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  • Relay Access Denied (State 13) Postfix + Dovecot + Mysql

    - by Pierre Jeptha
    So we have been scratching our heads for quite some time over this relay issue that has presented itself since we re-built our mail-server after a failed Webmin update. We are running Debian Karmic with postfix 2.6.5 and Dovecot 1.1.11, sourcing from a Mysql database and authenticating with SASL2 and PAM. Here are the symptoms of our problem: 1) When users are on our local network they can send and receive 100% perfectly fine. 2) When users are off our local network and try to send to domains not of this mail server (ie. gmail) they get the "Relay Access Denied" error. However users can send to domains of this mail server when off the local network fine. 3) We host several virtual domains on this mailserver, the primary domain being airnet.ca. The rest of our virtual domains (ex. jeptha.ca) cannot receive email from domains not hosted by this mailserver (ie. gmail and such cannot send to them). They receive bounce backs of "Relay Access Denied (State 13)". This is regardless of whether they are on our local network or not, which is why it is so urgent for us to get this solved. Here is our main.cf from postfix: myhostname = mail.airnet.ca mydomain = airnet.ca smtpd_banner = $myhostname ESMTP $mail_name (Ubuntu) biff = no smtpd_sasl_type = dovecot queue_directory = /var/spool/postfix smtpd_sasl_path = private/auth smtpd_sender_restrictions = permit_mynetworks permit_sasl_authenticated smtp_sasl_auth_enable = yes smtpd_sasl_auth_enable = yes append_dot_mydomain = no readme_directory = no smtp_tls_security_level = may smtpd_tls_security_level = may smtp_tls_note_starttls_offer = yes smtpd_tls_key_file = /etc/ssl/private/ssl-cert-snakeoil.key smtpd_tls_cert_file = /etc/ssl/certs/ssl-cert-snakeoil.pem smtpd_tls_loglevel = 1 smtpd_tls_received_header = yes smtpd_tls_auth_only = no alias_maps = proxy:mysql:/etc/postfix/mysql/alias.cf hash:/etc/aliases alias_database = hash:/etc/aliases mydestination = mail.airnet.ca, airnet.ca, localhost.$mydomain mailbox_command = procmail -a "$EXTENSION" mailbox_size_limit = 0 recipient_delimiter = + local_recipient_maps = $alias_maps $virtual_mailbox_maps proxy:unix:passwd.byname home_mailbox = /var/virtual/ mail_spool_directory = /var/spool/mail mailbox_transport = maildrop smtpd_helo_required = yes disable_vrfy_command = yes smtpd_etrn_restrictions = reject smtpd_data_restrictions = reject_unauth_pipelining, permit show_user_unknown_table_name = no proxy_read_maps = $local_recipient_maps $mydestination $virtual_alias_maps $virtual_alias_domains $virtual_mailbox_maps $virtual_mailbox_domains $relay_recipient_maps $relay_domains $canonical_maps $sender_canonical_maps $recipient_canonical_maps $relocated_maps $transport_maps $mynetworks $virtual_mailbox_limit_maps $virtual_uid_maps $virtual_gid_maps virtual_alias_domains = message_size_limit = 20971520 transport_maps = proxy:mysql:/etc/postfix/mysql/vdomain.cf virtual_mailbox_maps = proxy:mysql:/etc/postfix/mysql/vmailbox.cf virtual_alias_maps = proxy:mysql:/etc/postfix/mysql/alias.cf hash:/etc/mailman/aliases virtual_uid_maps = proxy:mysql:/etc/postfix/mysql/vuid.cf virtual_gid_maps = proxy:mysql:/etc/postfix/mysql/vgid.cf virtual_mailbox_base = / virtual_mailbox_limit = 209715200 virtual_mailbox_extended = yes virtual_create_maildirsize = yes virtual_mailbox_limit_maps = proxy:mysql:/etc/postfix/mysql/vmlimit.cf virtual_mailbox_limit_override = yes virtual_mailbox_limit_inbox = no virtual_overquote_bounce = yes virtual_minimum_uid = 1 maximal_queue_lifetime = 1d bounce_queue_lifetime = 4h delay_warning_time = 1h append_dot_mydomain = no qmgr_message_active_limit = 500 broken_sasl_auth_clients = yes smtpd_sasl_path = private/auth smtpd_sasl_local_domain = $myhostname smtpd_sasl_security_options = noanonymous smtpd_sasl_authenticated_header = yes smtp_bind_address = 142.46.193.6 relay_domains = $mydestination mynetworks = 127.0.0.0, 142.46.193.0/25 inet_interfaces = all inet_protocols = all And here is the master.cf from postfix: # ========================================================================== # service type private unpriv chroot wakeup maxproc command + args # (yes) (yes) (yes) (never) (100) # ========================================================================== smtp inet n - - - - smtpd #submission inet n - - - - smtpd # -o smtpd_tls_security_level=encrypt # -o smtpd_sasl_auth_enable=yes # -o smtpd_client_restrictions=permit_sasl_authenticated,reject # -o milter_macro_daemon_name=ORIGINATING #smtps inet n - - - - smtpd # -o smtpd_tls_wrappermode=yes # -o smtpd_sasl_auth_enable=yes # -o smtpd_client_restrictions=permit_sasl_authenticated,reject # -o milter_macro_daemon_name=ORIGINATING #628 inet n - - - - qmqpd pickup fifo n - - 60 1 pickup cleanup unix n - - - 0 cleanup qmgr fifo n - n 300 1 qmgr #qmgr fifo n - - 300 1 oqmgr tlsmgr unix - - - 1000? 1 tlsmgr rewrite unix - - - - - trivial-rewrite bounce unix - - - - 0 bounce defer unix - - - - 0 bounce trace unix - - - - 0 bounce verify unix - - - - 1 verify flush unix n - - 1000? 0 flush proxymap unix - - n - - proxymap proxywrite unix - - n - 1 proxymap smtp unix - - - - - smtp # When relaying mail as backup MX, disable fallback_relay to avoid MX loops relay unix - - - - - smtp -o smtp_fallback_relay= # -o smtp_helo_timeout=5 -o smtp_connect_timeout=5 showq unix n - - - - showq error unix - - - - - error retry unix - - - - - error discard unix - - - - - discard local unix - n n - - local virtual unix - n n - - virtual lmtp unix - - - - - lmtp anvil unix - - - - 1 anvil scache unix - - - - 1 scache maildrop unix - n n - - pipe flags=DRhu user=vmail argv=/usr/bin/maildrop -d ${recipient} # # See the Postfix UUCP_README file for configuration details. # uucp unix - n n - - pipe flags=Fqhu user=uucp argv=uux -r -n -z -a$sender - $nexthop!rmail ($recipient) # # Other external delivery methods. # ifmail unix - n n - - pipe flags=F user=ftn argv=/usr/lib/ifmail/ifmail -r $nexthop ($recipient) bsmtp unix - n n - - pipe flags=Fq. user=bsmtp argv=/usr/lib/bsmtp/bsmtp -t$nexthop -f$sender $recipient scalemail-backend unix - n n - 2 pipe flags=R user=scalemail argv=/usr/lib/scalemail/bin/scalemail-store ${nexthop} ${user} ${extension} mailman unix - n n - - pipe flags=FR user=list argv=/usr/lib/mailman/bin/postfix-to-mailman.py ${nexthop} ${user} spfpolicy unix - n n - - spawn user=nobody argv=/usr/bin/perl /usr/sbin/postfix-policyd-spf-perl smtp-amavis unix - - n - 4 smtp -o smtp_data_done_timeout=1200 -o smtp_send_xforward_command=yes -o disable_dns_lookups=yes #127.0.0.1:10025 inet n - n - - smtpd dovecot unix - n n - - pipe flags=DRhu user=dovecot:21pever1lcha0s argv=/usr/lib/dovecot/deliver -d ${recipient Here is Dovecot.conf protocols = imap imaps pop3 pop3s disable_plaintext_auth = no log_path = /etc/dovecot/logs/err info_log_path = /etc/dovecot/logs/info log_timestamp = "%Y-%m-%d %H:%M:%S ". syslog_facility = mail ssl_listen = 142.46.193.6 ssl_disable = no ssl_cert_file = /etc/ssl/certs/ssl-cert-snakeoil.pem ssl_key_file = /etc/ssl/private/ssl-cert-snakeoil.key mail_location = mbox:~/mail:INBOX=/var/virtual/%d/mail/%u mail_privileged_group = mail mail_debug = yes protocol imap { login_executable = /usr/lib/dovecot/imap-login mail_executable = /usr/lib/dovecot/rawlog /usr/lib/dovecot/imap mail_executable = /usr/lib/dovecot/gdbhelper /usr/lib/dovecot/imap mail_executable = /usr/lib/dovecot/imap imap_max_line_length = 65536 mail_max_userip_connections = 20 mail_plugin_dir = /usr/lib/dovecot/modules/imap login_greeting_capability = yes } protocol pop3 { login_executable = /usr/lib/dovecot/pop3-login mail_executable = /usr/lib/dovecot/pop3 pop3_enable_last = no pop3_uidl_format = %08Xu%08Xv mail_max_userip_connections = 10 mail_plugin_dir = /usr/lib/dovecot/modules/pop3 } protocol managesieve { sieve=~/.dovecot.sieve sieve_storage=~/sieve } mail_plugin_dir = /usr/lib/dovecot/modules/lda auth_executable = /usr/lib/dovecot/dovecot-auth auth_process_size = 256 auth_cache_ttl = 3600 auth_cache_negative_ttl = 3600 auth_username_chars = abcdefghijklmnopqrstuvwxyzABCDEFGHIJKLMNOPQRSTUVWXYZ01234567890.-_@ auth_verbose = yes auth_debug = yes auth_debug_passwords = yes auth_worker_max_count = 60 auth_failure_delay = 2 auth default { mechanisms = plain login passdb sql { args = /etc/dovecot/dovecot-sql.conf } userdb sql { args = /etc/dovecot/dovecot-sql.conf } socket listen { client { path = /var/spool/postfix/private/auth mode = 0660 user = postfix group = postfix } master { path = /var/run/dovecot/auth-master mode = 0600 } } } Please, if you require anything do not hesistate, I will post it ASAP. Any help or suggestions are greatly appreciated! Thanks, Pierre

