.2) json_pure (1.2.2) launchy (0.3.5) libxml-ruby (1.1.3, 1.1.2) linecache (0.43) log4r (1.1.5) mime-types (1.16) mongrel (1.1.5) mysql (2.8.1) needle (1.3.0) net-scp (1.0.2, 1.0.1) net-sftp (2.0.4, 2.0.1, 1.1.1) net-ssh (2.0.20, 2.0.4, 1.1.4) net-ssh-gateway (1.0.1, 1.0.0) nifty-generators (0.3.2) nokogiri (1.4.1) oauth (0.3.6) oniguruma (1.1.0) plist (3.1.0) polyglot (0.2.9) rack (1.1.0, 1.0.1) rack-test (0.5.3) rails (2.3.5, 2.2.2, 1.2.6) rake (0.8.7, 0.8.3) RedCloth (4.2.2, 4.1.1) rest-client (1.4.0) rspec (1.3.0) rspec-rails (1.3.2) ruby-activeldap (0.8.3.1) ruby-debug-base (0.10.3) ruby-debug-ide (0.4.9) ruby-hmac (0.4.0) ruby-net-ldap (0.0.4) ruby-openid (2.1.7, 2.1.2) ruby-yadis (0.3.4) rubyforge (2.0.4) rubygems-update (1.3.6) rubynode (0.1.5) rubyzip (0.9.4) sanitize (1.2.0) sequel (3.0.0) sinatra (0.9.2) spork (0.7.5) sqlite3-ruby (1.2.5, 1.2.4) taps (0.2.26) term-ansicolor (1.0.4) termios (0.9.4) textpow (0.10.1) thor (0.9.9) treetop (1.4.2) twitter4r (0.3.2, 0.3.1) ultraviolet (0.10.2) webrat (0.7.0) will_paginate (2.3.12) xmpp4r (0.5, 0.4)

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  • MySQL Binary Storage using BLOB VS OS File System: large files, large quantities, large problems.

    - by Quantico773
    Hi Guys, Versions I am running (basically latest of everything): PHP: 5.3.1 MySQL: 5.1.41 Apache: 2.2.14 OS: CentOS (latest) Here is the situation. I have thousands of very important documents, ranging from customer contracts to voice signatures (recordings of customer authorisation for contracts), with file types including, but not limited to jpg, gif, png, tiff, doc, docx, xls, wav, mp3, pdf, etc. All of these documents are currently stored on several servers including Windows 32 bit, CentOS and Mac, among others. Some files are also stored on employees desktop computers and laptops, and some are still hard copies stored in hundreds of boxes and filing cabinets. Now because customers or lawyers could demand evidence of contracts at any time, my company has to be able to search and locate the correct document(s) effectively, for this reason ALL of these files have to be digitised (if not already) and correlated into some sort of order for searching and accessing. As the programmer, I have created a full Customer Relations Management tool that the whole company uses. This includes Customer Profiles management, Order and job Tracking tools, Job/sale creation and management modules, etc, and at the moment any file that is needed at a customer profile level (drivers licence, credit authority, etc) or at a job/sale level (contracts, voice signatures, etc) can be uploaded to the server and sits in a parent/child hierarchy structure, just like Windows Explorer or any other typical file managment model. The structure appears as such: drivers_license |- DL_123.jpg voice_signatures |- VS_123.wav |- VS_4567.wav contracts So the files are uplaoded using PHP and Apache, and are stored in the file system of the OS. At the time of uploading, certain information about the file(s) is stored in a MySQL database. Some of the information stored is: TABLE: FileUploads FileID CustomerID (the customer id that the file belongs to, they all have this.) JobID/SaleID (the id of the job/sale associated, if any.) FileSize FileType UploadedDateTime UploadedBy FilePath (the directory path the file is stored in.) FileName (current file name of uploaded file, combination of CustomerID and JobID/SaleID if applicable.) FileDescription OriginalFileName (original name of the source file when uploaded, including extension.) So as you can see, the file is linked to the database by the File Name. When I want to provide a customers' files for download to a user all I have to do is "SELECT * FROM FileUploads WHERE CustomerID = 123 OR JobID = 2345;" and this will output all the file details I require, and with the FilePath and FileName I can provide the link for download. http... server / FilePath / FileName There are a number of problems with this method: Storing files in this "database unconcious" environment means data integrity is not kept. If a record is deleted, the file may not be deleted also, or vice versa. Files are strewn all over the place, different servers, computers, etc. The file name is the ONLY thing matching the binary to the database and customer profile and customer records. etc, etc. There are so many reasons, some of which are described here: http://www.dreamwerx.net/site/article01 . Also there is an interesting article here too: sietch.net/ViewNewsItem.aspx?NewsItemID=124 . SO, after much research I have pretty much decided I am going to store ALL of these files in the database, as a BLOB or LONGBLOB, but there are still many considerations before I do this. I know that storing them in the database is a viable option, however there are a number of methods of storing them. I also know storing them is one thing; correlating and accessing them in a manageable way is another thing entirely. The article provided at this link: dreamwerx.net/site/article01 describes a way of splitting the uploaded binary files into 64kb chunks and storing each chunk with the FileID, and then streaming the actual binary file to the client using headers. This is a really cool idea since it alleviates preassure on the servers memory; instead of loading an entire 100mb file into the RAM and then sending it to the client, it is doing it 64kb at a time. I have tried this (and updated his scripts) and this is totally successful, in a very small frame of testing. So if you are in agreeance that this method is a viable, stable and robust long-term option to store moderately large files (1kb to couple hundred megs), and large quantities of these files, let me know what other considerations or ideas you have. Also, I am considering getting a current "File Management" PHP script that gives an interface for managing files stored in the File System and converting it to manage files stored in the database. If there is already any software out there that does this, please let me know. I guess there are many questions I could ask, and all the information is up there ^^ so please, discuss all aspects of this and we can pass ideas back and forth and teach each other. Cheers, Quantico773

