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  • Cannot get official CentOS 5.4 BIND package to start

    - by Brian Cline
    Yesterday I installed CentOS 5.4 on one of my servers, and it appears that the official BIND/named package has trouble starting for reasons I cannot deduce. Here is what happens: [root@hal init.d]# service named start Starting named: Error in named configuration: /etc/named.conf:57: open: named.root.hints: permission denied [FAILED] The line in question, with the directory option for context: // further up in the file: directory "/var/named"; // line 57: include "named.root.hints"; Like you, my first reaction was to check permissions on /var/named/named.root.hints, /var/named, and /var to make sure the named user would be able to read it. Here are the permissions at each level: drwxr-xr-x 19 root root 4096 Nov 3 02:05 var drwxr-x--- 5 root named 4096 Nov 3 02:36 named -rw-r--r-- 1 named named 524 Mar 29 2006 named.root.hints Everything appears to be fine permission-wise. The same error occurs if the /var/named directory is writable by the named user. I've even temporarily allowed the named user to log in via bash, su'ed from root to named, and checked that I was, in fact, able to cat /var/named/named.root.hints successfully. (Yes, don't worry: I changed the shell back to nologin). My last endeavor showed that BIND is able to run under the named user account and start up just fine, if done so manually: [root@hal ~]# named -u named -g 03-Nov-2009 16:31:02.021 starting BIND 9.3.6-P1-RedHat-9.3.6-4.P1.el5 -u named -g 03-Nov-2009 16:31:02.021 adjusted limit on open files from 1024 to 1048576 03-Nov-2009 16:31:02.021 found 2 CPUs, using 2 worker threads 03-Nov-2009 16:31:02.021 using up to 4096 sockets 03-Nov-2009 16:31:02.028 loading configuration from '/etc/named.conf' 03-Nov-2009 16:31:02.030 using default UDP/IPv4 port range: [1024, 65535] 03-Nov-2009 16:31:02.031 using default UDP/IPv6 port range: [1024, 65535] 03-Nov-2009 16:31:02.034 listening on IPv4 interface lo, 127.0.0.1#53 03-Nov-2009 16:31:02.034 listening on IPv4 interface eth0, 10.0.0.5#53 03-Nov-2009 16:31:02.034 listening on IPv4 interface eth1, ww.xx.yy.zz#53 03-Nov-2009 16:31:02.040 command channel listening on 127.0.0.1#953 03-Nov-2009 16:31:02.040 command channel listening on ::1#953 03-Nov-2009 16:31:02.040 ignoring config file logging statement due to -g option 03-Nov-2009 16:31:02.041 zone 0.in-addr.arpa/IN/localhost_resolver: loaded serial 42 03-Nov-2009 16:31:02.042 zone 0.0.127.in-addr.arpa/IN/localhost_resolver: loaded serial 1997022700 03-Nov-2009 16:31:02.042 zone 255.in-addr.arpa/IN/localhost_resolver: loaded serial 42 03-Nov-2009 16:31:02.042 zone 0.0.0.0.0.0.0.0.0.0.0.0.0.0.0.0.0.0.0.0.0.0.0.0.0.0.0.0.0.0.0.ip6.arpa/IN/localhost_resolver: loaded serial 1997022700 03-Nov-2009 16:31:02.043 zone localdomain/IN/localhost_resolver: loaded serial 42 03-Nov-2009 16:31:02.043 zone localhost/IN/localhost_resolver: loaded serial 42 03-Nov-2009 16:31:02.043 zone x.y.z.in-addr.arpa/IN/internal: loaded serial 1 03-Nov-2009 16:31:02.044 zone x.y.z/IN/internal: loaded serial 2 03-Nov-2009 16:31:02.045 running What type and size of firearm should I use to resolve this? I'd prefer something with automatic ammunition, and, at worst, it should be able to fit on my shoulder. Of course I am open to suggestions.

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  • htaccess on remote server issues - password prompt not accepting input

    - by pying saucepan
    EDIT: I will contact the university about my problem after labor day weekend, but I thought if someone knew a quick fix that I haven't tried, or if the problem has an obvious fix then I could hope to try my luck here, thanks! TLDR: Sorry its a long post, I thought I should be... thorough. I am having a common issue (found a dead thread through google with no solution to the same problem) with the prompt to enter in a username and password via htaccess rights, but this prompt will keep popping up asking for a username and password when trying to access my home directory on my university's server which has the .htaccess and .htpasswd files. It does not matter if I enter in correct or incorrect credentials, the prompt will keep asking me for input without displaying my home directory. Ever since I have included these ht files I have never once been able to get past the username/password no matter what I have tried, save for removing them from the directory I am trying to access (my top level directory that I own). This kind of served my original goal of making the top level directory inaccessible to casual users, but if I wanted to use this method on other places, I would want it to work as intended. And I also like it when computers do what I wish they would, so any help is appreciated. Some things I have tried: Changing the file/directory access rights: they told me to try these commands if people can't access my files cd ~/public_html find ./ -type d -exec chmod 755 {} \; find ./ -type f -exec chmod 644 {} \; enter in the single character name/pw at least twenty times in a row, no cheddar. so I changed directory with cd ~ in hopes that this would be my home directory, since my home directory contains the "public_html" directory, so logic tells me that the ~ tilde symbol is the top level directory that I have ownership of. Then I did those two commands to change the rights on the files inside, I am still having no luck. How I got to this point: I have been following the instructions given to me through my university's website for setting up my little directory. A link on how they describe how to password protect the home directory is given below: "Protect Web Directories" instructions I have everything in order except for one small detail that I feel probably does not matter. I am on windows and so I am using winSCP to remote control my allocated server space. The small detail is that as the instructions indicate (on step 3) that I should use the command htpasswd -c .htpasswd {username} where {username} is my folder that holds my allocated server space. But this command requires further input through the terminal, and unfortunately winSCP does not offer this kind of functionality. So I looked up some basic instructions on using htaccess and it is formatted correctly such that the .htaccess file appears as follows: AuthType Basic AuthName "Verify" AuthUserFile /correctpath/.htpasswd require valid-user and this file is in the root directory for my server space as well as the .htpasswd file which has only this data inside: username:password I know for sure that these two files must be formatted correctly, at least according to their tutorial, because before my path was incorrectly formatted via including some curly { braces } without knowing the correct way to do this at first. And the password prompt that shows up when accessing my directory responded by loading an error page indicating to contact OSU admin or something not important. But now that I have everything like it 'should' be. I know this because when I enter in my credentials "username and password" the prompt pops up for my username and password again and again whether or not I enter in correct information. The only exception is that if I click cancel it will direct me to a page saying that I need to enter in a username and password. Note that I am very inexperienced at server-related buisness, two days ago I couldn't have told you what a website actually consists of. So, if you use some technical jargon I may or may not need to look it up and get back to you before I actually understand what you mean, but I am a quick learner and it probably wont matter.

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  • apache webserver unresponsible with server-status showing all child processes waiting for connection

    - by Jeff
    My setup: i have 3 nearly identical webserver machines serving the same high loaded dynamic website with simple load balancing over dns. The service has been working for over two ears with the same apache config. apache2, php5, ubuntu 8.04 linux 2.6.24-29-server My problem: since about two weeks i'm experiencing problems with this config. Nearly every day i have one small moment about 5 minutes, in which the website is unreachable. I'm still able to login to the servers over ssh. If i run htop, i see the machine simply doing nothing. i have about 1000 apache processes running, but no cpu activity. i've used the apache mod_status to debug this situation. the process scoreboard looks like this: _C.___K_______________________R._______.__K_K____K___C_______.__ _______C__________.___________________________________.________C _.____K__________K___K_WK_____._K_____________________________._ W______K__________K________.____________________._______C_______ _C_.__K__K____.._.._____________________________________C_______ _R___________K___.______C________.C_________.______._____C______ ____________KKC____K_____K__WC_________________C_____.__.____.__ _____________________C_________K______.____C______._____________ _.___C____.___.___________________________.K______.____K________ W__.___________________C.__.____K________K_______R_._.__._______ __C__C_.__________C__C_______._____W______________C_.___C_______ ____.______C_____________C________.____C____________.________._K __.__________.K_____________K_________._____C____.K__________KW_ __K.W________R_________._______.___W___________.____.__K_____W__ W___.___..________W____K Scoreboard Key: "_" Waiting for Connection, "S" Starting up, "R" Reading Request, "W" Sending Reply, "K" Keepalive (read), "D" DNS Lookup, "C" Closing connection, "L" Logging, "G" Gracefully finishing, "I" Idle cleanup of worker, "." Open slot with no current process So the most of the processes are just waiting for connection. after about 5 minutes the situation will return to normal: i have lot least processes on every machine, the most workers have the "."-status (meaing they are open to process a request) and of course the website is reachable! so i'm trying to find something in the logs, but there is simply nothing... the apache access log is silent for about 4 minutes, the same is for the error log. i also can not figure out anything wrong in other system logs. the situation is the same on all 3 webservers (all of them have this load peak and unresposibility at the same time), so i do not thing this is hardware related. but i think, this might be related to some network (tcp) issue. any ideas? EDIT: some more information, that i have just discovered: it has just happened again. and i was able to verify that i'm also not able to connect locally when this problem occurs. i have made some connection statistics with the following command after it happend netstat -an|awk '/tcp/ {print $6}'|sort|uniq -c 109 CLOSE_WAIT 2652 ESTABLISHED 2 FIN_WAIT1 11 LAST_ACK 12 LISTEN 91 SYN_RECV 1 SYN_SENT 16 TIME_WAIT If i execute the same command some time later, i have something like this: 4 CLOSING 108 ESTABLISHED 18 FIN_WAIT1 182 FIN_WAIT2 37 LAST_ACK 12 LISTEN 50 SYN_RECV 11276 TIME_WAIT So in the normal situation i have only 100-200 open connections by clients beeing handled by apache in this moment. when i have this "crash", i have a lot more connections. what is the best way to analyse this? EDIT2: the important lines in apache2.conf are: KeepAlive On MaxKeepAliveRequests 20 KeepAliveTimeout 1 <IfModule mpm_prefork_module> ServerLimit 920 StartServers 30 MinSpareServers 80 MaxSpareServers 120 MaxClients 920 MaxRequestsPerChild 700 </IfModule> it is an apache2 prefork with php_mod. the server has 8GB ram and a 4gb swap partition.

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  • In what (small) ways can I modify Octave's compile options to enhance it without breaking it?

    - by irrational John
    If the title to this question seems a bit vague, I am sorry. But I wasn't sure how to distill what I am attempting to do into a single sentence. A few weeks back I learned that I could build and install recent releases of Octave on an Ubuntu 12.04 system by following the steps below. Install the tools needed to compile, link, and run octave. For Ubuntu the commands below have worked for me. sudo apt-get build-dep octave3.2 sudo apt-get install build-essential gnuplot gtk2-engines-pixbuf sudo apt-get install libfontconfig-dev bison Next, download the source code for an Octave release from the Gnu Project Archives for Octave and unpack the archive into a folder on your system. Use the commands below to build, check, and install octave. ./configure make make check sudo make install Unfortunately it turns out that the above builds an Octave that contains all the debugging symbol tables. The object files alone are huge taking up around 1.7 GB. The current Octave documentation suggests To compile without debugging symbols try the command make CFLAGS=-O CXXFLAGS=-O LDFLAGS= instead of just make. However, when I tried this it did not work. The -g option was still used for the compiles. For the heck of it I instead tried ./configure CFLAGS=-O CXXFLAGS=-O and this did work. (Instead of ~1.7GB the result of the build now takes up around 253MB). My questions are Is this actually the correct (recommended?) method to use to compile Octave without debugging symbols (i.e. without -g)? How would I compile Octave so it uses x86_64 rather than x86? Note: I am not asking how to compile Octave to use the (experimental) 64-bit integers for array dimensions. I just want to allow the compiler to use the extra registers and word sizes available when an app runs in 64-bit mode. Is a (more) complete list available for the directives used with the Octave Makefile? I have only seen make, make check, and make install documented. But apparently make distclean is also allowed. (It removes the compilation results so you can do a complete rebuild of everything.) I'm wondering what else might be available. FWIW, I have tried using ./configure CFLAGS="-O3 -mtune=core2 -m64" CXXFLAGS="-O3 -mtune=core2 -m64" and, surprisingly, it not only appeared to build, but also ran and passed the make check tests. The ./configure script even gave me the (deceptively?) reassuring message "Octave is now configured for x86_64-unknown-linux-gnu". But of course that's not the same thing as saying it actually "works". Is there a recommended way to enable Octave to run as an x86_64 app? I have also tried looking inside the Octave Makefile to see if I could decipher what command line directives it accepts. I got nowhere. I have not a single clue as to how that Makefile does whatever it is that it does.

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  • Oracle Virtual Server OEL vm fails to start - kernel panic on cpu identify

    - by Towndrunk
    I am in the process of following a guide to setup various oracle vm templates, so far I have installed OVS 2. 2 and got the OVM Manager working, imported the template for OEL5U5 and created a vm from it.. the problem comes when starting that vm. The log in the OVMM console shows the following; Update VM Status - Running Configure CPU Cap Set CPU Cap: failed:<Exception: failed:<Exception: ['xm', 'sched-credit', '-d', '32_EM11g_OVM', '-c', '0'] => Error: Domain '32_EM11g_OVM' does not exist. StackTrace: File "/opt/ovs-agent-2.3/OVSXXenVMConfig.py", line 2531, in xen_set_cpu_cap run_cmd(args=['xm', File "/opt/ovs-agent-2.3/OVSCommons.py", line 92, in run_cmd raise Exception('%s => %s' % (args, err)) The xend.log shows; [2012-11-12 16:42:01 7581] DEBUG (DevController:139) Waiting for devices vtpm [2012-11-12 16:42:01 7581] INFO (XendDomain:1180) Domain 32_EM11g_OVM (3) unpaused. [2012-11-12 16:42:03 7581] WARNING (XendDomainInfo:1907) Domain has crashed: name=32_EM11g_OVM id=3. [2012-11-12 16:42:03 7581] ERROR (XendDomainInfo:2041) VM 32_EM11g_OVM restarting too fast (Elapsed time: 11.377262 seconds). Refusing to restart to avoid loops .> [2012-11-12 16:42:03 7581] DEBUG (XendDomainInfo:2757) XendDomainInfo.destroy: domid=3 [2012-11-12 16:42:12 7581] DEBUG (XendDomainInfo:2230) Destroying device model [2012-11-12 16:42:12 7581] INFO (image:553) 32_EM11g_OVM device model terminated I have set_on_crash="preserve" in the vm.cfg and have then run xm create -c to get the console screen while booting and this is the log of what happens.. Started domain 32_EM11g_OVM (id=4) Bootdata ok (command line is ro root=LABEL=/ ) Linux version 2.6.18-194.0.0.0.3.el5xen ([email protected]) (gcc version 4.1.2 20080704 (Red Hat 4.1.2-48)) #1 SMP Mon Mar 29 18:27:00 EDT 2010 BIOS-provided physical RAM map: Xen: 0000000000000000 - 0000000180800000 (usable)> No mptable found. Built 1 zonelists. Total pages: 1574912 Kernel command line: ro root=LABEL=/ Initializing CPU#0 PID hash table entries: 4096 (order: 12, 32768 bytes) Xen reported: 1600.008 MHz processor. Console: colour dummy device 80x25 Dentry cache hash table entries: 1048576 (order: 11, 8388608 bytes) Inode-cache hash table entries: 524288 (order: 10, 4194304 bytes) Software IO TLB disabled Memory: 6155256k/6299648k available (2514k kernel code, 135548k reserved, 1394k data, 184k init) Calibrating delay using timer specific routine.. 4006.42 BogoMIPS (lpj=8012858) Security Framework v1.0.0 initialized SELinux: Initializing. selinux_register_security: Registering secondary module capability Capability LSM initialized as secondary Mount-cache hash table entries: 256 CPU: L1 I Cache: 64K (64 bytes/line), D cache 16K (64 bytes/line) CPU: L2 Cache: 2048K (64 bytes/line) general protection fault: 0000 [1] SMP last sysfs file: CPU 0 Modules linked in: Pid: 0, comm: swapper Not tainted 2.6.18-194.0.0.0.3.el5xen #1 RIP: e030:[ffffffff80271280] [ffffffff80271280] identify_cpu+0x210/0x494 RSP: e02b:ffffffff80643f70 EFLAGS: 00010212 RAX: 0040401000810008 RBX: 0000000000000000 RCX: 00000000c001001f RDX: 0000000000404010 RSI: 0000000000000001 RDI: 0000000000000005 RBP: ffffffff8063e980 R08: 0000000000000025 R09: ffff8800019d1000 R10: 0000000000000026 R11: ffff88000102c400 R12: 0000000000000000 R13: 0000000000000000 R14: 0000000000000000 R15: 0000000000000000 FS: 0000000000000000(0000) GS:ffffffff805d2000(0000) knlGS:0000000000000000 CS: e033 DS: 0000 ES: 0000 Process swapper (pid: 0, threadinfo ffffffff80642000, task ffffffff804f4b80) Stack: 0000000000000000 ffffffff802d09bb ffffffff804f4b80 0000000000000000 0000000021100800 0000000000000000 0000000000000000 ffffffff8064cb00 0000000000000000 0000000000000000 Call Trace: [ffffffff802d09bb] kmem_cache_zalloc+0x62/0x80 [ffffffff8064cb00] start_kernel+0x210/0x224 [ffffffff8064c1e5] _sinittext+0x1e5/0x1eb Code: 0f 30 b8 73 00 00 00 f0 0f ab 45 08 e9 f0 00 00 00 48 89 ef RIP [ffffffff80271280] identify_cpu+0x210/0x494 RSP ffffffff80643f70 0 Kernel panic - not syncing: Fatal exception clear as mud to me. are there any other logs that will help me? I have now deployed another vm from the same template and used the default vm settings rather than adding more memory etc - I get exactly the same error.