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  • Unusually high memory usage on a CentOS VPS with 512 guaranteed RAM

    - by Andrei Bârsan
    I'm working on a medium-sized web application written in PHP that's running on a VPS with 512mb ram. The webapp hasn't been officially launched yet, so there isn't too much traffic going on, just me and a few other people working on it. There is another slightly smaller webapp also hosted on this machine, among 4-5 other small static sites. We are running Centos 5 32-bit & cPanel/WHM. This is the result of running ps aux and, as you can see, it's not using 100% of the RAM. However, on the hypanel overview, it's always shown as using aroun 500MB ram, just for running apache, mysql, and the lowest-memory-footprint versions of the mail server, ftp server etc. -bash-3.2# ps aux USER PID %CPU %MEM VSZ RSS TTY STAT START TIME COMMAND root 1 0.0 0.0 2156 664 ? Ss 12:08 0:00 init [3] root 1123 0.0 0.0 2260 548 ? S<s 12:08 0:00 /sbin/udevd -d root 1462 0.0 0.0 1812 568 ? Ss 12:08 0:00 syslogd -m 0 named 1496 0.0 0.0 3808 820 ? Ss 12:08 0:00 nsd named 1497 0.0 0.0 10672 756 ? S 12:08 0:00 nsd named 1499 0.0 0.0 3880 584 ? S 12:08 0:00 nsd root 1514 0.0 0.1 7240 1064 ? Ss 12:08 0:00 /usr/sbin/sshd root 1522 0.0 0.0 2832 832 ? Ss 12:08 0:00 xinetd -stayalive -pidfile /var/run/xinetd.pid root 1534 0.0 0.1 3712 1328 ? S 12:08 0:00 /bin/sh /usr/bin/mysqld_safe --datadir=/var/lib/mysql - mysql 1667 0.0 2.9 225680 30884 ? Sl 12:08 0:00 /usr/sbin/mysqld --basedir=/ --datadir=/var/lib/mysql - mailnull 1766 0.0 0.1 9352 1100 ? Ss 12:08 0:00 /usr/sbin/exim -bd -q60m root 1797 0.0 0.0 2156 708 ? Ss 12:08 0:00 /usr/sbin/dovecot root 1798 0.0 0.0 2632 1012 ? S 12:08 0:00 dovecot-auth root 1816 0.0 3.0 38580 32456 ? Ss 12:08 0:01 /usr/local/bin/spamd -d --allowed-ips=127.0.0.1 --pidfi root 1839 0.0 1.6 63200 17496 ? Ss 12:08 0:00 /usr/local/apache/bin/httpd -k start -DSSL root 1846 0.0 0.1 5416 1468 ? Ss 12:08 0:00 pure-ftpd (SERVER) root 1848 0.0 0.1 6212 1244 ? S 12:08 0:00 /usr/sbin/pure-authd -s /var/run/ftpd.sock -r /usr/sbin root 1856 0.0 0.1 4492 1112 ? Ss 12:08 0:00 crond root 1864 0.0 0.0 2356 428 ? Ss 12:08 0:00 /usr/sbin/atd dovecot 1927 0.0 0.1 5196 1952 ? S 12:08 0:00 pop3-login dovecot 1928 0.0 0.1 5196 1948 ? S 12:08 0:00 pop3-login dovecot 1929 0.0 0.1 5316 2012 ? S 12:08 0:00 imap-login dovecot 1930 0.0 0.2 5416 2228 ? S 12:08 0:00 imap-login root 1939 0.0 0.1 3936 1964 ? S 12:08 0:00 cPhulkd - processor root 1963 0.0 0.8 15876 8564 ? S 12:08 0:00 cpsrvd (SSL) - waiting for connections root 1966 0.0 0.7 15172 7748 ? S 12:08 0:00 cpdavd - accepting connections on 2077 and 2078 root 1990 0.0 0.2 5008 3136 ? S 12:08 0:00 queueprocd - wait to process a task root 2017 0.0 2.9 38580 31020 ? S 12:08 0:00 spamd child root 2018 0.0 0.5 8904 5636 ? S 12:08 0:00 /usr/bin/perl /usr/local/cpanel/bin/leechprotect nobody 2021 0.0 3.2 66512 33724 ? S 12:08 0:00 /usr/local/apache/bin/httpd -k start -DSSL nobody 2022 0.0 3.1 67812 33024 ? S 12:08 0:00 /usr/local/apache/bin/httpd -k start -DSSL nobody 2024 0.0 1.9 64364 20680 ? S 12:08 0:00 /usr/local/apache/bin/httpd -k start -DSSL root 2027 0.0 0.4 9000 4540 ? S 12:08 0:00 tailwatchd root 2032 0.0 0.1 4176 1836 ? SN 12:08 0:00 cpanellogd - sleeping for logs nobody 3096 0.0 1.9 64572 20264 ? S 12:09 0:00 /usr/local/apache/bin/httpd -k start -DSSL nobody 3097 0.0 2.8 66008 30136 ? S 12:09 0:00 /usr/local/apache/bin/httpd -k start -DSSL nobody 3098 0.0 2.8 65704 29752 ? S 12:09 0:00 /usr/local/apache/bin/httpd -k start -DSSL nobody 3099 0.0 3.1 67260 32816 ? S 12:09 0:00 /usr/local/apache/bin/httpd -k start -DSSL andrei 3448 0.0 0.1 3204 1632 ? S 12:50 0:00 imap nobody 3537 0.0 1.9 64308 20108 ? S 13:01 0:00 /usr/local/apache/bin/httpd -k start -DSSL nobody 3614 0.0 1.9 64576 20628 ? S 13:10 0:00 /usr/local/apache/bin/httpd -k start -DSSL nobody 3615 0.0 1.3 63200 14672 ? S 13:10 0:00 /usr/local/apache/bin/httpd -k start -DSSL root 3626 0.0 0.2 10232 2964 ? Rs 13:14 0:00 sshd: root@pts/0 root 3648 0.0 0.1 3844 1600 pts/0 Ss 13:14 0:00 -bash root 3826 0.0 0.0 2532 908 pts/0 R+ 13:21 0:00 ps aux Lately, without any significant changes to the configuration, the memory usage started peaking and going over 512, causing the virtual server to kill apache, basically murdering our site in the process. Do you have any idea if this is normal and more resources should be acquired? I don't think... since there isn't too much data or traffic online yet.