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  • IIS Strategies for Accessing Secured Network Resources

    - by ErikE
    Problem: A user connects to a service on a machine, such as an IIS web site or a SQL Server database. The site or the database need to gain access to network resources such as file shares (the most common) or a database on a different server. Permission is denied. This is because the user the service is running under doesn't have network permissions in the first place, or if it does, it doesn't have rights to access the remote resource. I keep running into this problem over and over again and am tired of not having a really solid way of handling it. Here are some workarounds I'm aware of: Run IIS as a custom-created domain user who is granted high permissions If permissions are granted one file share at a time, then every time I want to read from a new share, I would have to ask a network admin to add it for me. Eventually, with many web sites reading from many shares, it is going to get really complicated. If permissions are just opened up wide for the user to access any file shares in our domain, then this seems like an unnecessary security surface area to present. This also applies to all the sites running on IIS, rather than just the selected site or virtual directory that needs the access, a further surface area problem. Still use the IUSR account but give it network permissions and set up the same user name on the remote resource (not a domain user, a local user) This also has its problems. For example, there's a file share I am using that I have full rights to for sharing, but I can't log in to the machine. So I have to find the right admin and ask him to do it for me. Any time something has to change, it's another request to an admin. Allow IIS users to connect as anonymous, but set the account used for anonymous access to a high-privilege one This is even worse than giving the IIS IUSR full privileges, because it means my web site can't use any kind of security in the first place. Connect using Kerberos, then delegate This sounds good in principle but has all sorts of problems. First of all, if you're using virtual web sites where the domain name you connect to the site with is not the base machine name (as we do frequently), then you have to set up a Service Principal Name on the webserver using Microsoft's SetSPN utility. It's complicated and apparently prone to errors. Also, you have to ask your network/domain admin to change security policy for both the web server and the domain account so they are "trusted for delegation." If you don't get everything perfectly right, suddenly your intended Kerberos authentication is NTLM instead, and you can only impersonate rather than delegate, and thus no reaching out over the network as the user. Also, this method can be problematic because sometimes you need the web site or database to have permissions that the connecting user doesn't have. Create a service or COM+ application that fetches the resource for the web site Services and COM+ packages are run with their own set of credentials. Running as a high-privilege user is okay since they can do their own security and deny requests that are not legitimate, putting control in the hands of the application developer instead of the network admin. Problems: I am using a COM+ package that does exactly this on Windows Server 2000 to deliver highly sensitive images to a secured web application. I tried moving the web site to Windows Server 2003 and was suddenly denied permission to instantiate the COM+ object, very likely registry permissions. I trolled around quite a bit and did not solve the problem, partly because I was reluctant to give the IUSR account full registry permissions. That seems like the same bad practice as just running IIS as a high-privilege user. Note: This is actually really simple. In a programming language of your choice, you create a class with a function that returns an instance of the object you want (an ADODB.Connection, for example), and build a dll, which you register as a COM+ object. In your web server-side code, you create an instance of the class and use the function, and since it is running under a different security context, calls to network resources work. Map drive letters to shares This could theoretically work, but in my mind it's not really a good long-term strategy. Even though mappings can be created with specific credentials, and this can be done by others than a network admin, this also is going to mean that there are either way too many shared drives (small granularity) or too much permission is granted to entire file servers (large granularity). Also, I haven't figured out how to map a drive so that the IUSR gets the drives. Mapping a drive is for the current user, I don't know the IUSR account password to log in as it and create the mappings. Move the resources local to the web server/database There are times when I've done this, especially with Access databases. Does the database have to live out on the file share? Sometimes, it was just easiest to move the database to the web server or to the SQL database server (so the linked server to it would work). But I don't think this is a great all-around solution, either. And it won't work when the resource is a service rather than a file. Move the service to the final web server/database I suppose I could run a web server on my SQL Server database, so the web site can connect to it using impersonation and make me happy. But do we really want random extra web servers on our database servers just so this is possible? No. Virtual directories in IIS I know that virtual directories can help make remote resources look as though they are local, and this supports using custom credentials for each virtual directory. I haven't been able to come up with, yet, how this would solve the problem for system calls. Users could reach file shares directly, but this won't help, say, classic ASP code access resources. I could use a URL instead of a file path to read remote data files in a web page, but this isn't going to help me make a connection to an Access database, a SQL server database, or any other resource that uses a connection library rather than being able to just read all the bytes and work with them. I wish there was some kind of "service tunnel" that I could create. Think about how a VPN makes remote resources look like they are local. With a richer aliasing mechanism, perhaps code-based, why couldn't even database connections occur under a defined security context? Why not a special Windows component that lets you specify, per user, what resources are available and what alternate credentials are used for the connection? File shares, databases, web sites, you name it. I guess I'm almost talking about a specialized local proxy server. Anyway, so there's my list. I may update it if I think of more. Does anyone have any ideas for me? My current problem today is, yet again, I need a web site to connect to an Access database on a file share. Here we go again...