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  • Installing vim7 on Solaris Sparc 2.6 as non-root

    - by Tobbe
    I'm trying to install vim to $HOME/bin by compiling the sources. ./configure --prefix=$home/bin seems to work, but when running make I get: > make Starting make in the src directory. If there are problems, cd to the src directory and run make there cd src && make first gcc -c -I. -Iproto -DHAVE_CONFIG_H -DFEAT_GUI_GTK -I/usr/include/gtk-2.0 -I/usr/lib/gtk-2.0/include -I/usr/include/atk-1.0 -I/usr/include/pango-1.0 -I/usr/openwin/include -I/usr/sfw/include -I/usr/sfw/include/freetype2 -I/usr/include/glib-2.0 -I/usr/lib/glib-2.0/include -g -O2 -I/usr/openwin/include -o objects/buffer.o buffer.c In file included from buffer.c:28: vim.h:41: error: syntax error before ':' token In file included from os_unix.h:29, from vim.h:245, from buffer.c:28: /usr/include/sys/stat.h:251: error: syntax error before "blksize_t" /usr/include/sys/stat.h:255: error: syntax error before '}' token /usr/include/sys/stat.h:309: error: syntax error before "blksize_t" /usr/include/sys/stat.h:310: error: conflicting types for 'st_blocks' /usr/include/sys/stat.h:252: error: previous declaration of 'st_blocks' was here /usr/include/sys/stat.h:313: error: syntax error before '}' token In file included from /opt/local/bin/../lib/gcc/sparc-sun-solaris2.6/3.4.6/include/sys/signal.h:132, from /usr/include/signal.h:26, from os_unix.h:163, from vim.h:245, from buffer.c:28: /usr/include/sys/siginfo.h:259: error: syntax error before "ctid_t" /usr/include/sys/siginfo.h:292: error: syntax error before '}' token /usr/include/sys/siginfo.h:294: error: syntax error before '}' token /usr/include/sys/siginfo.h:390: error: syntax error before "ctid_t" /usr/include/sys/siginfo.h:398: error: conflicting types for '__fault' /usr/include/sys/siginfo.h:267: error: previous declaration of '__fault' was here /usr/include/sys/siginfo.h:404: error: conflicting types for '__file' /usr/include/sys/siginfo.h:273: error: previous declaration of '__file' was here /usr/include/sys/siginfo.h:420: error: conflicting types for '__prof' /usr/include/sys/siginfo.h:287: error: previous declaration of '__prof' was here /usr/include/sys/siginfo.h:424: error: conflicting types for '__rctl' /usr/include/sys/siginfo.h:291: error: previous declaration of '__rctl' was here /usr/include/sys/siginfo.h:426: error: syntax error before '}' token /usr/include/sys/siginfo.h:428: error: syntax error before '}' token /usr/include/sys/siginfo.h:432: error: syntax error before "k_siginfo_t" /usr/include/sys/siginfo.h:437: error: syntax error before '}' token In file included from /usr/include/signal.h:26, from os_unix.h:163, from vim.h:245, from buffer.c:28: /opt/local/bin/../lib/gcc/sparc-sun-solaris2.6/3.4.6/include/sys/signal.h:173: error: syntax error before "siginfo_t" In file included from os_unix.h:163, from vim.h:245, from buffer.c:28: /usr/include/signal.h:111: error: syntax error before "siginfo_t" /usr/include/signal.h:113: error: syntax error before "siginfo_t" buffer.c: In function `buflist_new': buffer.c:1502: error: storage size of 'st' isn't known buffer.c: In function `buflist_findname': buffer.c:1989: error: storage size of 'st' isn't known buffer.c: In function `setfname': buffer.c:2578: error: storage size of 'st' isn't known buffer.c: In function `otherfile_buf': buffer.c:2836: error: storage size of 'st' isn't known buffer.c: In function `buf_setino': buffer.c:2874: error: storage size of 'st' isn't known buffer.c: In function `buf_same_ino': buffer.c:2894: error: dereferencing pointer to incomplete type buffer.c:2895: error: dereferencing pointer to incomplete type *** Error code 1 make: Fatal error: Command failed for target `objects/buffer.o' Current working directory /home/xluntor/vim72/src *** Error code 1 make: Fatal error: Command failed for target `first' How do I fix the make errors? Or is there another way to install vim as non-root? Thanks in advance

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  • md/raid:md2: cannot start dirty degraded array, kernel panic

    - by nl-x
    After having made use of a remote power switch, my server did not come back online. When I went to the datacenter and reboot the computer on the spot I see the server booting (I see the centos progress bar with running almost all the way to the end) and eventually giving the following messages: md/raid:md2: cannot start dirty degraded array. md/raid:md2: failed to run raid set. md: pers->run() failed ... md/raid:md2: cannot start dirty degraded array. md/raid:md2: failed to run raid set. md: pers->run() failed ... Kernel panic - not syncing: Attempted to kill init! Pid: 1, comm: init not tainted 2.6.32-279.1.1.el6.i686 #1 Call Trace: [<c083bfbc>] ? panic+0x68/0x11c [<c045a501>] ? do_exit+0x741/0x750 [<c045a54c>] ? do_group_exit+0x3c/0xa0 [<c045a5c1>] ? sys_exit_group+0x11/0x20 [<c083eba4>] ? syscall_call+0x7/0xb [<c083007b>] ? cmos_wake_setup+0x62/0x112 The server runs CentOS and has software raid, and I don't have backups of the raid settings. The only backup I have is of /home and the database dumps. (Glad to at least have those though.) Since the server is an old Dell PowerEdge 1750 with no CD-ROM drive, I have no way of booting the machine from a boot disk. I also remember in the past that the server also wouldn't boot from a bootable USB disk. So the only way I know how to boot the server is to go to the datacenter, pick up the server and take it to the office. Screw open the server. Attach a cdrom drive to an IDE slot on the motherboard. And then boot it. I am hoping you guys could help me avoid this. I have looked a bit through the boot options and I found the following boot options. When CentOS is about to boot and interrupt the boot-countdown: CentOS (2.6.32-279.1.1.el63.i686) CentOS Linux (2.6.32-71.29.1.el6.i686) centos (2.6.32-71.el6.i686) I think the first configuration is the default one, because choosing that gets me to the above mentioned kernel panic. The other ones end with something like "Sleeping forever". I can press 'e' to edit boot commands, press 'a' to modify kernel arguments and press 'c' for grub command line. The command line gives a grub prompt. But I have no idea how to get the system to boot without (trying to) access the dirty partitions. What I want to do is off course: - boot the machine - check hard drive for errors - mark the drive as clean

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  • Bugzilla : No SASL mechanism found

    - by niteshsinha
    I am using Bugzilla on windows 7. I am using the unofficial Bugzilla installer. I followed the steps accordingly and gave valid credentials wherever required. I open Bugzilla and try to create a new account , but i get the following error. Software error: No SASL mechanism found at C:/Program Files/Bugzilla/perl/perl/site/lib/Authen/SASL.pm line 77 at C:/Program Files/Bugzilla/perl/perl/lib/Net/SMTP.pm line 143 i ran checksetup.pl and found that Authen::SASL and SMTP both are available on my machine. The output of checksetup.pl is as follows. * This is Bugzilla 3.6.3 on perl 5.10.1 * Running on Win7 Build 7600 Checking perl modules... Checking for CGI.pm (v3.33) ok: found v3.49 Checking for Digest-SHA (any) ok: found v5.48 Checking for TimeDate (v2.21) ok: found v2.24 Checking for DateTime (v0.28) ok: found v0.53 Checking for DateTime-TimeZone (v0.79) ok: found v1.10 Checking for DBI (v1.41) ok: found v1.609 Checking for Template-Toolkit (v2.22) ok: found v2.22 Checking for Email-Send (v2.16) ok: found v2.198 Checking for Email-MIME (v1.861) ok: found v1.903 Checking for Email-MIME-Encodings (v1.313) ok: found v1.313 Checking for Email-MIME-Modifier (v1.442) ok: found v1.903 Checking for URI (any) ok: found v1.52 Checking available perl DBD modules... Checking for DBD-Pg (v1.45) ok: found v2.16.1 Checking for DBD-mysql (v4.00) ok: found v4.012 Checking for DBD-Oracle (v1.19) not found The following Perl modules are optional: Checking for GD (v1.20) ok: found v2.44 Checking for Chart (v2.1) ok: found v2.4.1 Checking for Template-GD (any) ok: found v1.56 Checking for GDTextUtil (any) ok: found v0.86 Checking for GDGraph (any) ok: found v1.44 Checking for XML-Twig (any) ok: found v3.34 Checking for MIME-tools (v5.406) ok: found v5.427 Checking for libwww-perl (any) ok: found v5.834 Checking for PatchReader (v0.9.4) ok: found v0.9.5 Checking for perl-ldap (any) ok: found v0.39 Checking for Authen-SASL (any) ok: found v2.15 Checking for RadiusPerl (any) ok: found v0.17 Checking for SOAP-Lite (v0.710.06) ok: found v0.710.10 Checking for JSON-RPC (any) ok: found v0.95 Checking for Test-Taint (any) ok: found v1.04 Checking for HTML-Parser (v3.40) ok: found v3.64 Checking for HTML-Scrubber (any) ok: found v0.08 Checking for Email-MIME-Attachment-Stripper (any) ok: found v1.316 Checking for Email-Reply (any) ok: found v1.202 Checking for TheSchwartz (any) not found Checking for Daemon-Generic (any) not found Checking for mod_perl (v1.999022) not found *********************************************************************** * OPTIONAL MODULES * *********************************************************************** * Certain Perl modules are not required by Bugzilla, but by * * installing the latest version you gain access to additional * * features. * * * * The optional modules you do not have installed are listed below, * * with the name of the feature they enable. Below that table are the * * commands to install each module. * *********************************************************************** * MODULE NAME * ENABLES FEATURE(S) * *********************************************************************** * TheSchwartz * Mail Queueing * * Daemon-Generic * Mail Queueing * * mod_perl * mod_perl * *********************************************************************** * Note For Windows Users * *********************************************************************** * In order to install the modules listed below, you first have to run * * the following command as an Administrator: * * * * ppm repo add theory58S http://cpan.uwinnipeg.ca/PPMPackages/10xx/ * * * Then you have to do (also as an Administrator): * * * * ppm repo up theory58S * * * * Do that last command over and over until you see "theory58S" at the * * top of the displayed list. * *********************************************************************** COMMANDS TO INSTALL OPTIONAL MODULES: TheSchwartz: ppm install TheSchwartz Daemon-Generic: ppm install Daemon-Generic mod_perl: ppm install mod_perl Reading ./localconfig... Checking for DBD-mysql (v4.00) ok: found v4.012 Checking for MySQL (v4.1.2) ok: found v5.1.44-community-log Removing existing compiled templates... Precompiling templates...done. Now that you have installed Bugzilla, you should visit the 'Parameters' page (linked in the footer of the Administrator account) to ensure it is set up as you wish - this includes setting the 'urlbase' option to the correct URL. Press any key to continue . . . Please tell me what should i do. Please note: i am running behind a corporate proxy , SSL/TLS is not used internally but i am giving the smtpUser and smtpPass also.

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  • Can't re-mount existing RAID10 on Ubuntu

    - by Zoran
    I saw similar questions, but didn't find what solution to my problem. After power-cut, one of RAID10 (4 disks were) appears to be malfunctioning. I make tha array active one, but can not mount it. Always the same error: mount: you must specify the filesystem type So, here is what I have when type mdadm --detail /dev/md0 /dev/md0: Version : 00.90.03 Creation Time : Tue Sep 1 11:00:40 2009 Raid Level : raid10 Array Size : 1465148928 (1397.27 GiB 1500.31 GB) Used Dev Size : 732574464 (698.64 GiB 750.16 GB) Raid Devices : 4 Total Devices : 3 Preferred Minor : 0 Persistence : Superblock is persistent Update Time : Mon Jun 11 09:54:27 2012 State : clean, degraded Active Devices : 3 Working Devices : 3 Failed Devices : 0 Spare Devices : 0 Layout : near=2, far=1 Chunk Size : 64K UUID : 1a02e789:c34377a1:2e29483d:f114274d Events : 0.166 Number Major Minor RaidDevice State 0 8 16 0 active sync /dev/sdb 1 0 0 1 removed 2 8 48 2 active sync /dev/sdd 3 8 64 3 active sync /dev/sde At the /etc/mdadm/mdadm.conf I have by default, scan all partitions (/proc/partitions) for MD superblocks. alternatively, specify devices to scan, using wildcards if desired. DEVICE partitions auto-create devices with Debian standard permissions CREATE owner=root group=disk mode=0660 auto=yes automatically tag new arrays as belonging to the local system HOMEHOST <system> instruct the monitoring daemon where to send mail alerts MAILADDR root definitions of existing MD arrays ARRAY /dev/md0 level=raid10 num-devices=4 UUID=1a02e789:c34377a1:2e29483d:f114274d ARRAY /dev/md1 level=raid1 num-devices=2 UUID=9b592be7:c6a2052f:2e29483d:f114274d This file was auto-generated... So, my question is, how can I mount md0 array (md1 has been mounted without problem) in order to preserve existing data? One more thing, fdisk -l command gives the following result: Disk /dev/sdb: 750.1 GB, 750156374016 bytes 255 heads, 63 sectors/track, 91201 cylinders Units = cylinders of 16065 * 512 = 8225280 bytes Disk identifier: 0x660a6799 Device Boot Start End Blocks Id System /dev/sdb1 * 1 88217 708603021 83 Linux /dev/sdb2 88218 91201 23968980 5 Extended /dev/sdb5 88218 91201 23968948+ 82 Linux swap / Solaris Disk /dev/sdc: 750.1 GB, 750156374016 bytes 255 heads, 63 sectors/track, 91201 cylinders Units = cylinders of 16065 * 512 = 8225280 bytes Disk identifier: 0x0008f8ae Device Boot Start End Blocks Id System /dev/sdc1 1 88217 708603021 83 Linux /dev/sdc2 88218 91201 23968980 5 Extended /dev/sdc5 88218 91201 23968948+ 82 Linux swap / Solaris Disk /dev/sdd: 750.1 GB, 750156374016 bytes 255 heads, 63 sectors/track, 91201 cylinders Units = cylinders of 16065 * 512 = 8225280 bytes Disk identifier: 0x4be1abdb Device Boot Start End Blocks Id System Disk /dev/sde: 750.1 GB, 750156374016 bytes 255 heads, 63 sectors/track, 91201 cylinders Units = cylinders of 16065 * 512 = 8225280 bytes Disk identifier: 0xa4d5632e Device Boot Start End Blocks Id System Disk /dev/sdf: 750.1 GB, 750156374016 bytes 255 heads, 63 sectors/track, 91201 cylinders Units = cylinders of 16065 * 512 = 8225280 bytes Disk identifier: 0xdacb141c Device Boot Start End Blocks Id System Disk /dev/sdg: 750.1 GB, 750156374016 bytes 255 heads, 63 sectors/track, 91201 cylinders Units = cylinders of 16065 * 512 = 8225280 bytes Disk identifier: 0xdacb141c Device Boot Start End Blocks Id System Disk /dev/md1: 750.1 GB, 750156251136 bytes 2 heads, 4 sectors/track, 183143616 cylinders Units = cylinders of 8 * 512 = 4096 bytes Disk identifier: 0xdacb141c Device Boot Start End Blocks Id System Warning: ignoring extra data in partition table 5 Warning: ignoring extra data in partition table 5 Warning: ignoring extra data in partition table 5 Warning: invalid flag 0x7b6e of partition table 5 will be corrected by w(rite) Disk /dev/md0: 1500.3 GB, 1500312502272 bytes 255 heads, 63 sectors/track, 182402 cylinders Units = cylinders of 16065 * 512 = 8225280 bytes Disk identifier: 0x660a6799 Device Boot Start End Blocks Id System /dev/md0p1 * 1 88217 708603021 83 Linux /dev/md0p2 88218 91201 23968980 5 Extended /dev/md0p5 ? 121767 155317 269488144 20 Unknown And one more thing. When using mdadm --examine command, here ise result: mdadm -v --examine --scan /dev/sdb /dev/sdc /dev/sdd /dev/sde /dev/sdf /dev/sd ARRAY /dev/md1 level=raid1 num-devices=2 UUID=9b592be7:c6a2052f:2e29483d:f114274d devices=/dev/sdf ARRAY /dev/md0 level=raid10 num-devices=4 UUID=1a02e789:c34377a1:2e29483d:f114274d devices=/dev/sdb,/dev/sdc,/dev/sdd,/dev/sde md0 has 3 devices which are active. Can someone instruct me how to solve this issue? If it is possible, I would like not to removing faulty HDD. Please advise