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  • Unable to access intel fake RAID 1 array in Fedora 14 after reboot

    - by Sim
    Hello everyone, 1st I am relatively new to linux (but not to *nix). I have 4 disks assembled in the following intel ahci bios fake raid arrays: 2x320GB RAID1 - used for operating systems md126 2x1TB RAID1 - used for data md125 I have used the raid of size 320GB to install my operating system and the second raid I didn't even select during the installation of Fedora 14. After successful partitioning and installation of Fedora, I tried to make the second array available, it was possible to make it visible in linux with mdadm --assembe --scan , after that I created one maximum size partition and 1 maximum size ext4 filesystem in it. Mounted, and used it. After restart - a few I/O errors during boot regarding md125 + inability to mount the filesystem on it and dropped into repair shell. I commented the filesystem in fstab and it booted. To my surprise, the array was marked as "auto read only": [root@localhost ~]# cat /proc/mdstat Personalities : [raid1] md125 : active (auto-read-only) raid1 sdc[1] sdd[0] 976759808 blocks super external:/md127/0 [2/2] [UU] md127 : inactive sdc[1](S) sdd[0](S) 4514 blocks super external:imsm md126 : active raid1 sda[1] sdb[0] 312566784 blocks super external:/md1/0 [2/2] [UU] md1 : inactive sdb[1](S) sda[0](S) 4514 blocks super external:imsm unused devices: <none> [root@localhost ~]# and the partition in it was not available as device special file in /dev: [root@localhost ~]# ls -l /dev/md125* brw-rw---- 1 root disk 9, 125 Jan 6 15:50 /dev/md125 [root@localhost ~]# But the partition is there according to fdisk: [root@localhost ~]# fdisk -l /dev/md125 Disk /dev/md125: 1000.2 GB, 1000202043392 bytes 19 heads, 10 sectors/track, 10281682 cylinders, total 1953519616 sectors Units = sectors of 1 * 512 = 512 bytes Sector size (logical/physical): 512 bytes / 512 bytes I/O size (minimum/optimal): 512 bytes / 512 bytes Disk identifier: 0x1b238ea9 Device Boot Start End Blocks Id System /dev/md125p1 2048 1953519615 976758784 83 Linux [root@localhost ~]# I tried to "activate" the array in different ways (I'm not experienced with mdadm and the man page is gigantic so I was only browsing it looking for my answer) but it was impossible - the array would still stay in "auto read only" and the device special file for the partition it will not be in /dev. It was only after I recreated the partition via fdisk that it reappeared in /dev... until next reboot. So, my question is - How do I make the array automatically available after reboot? Here is some additional information: 1st I am able to see the UUID of the array in blkid: [root@localhost ~]# blkid /dev/sdc: UUID="b9a1149f-ae11-4fc8-a600-0d77354dc42a" SEC_TYPE="ext2" TYPE="ext3" /dev/sdd: UUID="b9a1149f-ae11-4fc8-a600-0d77354dc42a" SEC_TYPE="ext2" TYPE="ext3" /dev/md126p1: UUID="60C8D9A7C8D97C2A" TYPE="ntfs" /dev/md126p2: UUID="3d1b38a3-b469-4b7c-b016-8abfb26a5d7d" TYPE="ext4" /dev/md126p3: UUID="1Msqqr-AAF8-k0wi-VYnq-uWJU-y0OD-uIFBHL" TYPE="LVM2_member" /dev/mapper/vg00-rootlv: LABEL="_Fedora-14-x86_6" UUID="34cc1cf5-6845-4489-8303-7a90c7663f0a" TYPE="ext4" /dev/mapper/vg00-swaplv: UUID="4644d857-e13b-456c-ac03-6f26299c1046" TYPE="swap" /dev/mapper/vg00-homelv: UUID="82bd58b2-edab-4b4b-aec4-b79595ecd0e3" TYPE="ext4" /dev/mapper/vg00-varlv: UUID="1b001444-5fdd-41b6-a59a-9712ec6def33" TYPE="ext4" /dev/mapper/vg00-tmplv: UUID="bf7d2459-2b35-4a1c-9b81-d4c4f24a9842" TYPE="ext4" /dev/md125: UUID="b9a1149f-ae11-4fc8-a600-0d77354dc42a" SEC_TYPE="ext2" TYPE="ext3" /dev/sda: TYPE="isw_raid_member" /dev/md125p1: UUID="420adfdd-6c4e-4552-93f0-2608938a4059" TYPE="ext4" [root@localhost ~]# Here is how /etc/mdadm.conf looks like: [root@localhost ~]# cat /etc/mdadm.conf # mdadm.conf written out by anaconda MAILADDR root AUTO +imsm +1.x -all ARRAY /dev/md1 UUID=89f60dee:e46a251f:7475814b:d4cc19a9 ARRAY /dev/md126 UUID=a8775c90:cee66376:5310fc13:63bcba5b ARRAY /dev/md125 UUID=b9a1149f:ae114fc8:a6000d77:354dc42a [root@localhost ~]# here is how /proc/mdstat looks like after I recreate the partition in the array so that it becomes available: [root@localhost ~]# cat /proc/mdstat Personalities : [raid1] md125 : active raid1 sdc[1] sdd[0] 976759808 blocks super external:/md127/0 [2/2] [UU] md127 : inactive sdc[1](S) sdd[0](S) 4514 blocks super external:imsm md126 : active raid1 sda[1] sdb[0] 312566784 blocks super external:/md1/0 [2/2] [UU] md1 : inactive sdb[1](S) sda[0](S) 4514 blocks super external:imsm unused devices: <none> [root@localhost ~]# Detailed output regarding the array in subject: [root@localhost ~]# mdadm --detail /dev/md125 /dev/md125: Container : /dev/md127, member 0 Raid Level : raid1 Array Size : 976759808 (931.51 GiB 1000.20 GB) Used Dev Size : 976759940 (931.51 GiB 1000.20 GB) Raid Devices : 2 Total Devices : 2 Update Time : Fri Jan 7 00:38:00 2011 State : clean Active Devices : 2 Working Devices : 2 Failed Devices : 0 Spare Devices : 0 UUID : 30ebc3c2:b6a64751:4758d05c:fa8ff782 Number Major Minor RaidDevice State 1 8 32 0 active sync /dev/sdc 0 8 48 1 active sync /dev/sdd [root@localhost ~]# and /etc/fstab, with /data commented (the filesystem that is on this array): # # /etc/fstab # Created by anaconda on Thu Jan 6 03:32:40 2011 # # Accessible filesystems, by reference, are maintained under '/dev/disk' # See man pages fstab(5), findfs(8), mount(8) and/or blkid(8) for more info # /dev/mapper/vg00-rootlv / ext4 defaults 1 1 UUID=3d1b38a3-b469-4b7c-b016-8abfb26a5d7d /boot ext4 defaults 1 2 #UUID=420adfdd-6c4e-4552-93f0-2608938a4059 /data ext4 defaults 0 1 /dev/mapper/vg00-homelv /home ext4 defaults 1 2 /dev/mapper/vg00-tmplv /tmp ext4 defaults 1 2 /dev/mapper/vg00-varlv /var ext4 defaults 1 2 /dev/mapper/vg00-swaplv swap swap defaults 0 0 tmpfs /dev/shm tmpfs defaults 0 0 devpts /dev/pts devpts gid=5,mode=620 0 0 sysfs /sys sysfs defaults 0 0 proc /proc proc defaults 0 0 [root@localhost ~]# Thanks in advance to everyone that even read this whole issue :-)