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  • What strategy do you use for package naming in Java projects and why?

    - by Tim Visher
    I thought about this awhile ago and it recently resurfaced as my shop is doing its first real Java web app. As an intro, I see two main package naming strategies. (To be clear, I'm not referring to the whole 'domain.company.project' part of this, I'm talking about the package convention beneath that.) Anyway, the package naming conventions that I see are as follows: Functional: Naming your packages according to their function architecturally rather than their identity according to the business domain. Another term for this might be naming according to 'layer'. So, you'd have a *.ui package and a *.domain package and a *.orm package. Your packages are horizontal slices rather than vertical. This is much more common than logical naming. In fact, I don't believe I've ever seen or heard of a project that does this. This of course makes me leery (sort of like thinking that you've come up with a solution to an NP problem) as I'm not terribly smart and I assume everyone must have great reasons for doing it the way they do. On the other hand, I'm not opposed to people just missing the elephant in the room and I've never heard a an actual argument for doing package naming this way. It just seems to be the de facto standard. Logical: Naming your packages according to their business domain identity and putting every class that has to do with that vertical slice of functionality into that package. I have never seen or heard of this, as I mentioned before, but it makes a ton of sense to me. I tend to approach systems vertically rather than horizontally. I want to go in and develop the Order Processing system, not the data access layer. Obviously, there's a good chance that I'll touch the data access layer in the development of that system, but the point is that I don't think of it that way. What this means, of course, is that when I receive a change order or want to implement some new feature, it'd be nice to not have to go fishing around in a bunch of packages in order to find all the related classes. Instead, I just look in the X package because what I'm doing has to do with X. From a development standpoint, I see it as a major win to have your packages document your business domain rather than your architecture. I feel like the domain is almost always the part of the system that's harder to grok where as the system's architecture, especially at this point, is almost becoming mundane in its implementation. The fact that I can come to a system with this type of naming convention and instantly from the naming of the packages know that it deals with orders, customers, enterprises, products, etc. seems pretty darn handy. It seems like this would allow you to take much better advantage of Java's access modifiers. This allows you to much more cleanly define interfaces into subsystems rather than into layers of the system. So if you have an orders subsystem that you want to be transparently persistent, you could in theory just never let anything else know that it's persistent by not having to create public interfaces to its persistence classes in the dao layer and instead packaging the dao class in with only the classes it deals with. Obviously, if you wanted to expose this functionality, you could provide an interface for it or make it public. It just seems like you lose a lot of this by having a vertical slice of your system's features split across multiple packages. I suppose one disadvantage that I can see is that it does make ripping out layers a little bit more difficult. Instead of just deleting or renaming a package and then dropping a new one in place with an alternate technology, you have to go in and change all of the classes in all of the packages. However, I don't see this is a big deal. It may be from a lack of experience, but I have to imagine that the amount of times you swap out technologies pales in comparison to the amount of times you go in and edit vertical feature slices within your system. So I guess the question then would go out to you, how do you name your packages and why? Please understand that I don't necessarily think that I've stumbled onto the golden goose or something here. I'm pretty new to all this with mostly academic experience. However, I can't spot the holes in my reasoning so I'm hoping you all can so that I can move on. Thanks in advance!