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  • Installation of Access Database Engine 32-bit Fails

    - by Rayzor78
    I am trying to install Access Database Engine 2007 32-bit. The splash screen comes up, you click "Next", then it fails with the error: Installation ended prematurely because of an error You click "OK" and another error window says: The installation of the package failed. The exact same situation happens when I try this with Access Database Engine 2010 32-bit. This production server is running Windows Server 2008 R2 SP1 64-bit. Before I tried installing Access Database Engine 32-bit, I first needed to install Microsoft Office 2010 Pro (Excel and Office Tools only). I tried the 32-bit version on the production server since that is how I set it up in our Dev environment. No luck. The 32-bit version would not install. I did NOT get the error "You have 64-bit components of Office installed". I simply received the exact same two errors listed above. So, I knew that 32-bit/64-bit did not really matter for the Office install for my project, so I installed 64-bit of Office Pro 2010 (Excel and Office Tools only) with no problems. I have a requirement that I need to have the 32-bit version of the Access Database Engine installed. 2007 or 2010, doesn't matter. I cannot use the 64-bit version of Access Database Engine 2010 because my SSIS package will not work with it. I require the 32-bit version. I've tried several steps to try to get it installed. I seriously think that the production server has some aversion to installing 32-bit applications. Here's what I've tried: Tried installing via command line with the "/passive" switch....no luck. Tried numerous iterations to copy the install file to the server (downloaded a fresh copy directly to the server, downloaded a fresh copy to my local machine then copied it over, copied it over zipped up) (http://social.msdn.microsoft.com/Forums/en-US/sqldataaccess/thread/efd3c1f0-07cd-45ca-a626-2dd0c7ac3e9f). Tried Method 1 from this link. Could not try Method 2 because it requires a server reboot and in my environment that requires a long change management process. I've verified that I am a local administrator on the server. (Evidence, I am able to install other applications (office 64-bit per above)). Verified that there are no other office products that should be blocking the installation. The fore-mentioned install of Excel 2010 64-bit was the first Office product installed on the server. VERY ODD: To test my theory that the production server does not like 32-bit applications, I installed something lightweight. I installed 7-Zip 32-bit on the production server with no problems whatsoever. Here are some things that I have not tried (i will follow-up once I do): Method 2 (as mentioned above). Requires a server reboot. Have not verified that the Dev and Production environments are 100% identical. I've done a cursory check and on the surface they appear to be the same (same OS and SP version). I need to do a deeper dive to be 100% certain. I had no problems in my Dev environment. In Dev, I installed Office 2010 Pro 64-bit (Excel & Office Tools only) then via command line w/ the "/passive" switch, installed Access Database Engine 2010 32-bit. I don't know what else to try. Any suggestions or comments?

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  • Mongodb Slave replication lag

    - by Leonid Bugaev
    We using standard mongo setup: 2 replicas + 1 arbiter. Both replica servers use same AWS m1.medium with RAID10 EBS. We experiencing constantly growing replication lag on secondary replica. I tried to do full-resync, you can see it on graph, but it helped only for some hours. Our mongo usage is really low now, and frankly i can't understan why it can be. iostat 1 for secondary: avg-cpu: %user %nice %system %iowait %steal %idle 80.39 0.00 2.94 0.00 16.67 0.00 Device: tps kB_read/s kB_wrtn/s kB_read kB_wrtn xvdap1 0.00 0.00 0.00 0 0 xvdb 0.00 0.00 0.00 0 0 xvdfp4 12.75 0.00 189.22 0 193 xvdfp3 12.75 0.00 189.22 0 193 xvdfp2 7.84 0.00 40.20 0 41 xvdfp1 7.84 0.00 40.20 0 41 md127 19.61 0.00 219.61 0 224 mongostat for secondary (why 100% locks? i guess its the problem): insert query update delete getmore command flushes mapped vsize res faults locked % idx miss % qr|qw ar|aw netIn netOut conn set repl time *10 *0 *16 *0 0 2|4 0 30.9g 62.4g 1.65g 0 107 0 0|0 0|0 198b 1k 16 replset-01 SEC 06:55:37 *4 *0 *8 *0 0 12|0 0 30.9g 62.4g 1.65g 0 91.7 0 0|0 0|0 837b 5k 16 replset-01 SEC 06:55:38 *4 *0 *7 *0 0 3|0 0 30.9g 62.4g 1.64g 0 110 0 0|0 0|0 342b 1k 16 replset-01 SEC 06:55:39 *4 *0 *8 *0 0 1|0 0 30.9g 62.4g 1.64g 0 82.9 0 0|0 0|0 62b 1k 16 replset-01 SEC 06:55:40 *3 *0 *7 *0 0 5|0 0 30.9g 62.4g 1.6g 0 75.2 0 0|0 0|0 466b 2k 16 replset-01 SEC 06:55:41 *4 *0 *7 *0 0 1|0 0 30.9g 62.4g 1.64g 0 138 0 0|0 0|1 62b 1k 16 replset-01 SEC 06:55:42 *7 *0 *15 *0 0 3|0 0 30.9g 62.4g 1.64g 0 95.4 0 0|0 0|0 342b 1k 16 replset-01 SEC 06:55:43 *7 *0 *14 *0 0 1|0 0 30.9g 62.4g 1.64g 0 98 0 0|0 0|0 62b 1k 16 replset-01 SEC 06:55:44 *8 *0 *17 *0 0 3|0 0 30.9g 62.4g 1.64g 0 96.3 0 0|0 0|0 342b 1k 16 replset-01 SEC 06:55:45 *7 *0 *14 *0 0 3|0 0 30.9g 62.4g 1.64g 0 96.1 0 0|0 0|0 186b 2k 16 replset-01 SEC 06:55:46 mongostat for primary insert query update delete getmore command flushes mapped vsize res faults locked % idx miss % qr|qw ar|aw netIn netOut conn set repl time 12 30 20 0 0 3 0 30.9g 62.6g 641m 0 0.9 0 0|0 0|0 212k 619k 48 replset-01 M 06:56:41 5 17 10 0 0 2 0 30.9g 62.6g 641m 0 0.5 0 0|0 0|0 159k 429k 48 replset-01 M 06:56:42 9 22 16 0 0 3 0 30.9g 62.6g 642m 0 0.7 0 0|0 0|0 158k 276k 48 replset-01 M 06:56:43 6 18 12 0 0 2 0 30.9g 62.6g 640m 0 0.7 0 0|0 0|0 93k 231k 48 replset-01 M 06:56:44 6 12 8 0 0 3 0 30.9g 62.6g 640m 0 0.3 0 0|0 0|0 80k 125k 48 replset-01 M 06:56:45 8 21 14 0 0 9 0 30.9g 62.6g 641m 0 0.6 0 0|0 0|0 118k 419k 48 replset-01 M 06:56:46 10 34 20 0 0 6 0 30.9g 62.6g 640m 0 1.3 0 0|0 0|0 164k 527k 48 replset-01 M 06:56:47 6 21 13 0 0 2 0 30.9g 62.6g 641m 0 0.7 0 0|0 0|0 111k 477k 48 replset-01 M 06:56:48 8 21 15 0 0 2 0 30.9g 62.6g 641m 0 0.7 0 0|0 0|0 204k 336k 48 replset-01 M 06:56:49 4 12 8 0 0 8 0 30.9g 62.6g 641m 0 0.5 0 0|0 0|0 156k 530k 48 replset-01 M 06:56:50 Mongo version: 2.0.6

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  • Where is my VMware-ws FreeNAS CIFS(ZFS) bottle-neck?

    - by maka
    Background: I'm building a quiet HTPC + NAS that is also supposed to be used for general computer usage. I'm so far generally happy with things, it was just that I was expecting a little better IO performance. I have no clue if my expectations are unreal. The NAS is there as a general purpose file storage and as a media server for XBMC and other devices. ZFS is a requirement. Question: Where is my bottle-neck, and is there anything I can do config wise, to improve my performance? I'm thinking VM-disk settings could be something but I really have no idea where to go since I'm neither experienced with FreeNAS nor VMware-WS. Tests: When I'm on the host OS and copy files (from the SSD) to the CIFS share, I get around 30 Mbytes/sec read and write. When I'm on my laptop laptop, wired to the network, I get about the same specs. The test I've done are with a 16 GB ISO, and with about 200 MB of RARs and I've tried avoiding the RAM-cache by reading different files than the ones I'm writing ( 10 GB). It feels like having less CPU cores is a lot more efficient, since the resource manager in Windows reports less CPU-usage. With 4 cores in VMware, CPU usage was 50-80%, with 1 core it was 25-60%. EDIT: HD ActiveTime was quite high on SSD so I moved the page file, disabled hibernate and enabled Win DiskCache both on SSD and RAID. This resulted in no real performance difference for one file, but if i transferred 2 files the total speed went up to 50 Mbytes/s vs ~40. The ActiveTime avg also went down a lot (to ~20%) but has now higher bursts. DiskIO is on ~ 30-35 Mbytes/s avgs, with ~100Mb bursts. Network is on 200-250Mbits/s with ~45 active TCP connections. Hardware Asus F2A85-M Pro A10-5700 16GB DDR3 1600 OCZ Vertex 2 128GB SSD 2x Generic 1tb 7200 RPM drives as RAID0 (in win7) Intel Gigabit Desktop CT Software Host OS: Win7 (SSD) VMware Worksation 9 (SSD) FreeNAS 8.3 VM (20GB VDisk on SSD) CPU: I've tried 1, 2 and 4 cores. Virtualisation engine, Preferred mode: Automatic 10,24Gb ram 50Gb SCSI VDisk on the RAID0, VDisk is formatted as ZFS and exposed through CIFS through FreeNAS. NIC Bridge, Replicate physical network state Below are two typical process print-outs while I'm transfering one file to the CIFS share. last pid: 2707; load averages: 0.60, 0.43, 0.24 up 0+00:07:05 00:34:26 32 processes: 2 running, 30 sleeping Mem: 101M Active, 53M Inact, 1620M Wired, 2188K Cache, 149M Buf, 8117M Free Swap: 4096M Total, 4096M Free PID USERNAME THR PRI NICE SIZE RES STATE TIME WCPU COMMAND 2640 root 1 102 0 50164K 10364K RUN 0:25 25.98% smbd 1897 root 6 44 0 168M 74808K uwait 0:02 0.00% python last pid: 2746; load averages: 0.93, 0.60, 0.33 up 0+00:08:53 00:36:14 33 processes: 2 running, 31 sleeping Mem: 101M Active, 53M Inact, 4722M Wired, 2188K Cache, 152M Buf, 5015M Free Swap: 4096M Total, 4096M Free PID USERNAME THR PRI NICE SIZE RES STATE TIME WCPU COMMAND 2640 root 1 76 0 50164K 10364K RUN 0:52 16.99% smbd 1897 root 6 44 0 168M 74816K uwait 0:02 0.00% python I'm sorry if my question isn't phrased right, I'm really bad at these kind of things, and it is the first time I post here at SU. I also appreciate any other suggestions to something, I could have missed.

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  • Capistrano + Nginx + Passenger = 403

    - by slimchrisp
    I asked this over at stackoverflow as well, but still haven't received any answers that have helped me to solve this problem. I have spent almost a week at this point trying to solve the issue, and I'm just not making any headway. It seems that this issue is pretty common, but none of the solutions I found online work for me. A buddy of mine is actually creating the same setup, and he is having the same issue. After a few days stuck with the 403 error I started over using this tutorial: http://blog.ninjahideout.com/posts/a-guide-to-a-nginx-passenger-and-rvm-server I had hoped starting from scratch using this tutorial would work, but no dice. Either way, if you view the tutorial you can see what steps I have taken. Here is essentially what I have going on. I have a VPS account on linode.com Server OS is Ubuntu 10.04 Local OS (shouldn't matter, but just so you know) used to deploy with Capistrano is Snow Leopard 10.6.6 I use RVM on the server. Version is 1.2.2 I was previously on ruby-1.9.2-p0 [ i386 ], but per the tutorial listed above I switched to ree-1.8.7-2010.02 [ i386 ]. Running 'which ruby' from the command line verifies that I am using 1.8.7 with the following output: /usr/local/rvm/rubies/ree-1.8.7-2010.02/bin/ruby passenger -v prints the following: Phusion Passenger version 3.0.2 Running 'nginx -v' gives me a message that the command nginx could not be found. The server is definitely there and running as I can use nginx to serve static files, but this could have something to do with my problem. I have two users dealing with the install. root which I used to install everything, and deployer which is a user I created specifically to for deploying my applications My web app directory is in the deployer user's home directory as follows: /home/deployer/webapps/mysite.com/public Per Capistrano default deploy, a symbolic link called current is created in the public folder, and points to /home/deployer/webapps/mysite.com/public/releases/most_current_release I have chmodded the deployer directory recursively to 777 /opt/nginx permissions: rwxr-xr-x /usr/local/rvm/gems/ree-1.8.7-2010.02/gems/passenger-3.0.2 permissions: rwxrwsrwx My nginx config file has gone through just short of eternity variations, but currently looks like this: ================================================================================== worker_processes 1; events { worker_connections 1024; } http { passenger_root /usr/local/rvm/gems/ree-1.8.7-2010.02/gems/passenger-3.0.2; passenger_ruby /usr/local/rvm/bin/passenger_ruby; include mime.types; default_type application/octet-stream; sendfile on; keepalive_timeout 65; server { # listen *:80; server_name mysite.com www.mysite.com; root /home/deployer/webapps/mysite.com/public/current; passenger_enabled on; passenger_friendly_error_pages on; access_log logs/mysite.com/server.log; error_log logs/mysite.com/error.log info; error_page 500 502 503 504 /50x.html; location = /50x.html { root html; } } } ================================================================================== I bounce nginx, hit the site, and boom. 403, and logs say directory index of /home/deployer... is forbidden As others with a similar problem have said, you can drop an index.html into the public/releases/current_release and it will render. But rails no worky. That's basically it. At this point I have just about completely exhausted every possible solution attempt I can think of. I am a programmer and definitely not a sysadmin, so I am 99% sure this has something to do with permissions that I have hosed, but for the life of me I just can't figure out where. If anyone can help I would really really appreciate it. If there's any specific permission things you want me to check (ie groups/permissions), can you please include the commands to do so as well. Hopefully this will help others in the future who read this post. Let me know if there is any other information I can provide, and thanks in advance!!!