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  • vsftpd not allowing uploads. 550 response.

    - by Josh
    I've set vsftpd up on a centos box. I keep trying to upload files but I keep getting "550 Failed to change directory" and "550 Could not get file size." Here's my vsftpd.conf # The default compiled in settings are fairly paranoid. This sample file # loosens things up a bit, to make the ftp daemon more usable. # Please see vsftpd.conf.5 for all compiled in defaults. # # READ THIS: This example file is NOT an exhaustive list of vsftpd options. # Please read the vsftpd.conf.5 manual page to get a full idea of vsftpd's # capabilities. # # Allow anonymous FTP? (Beware - allowed by default if you comment this out). anonymous_enable=YES # # Uncomment this to allow local users to log in. local_enable=YES # # Uncomment this to enable any form of FTP write command. write_enable=YES # # Default umask for local users is 077. You may wish to change this to 022, # if your users expect that (022 is used by most other ftpd's) local_umask=022 # # Uncomment this to allow the anonymous FTP user to upload files. This only # has an effect if the above global write enable is activated. Also, you will # obviously need to create a directory writable by the FTP user. anon_upload_enable=YES # # Uncomment this if you want the anonymous FTP user to be able to create # new directories. anon_mkdir_write_enable=YES anon_other_write_enable=YES # # Activate directory messages - messages given to remote users when they # go into a certain directory. dirmessage_enable=YES # # The target log file can be vsftpd_log_file or xferlog_file. # This depends on setting xferlog_std_format parameter xferlog_enable=YES # # Make sure PORT transfer connections originate from port 20 (ftp-data). connect_from_port_20=YES # # If you want, you can arrange for uploaded anonymous files to be owned by # a different user. Note! Using "root" for uploaded files is not # recommended! #chown_uploads=YES #chown_username=whoever # # The name of log file when xferlog_enable=YES and xferlog_std_format=YES # WARNING - changing this filename affects /etc/logrotate.d/vsftpd.log #xferlog_file=/var/log/xferlog # # Switches between logging into vsftpd_log_file and xferlog_file files. # NO writes to vsftpd_log_file, YES to xferlog_file xferlog_std_format=NO # # You may change the default value for timing out an idle session. #idle_session_timeout=600 # # You may change the default value for timing out a data connection. #data_connection_timeout=120 # # It is recommended that you define on your system a unique user which the # ftp server can use as a totally isolated and unprivileged user. #nopriv_user=ftpsecure # # Enable this and the server will recognise asynchronous ABOR requests. Not # recommended for security (the code is non-trivial). Not enabling it, # however, may confuse older FTP clients. #async_abor_enable=YES # # By default the server will pretend to allow ASCII mode but in fact ignore # the request. Turn on the below options to have the server actually do ASCII # mangling on files when in ASCII mode. # Beware that on some FTP servers, ASCII support allows a denial of service # attack (DoS) via the command "SIZE /big/file" in ASCII mode. vsftpd # predicted this attack and has always been safe, reporting the size of the # raw file. # ASCII mangling is a horrible feature of the protocol. #ascii_upload_enable=YES #ascii_download_enable=YES # # You may fully customise the login banner string: #ftpd_banner=Welcome to blah FTP service. # # You may specify a file of disallowed anonymous e-mail addresses. Apparently # useful for combatting certain DoS attacks. #deny_email_enable=YES # (default follows) #banned_email_file=/etc/vsftpd/banned_emails # # You may specify an explicit list of local users to chroot() to their home # directory. If chroot_local_user is YES, then this list becomes a list of # users to NOT chroot(). #chroot_list_enable=YES # (default follows) #chroot_list_file=/etc/vsftpd/chroot_list # # You may activate the "-R" option to the builtin ls. This is disabled by # default to avoid remote users being able to cause excessive I/O on large # sites. However, some broken FTP clients such as "ncftp" and "mirror" assume # the presence of the "-R" option, so there is a strong case for enabling it. #ls_recurse_enable=YES # # When "listen" directive is enabled, vsftpd runs in standalone mode and # listens on IPv4 sockets. This directive cannot be used in conjunction # with the listen_ipv6 directive. listen=YES # This directive enables listening on IPv6 sockets. To listen on IPv4 and IPv6 # sockets, you must run two copies of vsftpd whith two configuration files. # Make sure, that one of the listen options is commented !! #listen_ipv6=YES pam_service_name=vsftpd userlist_enable=YES tcp_wrappers=YES log_ftp_protocol=YES banner_file=/etc/vsftpd/issue local_root=/var/www guest_enable=YES guest_username=ftpusr ftp_username=nobody

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  • Can't deploy rails 4 app on Bluehost with Passenger 4 and nginx