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  • Drupal: Create custom search

    - by Dr. Hfuhruhurr
    I'm trying to create a custom search but getting stuck. What I want is to have a dropdownbox so the user can choose where to search in. These options can mean 1 or more content types. So if he chooses options A, then the search will look in node-type P,Q,R. But he may not give those results, but only the uid's which will be then themed to gather specific data for that user. To make it a little bit clearer, Suppose I want to llok for people. The what I'm searching in is 2 content profile types. But ofcourse you dont want to display those as a result, but a nice picture of the user and some data. I started with creating a form with a textfield and the dropdown box. Then, in the submit handler, i created the keys and redirected to another pages with those keys as a tail. This page has been defined in the menu hook, just like how search does it. After that I want to call hook_view to do the actual search by calling node_search, and give back the results. Unfortunately, it goes wrong. When i click the Search button, it gives me a 404. But am I on the right track? Is this the way to create a custom search? Thx for your help. Here's the code for some clarity: <?php // $Id$ /* * @file * Searches on Project, Person, Portfolio or Group. */ /** * returns an array of menu items * @return array of menu items */ function vm_search_menu() { $subjects = _vm_search_get_subjects(); foreach ($subjects as $name => $description) { $items['zoek/'. $name .'/%menu_tail'] = array( 'page callback' => 'vm_search_view', 'page arguments' => array($name), 'type' => MENU_LOCAL_TASK, ); } return $items; } /** * create a block to put the form into. * @param $op * @param $delta * @param $edit * @return mixed */ function vm_search_block($op = 'list', $delta = 0, $edit = array()) { switch ($op) { case 'list': $blocks[0]['info'] = t('Algemene zoek'); return $blocks; case 'view': if (0 == $delta) { $block['subject'] = t(''); $block['content'] = drupal_get_form('vm_search_general_form'); } return $block; } } /** * Define the form. */ function vm_search_general_form() { $subjects = _vm_search_get_subjects(); foreach ($subjects as $key => $subject) { $options[$key] = $subject['desc']; } $form['subjects'] = array( '#type' => 'select', '#options' => $options, '#required' => TRUE, ); $form['keys'] = array( '#type' => 'textfield', '#required' => TRUE, ); $form['submit'] = array( '#type' => 'submit', '#value' => t('Zoek'), ); return $form; } function vm_search_general_form_submit($form, &$form_state) { $subjects = _vm_search_get_subjects(); $keys = $form_state['values']['keys']; //the search keys //the content types to search in $keys .= ' type:' . implode(',', $subjects[$form_state['values']['subjects']]['types']); //redirect to the page, where vm_search_view will handle the actual search $form_state['redirect'] = 'zoek/'. $form_state['values']['subjects'] .'/'. $keys; } /** * Menu callback; presents the search results. */ function vm_search_view($type = 'node') { // Search form submits with POST but redirects to GET. This way we can keep // the search query URL clean as a whistle: // search/type/keyword+keyword if (!isset($_POST['form_id'])) { if ($type == '') { // Note: search/node can not be a default tab because it would take on the // path of its parent (search). It would prevent remembering keywords when // switching tabs. This is why we drupal_goto to it from the parent instead. drupal_goto($front_page); } $keys = search_get_keys(); // Only perform search if there is non-whitespace search term: $results = ''; if (trim($keys)) { // Log the search keys: watchdog('vm_search', '%keys (@type).', array('%keys' => $keys, '@type' => $type)); // Collect the search results: $results = node_search('search', $type); if ($results) { $results = theme('box', t('Zoek resultaten'), $results); } else { $results = theme('box', t('Je zoek heeft geen resultaten opgeleverd.')); } } } return $results; } /** * returns array where to look for * @return array */ function _vm_search_get_subjects() { $subjects['opdracht'] = array('desc' => t('Opdracht'), 'types' => array('project') ); $subjects['persoon'] = array('desc' => t('Persoon'), 'types' => array('types_specialisatie', 'smaak_en_interesses') ); $subjects['groep'] = array('desc' => t('Groep'), 'types' => array('Villamedia_groep') ); $subjects['portfolio'] = array('desc' => t('Portfolio'), 'types' => array('artikel') ); return $subjects; }

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