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  • Debian dependency problems / partially installed

    - by Michael
    I tried to install curl support for php 5 on my debian squeeze machine and since I'm having problems. After trying to install curl I got dependency issues which I tried to solve by removing what started the issues. From one thing came another and I'm currently looking at ~29 issues when I try to do an apt-get upgrade. These issues vary from unable to config, dependency and unable to remove errors. I tried apt-get upgrade -f and installing packages using dpkg command. I tried removing using purge and force. I manually removed stuff to try and fix it. I tried running dpkg --configure -a. I've to say I'm still pretty new to linux so I'm out of idea's and cant seem to find an answer online that matches my problems. Here's a part of the apt-get upgrade command output: Reading package lists... Building dependency tree... Reading state information... 0 upgraded, 0 newly installed, 0 to remove and 0 not upgraded. 29 not fully installed or removed. After this operation, 0 B of additional disk space will be used. Setting up libgeoip1 (1.4.7~beta6+dfsg-1) ... Bus error dpkg: error processing libgeoip1 (--configure): subprocess installed post-installation script returned error exit status 135 Setting up libisc62 (1:9.7.3.dfsg-1~squeeze3) ... Bus error dpkg: error processing libisc62 (--configure): subprocess installed post-installation script returned error exit status 135 dpkg: dependency problems prevent configuration of libdns69: libdns69 depends on libgeoip1 (>= 1.4.7~beta6+dfsg); however: Package libgeoip1 is not configured yet. libdns69 depends on libisc62; however: Package libisc62 is not configured yet. dpkg: error processing libdns69 (--configure): dependency problems - leaving unconfigured dpkg: dependency problems prevent configuration of libisccc60: libisccc60 depends on libisc62; however: Package libisc62 is not configured yet. dpkg: error processing libisccc60 (--configure): dependency problems - leaving unconfigured dpkg: dependency problems prevent configuration of libisccfg62: libisccfg62 depends on libdns69; however: Package libdns69 is not configured yet. .. continues Errors were encountered while processing: libgeoip1 libisc62 libdns69 libisccc60 libisccfg62 libbind9-60 liblwres60 bind9-host libavahi-core7 libdaemon0 avahi-daemon libexif12 libffi5 libgomp1 libgphoto2-port0 libgphoto2-2 libperl5.10 libsensors4 libsnmp15 libhpmud0 libieee1284-3 libnss-mdns libossp-uuid16 libpq5 libv4l-0 libsane libsane-hpaio libssh2-1 python-gobject dpkg --configure -a Setting up libpq5 (8.4.8-0squeeze2) ... Bus error dpkg: error processing libpq5 (--configure): subprocess installed post-installation script returned error exit status 135 Setting up libperl5.10 (5.10.1-17squeeze2) ... Bus error dpkg: error processing libperl5.10 (--configure): subprocess installed post-installation script returned error exit status 135 Setting up libffi5 (3.0.9-3) ... Bus error dpkg: error processing libffi5 (--configure): subprocess installed post-installation script returned error exit status 135 Setting up libexif12 (0.6.19-1) ... .. continues Suggestions are really welcome I really don't know what to do. Michael.

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  • Windows 7 disk errors after a few hours of runtime

    - by GFK
    I'm having trouble understanding what is going on with my work PC. Whenever I boot it, it runs fine for a while, then starts to randomly show disk errors. The displayed error often contains the message "not enough storage is available to process this command", although depending on the application that fails it can be different. This has happened for weeks now and is getting worse. This is what troubles me: It never seems to impact critical parts of the system (no BSOD, no freeze). Only some applications seem impacted, refusing to function correctly after a while: Outlook 2010 cannot download RSS feeds anymore, Firefox 6 or IE9 cannot download anything bigger than 3MB without failing, Windows Update fails, all msi installers fail, Visual Studio 2010 starts failing in weird manners... It only happens after a while using it (typically 3 hours, but it seems that installing a program or compiling several times makes it shorter) Rebooting solves it (temporarily). The system: The OS is Windows 7 Pro Spanish SP1, 32 bits The system is an HP Compaq 6000 Pro with 4 GB memory (only 3.4GB usable since the system is 32bit), one 500GB hard drive. Installed applications include: Visual Studio 2010, SQL Server 2008 R2, VMWare Workstation 7, Microsoft Security Essentials, Office 2010. Shutting down all related services and processes doesn't seem to change anything. The diagnostics I've run so far: Hard drive : 465GB, 165GB free Process Explorer : physical and virtual memory seem ok (pagefile is 5.3GB, physical memory usage 70%, system commit 39%) Windows Memory diagnostic tool: OK CHKDSK returned: 488282111 KB total disk space. 281668248 KB in 265779 files. 150188 KB in 62949 indexes. 0 KB in bad sectors. 571755 KB in use by the system. The log file has occupied 65536 kilobytes. 205891920 KB available on disk. For non-spanish speakers, that means all ok. SMART diagnostic tools (DiskCheckup) report all values normal. temperatures are in the normal range (HWinfo). The event viewer doesn't seem to contain any significant message. ran CCleaner 3, without any noticeable effect. I was thinking about some file number limit (between Visual Studio projects and other applications, there are around 300.000 files on the hard drive), but I couldn't find any. It's possible there is something related with the use of the temporary folders (it's the only explanation I have for why applications fail but Windows doesn't), but I cannot confirm that. Only thing I cannot find out is if chkdsk reporting 65MB for the log is normal. It seems since Vista it always reports this. Any other cleaning/diagnostic tool you might know of? Edit: I ran several other tools since I first published the question: Seagate SeaTools (the HD manufacturer's analysis tool): complete test run OK. Intel Rapid 10.1 (the HD controller manufacturer's troubleshooting tool): the HD's ok. Microsoft Desktop Heap Monitor: Desktop Heap Information Monitor Tool (Version 8.1.2925.0) Copyright (c) Microsoft Corporation. All rights reserved. Session ID: 1 Total Desktop: ( 46464 KB - 11 desktops) WinStation\Desktop Heap Size(KB) Used Rate(%) WinSta0\Winlogon (s1) 128 3.6 WinSta0\Disconnect (s1) 64 3.8 WinSta0\Default (s1) 20480 3.0 msswindowstation\mssrestricteddesk (s0) 1024 0.2 __X78B95_89_IW__A8D9S1_42_ID (s0) 1024 0.2 Service-0x0-3e5$\Default (s0) 1024 0.6 Service-0x0-3e4$\Default (s0) 1024 0.3 Service-0x0-3e7$\Default (s0) 1024 2.1 WinSta0\Winlogon (s0) 128 1.9 WinSta0\Disconnect (s0) 64 3.8 WinSta0\Default (s0) 20480 0.0 All ok, desktop heap usage < 5% Edit 2: I tried totally resetting my account by creating a new one, logging under this new one and delete the first one (local rights and files), then logging back with this deleted account (it is a domain account). No luck. Also, I found out often the error is "not enough storage is available to process this command". Searching on the internet, I found an old troubleshooting tip (setting a registry key to raise the IRP stack limit, whatever it is) which did not change anything.

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  • RedHat 5.5 server does not show per processor memory utilization

    - by Mike S
    I have been searching all over internet but not finding any leads. I have a system with a memory leak that I am trying to troubleshoot. Unfortunately I am not able to see per processor memory utilization. Here are the outputs of TOP and PS commands. Linux SERVER_NAME 2.6.18-194.8.1.el5 #1 SMP Wed Jun 23 10:52:51 EDT 2010 x86_64 x86_64 x86_64 GNU/Linux top - 09:17:13 up 18:43, 3 users, load average: 0.00, 0.00, 0.00 Tasks: 375 total, 1 running, 373 sleeping, 0 stopped, 1 zombie Cpu(s): 0.0%us, 0.0%sy, 0.0%ni,100.0%id, 0.0%wa, 0.0%hi, 0.0%si, 0.0%st Mem: 32922828k total, 32776712k used, 146116k free, 267128k buffers Swap: 5245212k total, 0k used, 5245212k free, 32141044k cached PID USER PR NI VIRT RES SHR S %CPU %MEM TIME+ COMMAND 1 root 15 0 10348 744 620 S 0.0 0.0 0:05.65 init 2 root RT -5 0 0 0 S 0.0 0.0 0:00.05 migration/0 3 root 34 19 0 0 0 S 0.0 0.0 0:00.00 ksoftirqd/0 4 root RT -5 0 0 0 S 0.0 0.0 0:00.00 watchdog/0 5 root RT -5 0 0 0 S 0.0 0.0 0:00.00 migration/1 6 root 34 19 0 0 0 S 0.0 0.0 0:00.00 ksoftirqd/1 7 root RT -5 0 0 0 S 0.0 0.0 0:00.00 watchdog/1 8 root RT -5 0 0 0 S 0.0 0.0 0:00.00 migration/2 9 root 34 19 0 0 0 S 0.0 0.0 0:00.00 ksoftirqd/2 10 root RT -5 0 0 0 S 0.0 0.0 0:00.00 watchdog/2 11 root RT -5 0 0 0 S 0.0 0.0 0:00.01 migration/3 12 root 34 19 0 0 0 S 0.0 0.0 0:00.01 ksoftirqd/3 13 root RT -5 0 0 0 S 0.0 0.0 0:00.00 watchdog/3 14 root RT -5 0 0 0 S 0.0 0.0 0:00.00 migration/4 15 root 34 19 0 0 0 S 0.0 0.0 0:00.01 ksoftirqd/4 16 root RT -5 0 0 0 S 0.0 0.0 0:00.00 watchdog/4 17 root RT -5 0 0 0 S 0.0 0.0 0:00.00 migration/5 18 root 34 19 0 0 0 S 0.0 0.0 0:00.00 ksoftirqd/5 19 root RT -5 0 0 0 S 0.0 0.0 0:00.00 watchdog/5 20 root RT -5 0 0 0 S 0.0 0.0 0:00.00 migration/6 % ps -auxf | sort -nr -k 4 | head -10 Warning: bad syntax, perhaps a bogus '-'? See /usr/share/doc/procps-3.2.7/FAQ xfs 6205 0.0 0.0 23316 3892 ? Ss Aug19 0:00 xfs -droppriv -daemon uuidd 6101 0.0 0.0 60976 224 ? Ss Aug19 0:00 /usr/sbin/uuidd USER PID %CPU %MEM VSZ RSS TTY STAT START TIME COMMAND smmsp 6130 0.0 0.0 57900 1784 ? Ss Aug19 0:00 sendmail: Queue runner@01:00:00 for /var/spool/clientmqueue rpc 5126 0.0 0.0 8052 632 ? Ss Aug19 0:00 portmap root 99 0.0 0.0 0 0 ? S< Aug19 0:00 [events/1] root 98 0.0 0.0 0 0 ? S< Aug19 0:00 [events/0] root 97 0.0 0.0 0 0 ? S< Aug19 0:00 [watchdog/31] root 96 0.0 0.0 0 0 ? SN Aug19 0:00 [ksoftirqd/31] root 95 0.0 0.0 0 0 ? S< Aug19 0:00 [migration/31] Any help with this is appretiate.

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  • Using TypeScript in ASP.NET MVC Projects

    - by shiju
    In the previous blog post Microsoft TypeScript : A Typed Superset of JavaScript, I have given a brief introduction on TypeScript. In this post, I will demonstrate how to use TypeScript with ASP.NET MVC projects and how we can compile TypeScript within the ASP.NET MVC projects. Using TypeScript with ASP.NET MVC 3 Projects The Visual Studio plug-in for TypeScript provides an ASP.NET MVC 3 project template for TypeScript that lets you to compile TypeScript from the Visual Studio. The following screen shot shows the TypeScript template for ASP.NET MVC 3 project The “TypeScript Internet Application” template is just a ASP.NET MVC 3 internet application project template which will allows to compile TypeScript programs to JavaScript when you are building your ASP.NET MVC projects. This project template will have the following section in the .csproject file <None Include="Scripts\jquery.d.ts" /> <TypeScriptCompile Include="Scripts\site.ts" /> <Content Include="Scripts\site.js"> <DependentUpon>site.ts</DependentUpon> </Content> .csharpcode, .csharpcode pre { font-size: small; color: black; font-family: consolas, "Courier New", courier, monospace; background-color: #ffffff; /*white-space: pre;*/ } .csharpcode pre { margin: 0em; } .csharpcode .rem { color: #008000; } .csharpcode .kwrd { color: #0000ff; } .csharpcode .str { color: #006080; } .csharpcode .op { color: #0000c0; } .csharpcode .preproc { color: #cc6633; } .csharpcode .asp { background-color: #ffff00; } .csharpcode .html { color: #800000; } .csharpcode .attr { color: #ff0000; } .csharpcode .alt { background-color: #f4f4f4; width: 100%; margin: 0em; } .csharpcode .lnum { color: #606060; } <Target Name="BeforeBuild"> <Exec Command="&amp;quot;$(PROGRAMFILES)\ Microsoft SDKs\TypeScript\0.8.0.0\tsc&amp;quot; @(TypeScriptCompile ->'&quot;%(fullpath)&quot;', ' ')" /> </Target> .csharpcode, .csharpcode pre { font-size: small; color: black; font-family: consolas, "Courier New", courier, monospace; background-color: #ffffff; /*white-space: pre;*/ } .csharpcode pre { margin: 0em; } .csharpcode .rem { color: #008000; } .csharpcode .kwrd { color: #0000ff; } .csharpcode .str { color: #006080; } .csharpcode .op { color: #0000c0; } .csharpcode .preproc { color: #cc6633; } .csharpcode .asp { background-color: #ffff00; } .csharpcode .html { color: #800000; } .csharpcode .attr { color: #ff0000; } .csharpcode .alt { background-color: #f4f4f4; width: 100%; margin: 0em; } .csharpcode .lnum { color: #606060; } The “BeforeBuild” target will allows you to compile TypeScript programs when you are building your ASP.NET MVC projects. The TypeScript project template will provide a typing reference file for the jQuery library named “jquery.d.ts”. The following default app.ts file referenced to jquery.d.ts 1: ///<reference path='jquery.d.ts' /> 2:   3: $(document).ready(function () { 4:   5: $(".btn-slide").click(function () { 6: $("#main").slideToggle("slow"); 7: $(this).toggleClass("active"); 8: }); 9:   10: }); .csharpcode, .csharpcode pre { font-size: small; color: black; font-family: consolas, "Courier New", courier, monospace; background-color: #ffffff; /*white-space: pre;*/ } .csharpcode pre { margin: 0em; } .csharpcode .rem { color: #008000; } .csharpcode .kwrd { color: #0000ff; } .csharpcode .str { color: #006080; } .csharpcode .op { color: #0000c0; } .csharpcode .preproc { color: #cc6633; } .csharpcode .asp { background-color: #ffff00; } .csharpcode .html { color: #800000; } .csharpcode .attr { color: #ff0000; } .csharpcode .alt { background-color: #f4f4f4; width: 100%; margin: 0em; } .csharpcode .lnum { color: #606060; } Using TypeScript with ASP.NET MVC 4 Projects The current preview version of TypeScript is not providing a project template for ASP.NET MVC 4 projects. But you can use TypeScript with ASP.NET MVC 4 projects by editing the project’s .csproject file. You can take the necessary settings from ASP.NET MVC 3 project file. I have just added the following section in the end of the .csproj file of a ASP.NET MVC 4 project, which will allows to compile all TypeScript when building ASP.NET MVC 4 project. <ItemGroup> <TypeScriptCompile Include="$(ProjectDir)\**\*.ts" /> </ItemGroup> <Target Name="BeforeBuild"> <Exec Command="&amp;quot;$(PROGRAMFILES)\ Microsoft SDKs\TypeScript\0.8.0.0\tsc&amp;quot; @(TypeScriptCompile ->'&quot;%(fullpath)&quot;', ' ')" /> </Target> .csharpcode, .csharpcode pre { font-size: small; color: black; font-family: consolas, "Courier New", courier, monospace; background-color: #ffffff; /*white-space: pre;*/ } .csharpcode pre { margin: 0em; } .csharpcode .rem { color: #008000; } .csharpcode .kwrd { color: #0000ff; } .csharpcode .str { color: #006080; } .csharpcode .op { color: #0000c0; } .csharpcode .preproc { color: #cc6633; } .csharpcode .asp { background-color: #ffff00; } .csharpcode .html { color: #800000; } .csharpcode .attr { color: #ff0000; } .csharpcode .alt { background-color: #f4f4f4; width: 100%; margin: 0em; } .csharpcode .lnum { color: #606060; }