    - by user2205763
    I am at Bluehost (dedicated server) trying to run a rails 4 app. I asked to have my server re-imaged, specifying that I do not want rails, ruby, or passenger install automatically as I wanted to install the latest versions myself using a version manager (Bluehost by default offers rails 2.3, ruby 1.8, and passenger 3, which won't work with my app). I installed ruby 1.9.3p327, rails 4.0.0, and passenger 4.0.5. I can verify this by typing, "ruby -v", "rails -v", and "passenger -v" (also "gem -v"). I made sure to install these not as root, so that I don't get a 403 forbidden error when trying to deploy the app. I installed passenger by typing "gem install passenger", and then installed the nginx passenger module (into "/nginx") with "passenger-install-nginx-module". I am trying to run my rails app on a subdomain, http://development.thegraduate.hk (I am using the subdomain to show my client progress on the website). In bluehost I created that subdomain, and had it point to "public_html/thegraduate". I then created a symlink from "rails_apps/thegraduate/public" to "public_html/thegraduate" and verified that the symlink exists. The problem is: when I go to http://development.thegraduate.hk, I get a directory listing. There is nothing resembling a rails app. I have not added a .htaccess file to /rails_apps/thegraduate/public, as that was never specified in the installation of passenger. It was meant to be 'install and go'. When I type "passenger-memory-status", I get 3 things: - Apache processes (7) - Nginx processes (0) - Passenger processes (0) So it appears that nginx and passenger are not running, and I can't figure out how to get it to run (I'm not looking to have it run as a standalone server). Here is my nginx.conf file (/nginx/conf/nginx.conf): #user nobody; worker_processes 1; #error_log logs/error.log; #error_log logs/error.log notice; #error_log logs/error.log info; #pid logs/nginx.pid; events { worker_connections 1024; } http { passenger_root /home/thegrad4/.rbenv/versions/1.9.3-p327/lib/ruby/gems/1.9.1/gems/passenger-4.0.5; passenger_ruby /home/thegrad4/.rbenv/versions/1.9.3-p327/bin/ruby; include mime.types; default_type application/octet-stream; #log_format main '$remote_addr - $remote_user [$time_local] "$request" ' # '$status $body_bytes_sent "$http_referer" ' # '"$http_user_agent" "$http_x_forwarded_for"'; #access_log logs/access.log main; sendfile on; #tcp_nopush on; #keepalive_timeout 0; keepalive_timeout 65; #gzip on; server { listen 80; server_name development.thegraduate.hk; root ~/rails_apps/thegraduate/public; passenger_enabled on; #charset koi8-r; #access_log logs/host.access.log main; location / { root html; index index.html index.htm; } #error_page 404 /404.html; # redirect server error pages to the static page /50x.html # error_page 500 502 503 504 /50x.html; location = /50x.html { root html; } # proxy the PHP scripts to Apache listening on 127.0.0.1:80 # #location ~ \.php$ { # proxy_pass http://127.0.0.1; #} # pass the PHP scripts to FastCGI server listening on 127.0.0.1:9000 # #location ~ \.php$ { # root html; # fastcgi_pass 127.0.0.1:9000; # fastcgi_index index.php; # fastcgi_param SCRIPT_FILENAME /scripts$fastcgi_script_name; # include fastcgi_params; #} # deny access to .htaccess files, if Apache's document root # concurs with nginx's one # #location ~ /\.ht { # deny all; #} } # another virtual host using mix of IP-, name-, and port-based configuration # #server { # listen 8000; # listen somename:8080; # server_name somename alias another.alias; # location / { # root html; # index index.html index.htm; # } #} # HTTPS server # #server { # listen 443; # server_name localhost; # ssl on; # ssl_certificate cert.pem; # ssl_certificate_key cert.key; # ssl_session_timeout 5m; # ssl_protocols SSLv2 SSLv3 TLSv1; # ssl_ciphers HIGH:!aNULL:!MD5; # ssl_prefer_server_ciphers on; # location / { # root html; # index index.html index.htm; # } #} } I don't get any errors, just the directory listing. I've tried to be as detailed as possible. Any help on this issue would be greatly appreciated as I've been stumped for the past 3 days. Scouring the web has not helped as my issue seems to be specific to me. Thanks so much. If there are any potential details I forgot to specify, just ask. ** ADDITIONAL INFORMATION ** Going to development.thegraduate.hk/public/ will correctly display the index.html page in /rails_apps/thegraduate/public. However, changing root in the routes.rb file to "root = 'home#index'" does nothing.

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  • fd partitions gone from 2 discs, md happy with it and resyncs. How to recover ?