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  • Adding the New HTML Editor Extender to a Web Forms Application using NuGet

    - by Stephen Walther
    The July 2011 release of the Ajax Control Toolkit includes a new, lightweight, HTML5 compatible HTML Editor extender. In this blog entry, I explain how you can take advantage of NuGet to quickly add the new HTML Editor control extender to a new or existing ASP.NET Web Forms application. Installing the Latest Version of the Ajax Control Toolkit with NuGet NuGet is a package manager. It enables you to quickly install new software directly from within Visual Studio 2010. You can use NuGet to install additional software when building any type of .NET application including ASP.NET Web Forms and ASP.NET MVC applications. If you have not already installed NuGet then you can install NuGet by navigating to the following address and clicking the giant install button: http://nuget.org/ After you install NuGet, you can add the Ajax Control Toolkit to a new or existing ASP.NET Web Forms application by selecting the Visual Studio menu option Tools, Library Package Manager, Package Manager Console: Selecting this menu option opens the Package Manager Console. You can enter the command Install-Package AjaxControlToolkit in the console to install the Ajax Control Toolkit: After you install the Ajax Control Toolkit with NuGet, your application will include an assembly reference to the AjaxControlToolkit.dll and SanitizerProviders.dll assemblies: Furthermore, your Web.config file will be updated to contain a new tag prefix for the Ajax Control Toolkit controls: <configuration> <system.web> <compilation debug="true" targetFramework="4.0" /> <pages> <controls> <add tagPrefix="ajaxToolkit" assembly="AjaxControlToolkit" namespace="AjaxControlToolkit" /> </controls> </pages> </system.web> </configuration> The configuration file installed by NuGet adds the prefix ajaxToolkit for all of the Ajax Control Toolkit controls. You can type ajaxToolkit: in source view to get auto-complete in Source view. You can, of course, change this prefix to anything you want. Using the HTML Editor Extender After you install the Ajax Control Toolkit, you can use the HTML Editor Extender with the standard ASP.NET TextBox control to enable users to enter rich formatting such as bold, underline, italic, different fonts, and different background and foreground colors. For example, the following page can be used for entering comments. The page contains a standard ASP.NET TextBox, Button, and Label control. When you click the button, any text entered into the TextBox is displayed in the Label control. It is a pretty boring page: Let’s make this page fancier by extending the standard ASP.NET TextBox with the HTML Editor extender control: Notice that the ASP.NET TextBox now has a toolbar which includes buttons for performing various kinds of formatting. For example, you can change the size and font used for the text. You also can change the foreground and background color – and make many other formatting changes. You can customize the toolbar buttons which the HTML Editor extender displays. To learn how to customize the toolbar, see the HTML Editor Extender sample page here: http://www.asp.net/ajaxLibrary/AjaxControlToolkitSampleSite/HTMLEditorExtender/HTMLEditorExtender.aspx Here’s the source code for the ASP.NET page: <%@ Page Language="C#" AutoEventWireup="true" CodeBehind="Default.aspx.cs" Inherits="WebApplication1.Default" %> <!DOCTYPE html PUBLIC "-//W3C//DTD XHTML 1.0 Transitional//EN" "http://www.w3.org/TR/xhtml1/DTD/xhtml1-transitional.dtd"> <html xmlns="http://www.w3.org/1999/xhtml"> <head runat="server"> <title>Add Comments</title> </head> <body> <form id="form1" runat="server"> <div> <ajaxToolkit:ToolkitScriptManager ID="TSM1" runat="server" /> <asp:TextBox ID="txtComments" TextMode="MultiLine" Columns="50" Rows="8" Runat="server" /> <ajaxToolkit:HtmlEditorExtender ID="hee" TargetControlID="txtComments" Runat="server" /> <br /><br /> <asp:Button ID="btnSubmit" Text="Add Comment" Runat="server" onclick="btnSubmit_Click" /> <hr /> <asp:Label ID="lblComment" Runat="server" /> </div> </form> </body> </html> Notice that the page above contains 5 controls. The page contains a standard ASP.NET TextBox, Button, and Label control. However, the page also contains an Ajax Control Toolkit ToolkitScriptManager control and HtmlEditorExtender control. The HTML Editor extender control extends the standard ASP.NET TextBox control. The HTML Editor TargetID attribute points at the TextBox control. Here’s the code-behind for the page above:   using System; namespace WebApplication1 { public partial class Default : System.Web.UI.Page { protected void btnSubmit_Click(object sender, EventArgs e) { lblComment.Text = txtComments.Text; } } }   Preventing XSS/JavaScript Injection Attacks If you use an HTML Editor -- any HTML Editor -- in a public facing web page then you are opening your website up to Cross-Site Scripting (XSS) attacks. An evil hacker could submit HTML using the HTML Editor which contains JavaScript that steals private information such as other user’s passwords. Imagine, for example, that you create a web page which enables your customers to post comments about your website. Furthermore, imagine that you decide to redisplay the comments so every user can see them. In that case, a malicious user could submit JavaScript which displays a dialog asking for a user name and password. When an unsuspecting customer enters their secret password, the script could transfer the password to the hacker’s website. So how do you accept HTML content without opening your website up to JavaScript injection attacks? The Ajax Control Toolkit HTML Editor supports the Anti-XSS library. You can use the Anti-XSS library to sanitize any HTML content. The Anti-XSS library, for example, strips away all JavaScript automatically. You can download the Anti-XSS library from NuGet. Open the Package Manager Console and execute the command Install-Package AntiXSS: Adding the Anti-XSS library to your application adds two assemblies to your application named AntiXssLibrary.dll and HtmlSanitizationLibrary.dll. After you install the Anti-XSS library, you can configure the HTML Editor extender to use the Anti-XSS library your application’s web.config file: <?xml version="1.0" encoding="utf-8"?> <configuration> <configSections> <sectionGroup name="system.web"> <section name="sanitizer" requirePermission="false" type="AjaxControlToolkit.Sanitizer.ProviderSanitizerSection, AjaxControlToolkit"/> </sectionGroup> </configSections> <system.web> <sanitizer defaultProvider="AntiXssSanitizerProvider"> <providers> <add name="AntiXssSanitizerProvider" type="AjaxControlToolkit.Sanitizer.AntiXssSanitizerProvider"></add> </providers> </sanitizer> <compilation debug="true" targetFramework="4.0" /> <pages> <controls> <add tagPrefix="ajaxToolkit" assembly="AjaxControlToolkit" namespace="AjaxControlToolkit" /> </controls> </pages> </system.web> </configuration> Summary In this blog entry, I described how you can quickly get started using the new HTML Editor extender – included with the July 2011 release of the Ajax Control Toolkit – by installing the Ajax Control Toolkit with NuGet. If you want to learn more about the HTML Editor then please take a look at the Ajax Control Toolkit sample site: http://www.asp.net/ajaxLibrary/AjaxControlToolkitSampleSite/HTMLEditorExtender/HTMLEditorExtender.aspx

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  • TFS 2010 Build Custom Activity for Merging Assemblies

    - by Jakob Ehn
    *** The sample build process template discussed in this post is available for download from here: http://cid-ee034c9f620cd58d.office.live.com/self.aspx/BlogSamples/ILMerge.xaml ***   In my previous post I talked about library builds that we use to build and replicate dependencies between applications in TFS. This is typically used for common libraries and tools that several other application need to reference. When the libraries grow in size over time, so does the number of assemblies. So all solutions that uses the common library must reference all the necessary assemblies that they need, and if we for example do a refactoring and extract some code into a new assembly, all the clients must update their references to reflect these changes, otherwise it won’t compile. To improve on this, we use a tool from Microsoft Research called ILMerge (Download from here). It can be used to merge several assemblies into one assembly that contains all types. If you haven’t used this tool before, you should check it out. Previously I have implemented this in builds using a simple batch file that contains the full command, something like this: "%ProgramFiles(x86)%\microsoft\ilmerge\ilmerge.exe" /target:library /attr:ClassLibrary1.bl.dll /out:MyNewLibrary.dll ClassLibrary1.dll ClassLibrar2.dll ClassLibrary3.dll This merges 3 assemblies (ClassLibrary1, 2 and 3) into a new assembly called MyNewLibrary.dll. It will copy the attributes (file version, product version etc..) from ClassLibrary1.dll, using the /attr switch. For more info on ILMerge command line tool, see the above link. This approach works, but requires a little bit too much knowledge for the developers creating builds, therefor I have implemented a custom activity that wraps the use of ILMerge. This makes it much simpler to setup a new build definition and have the build automatically do the merging. The usage of the activity is then implemented as part of the Library Build process template mentioned in the previous post. For this article I have just created a simple build process template that only performs the ILMerge operation.   Below is the code for the custom activity. To make it compile, you need to reference the ILMerge.exe assembly. /// <summary> /// Activity for merging a list of assembies into one, using ILMerge /// </summary> public sealed class ILMergeActivity : BaseCodeActivity { /// <summary> /// A list of file paths to the assemblies that should be merged /// </summary> [RequiredArgument] public InArgument<IEnumerable<string>> InputAssemblies { get; set; } /// <summary> /// Full path to the generated assembly /// </summary> [RequiredArgument] public InArgument<string> OutputFile { get; set; } /// <summary> /// Which input assembly that the attibutes for the generated assembly should be copied from. /// Optional. If not specified, the first input assembly will be used /// </summary> public InArgument<string> AttributeFile { get; set; } /// <summary> /// Kind of assembly to generate, dll or exe /// </summary> public InArgument<TargetKindEnum> TargetKind { get; set; } // If your activity returns a value, derive from CodeActivity<TResult> // and return the value from the Execute method. protected override void Execute(CodeActivityContext context) { string message = InputAssemblies.Get(context).Aggregate("", (current, assembly) => current + (assembly + " ")); TrackMessage(context, "Merging " + message + " into " + OutputFile.Get(context)); ILMerge m = new ILMerge(); m.SetInputAssemblies(InputAssemblies.Get(context).ToArray()); m.TargetKind = TargetKind.Get(context) == TargetKindEnum.Dll ? ILMerge.Kind.Dll : ILMerge.Kind.Exe; m.OutputFile = OutputFile.Get(context); m.AttributeFile = !String.IsNullOrEmpty(AttributeFile.Get(context)) ? AttributeFile.Get(context) : InputAssemblies.Get(context).First(); m.SetTargetPlatform(RuntimeEnvironment.GetSystemVersion().Substring(0,2), RuntimeEnvironment.GetRuntimeDirectory()); m.Merge(); TrackMessage(context, "Generated " + m.OutputFile); } } [Browsable(true)] public enum TargetKindEnum { Dll, Exe } NB: The activity inherits from a BaseCodeActivity class which is an internal helper class which contains some methods and properties useful for moste custom activities. In this case, it uses the TrackeMessage method for writing to the build log. You either need to remove the TrackMessage method calls, or implement this yourself (which is not very hard… ) The custom activity has the following input arguments: InputAssemblies A list with the (full) paths to the assemblies to merge OutputFile The name of the resulting merged assembly AttributeFile Which assembly to use as the template for the attribute of the merged assembly. This argument is optional and if left blank, the first assembly in the input list is used TargetKind Decides what type of assembly to create, can be either a dll or an exe Of course, there are more switches to the ILMerge.exe, and these can be exposed as input arguments as well if you need it. To show how the custom activity can be used, I have attached a build process template (see link at the top of this post) that merges the output of the projects being built (CommonLibrary.dll and CommonLibrary2.dll) into a merged assembly (NewLibrary.dll). The build process template has the following custom process parameters:   The Assemblies To Merge argument is passed into a FindMatchingFiles activity to located all assemblies that are located in the BinariesDirectory folder after the compilation has been performed by Team Build. Here is the complete sequence of activities that performs the merge operation. It is located at the end of the Try, Compile, Test and Associate… sequence: It splits the AssembliesToMerge parameter and appends the full path (using the BinariesDirectory variable) and then enumerates the matching files using the FindMatchingFiles activity. When running the build, you can see that it merges two assemblies into a new one:     And the merged assembly (and associated pdb file) is copied to the drop location together with the rest of the assemblies:

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  • ActiveX component can't create Object Error? Check 64 bit Status

    - by Rick Strahl
    If you're running on IIS 7 and a 64 bit operating system you might run into the following error using ASP classic or ASP.NET with COM interop. In classic ASP applications the error will show up as: ActiveX component can't create object   (Error 429) (actually without error handling the error just shows up as 500 error page) In my case the code that's been giving me problems has been a FoxPro COM object I'd been using to serve banner ads to some of my pages. The code basically looks up banners from a database table and displays them at random. The ASP classic code that uses it looks like this: <% Set banner = Server.CreateObject("wwBanner.aspBanner") banner.BannerFile = "wwsitebanners" Response.Write(banner.GetBanner(-1)) %> Originally this code had no specific error checking as above so the ASP pages just failed with 500 error pages from the Web server. To find out what the problem is this code is more useful at least for debugging: <% ON ERROR RESUME NEXT Set banner = Server.CreateObject("wwBanner.aspBanner") Response.Write(err.Number & " - " & err.Description) banner.BannerFile = "wwsitebanners" Response.Write(banner.GetBanner(-1)) %> which results in: 429 - ActiveX component can't create object which at least gives you a slight clue. In ASP.NET invoking the same COM object with code like this: <% dynamic banner = wwUtils.CreateComInstance("wwBanner.aspBanner") as dynamic; banner.cBANNERFILE = "wwsitebanners"; Response.Write(banner.getBanner(-1)); %> results in: Retrieving the COM class factory for component with CLSID {B5DCBB81-D5F5-11D2-B85E-00600889F23B} failed due to the following error: 80040154 Class not registered (Exception from HRESULT: 0x80040154 (REGDB_E_CLASSNOTREG)). The class is in fact registered though and the COM server loads fine from a command prompt or other COM client. This error can be caused by a COM server that doesn't load. It looks like a COM registration error. There are a number of traditional reasons why this error can crop up of course. The server isn't registered (run regserver32 to register a DLL server or /regserver on an EXE server) Access permissions aren't set on the COM server (Web account has to be able to read the DLL ie. Network service) The COM server fails to load during initialization ie. failing during startup One thing I always do to check for COM errors fire up the server in a COM client outside of IIS and ensure that it works there first - it's almost always easier to debug a server outside of the Web environment. In my case I tried the server in Visual FoxPro on the server with: loBanners = CREATEOBJECT("wwBanner.aspBanner") loBanners.cBannerFile = "wwsitebanners" ? loBanners.GetBanner(-1) and it worked just fine. If you don't have a full dev environment on the server you can also use VBScript do the same thing and run the .vbs file from the command prompt: Set banner = Server.CreateObject("wwBanner.aspBanner") banner.BannerFile = "wwsitebanners" MsgBox(banner.getBanner(-1)) Since this both works it tells me the server is registered and working properly. This leaves startup failures or permissions as the problem. I double checked permissions for the Application Pool and the permissions of the folder where the DLL lives and both are properly set to allow access by the Application Pool impersonated user. Just to be sure I assigned an Admin user to the Application Pool but still no go. So now what? 64 bit Servers Ahoy A couple of weeks back I had set up a few of my Application pools to 64 bit mode. My server is Server 2008 64 bit and by default Application Pools run 64 bit. Originally when I installed the server I set up most of my Application Pools to 32 bit mainly for backwards compatibility. But as more of my code migrates to 64 bit OS's I figured it'd be a good idea to see how well code runs under 64 bit code. The transition has been mostly painless. Until today when I noticed the problem with the code above when scrolling to my IIS logs and noticing a lot of 500 errors on many of my ASP classic pages. The code in question in most of these pages deals with this single simple COM object. It took a while to figure out that the problem is caused by the Application Pool running in 64 bit mode. The issue is that 32 bit COM objects (ie. my old Visual FoxPro COM component) cannot be loaded in a 64 bit Application Pool. The ASP pages using this COM component broke on the day I switched my main Application Pool into 64 bit mode but I didn't find the problem until I searched my logs for errors by pure chance. To fix this is easy enough once you know what the problem is by switching the Application Pool to Enable 32-bit Applications: Once this is done the COM objects started working correctly again. 64 bit ASP and ASP.NET with DCOM Servers This is kind of off topic, but incidentally it's possible to load 32 bit DCOM (out of process) servers from ASP.NET and ASP classic even if those applications run in 64 bit application pools. In fact, in West Wind Web Connection I use this capability to run a 64 bit ASP.NET handler that talks to a 32 bit FoxPro COM server which allows West Wind Web Connection to run in native 64 bit mode without custom configuration (which is actually quite useful). It's probably not a common usage scenario but it's good to know that you can actually access 32 bit COM objects this way from ASP.NET. For West Wind Web Connection this works out well as the DCOM interface only makes one non-chatty call to the backend server that handles all the rest of the request processing. Application Pool Isolation is your Friend For me the recent incident of failure in the classic ASP pages has just been another reminder to be very careful with moving applications to 64 bit operation. There are many little traps when switching to 64 bit that are very difficult to track and test for. I described one issue I had a couple of months ago where one of the default ASP.NET filters was loading the wrong version (32bit instead of 64bit) which was extremely difficult to track down and was caused by a very sneaky configuration switch error (basically 3 different entries for the same ISAPI filter all with different bitness settings). It took me almost a full day to track this down). Recently I've been taken to isolate individual applications into separate Application Pools rather than my past practice of combining many apps into shared AppPools. This is a good practice assuming you have enough memory to make this work. Application Pool isolate provides more modularity and allows me to selectively move applications to 64 bit. The error above came about precisely because I moved one of my most populous app pools to 64 bit and forgot about the minimal COM object use in some of my old pages. It's easy to forget. To 64bit or Not Is it worth it to move to 64 bit? Currently I'd say -not really. In my - admittedly limited - testing I don't see any significant performance increases. In fact 64 bit apps just seem to consume considerably more memory (30-50% more in my pools on average) and performance is minimally improved (less than 5% at the very best) in the load testing I've performed on a couple of sites in both modes. The only real incentive for 64 bit would be applications that require huge data spaces that exceed the 32 bit 4 gigabyte memory limit. However I have a hard time imagining an application that needs 4 gigs of memory in a single Application Pool :-). Curious to hear other opinions on benefits of 64 bit operation. © Rick Strahl, West Wind Technologies, 2005-2011Posted in COM   ASP.NET  FoxPro  

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  • VS 2010 SP1 (Beta) and IIS Express

    - by ScottGu
    Last month we released the VS 2010 Service Pack 1 (SP1) Beta.  You can learn more about the VS 2010 SP1 Beta from Jason Zander’s two blog posts about it, and from Scott Hanselman’s blog post that covers some of the new capabilities enabled with it.  You can download and install the VS 2010 SP1 Beta here. IIS Express Earlier this summer I blogged about IIS Express.  IIS Express is a free version of IIS 7.5 that is optimized for developer scenarios.  We think it combines the ease of use of the ASP.NET Web Server (aka Cassini) currently built-into VS today with the full power of IIS.  Specifically: It’s lightweight and easy to install (less than 5Mb download and a quick install) It does not require an administrator account to run/debug applications from Visual Studio It enables a full web-server feature set – including SSL, URL Rewrite, and other IIS 7.x modules It supports and enables the same extensibility model and web.config file settings that IIS 7.x support It can be installed side-by-side with the full IIS web server as well as the ASP.NET Development Server (they do not conflict at all) It works on Windows XP and higher operating systems – giving you a full IIS 7.x developer feature-set on all Windows OS platforms IIS Express (like the ASP.NET Development Server) can be quickly launched to run a site from a directory on disk.  It does not require any registration/configuration steps. This makes it really easy to launch and run for development scenarios. Visual Studio 2010 SP1 adds support for IIS Express – and you can start to take advantage of this starting with last month’s VS 2010 SP1 Beta release. Downloading and Installing IIS Express IIS Express isn’t included as part of the VS 2010 SP1 Beta.  Instead it is a separate ~4MB download which you can download and install using this link (it uses WebPI to install it).  Once IIS Express is installed, VS 2010 SP1 will enable some additional IIS Express commands and dialog options that allow you to easily use it. Enabling IIS Express for Existing Projects Visual Studio today defaults to using the built-in ASP.NET Development Server (aka Cassini) when running ASP.NET Projects: Converting your existing projects to use IIS Express is really easy.  You can do this by opening up the project properties dialog of an existing project, and then by clicking the “web” tab within it and selecting the “Use IIS Express” checkbox. Or even simpler, just right-click on your existing project, and select the “Use IIS Express…” menu command: And now when you run or debug your project you’ll see that IIS Express now starts up and runs automatically as your web-server: You can optionally right-click on the IIS Express icon within your system tray to see/browse all of sites and applications running on it: Note that if you ever want to revert back to using the ASP.NET Development Server you can do this by right-clicking the project again and then select the “Use Visual Studio Development Server” option (or go into the project properties, click the web tab, and uncheck IIS Express).  This will revert back to the ASP.NET Development Server the next time you run the project. IIS Express Properties Visual Studio 2010 SP1 exposes several new IIS Express configuration options that you couldn’t previously set with the ASP.NET Development Server.  Some of these are exposed via the property grid of your project (select the project node in the solution explorer and then change them via the property window): For example, enabling something like SSL support (which is not possible with the ASP.NET Development Server) can now be done simply by changing the “SSL Enabled” property to “True”: Once this is done IIS Express will expose both an HTTP and HTTPS endpoint for the project that we can use: SSL Self Signed Certs IIS Express ships with a self-signed SSL cert that it installs as part of setup – which removes the need for you to install your own certificate to use SSL during development.  Once you change the above drop-down to enable SSL, you’ll be able to browse to your site with the appropriate https:// URL prefix and it will connect via SSL. One caveat with self-signed certificates, though, is that browsers (like IE) will go out of their way to warn you that they aren’t to be trusted: You can mark the certificate as trusted to avoid seeing dialogs like this – or just keep the certificate un-trusted and press the “continue” button when the browser warns you not to trust your local web server. Additional IIS Settings IIS Express uses its own per-user ApplicationHost.config file to configure default server behavior.  Because it is per-user, it can be configured by developers who do not have admin credentials – unlike the full IIS.  You can customize all IIS features and settings via it if you want ultimate server customization (for example: to use your own certificates for SSL instead of self-signed ones). We recommend storing all app specific settings for IIS and ASP.NET within the web.config file which is part of your project – since that makes deploying apps easier (since the settings can be copied with the application content).  IIS (since IIS 7) no longer uses the metabase, and instead uses the same web.config configuration files that ASP.NET has always supported – which makes xcopy/ftp based deployment much easier. Making IIS Express your Default Web Server Above we looked at how we can convert existing sites that use the ASP.NET Developer Web Server to instead use IIS Express.  You can configure Visual Studio to use IIS Express as the default web server for all new projects by clicking the Tools->Options menu  command and opening up the Projects and Solutions->Web Projects node with the Options dialog: Clicking the “Use IIS Express for new file-based web site and projects” checkbox will cause Visual Studio to use it for all new web site and projects. Summary We think IIS Express makes it even easier to build, run and test web applications.  It works with all versions of ASP.NET and supports all ASP.NET application types (including obviously both ASP.NET Web Forms and ASP.NET MVC applications).  Because IIS Express is based on the IIS 7.5 codebase, you have a full web-server feature-set that you can use.  This means you can build and run your applications just like they’ll work on a real production web-server.  In addition to supporting ASP.NET, IIS Express also supports Classic ASP and other file-types and extensions supported by IIS – which also makes it ideal for sites that combine a variety of different technologies. Best of all – you do not need to change any code to take advantage of it.  As you can see above, updating existing Visual Studio web projects to use it is trivial.  You can begin to take advantage of IIS Express today using the VS 2010 SP1 Beta. Hope this helps, Scott

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  • The Incremental Architect&rsquo;s Napkin - #5 - Design functions for extensibility and readability