    - by d0nd
    Hey gurus, need some help badly with this one. I run a server with a 6Tb md raid5 volume built over 7*1Tb disks. I've had to shut down the server lately and when it went back up, 2 out of the 7 disks used for the raid volume had lost its conf : dmesg : [ 10.184167] sda: sda1 sda2 sda3 // System disk [ 10.202072] sdb: sdb1 [ 10.210073] sdc: sdc1 [ 10.222073] sdd: sdd1 [ 10.229330] sde: sde1 [ 10.239449] sdf: sdf1 [ 11.099896] sdg: unknown partition table [ 11.255641] sdh: unknown partition table All 7 disks have same geometry and were configured alike : dmesg : Disk /dev/sdb: 1000.2 GB, 1000204886016 bytes 255 heads, 63 sectors/track, 121601 cylinders Units = cylinders of 16065 * 512 = 8225280 bytes Disk identifier: 0x1e7481a5 Device Boot Start End Blocks Id System /dev/sdb1 1 121601 976760001 fd Linux raid autodetect All 7 disks (sdb1, sdc1, sdd1, sde1, sdf1, sdg1, sdh1) were used in a md raid5 xfs volume. When booting, md, which was (obviously) out of sync kicked in and automatically started rebuilding over the 7 disks, including the two "faulty" ones; xfs tried to do some shenanigans as well: dmesg : [ 19.566941] md: md0 stopped. [ 19.817038] md: bind<sdc1> [ 19.817339] md: bind<sdd1> [ 19.817465] md: bind<sde1> [ 19.817739] md: bind<sdf1> [ 19.817917] md: bind<sdh> [ 19.818079] md: bind<sdg> [ 19.818198] md: bind<sdb1> [ 19.818248] md: md0: raid array is not clean -- starting background reconstruction [ 19.825259] raid5: device sdb1 operational as raid disk 0 [ 19.825261] raid5: device sdg operational as raid disk 6 [ 19.825262] raid5: device sdh operational as raid disk 5 [ 19.825264] raid5: device sdf1 operational as raid disk 4 [ 19.825265] raid5: device sde1 operational as raid disk 3 [ 19.825267] raid5: device sdd1 operational as raid disk 2 [ 19.825268] raid5: device sdc1 operational as raid disk 1 [ 19.825665] raid5: allocated 7334kB for md0 [ 19.825667] raid5: raid level 5 set md0 active with 7 out of 7 devices, algorithm 2 [ 19.825669] RAID5 conf printout: [ 19.825670] --- rd:7 wd:7 [ 19.825671] disk 0, o:1, dev:sdb1 [ 19.825672] disk 1, o:1, dev:sdc1 [ 19.825673] disk 2, o:1, dev:sdd1 [ 19.825675] disk 3, o:1, dev:sde1 [ 19.825676] disk 4, o:1, dev:sdf1 [ 19.825677] disk 5, o:1, dev:sdh [ 19.825679] disk 6, o:1, dev:sdg [ 19.899787] PM: Starting manual resume from disk [ 28.663228] Filesystem "md0": Disabling barriers, not supported by the underlying device [ 28.663228] XFS mounting filesystem md0 [ 28.884433] md: resync of RAID array md0 [ 28.884433] md: minimum _guaranteed_ speed: 1000 KB/sec/disk. [ 28.884433] md: using maximum available idle IO bandwidth (but not more than 200000 KB/sec) for resync. [ 28.884433] md: using 128k window, over a total of 976759936 blocks. [ 29.025980] Starting XFS recovery on filesystem: md0 (logdev: internal) [ 32.680486] XFS: xlog_recover_process_data: bad clientid [ 32.680495] XFS: log mount/recovery failed: error 5 [ 32.682773] XFS: log mount failed I ran fdisk and flagged sdg1 and sdh1 as fd. I tried to reassemble the array but it didnt work: no matter what was in mdadm.conf, it still uses sdg and sdh instead of sdg1 and sdh1. I checked in /dev and I see no sdg1 and and sdh1, shich explains why it wont use it. I just don't know why those partitions are gone from /dev and how to readd those... blkid : /dev/sda1: LABEL="boot" UUID="519790ae-32fe-4c15-a7f6-f1bea8139409" TYPE="ext2" /dev/sda2: TYPE="swap" /dev/sda3: LABEL="root" UUID="91390d23-ed31-4af0-917e-e599457f6155" TYPE="ext3" /dev/sdb1: UUID="2802e68a-dd11-c519-e8af-0d8f4ed72889" TYPE="mdraid" /dev/sdc1: UUID="2802e68a-dd11-c519-e8af-0d8f4ed72889" TYPE="mdraid" /dev/sdd1: UUID="2802e68a-dd11-c519-e8af-0d8f4ed72889" TYPE="mdraid" /dev/sde1: UUID="2802e68a-dd11-c519-e8af-0d8f4ed72889" TYPE="mdraid" /dev/sdf1: UUID="2802e68a-dd11-c519-e8af-0d8f4ed72889" TYPE="mdraid" /dev/sdg: UUID="2802e68a-dd11-c519-e8af-0d8f4ed72889" TYPE="mdraid" /dev/sdh: UUID="2802e68a-dd11-c519-e8af-0d8f4ed72889" TYPE="mdraid" fdisk -l : Disk /dev/sda: 40.0 GB, 40020664320 bytes 255 heads, 63 sectors/track, 4865 cylinders Units = cylinders of 16065 * 512 = 8225280 bytes Disk identifier: 0x8c878c87 Device Boot Start End Blocks Id System /dev/sda1 * 1 12 96358+ 83 Linux /dev/sda2 13 134 979965 82 Linux swap / Solaris /dev/sda3 135 4865 38001757+ 83 Linux Disk /dev/sdb: 1000.2 GB, 1000204886016 bytes 255 heads, 63 sectors/track, 121601 cylinders Units = cylinders of 16065 * 512 = 8225280 bytes Disk identifier: 0x1e7481a5 Device Boot Start End Blocks Id System /dev/sdb1 1 121601 976760001 fd Linux raid autodetect Disk /dev/sdc: 1000.2 GB, 1000204886016 bytes 255 heads, 63 sectors/track, 121601 cylinders Units = cylinders of 16065 * 512 = 8225280 bytes Disk identifier: 0xc9bdc1e9 Device Boot Start End Blocks Id System /dev/sdc1 1 121601 976760001 fd Linux raid autodetect Disk /dev/sdd: 1000.2 GB, 1000204886016 bytes 255 heads, 63 sectors/track, 121601 cylinders Units = cylinders of 16065 * 512 = 8225280 bytes Disk identifier: 0xcc356c30 Device Boot Start End Blocks Id System /dev/sdd1 1 121601 976760001 fd Linux raid autodetect Disk /dev/sde: 1000.2 GB, 1000204886016 bytes 255 heads, 63 sectors/track, 121601 cylinders Units = cylinders of 16065 * 512 = 8225280 bytes Disk identifier: 0xe87f7a3d Device Boot Start End Blocks Id System /dev/sde1 1 121601 976760001 fd Linux raid autodetect Disk /dev/sdf: 1000.2 GB, 1000204886016 bytes 255 heads, 63 sectors/track, 121601 cylinders Units = cylinders of 16065 * 512 = 8225280 bytes Disk identifier: 0xb17a2d22 Device Boot Start End Blocks Id System /dev/sdf1 1 121601 976760001 fd Linux raid autodetect Disk /dev/sdg: 1000.2 GB, 1000204886016 bytes 255 heads, 63 sectors/track, 121601 cylinders Units = cylinders of 16065 * 512 = 8225280 bytes Disk identifier: 0x8f3bce61 Device Boot Start End Blocks Id System /dev/sdg1 1 121601 976760001 fd Linux raid autodetect Disk /dev/sdh: 1000.2 GB, 1000204886016 bytes 255 heads, 63 sectors/track, 121601 cylinders Units = cylinders of 16065 * 512 = 8225280 bytes Disk identifier: 0xa98062ce Device Boot Start End Blocks Id System /dev/sdh1 1 121601 976760001 fd Linux raid autodetect I really dont know what happened nor how to recover from this mess. Needless to say the 5TB or so worth of data sitting on those disks are very valuable to me... Any idea any one? Did anybody ever experienced a similar situation or know how to recover from it ? Can someone help me? I'm really desperate... :x

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  • With dnsmasq as the DNS server, 'dig' and 'ping' succeed while 'nslookup' fails

    - by einpoklum
    I installed dnsmasq on a machine of mine (It's a Kubuntu 12.04 LTS), backed only by /etc/hosts (no connection to the Internet until later). Now, if I dig mymachine, I get 192.168.0.1, but if I try to nslookup mymachine, I get: >> connection timed out; no servers could be reached Tried also nslookup mymachine.mynicedomain.org - didn't work either. pinging (Edit:) succeeds. This happens both on the server machine itself and on other machines on the network. How can I the DNS lookups to work? What problem is preventing nslookup from succeeding? Additional Information In the server's /etc/hosts: 192.168.0.1 mymachine In the server's nsswitch.conf: hosts: files mdns4_mininal [NOTFOUND=return] dns mdns4 (admittedly, this is a bit weird; but I also tried: hosts: files dns instead, with the same effect) In resolv.conf (which is generated by dnsmasq): nameserver 127.0.0.1 search mynicedomain.org In the server's /etc/hosts.allow: domain: ALL In the other machines' /etc/resolv.conf (this is set by the DHCP client): nameserver 192.168.0.1 search mynicedomain.org Relevant netstat output on the server: Proto Recv-Q Send-Q Local Address Foreign Address State tcp 0 0 127.0.0.1:53 0.0.0.0:* LISTEN tcp 0 0 192.168.0.1:53 0.0.0.0:* LISTEN Finally, here's the ipconfig output from one of the client machines on the network (running Windows 7): Connection-specific DNS Suffix . : mynicedomain.org Description . . . . . . . . . . . : Intel(R) 82579LM Gigabit Network Connection Physical Address. . . . . . . . . : 12-34-56-78-9A-BC DHCP Enabled. . . . . . . . . . . : Yes Autoconfiguration Enabled . . . . : Yes IPv4 Address. . . . . . . . . . . : 192.168.0.50(Preferred) Subnet Mask . . . . . . . . . . . : 255.255.255.0 Lease Obtained. . . . . . . . . . : Sunday, October 20th 2013 16:20:25 Lease Expires . . . . . . . . . . : Sunday, October 20th 2013 18:20:24 Default Gateway . . . . . . . . . : 192.168.0.1 DHCP Server . . . . . . . . . . . : 192.168.0.1 DNS Servers . . . . . . . . . . . : 192.168.0.1 NetBIOS over Tcpip. . . . . . . . : Enabled Notes: May be related to this question.

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  • With dnsmasq as the DNS server, 'dig' succeeds, but 'nslookup' and 'ping' fail

    - by einpoklum
    I installed dnsmasq on a machine of mine (It's a Kubuntu 12.04 LTS), backed only by /etc/hosts (no connection to the Internet until later). Now, when I'm on the same machine as the dnsmasq - or any other machine on the server, I can dig mymachine and get 192.168.0.1, but if I try to nslookup mymachine, I get: >> connection timed out; no servers could be reached Tried also nslookup mymachine.mynicedomain.org - didn't work either. pinging fails. How can I the DNS lookups to work? Is the problem with the nsswitch entries? The dnsmasq configuration? Additional Information In the server's /etc/hosts: 192.168.0.1 mymachine In the server's nsswitch.conf: hosts: files mdns4_mininal [NOTFOUND=return] dns mdns4 (admittedly, this is a bit weird) In resolv.conf (which is generated by dnsmasq): nameserver 127.0.0.1 search mynicedomain.org In the server's /etc/hosts.allow: domain: ALL In the other machines' /etc/resolv.conf (this is set by the DHCP client): nameserver 192.168.0.1 search mynicedomain.org Finally, here's the ipconfig output from one of the client machines on the network (running Windows 7): Connection-specific DNS Suffix . : mynicedomain.org Description . . . . . . . . . . . : Intel(R) 82579LM Gigabit Network Connection Physical Address. . . . . . . . . : 12-34-56-78-9A-BC DHCP Enabled. . . . . . . . . . . : Yes Autoconfiguration Enabled . . . . : Yes IPv4 Address. . . . . . . . . . . : 192.168.0.50(Preferred) Subnet Mask . . . . . . . . . . . : 255.255.255.0 Lease Obtained. . . . . . . . . . : Sunday, October 20th 2013 16:20:25 Lease Expires . . . . . . . . . . : Sunday, October 20th 2013 18:20:24 Default Gateway . . . . . . . . . : 192.168.0.1 DHCP Server . . . . . . . . . . . : 192.168.0.1 DNS Servers . . . . . . . . . . . : 192.168.0.1 NetBIOS over Tcpip. . . . . . . . : Enabled Notes: May be related to this question.

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  • Nginx no static files after update

    - by SomeoneS
    First, i must say that i am not expert in server administration, my site was setup by hosting admins (that i cannot contact anymore). Few days ago, i updated Nginx to latest version (admin told me that it is safe to do). But after that, my site serves only html content, no CSS, images, JS. If i try to open some image i get message "Wellcome to Nginx" (same thin if i try to open static.mysitedomain.com). More details: Site has static. subdomain, but static files are in same directory as they used to be before setting up static files. I was googling for some solutions, i tried to change something in /etc/nginx/, but no luck. I feel that this is some minor configuration problem, any ideas? EDIT: Here is /etc/nginx/nginx.conf file content: user www-data; worker_processes 4; pid /var/run/nginx.pid; events { worker_connections 768; # multi_accept on; } http { ## # Basic Settings ## sendfile on; tcp_nopush on; tcp_nodelay on; keepalive_timeout 65; types_hash_max_size 2048; # server_tokens off; # server_names_hash_bucket_size 64; # server_name_in_redirect off; include /etc/nginx/mime.types; default_type application/octet-stream; ## # Logging Settings ## access_log /var/log/nginx/access.log; error_log /var/log/nginx/error.log; ## # Gzip Settings ## gzip on; gzip_disable "msie6"; # gzip_vary on; # gzip_proxied any; # gzip_comp_level 6; # gzip_buffers 16 8k; # gzip_http_version 1.1; # gzip_types text/plain text/css application/json application/x-javascript text/xml application/xml application/xml+rss text/javascript; ## # nginx-naxsi config ## # Uncomment it if you installed nginx-naxsi ## #include /etc/nginx/naxsi_core.rules; ## # nginx-passenger config ## # Uncomment it if you installed nginx-passenger ## #passenger_root /usr; #passenger_ruby /usr/bin/ruby; ## # Virtual Host Configs ## include /etc/nginx/conf.d/*.conf; include /etc/nginx/sites-enabled/*; } Here is /etc/nginx/sites-enabled/default file content: server { #listen 80; ## listen for ipv4; this line is default and implied #listen [::]:80 default ipv6only=on; ## listen for ipv6 root /usr/share/nginx/www; index index.html index.htm; # Make site accessible from http://localhost/ server_name localhost; location / { # First attempt to serve request as file, then # as directory, then fall back to index.html try_files $uri $uri/ /index.html; # Uncomment to enable naxsi on this location # include /etc/nginx/naxsi.rules } location /doc/ { alias /usr/share/doc/; autoindex on; allow 127.0.0.1; deny all; } # Only for nginx-naxsi : process denied requests #location /RequestDenied { # For example, return an error code #return 418; #} #error_page 404 /404.html; # redirect server error pages to the static page /50x.html # #error_page 500 502 503 504 /50x.html; #location = /50x.html { # root /usr/share/nginx/www; #} # pass the PHP scripts to FastCGI server listening on 127.0.0.1:9000 # #location ~ \.php$ { # fastcgi_split_path_info ^(.+\.php)(/.+)$; # # NOTE: You should have "cgi.fix_pathinfo = 0;" in php.ini # # # With php5-cgi alone: # fastcgi_pass 127.0.0.1:9000; # # With php5-fpm: # fastcgi_pass unix:/var/run/php5-fpm.sock; # fastcgi_index index.php; # include fastcgi_params; #} # deny access to .htaccess files, if Apache's document root # concurs with nginx's one # #location ~ /\.ht { # deny all; #} } # another virtual host using mix of IP-, name-, and port-based configuration # #server { # listen 8000; # listen somename:8080; # server_name somename alias another.alias; # root html; # index index.html index.htm; # # location / { # try_files $uri $uri/ /index.html; # } #} # HTTPS server # #server { # listen 443; # server_name localhost; # # root html; # index index.html index.htm; # # ssl on; # ssl_certificate cert.pem; # ssl_certificate_key cert.key; # # ssl_session_timeout 5m; # # ssl_protocols SSLv3 TLSv1; # ssl_ciphers ALL:!ADH:!EXPORT56:RC4+RSA:+HIGH:+MEDIUM:+LOW:+SSLv3:+EXP; # ssl_prefer_server_ciphers on; # # location / { # try_files $uri $uri/ /index.html; # } #}

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  • What is the best way to solve an Objective-C namespace collision?