    - by Ralf Westphal
    Originally posted on: http://geekswithblogs.net/theArchitectsNapkin/archive/2014/08/24/the-incremental-architectrsquos-napkin---5---design-functions-for.aspx The functionality of programs is entered via Entry Points. So what we´re talking about when designing software is a bunch of functions handling the requests represented by and flowing in through those Entry Points. Designing software thus consists of at least three phases: Analyzing the requirements to find the Entry Points and their signatures Designing the functionality to be executed when those Entry Points get triggered Implementing the functionality according to the design aka coding I presume, you´re familiar with phase 1 in some way. And I guess you´re proficient in implementing functionality in some programming language. But in my experience developers in general are not experienced in going through an explicit phase 2. “Designing functionality? What´s that supposed to mean?” you might already have thought. Here´s my definition: To design functionality (or functional design for short) means thinking about… well, functions. You find a solution for what´s supposed to happen when an Entry Point gets triggered in terms of functions. A conceptual solution that is, because those functions only exist in your head (or on paper) during this phase. But you may have guess that, because it´s “design” not “coding”. And here is, what functional design is not: It´s not about logic. Logic is expressions (e.g. +, -, && etc.) and control statements (e.g. if, switch, for, while etc.). Also I consider calling external APIs as logic. It´s equally basic. It´s what code needs to do in order to deliver some functionality or quality. Logic is what´s doing that needs to be done by software. Transformations are either done through expressions or API-calls. And then there is alternative control flow depending on the result of some expression. Basically it´s just jumps in Assembler, sometimes to go forward (if, switch), sometimes to go backward (for, while, do). But calling your own function is not logic. It´s not necessary to produce any outcome. Functionality is not enhanced by adding functions (subroutine calls) to your code. Nor is quality increased by adding functions. No performance gain, no higher scalability etc. through functions. Functions are not relevant to functionality. Strange, isn´t it. What they are important for is security of investment. By introducing functions into our code we can become more productive (re-use) and can increase evolvability (higher unterstandability, easier to keep code consistent). That´s no small feat, however. Evolvable code can hardly be overestimated. That´s why to me functional design is so important. It´s at the core of software development. To sum this up: Functional design is on a level of abstraction above (!) logical design or algorithmic design. Functional design is only done until you get to a point where each function is so simple you are very confident you can easily code it. Functional design an logical design (which mostly is coding, but can also be done using pseudo code or flow charts) are complementary. Software needs both. If you start coding right away you end up in a tangled mess very quickly. Then you need back out through refactoring. Functional design on the other hand is bloodless without actual code. It´s just a theory with no experiments to prove it. But how to do functional design? An example of functional design Let´s assume a program to de-duplicate strings. The user enters a number of strings separated by commas, e.g. a, b, a, c, d, b, e, c, a. And the program is supposed to clear this list of all doubles, e.g. a, b, c, d, e. There is only one Entry Point to this program: the user triggers the de-duplication by starting the program with the string list on the command line C:\>deduplicate "a, b, a, c, d, b, e, c, a" a, b, c, d, e …or by clicking on a GUI button. This leads to the Entry Point function to get called. It´s the program´s main function in case of the batch version or a button click event handler in the GUI version. That´s the physical Entry Point so to speak. It´s inevitable. What then happens is a three step process: Transform the input data from the user into a request. Call the request handler. Transform the output of the request handler into a tangible result for the user. Or to phrase it a bit more generally: Accept input. Transform input into output. Present output. This does not mean any of these steps requires a lot of effort. Maybe it´s just one line of code to accomplish it. Nevertheless it´s a distinct step in doing the processing behind an Entry Point. Call it an aspect or a responsibility - and you will realize it most likely deserves a function of its own to satisfy the Single Responsibility Principle (SRP). Interestingly the above list of steps is already functional design. There is no logic, but nevertheless the solution is described - albeit on a higher level of abstraction than you might have done yourself. But it´s still on a meta-level. The application to the domain at hand is easy, though: Accept string list from command line De-duplicate Present de-duplicated strings on standard output And this concrete list of processing steps can easily be transformed into code:static void Main(string[] args) { var input = Accept_string_list(args); var output = Deduplicate(input); Present_deduplicated_string_list(output); } Instead of a big problem there are three much smaller problems now. If you think each of those is trivial to implement, then go for it. You can stop the functional design at this point. But maybe, just maybe, you´re not so sure how to go about with the de-duplication for example. Then just implement what´s easy right now, e.g.private static string Accept_string_list(string[] args) { return args[0]; } private static void Present_deduplicated_string_list( string[] output) { var line = string.Join(", ", output); Console.WriteLine(line); } Accept_string_list() contains logic in the form of an API-call. Present_deduplicated_string_list() contains logic in the form of an expression and an API-call. And then repeat the functional design for the remaining processing step. What´s left is the domain logic: de-duplicating a list of strings. How should that be done? Without any logic at our disposal during functional design you´re left with just functions. So which functions could make up the de-duplication? Here´s a suggestion: De-duplicate Parse the input string into a true list of strings. Register each string in a dictionary/map/set. That way duplicates get cast away. Transform the data structure into a list of unique strings. Processing step 2 obviously was the core of the solution. That´s where real creativity was needed. That´s the core of the domain. But now after this refinement the implementation of each step is easy again:private static string[] Parse_string_list(string input) { return input.Split(',') .Select(s => s.Trim()) .ToArray(); } private static Dictionary<string,object> Compile_unique_strings(string[] strings) { return strings.Aggregate( new Dictionary<string, object>(), (agg, s) => { agg[s] = null; return agg; }); } private static string[] Serialize_unique_strings( Dictionary<string,object> dict) { return dict.Keys.ToArray(); } With these three additional functions Main() now looks like this:static void Main(string[] args) { var input = Accept_string_list(args); var strings = Parse_string_list(input); var dict = Compile_unique_strings(strings); var output = Serialize_unique_strings(dict); Present_deduplicated_string_list(output); } I think that´s very understandable code: just read it from top to bottom and you know how the solution to the problem works. It´s a mirror image of the initial design: Accept string list from command line Parse the input string into a true list of strings. Register each string in a dictionary/map/set. That way duplicates get cast away. Transform the data structure into a list of unique strings. Present de-duplicated strings on standard output You can even re-generate the design by just looking at the code. Code and functional design thus are always in sync - if you follow some simple rules. But about that later. And as a bonus: all the functions making up the process are small - which means easy to understand, too. So much for an initial concrete example. Now it´s time for some theory. Because there is method to this madness ;-) The above has only scratched the surface. Introducing Flow Design Functional design starts with a given function, the Entry Point. Its goal is to describe the behavior of the program when the Entry Point is triggered using a process, not an algorithm. An algorithm consists of logic, a process on the other hand consists just of steps or stages. Each processing step transforms input into output or a side effect. Also it might access resources, e.g. a printer, a database, or just memory. Processing steps thus can rely on state of some sort. This is different from Functional Programming, where functions are supposed to not be stateful and not cause side effects.[1] In its simplest form a process can be written as a bullet point list of steps, e.g. Get data from user Output result to user Transform data Parse data Map result for output Such a compilation of steps - possibly on different levels of abstraction - often is the first artifact of functional design. It can be generated by a team in an initial design brainstorming. Next comes ordering the steps. What should happen first, what next etc.? Get data from user Parse data Transform data Map result for output Output result to user That´s great for a start into functional design. It´s better than starting to code right away on a given function using TDD. Please get me right: TDD is a valuable practice. But it can be unnecessarily hard if the scope of a functionn is too large. But how do you know beforehand without investing some thinking? And how to do this thinking in a systematic fashion? My recommendation: For any given function you´re supposed to implement first do a functional design. Then, once you´re confident you know the processing steps - which are pretty small - refine and code them using TDD. You´ll see that´s much, much easier - and leads to cleaner code right away. For more information on this approach I call “Informed TDD” read my book of the same title. Thinking before coding is smart. And writing down the solution as a bunch of functions possibly is the simplest thing you can do, I´d say. It´s more according to the KISS (Keep It Simple, Stupid) principle than returning constants or other trivial stuff TDD development often is started with. So far so good. A simple ordered list of processing steps will do to start with functional design. As shown in the above example such steps can easily be translated into functions. Moving from design to coding thus is simple. However, such a list does not scale. Processing is not always that simple to be captured in a list. And then the list is just text. Again. Like code. That means the design is lacking visuality. Textual representations need more parsing by your brain than visual representations. Plus they are limited in their “dimensionality”: text just has one dimension, it´s sequential. Alternatives and parallelism are hard to encode in text. In addition the functional design using numbered lists lacks data. It´s not visible what´s the input, output, and state of the processing steps. That´s why functional design should be done using a lightweight visual notation. No tool is necessary to draw such designs. Use pen and paper; a flipchart, a whiteboard, or even a napkin is sufficient. Visualizing processes The building block of the functional design notation is a functional unit. I mostly draw it like this: Something is done, it´s clear what goes in, it´s clear what comes out, and it´s clear what the processing step requires in terms of state or hardware. Whenever input flows into a functional unit it gets processed and output is produced and/or a side effect occurs. Flowing data is the driver of something happening. That´s why I call this approach to functional design Flow Design. It´s about data flow instead of control flow. Control flow like in algorithms is of no concern to functional design. Thinking about control flow simply is too low level. Once you start with control flow you easily get bogged down by tons of details. That´s what you want to avoid during design. Design is supposed to be quick, broad brush, abstract. It should give overview. But what about all the details? As Robert C. Martin rightly said: “Programming is abot detail”. Detail is a matter of code. Once you start coding the processing steps you designed you can worry about all the detail you want. Functional design does not eliminate all the nitty gritty. It just postpones tackling them. To me that´s also an example of the SRP. Function design has the responsibility to come up with a solution to a problem posed by a single function (Entry Point). And later coding has the responsibility to implement the solution down to the last detail (i.e. statement, API-call). TDD unfortunately mixes both responsibilities. It´s just coding - and thereby trying to find detailed implementations (green phase) plus getting the design right (refactoring). To me that´s one reason why TDD has failed to deliver on its promise for many developers. Using functional units as building blocks of functional design processes can be depicted very easily. Here´s the initial process for the example problem: For each processing step draw a functional unit and label it. Choose a verb or an “action phrase” as a label, not a noun. Functional design is about activities, not state or structure. Then make the output of an upstream step the input of a downstream step. Finally think about the data that should flow between the functional units. Write the data above the arrows connecting the functional units in the direction of the data flow. Enclose the data description in brackets. That way you can clearly see if all flows have already been specified. Empty brackets mean “no data is flowing”, but nevertheless a signal is sent. A name like “list” or “strings” in brackets describes the data content. Use lower case labels for that purpose. A name starting with an upper case letter like “String” or “Customer” on the other hand signifies a data type. If you like, you also can combine descriptions with data types by separating them with a colon, e.g. (list:string) or (strings:string[]). But these are just suggestions from my practice with Flow Design. You can do it differently, if you like. Just be sure to be consistent. Flows wired-up in this manner I call one-dimensional (1D). Each functional unit just has one input and/or one output. A functional unit without an output is possible. It´s like a black hole sucking up input without producing any output. Instead it produces side effects. A functional unit without an input, though, does make much sense. When should it start to work? What´s the trigger? That´s why in the above process even the first processing step has an input. If you like, view such 1D-flows as pipelines. Data is flowing through them from left to right. But as you can see, it´s not always the same data. It get´s transformed along its passage: (args) becomes a (list) which is turned into (strings). The Principle of Mutual Oblivion A very characteristic trait of flows put together from function units is: no functional units knows another one. They are all completely independent of each other. Functional units don´t know where their input is coming from (or even when it´s gonna arrive). They just specify a range of values they can process. And they promise a certain behavior upon input arriving. Also they don´t know where their output is going. They just produce it in their own time independent of other functional units. That means at least conceptually all functional units work in parallel. Functional units don´t know their “deployment context”. They now nothing about the overall flow they are place in. They are just consuming input from some upstream, and producing output for some downstream. That makes functional units very easy to test. At least as long as they don´t depend on state or resources. I call this the Principle of Mutual Oblivion (PoMO). Functional units are oblivious of others as well as an overall context/purpose. They are just parts of a whole focused on a single responsibility. How the whole is built, how a larger goal is achieved, is of no concern to the single functional units. By building software in such a manner, functional design interestingly follows nature. Nature´s building blocks for organisms also follow the PoMO. The cells forming your body do not know each other. Take a nerve cell “controlling” a muscle cell for example:[2] The nerve cell does not know anything about muscle cells, let alone the specific muscel cell it is “attached to”. Likewise the muscle cell does not know anything about nerve cells, let a lone a specific nerve cell “attached to” it. Saying “the nerve cell is controlling the muscle cell” thus only makes sense when viewing both from the outside. “Control” is a concept of the whole, not of its parts. Control is created by wiring-up parts in a certain way. Both cells are mutually oblivious. Both just follow a contract. One produces Acetylcholine (ACh) as output, the other consumes ACh as input. Where the ACh is going, where it´s coming from neither cell cares about. Million years of evolution have led to this kind of division of labor. And million years of evolution have produced organism designs (DNA) which lead to the production of these different cell types (and many others) and also to their co-location. The result: the overall behavior of an organism. How and why this happened in nature is a mystery. For our software, though, it´s clear: functional and quality requirements needs to be fulfilled. So we as developers have to become “intelligent designers” of “software cells” which we put together to form a “software organism” which responds in satisfying ways to triggers from it´s environment. My bet is: If nature gets complex organisms working by following the PoMO, who are we to not apply this recipe for success to our much simpler “machines”? So my rule is: Wherever there is functionality to be delivered, because there is a clear Entry Point into software, design the functionality like nature would do it. Build it from mutually oblivious functional units. That´s what Flow Design is about. In that way it´s even universal, I´d say. Its notation can also be applied to biology: Never mind labeling the functional units with nouns. That´s ok in Flow Design. You´ll do that occassionally for functional units on a higher level of abstraction or when their purpose is close to hardware. Getting a cockroach to roam your bedroom takes 1,000,000 nerve cells (neurons). Getting the de-duplication program to do its job just takes 5 “software cells” (functional units). Both, though, follow the same basic principle. Translating functional units into code Moving from functional design to code is no rocket science. In fact it´s straightforward. There are two simple rules: Translate an input port to a function. Translate an output port either to a return statement in that function or to a function pointer visible to that function. The simplest translation of a functional unit is a function. That´s what you saw in the above example. Functions are mutually oblivious. That why Functional Programming likes them so much. It makes them composable. Which is the reason, nature works according to the PoMO. Let´s be clear about one thing: There is no dependency injection in nature. For all of an organism´s complexity no DI container is used. Behavior is the result of smooth cooperation between mutually oblivious building blocks. Functions will often be the adequate translation for the functional units in your designs. But not always. Take for example the case, where a processing step should not always produce an output. Maybe the purpose is to filter input. Here the functional unit consumes words and produces words. But it does not pass along every word flowing in. Some words are swallowed. Think of a spell checker. It probably should not check acronyms for correctness. There are too many of them. Or words with no more than two letters. Such words are called “stop words”. In the above picture the optionality of the output is signified by the astrisk outside the brackets. It means: Any number of (word) data items can flow from the functional unit for each input data item. It might be none or one or even more. This I call a stream of data. Such behavior cannot be translated into a function where output is generated with return. Because a function always needs to return a value. So the output port is translated into a function pointer or continuation which gets passed to the subroutine when called:[3]void filter_stop_words( string word, Action<string> onNoStopWord) { if (...check if not a stop word...) onNoStopWord(word); } If you want to be nitpicky you might call such a function pointer parameter an injection. And technically you´re right. Conceptually, though, it´s not an injection. Because the subroutine is not functionally dependent on the continuation. Firstly continuations are procedures, i.e. subroutines without a return type. Remember: Flow Design is about unidirectional data flow. Secondly the name of the formal parameter is chosen in a way as to not assume anything about downstream processing steps. onNoStopWord describes a situation (or event) within the functional unit only. Translating output ports into function pointers helps keeping functional units mutually oblivious in cases where output is optional or produced asynchronically. Either pass the function pointer to the function upon call. Or make it global by putting it on the encompassing class. Then it´s called an event. In C# that´s even an explicit feature.class Filter { public void filter_stop_words( string word) { if (...check if not a stop word...) onNoStopWord(word); } public event Action<string> onNoStopWord; } When to use a continuation and when to use an event dependens on how a functional unit is used in flows and how it´s packed together with others into classes. You´ll see examples further down the Flow Design road. Another example of 1D functional design Let´s see Flow Design once more in action using the visual notation. How about the famous word wrap kata? Robert C. Martin has posted a much cited solution including an extensive reasoning behind his TDD approach. So maybe you want to compare it to Flow Design. The function signature given is:string WordWrap(string text, int maxLineLength) {...} That´s not an Entry Point since we don´t see an application with an environment and users. Nevertheless it´s a function which is supposed to provide a certain functionality. The text passed in has to be reformatted. The input is a single line of arbitrary length consisting of words separated by spaces. The output should consist of one or more lines of a maximum length specified. If a word is longer than a the maximum line length it can be split in multiple parts each fitting in a line. Flow Design Let´s start by brainstorming the process to accomplish the feat of reformatting the text. What´s needed? Words need to be assembled into lines Words need to be extracted from the input text The resulting lines need to be assembled into the output text Words too long to fit in a line need to be split Does sound about right? I guess so. And it shows a kind of priority. Long words are a special case. So maybe there is a hint for an incremental design here. First let´s tackle “average words” (words not longer than a line). Here´s the Flow Design for this increment: The the first three bullet points turned into functional units with explicit data added. As the signature requires a text is transformed into another text. See the input of the first functional unit and the output of the last functional unit. In between no text flows, but words and lines. That´s good to see because thereby the domain is clearly represented in the design. The requirements are talking about words and lines and here they are. But note the asterisk! It´s not outside the brackets but inside. That means it´s not a stream of words or lines, but lists or sequences. For each text a sequence of words is output. For each sequence of words a sequence of lines is produced. The asterisk is used to abstract from the concrete implementation. Like with streams. Whether the list of words gets implemented as an array or an IEnumerable is not important during design. It´s an implementation detail. Does any processing step require further refinement? I don´t think so. They all look pretty “atomic” to me. And if not… I can always backtrack and refine a process step using functional design later once I´ve gained more insight into a sub-problem. Implementation The implementation is straightforward as you can imagine. The processing steps can all be translated into functions. Each can be tested easily and separately. Each has a focused responsibility. And the process flow becomes just a sequence of function calls: Easy to understand. It clearly states how word wrapping works - on a high level of abstraction. And it´s easy to evolve as you´ll see. Flow Design - Increment 2 So far only texts consisting of “average words” are wrapped correctly. Words not fitting in a line will result in lines too long. Wrapping long words is a feature of the requested functionality. Whether it´s there or not makes a difference to the user. To quickly get feedback I decided to first implement a solution without this feature. But now it´s time to add it to deliver the full scope. Fortunately Flow Design automatically leads to code following the Open Closed Principle (OCP). It´s easy to extend it - instead of changing well tested code. How´s that possible? Flow Design allows for extension of functionality by inserting functional units into the flow. That way existing functional units need not be changed. The data flow arrow between functional units is a natural extension point. No need to resort to the Strategy Pattern. No need to think ahead where extions might need to be made in the future. I just “phase in” the remaining processing step: Since neither Extract words nor Reformat know of their environment neither needs to be touched due to the “detour”. The new processing step accepts the output of the existing upstream step and produces data compatible with the existing downstream step. Implementation - Increment 2 A trivial implementation checking the assumption if this works does not do anything to split long words. The input is just passed on: Note how clean WordWrap() stays. The solution is easy to understand. A developer looking at this code sometime in the future, when a new feature needs to be build in, quickly sees how long words are dealt with. Compare this to Robert C. Martin´s solution:[4] How does this solution handle long words? Long words are not even part of the domain language present in the code. At least I need considerable time to understand the approach. Admittedly the Flow Design solution with the full implementation of long word splitting is longer than Robert C. Martin´s. At least it seems. Because his solution does not cover all the “word wrap situations” the Flow Design solution handles. Some lines would need to be added to be on par, I guess. But even then… Is a difference in LOC that important as long as it´s in the same ball park? I value understandability and openness for extension higher than saving on the last line of code. Simplicity is not just less code, it´s also clarity in design. But don´t take my word for it. Try Flow Design on larger problems and compare for yourself. What´s the easier, more straightforward way to clean code? And keep in mind: You ain´t seen all yet ;-) There´s more to Flow Design than described in this chapter. In closing I hope I was able to give you a impression of functional design that makes you hungry for more. To me it´s an inevitable step in software development. Jumping from requirements to code does not scale. And it leads to dirty code all to quickly. Some thought should be invested first. Where there is a clear Entry Point visible, it´s functionality should be designed using data flows. Because with data flows abstraction is possible. For more background on why that´s necessary read my blog article here. For now let me point out to you - if you haven´t already noticed - that Flow Design is a general purpose declarative language. It´s “programming by intention” (Shalloway et al.). Just write down how you think the solution should work on a high level of abstraction. This breaks down a large problem in smaller problems. And by following the PoMO the solutions to those smaller problems are independent of each other. So they are easy to test. Or you could even think about getting them implemented in parallel by different team members. Flow Design not only increases evolvability, but also helps becoming more productive. All team members can participate in functional design. This goes beyon collective code ownership. We´re talking collective design/architecture ownership. Because with Flow Design there is a common visual language to talk about functional design - which is the foundation for all other design activities.   PS: If you like what you read, consider getting my ebook “The Incremental Architekt´s Napkin”. It´s where I compile all the articles in this series for easier reading. I like the strictness of Function Programming - but I also find it quite hard to live by. And it certainly is not what millions of programmers are used to. Also to me it seems, the real world is full of state and side effects. So why give them such a bad image? That´s why functional design takes a more pragmatic approach. State and side effects are ok for processing steps - but be sure to follow the SRP. Don´t put too much of it into a single processing step. ? Image taken from www.physioweb.org ? My code samples are written in C#. C# sports typed function pointers called delegates. Action is such a function pointer type matching functions with signature void someName(T t). Other languages provide similar ways to work with functions as first class citizens - even Java now in version 8. I trust you find a way to map this detail of my translation to your favorite programming language. I know it works for Java, C++, Ruby, JavaScript, Python, Go. And if you´re using a Functional Programming language it´s of course a no brainer. ? Taken from his blog post “The Craftsman 62, The Dark Path”. ?

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  • Use an Ubuntu Live CD to Securely Wipe Your PC’s Hard Drive

    - by Trevor Bekolay
    Deleting files or quickly formatting a drive isn’t enough for sensitive personal information. We’ll show you how to get rid of it for good using a Ubuntu Live CD. When you delete a file in Windows, Ubuntu, or any other operating system, it doesn’t actually destroy the data stored on your hard drive, it just marks that data as “deleted.” If you overwrite it later, then that data is generally unrecoverable, but if the operating system don’t happen to overwrite it, then your data is still stored on your hard drive, recoverable by anyone who has the right software. By securely delete files or entire hard drives, your data will be gone for good. Note: Modern hard drives are extremely sophisticated, as are the experts who recover data for a living. There is no guarantee that the methods covered in this article will make your data completely unrecoverable; however, they will make your data unrecoverable to the majority of recovery methods, and all methods that are readily available to the general public. Shred individual files Most of the data stored on your hard drive is harmless, and doesn’t reveal anything about you. If there are just a few files that you know you don’t want someone else to see, then the easiest way to get rid of them is a built-in Linux utility called shred. Open a terminal window by clicking on Applications at the top-left of the screen, then expanding the Accessories menu and clicking on Terminal. Navigate to the file that you want to delete using cd to change directories and ls to list the files and folders in the current directory. As an example, we’ve got a file called BankInfo.txt on a Windows NTFS-formatted hard drive. We want to delete it securely, so we’ll call shred by entering the following in the terminal window: shred <file> which is, in our example: shred BankInfo.txt Notice that our BankInfo.txt file still exists, even though we’ve shredded it. A quick look at the contents of BankInfo.txt make it obvious that the file has indeed been securely overwritten. We can use some command-line arguments to make shred delete the file from the hard drive as well. We can also be extra-careful about the shredding process by upping the number of times shred overwrites the original file. To do this, in the terminal, type in: shred –remove –iterations=<num> <file> By default, shred overwrites the file 25 times. We’ll double this, giving us the following command: shred –remove –iterations=50 BankInfo.txt BankInfo.txt has now been securely wiped on the physical disk, and also no longer shows up in the directory listing. Repeat this process for any sensitive files on your hard drive! Wipe entire hard drives If you’re disposing of an old hard drive, or giving it to someone else, then you might instead want to wipe your entire hard drive. shred can be invoked on hard drives, but on modern file systems, the shred process may be reversible. We’ll use the program wipe to securely delete all of the data on a hard drive. Unlike shred, wipe is not included in Ubuntu by default, so we have to install it. Open up the Synaptic Package Manager by clicking on System in the top-left corner of the screen, then expanding the Administration folder and clicking on Synaptic Package Manager. wipe is part of the Universe repository, which is not enabled by default. We’ll enable it by clicking on Settings > Repositories in the Synaptic Package Manager window. Check the checkbox next to “Community-maintained Open Source software (universe)”. Click Close. You’ll need to reload Synaptic’s package list. Click on the Reload button in the main Synaptic Package Manager window. Once the package list has been reloaded, the text over the search field will change to “Rebuilding search index”. Wait until it reads “Quick search,” and then type “wipe” into the search field. The wipe package should come up, along with some other packages that perform similar functions. Click on the checkbox to the left of the label “wipe” and select “Mark for Installation”. Click on the Apply button to start the installation process. Click the Apply button on the Summary window that pops up. Once the installation is done, click the Close button and close the Synaptic Package Manager window. Open a terminal window by clicking on Applications in the top-left of the screen, then Accessories > Terminal. You need to figure our the correct hard drive to wipe. If you wipe the wrong hard drive, that data will not be recoverable, so exercise caution! In the terminal window, type in: sudo fdisk -l A list of your hard drives will show up. A few factors will help you identify the right hard drive. One is the file system, found in the System column of  the list – Windows hard drives are usually formatted as NTFS (which shows up as HPFS/NTFS). Another good identifier is the size of the hard drive, which appears after its identifier (highlighted in the following screenshot). In our case, the hard drive we want to wipe is only around 1 GB large, and is formatted as NTFS. We make a note of the label found under the the Device column heading. If you have multiple partitions on this hard drive, then there will be more than one device in this list. The wipe developers recommend wiping each partition separately. To start the wiping process, type the following into the terminal: sudo wipe <device label> In our case, this is: sudo wipe /dev/sda1 Again, exercise caution – this is the point of no return! Your hard drive will be completely wiped. It may take some time to complete, depending on the size of the drive you’re wiping. Conclusion If you have sensitive information on your hard drive – and chances are you probably do – then it’s a good idea to securely delete sensitive files before you give away or dispose of your hard drive. The most secure way to delete your data is with a few swings of a hammer, but shred and wipe from a Ubuntu Live CD is a good alternative! Similar Articles Productive Geek Tips Reset Your Ubuntu Password Easily from the Live CDScan a Windows PC for Viruses from a Ubuntu Live CDRecover Deleted Files on an NTFS Hard Drive from a Ubuntu Live CDCreate a Bootable Ubuntu 9.10 USB Flash DriveCreate a Bootable Ubuntu USB Flash Drive the Easy Way TouchFreeze Alternative in AutoHotkey The Icy Undertow Desktop Windows Home Server – Backup to LAN The Clear & Clean Desktop Use This Bookmarklet to Easily Get Albums Use AutoHotkey to Assign a Hotkey to a Specific Window Latest Software Reviews Tinyhacker Random Tips DVDFab 6 Revo Uninstaller Pro Registry Mechanic 9 for Windows PC Tools Internet Security Suite 2010 Office 2010 Product Guides Google Maps Place marks – Pizza, Guns or Strip Clubs Monitor Applications With Kiwi LocPDF is a Visual PDF Search Tool Download Free iPad Wallpapers at iPad Decor Get Your Delicious Bookmarks In Firefox’s Awesome Bar

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  • Turn A Flash Drive Into a Portable Web Server

    - by Matthew Guay
    Portable applications are very useful for getting work done on the go, but how about portable servers?  Here’s how you can turn your flash drive into a portable web server. Getting Started To put a full web server on our flash drive, we’re going to use XAMPP Lite.  This lightweight, preconfigured server includes recent versions of Apache, MySQL, and PHP so you can run most websites and webapps directly from it.  You could use the full XAMPP, which includes more features such as a FileZilla FTP server and OpenSSL, but for most purposes, the light version is plenty for a portable server. Download the latest version of XAMPP Lite (link below).  In this tutorial, we used the self-extracting EXE version; you could choose the ZIP file and extract the files yourself, but we found it easier to use the executable. Run the installer, and click Browse choose where to install your server. Select your flash drive, or a folder in it, and click Ok.  Make sure your flash drive has at least 250MB of available storage space.  XAMPP will create an xampplite folder and store all the files in it during the installation.   Click Install, and all of the files will be extracted to your flash drive.  This may take a few moments depending on your flash drive’s speed. When the extraction process is finished, a Command Prompt window will open to finish the installation.  The first prompt will ask if you want to add shortcuts to the start menu and desktop; enter “n” since we don’t want to create start menu links to our portable server. Now enter “y” to configure XAMPP’s directories automatically. Finally, enter “y” to make XAMPP fully portable.  It will set up the servers to run without specific drive letters so your server will run from any computer. XAMPP will finalize your changes; press Enter when everything is completed. Setup will automatically launch the command line version of XAMPP.  On first run, confirm that your time zone is correct. And that’s it!  You can now run XAMPP’s control panel by entering 1, or you can exit and run XAMPP from any other computer with your flash drive. To complete your portable webserver kit, you may want to install Portable Firefox or Iron Browser on your flash drive so you always have your favorite browser ready to use. Running your portable server Using your portable server is very simple.  Open the xampplite folder on your flash drive and launch xampp-control.exe. Click Start beside Apache and MySql to get your webserver running. Please note: Do not check the Svc box, as this will run the server as a Windows service.  To keep XAMPP portable, you do not want it running as a service! Windows Firewall may prompt you that it blocked the server; click Allow access to let your server run. Once they’re running, you can click Admin to open the default XAMPP admin page running from your local webserver.  Or, you can view it by browsing to http://localhost/ or http://127.0.0.1/ in your browser. If everything is working correctly, you should see this page in your browser.  Choose your default language… And then you’ll see the default XAMPP admin page.   Click the Status link on the left sidebar to make sure everything is running correctly. If you click the Admin button for MySql in the XAMPP Control Panel, it will open phpMyAdmin in your default browser.  Alternately, you can open the MySql admin page by entering http://localhost/phpmyadmin/ or http://127.0.0.1/phpmyadmin/ in your favorite browser. Now you can add your own webpages to your webserver.  Save all of your web files in the \xampplight\htdocs\ folder on your flash drive. Install WordPress in your portable server Since XAMPP Lite includes MySql and PHP, you can even run webapps such as WordPress, the popular CMS and blogging platform.  Download WordPress (link below), and extract the files to the \xampplite\htdocs folder on your flash drive. Now all of the WordPress files are stored in \xampplite\htdocs\wordpress on your flash drive. We still need to setup WordPress on our portable server.  Open your MySql admin page http://localhost/phpmyadmin/ to create a new database for WordPress.  Enter a name for your database in the “Create new database” box, and click Create. Click the Privileges tab on the top, and the select “Add a new User”.   Enter a username and password for the database, and then click the Go button on the bottom of the page. Using WordPress Now, in your browser, enter http://localhost/wordpress/wp-admin/install.php.  Click Create a Configuration File to continue. Make sure you have your Database name, username, and password we created previously, and click “Let’s Go!” Enter your WordPress database name, username, and password, leave the other two entries as default, and click Submit. You should now have the database all ready to go.  Click “Run the install” to finish installing WordPress. Enter a title, username, and password for your test blog, as well as your email address, and then click “Install WordPress”. You now have a portable install of WordPress.  Click “Log In” to  access your WordPress admin page. Enter your username and password, and click Log In. Here you can add pages, posts, themes, extensions, and anything else just like you would on a normal WordPress site.  This is a great way to experiment with WordPress without messing up your real website. You can view your portable WordPress site by entering http://localhost/wordpress/ in your address bar. Closing your server When you’re done running your test server, click the Stop button on each of the services and then click the Exit button in the XAMPP control panel.  If you press the exit button on the top of the window, it will just minimize the control panel to the tray.   Alternately, you can shutdown your server by running xampp_stop.exe from your xampplite folder. Conclusion XAMPP Lite gives you a great way to run a full webserver directly from your flash drive.  Now, anywhere you go, you can test and tweak your webpages and webapps from any Windows computer.  Links Download XAMPP Lite Download WordPress Similar Articles Productive Geek Tips BitLocker To Go Encrypts Portable Flash Drives in Windows 7How To Use BitLocker on Drives without TPMSpeed up Your Windows Vista Computer with ReadyBoostView and Manage Flash Cookies the Easy WayInstall and Run Applications from Your iPod, Flash Drive or Mp3 Player TouchFreeze Alternative in AutoHotkey The Icy Undertow Desktop Windows Home Server – Backup to LAN The Clear & Clean Desktop Use This Bookmarklet to Easily Get Albums Use AutoHotkey to Assign a Hotkey to a Specific Window Latest Software Reviews Tinyhacker Random Tips DVDFab 6 Revo Uninstaller Pro Registry Mechanic 9 for Windows PC Tools Internet Security Suite 2010 OutlookStatView Scans and Displays General Usage Statistics How to Add Exceptions to the Windows Firewall Office 2010 reviewed in depth by Ed Bott FoxClocks adds World Times in your Statusbar (Firefox) Have Fun Editing Photo Editing with Citrify Outlook Connector Upgrade Error

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  • Interesting articles and blogs on SPARC T4

    - by mv
    Interesting articles and blogs on SPARC T4 processor   I have consolidated all the interesting information I could get on SPARC T4 processor and its hardware cryptographic capabilities.  Hope its useful. 1. Advantages of SPARC T4 processor  Most important points in this T4 announcement are : "The SPARC T4 processor was designed from the ground up for high speed security and has a cryptographic stream processing unit (SPU) integrated directly into each processor core. These accelerators support 16 industry standard security ciphers and enable high speed encryption at rates 3 to 5 times that of competing processors. By integrating encryption capabilities directly inside the instruction pipeline, the SPARC T4 processor eliminates the performance and cost barriers typically associated with secure computing and makes it possible to deliver high security levels without impacting the user experience." Data Sheet has more details on these  : "New on-chip Encryption Instruction Accelerators with direct non-privileged support for 16 industry-standard cryptographic algorithms plus random number generation in each of the eight cores: AES, Camellia, CRC32c, DES, 3DES, DH, DSA, ECC, Kasumi, MD5, RSA, SHA-1, SHA-224, SHA-256, SHA-384, SHA-512" I ran "isainfo -v" command on Solaris 11 Sparc T4-1 system. It shows the new instructions as expected  : $ isainfo -v 64-bit sparcv9 applications crc32c cbcond pause mont mpmul sha512 sha256 sha1 md5 camellia kasumi des aes ima hpc vis3 fmaf asi_blk_init vis2 vis popc 32-bit sparc applications crc32c cbcond pause mont mpmul sha512 sha256 sha1 md5 camellia kasumi des aes ima hpc vis3 fmaf asi_blk_init vis2 vis popc v8plus div32 mul32  2.  Dan Anderson's Blog have some interesting points about how these can be used : "New T4 crypto instructions include: aes_kexpand0, aes_kexpand1, aes_kexpand2,         aes_eround01, aes_eround23, aes_eround01_l, aes_eround_23_l, aes_dround01, aes_dround23, aes_dround01_l, aes_dround_23_l.       Having SPARC T4 hardware crypto instructions is all well and good, but how do we access it ?      The software is available with Solaris 11 and is used automatically if you are running Solaris a SPARC T4.  It is used internally in the kernel through kernel crypto modules.  It is available in user space through the PKCS#11 library." 3.   Dans' Blog on Where's the Crypto Libraries? Although this was written in 2009 but still is very useful  "Here's a brief tour of the major crypto libraries shown in the digraph:   The libpkcs11 library contains the PKCS#11 API (C_\*() functions, such as C_Initialize()). That in turn calls library pkcs11_softtoken or pkcs11_kernel, for userland or kernel crypto providers. The latter is used mostly for hardware-assisted cryptography (such as n2cp for Niagara2 SPARC processors), as that is performed more efficiently in kernel space with the "kCF" module (Kernel Crypto Framework). Additionally, for Solaris 10, strong crypto algorithms were split off in separate libraries, pkcs11_softtoken_extra libcryptoutil contains low-level utility functions to help implement cryptography. libsoftcrypto (OpenSolaris and Solaris Nevada only) implements several symmetric-key crypto algorithms in software, such as AES, RC4, and DES3, and the bignum library (used for RSA). libmd implements MD5, SHA, and SHA2 message digest algorithms" 4. Difference in T3 and T4 Diagram in this blog is good and self explanatory. Jeff's blog also highlights the differences  "The T4 servers have improved crypto acceleration, described at https://blogs.oracle.com/DanX/entry/sparc_t4_openssl_engine. It is "just built in" so administrators no longer have to assign crypto accelerator units to domains - it "just happens". Every physical or virtual CPU on a SPARC-T4 has full access to hardware based crypto acceleration at all times. .... For completeness sake, it's worth noting that the T4 adds more crypto algorithms, and accelerates Camelia, CRC32c, and more SHA-x." 5. About performance counters In this blog, performance counters are explained : "Note that unlike T3 and before, T4 crypto doesn't require kernel modules like ncp or n2cp, there is no visibility of crypto hardware with kstats or cryptoadm. T4 does provide hardware counters for crypto operations.  You can see these using cpustat: cpustat -c pic0=Instr_FGU_crypto 5 You can check the general crypto support of the hardware and OS with the command "isainfo -v". Since T4 crypto's implementation now allows direct userland access, there are no "crypto units" visible to cryptoadm.  " For more details refer Martin's blog as well. 6. How to turn off  SPARC T4 or Intel AES-NI crypto acceleration  I found this interesting blog from Darren about how to turn off  SPARC T4 or Intel AES-NI crypto acceleration. "One of the new Solaris 11 features of the linker/loader is the ability to have a single ELF object that has multiple different implementations of the same functions that are selected at runtime based on the capabilities of the machine.   The alternate to this is having the application coded to call getisax(2) system call and make the choice itself.  We use this functionality of the linker/loader when we build the userland libraries for the Solaris Cryptographic Framework (specifically libmd.so and libsoftcrypto.so) The Solaris linker/loader allows control of a lot of its functionality via environment variables, we can use that to control the version of the cryptographic functions we run.  To do this we simply export the LD_HWCAP environment variable with values that tell ld.so.1 to not select the HWCAP section matching certain features even if isainfo says they are present.  This will work for consumers of the Solaris Cryptographic Framework that use the Solaris PKCS#11 libraries or use libmd.so interfaces directly.  For SPARC T4 : export LD_HWCAP="-aes -des -md5 -sha256 -sha512 -mont -mpul" .. For Intel systems with AES-NI support: export LD_HWCAP="-aes"" Note that LD_HWCAP is explained in  http://docs.oracle.com/cd/E23823_01/html/816-5165/ld.so.1-1.html "LD_HWCAP, LD_HWCAP_32, and LD_HWCAP_64 -  Identifies an alternative hardware capabilities value... A “-” prefix results in the capabilities that follow being removed from the alternative capabilities." 7. Whitepaper on SPARC T4 Servers—Optimized for End-to-End Data Center Computing This Whitepaper on SPARC T4 Servers—Optimized for End-to-End Data Center Computing explains more details.  It has DTrace scripts which may come in handy : "To ensure the hardware-assisted cryptographic acceleration is configured to use and working with the security scenarios, it is recommended to use the following Solaris DTrace script. #!/usr/sbin/dtrace -s pid$1:libsoftcrypto:yf*:entry, pid$target:libsoftcrypto:rsa*:entry, pid$1:libmd:yf*:entry { @[probefunc] = count(); } tick-1sec { printa(@ops); trunc(@ops); }" Note that I have slightly modified the D Script to have RSA "libsoftcrypto:rsa*:entry" as well as per recommendations from Chi-Chang Lin. 8. References http://www.oracle.com/us/corporate/features/sparc-t4-announcement-494846.html http://www.oracle.com/us/products/servers-storage/servers/sparc-enterprise/t-series/sparc-t4-1-ds-487858.pdf https://blogs.oracle.com/DanX/entry/sparc_t4_openssl_engine https://blogs.oracle.com/DanX/entry/where_s_the_crypto_libraries https://blogs.oracle.com/darren/entry/howto_turn_off_sparc_t4 http://docs.oracle.com/cd/E23823_01/html/816-5165/ld.so.1-1.html   https://blogs.oracle.com/hardware/entry/unleash_the_power_of_cryptography https://blogs.oracle.com/cmt/entry/t4_crypto_cheat_sheet https://blogs.oracle.com/martinm/entry/t4_performance_counters_explained  https://blogs.oracle.com/jsavit/entry/no_mau_required_on_a http://www.oracle.com/us/products/servers-storage/servers/sparc-enterprise/t-series/sparc-t4-business-wp-524472.pdf

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