    - by Mecki
    Objective-C has no namespaces; it's much like C, everything is within one global namespace. Common practice is to prefix classes with initials, e.g. if you are working at IBM, you could prefix them with "IBM"; if you work for Microsoft, you could use "MS"; and so on. Sometimes the initials refer to the project, e.g. Adium prefixes classes with "AI" (as there is no company behind it of that you could take the initials). Apple prefixes classes with NS and says this prefix is reserved for Apple only. So far so well. But appending 2 to 4 letters to a class name in front is a very, very limited namespace. E.g. MS or AI could have an entirely different meanings (AI could be Artificial Intelligence for example) and some other developer might decide to use them and create an equally named class. Bang, namespace collision. Okay, if this is a collision between one of your own classes and one of an external framework you are using, you can easily change the naming of your class, no big deal. But what if you use two external frameworks, both frameworks that you don't have the source to and that you can't change? Your application links with both of them and you get name conflicts. How would you go about solving these? What is the best way to work around them in such a way that you can still use both classes? In C you can work around these by not linking directly to the library, instead you load the library at runtime, using dlopen(), then find the symbol you are looking for using dlsym() and assign it to a global symbol (that you can name any way you like) and then access it through this global symbol. E.g. if you have a conflict because some C library has a function named open(), you could define a variable named myOpen and have it point to the open() function of the library, thus when you want to use the system open(), you just use open() and when you want to use the other one, you access it via the myOpen identifier. Is something similar possible in Objective-C and if not, is there any other clever, tricky solution you can use resolve namespace conflicts? Any ideas? Update: Just to clarify this: answers that suggest how to avoid namespace collisions in advance or how to create a better namespace are certainly welcome; however, I will not accept them as the answer since they don't solve my problem. I have two libraries and their class names collide. I can't change them; I don't have the source of either one. The collision is already there and tips on how it could have been avoided in advance won't help anymore. I can forward them to the developers of these frameworks and hope they choose a better namespace in the future, but for the time being I'm searching a solution to work with the frameworks right now within a single application. Any solutions to make this possible?

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  • DNS problems on CentOS fresh install

    - by Rick Koshi
    I'm having some DNS issues on a new box I'm installing with CentOS 6.2. I am able to look up names using nslookup, dig, or host. I am able to ping machines by name or by IP address. However, when I try other tools, such as ssh, wget, or yum, they are unable to resolve names. For example: # wget http://www.google.com --2012-03-08 14:48:06-- http://www.google.com/ Resolving www.google.com... failed: Name or service not known. wget: unable to resolve host address `www.google.com' # ssh www.google.com ssh: Could not resolve hostname www.google.com: Name or service not known # ping -c 1 www.google.com PING www.l.google.com (74.125.113.106) 56(84) bytes of data. 64 bytes from vw-in-f106.1e100.net (74.125.113.106): icmp_seq=1 ttl=46 time=43.6 ms --- www.l.google.com ping statistics --- 1 packets transmitted, 1 received, 0% packet loss, time 59ms rtt min/avg/max/mdev = 43.665/43.665/43.665/0.000 ms # host www.google.com www.google.com is an alias for www.l.google.com. www.l.google.com has address 74.125.113.99 www.l.google.com has address 74.125.113.103 www.l.google.com has address 74.125.113.104 www.l.google.com has address 74.125.113.105 www.l.google.com has address 74.125.113.106 www.l.google.com has address 74.125.113.147 My /etc/nsswitch.conf file is the default, including this (standard) line: hosts: files dns /etc/resolv.conf is as set up by DHCP: ; generated by /sbin/dhclient-script nameserver 192.168.1.254 192.168.1.254 is a working DNS server (my DSL modem, working for years with other machines) Anyone know why ping would work, but ssh/wget would fail? Per NcA's suggestion, I tried changing /etc/resolv.conf to point to 8.8.8.8. Oddly enough, this does make it work. Obviously, my DSL modem is responding to DNS requests in some way that some parts of Linux's resolution system don't like. Looking at the tcpdump, I am unable to see what the difference is. Certainly, both servers are sending the same addresses. Here's the output from tcpdump -nn -X with the server set to the DNS server on the DSL modem. It's clearly replying with the correct addresses, but ssh/wget don't seem happy with it for some reason: 15:53:52.133580 IP 192.168.1.254.53 > 192.168.1.2.54836: 33157 7/0/0 CNAME www.l.google.com., A 74.125.115.105, A 74.125.115.106, A 74.125.115.147, A 74.125.115.99, A 74.125.115.103, A 74.125.115.104 (148) 0x0000: 4500 00b0 e33a 0000 ff11 53b1 c0a8 01fe E....:....S..... 0x0010: c0a8 0102 0035 d634 009c 7528 8185 8180 .....5.4..u(.... 0x0020: 0001 0007 0000 0000 0377 7777 0667 6f6f .........www.goo 0x0030: 676c 6503 636f 6d00 0001 0001 c00c 0005 gle.com......... 0x0040: 0001 0007 acd0 0008 0377 7777 016c c010 .........www.l.. 0x0050: c02c 0001 0001 0000 0001 0004 4a7d 7369 .,..........J}si 0x0060: c02c 0001 0001 0000 0001 0004 4a7d 736a .,..........J}sj 0x0070: c02c 0001 0001 0000 0001 0004 4a7d 7393 .,..........J}s. 0x0080: c02c 0001 0001 0000 0001 0004 4a7d 7363 .,..........J}sc 0x0090: c02c 0001 0001 0000 0001 0004 4a7d 7367 .,..........J}sg 0x00a0: c02c 0001 0001 0000 0001 0004 4a7d 7368 .,..........J}sh 15:53:52.135669 IP 192.168.1.254.53 > 192.168.1.2.54836: 65062- 0/0/0 (32) 0x0000: 4500 003c e33b 0000 ff11 5424 c0a8 01fe E..<.;....T$.... 0x0010: c0a8 0102 0035 d634 0028 98f9 fe26 8000 .....5.4.(...&.. 0x0020: 0001 0000 0000 0000 0377 7777 0667 6f6f .........www.goo 0x0030: 676c 6503 636f 6d00 001c 0001 gle.com..... I'm not enough of an expert to know if this is malformed in some way, but ping seems to do the right thing with it. For comparison, here's the same thing when querying 8.8.8.8: 15:57:27.990270 IP 8.8.8.8.53 > 192.168.1.2.49028: 59114 7/0/0 CNAME www.l.google.com., A 74.125.113.105, A 74.125.113.103, A 74.125.113.106, A 74.125.113.147, A 74.125.113.104, A 74.125.113.99 (148) 0x0000: 4500 00b0 5530 0000 2f11 6453 0808 0808 E...U0../.dS.... 0x0010: c0a8 0102 0035 bf84 009c 39f8 e6ea 8180 .....5....9..... 0x0020: 0001 0007 0000 0000 0377 7777 0667 6f6f .........www.goo 0x0030: 676c 6503 636f 6d00 0001 0001 c00c 0005 gle.com......... 0x0040: 0001 0001 516a 0008 0377 7777 016c c010 ....Qj...www.l.. 0x0050: c02c 0001 0001 0000 0116 0004 4a7d 7169 .,..........J}qi 0x0060: c02c 0001 0001 0000 0116 0004 4a7d 7167 .,..........J}qg 0x0070: c02c 0001 0001 0000 0116 0004 4a7d 716a .,..........J}qj 0x0080: c02c 0001 0001 0000 0116 0004 4a7d 7193 .,..........J}q. 0x0090: c02c 0001 0001 0000 0116 0004 4a7d 7168 .,..........J}qh 0x00a0: c02c 0001 0001 0000 0116 0004 4a7d 7163 .,..........J}qc 15:57:28.018909 IP 8.8.8.8.53 > 192.168.1.2.49028: 31984 1/1/0 CNAME www.l.google.com. (102) 0x0000: 4500 0082 7b1b 0000 2f11 3e96 0808 0808 E...{.../.>..... 0x0010: c0a8 0102 0035 bf84 006e c67e 7cf0 8180 .....5...n.~|... 0x0020: 0001 0001 0001 0000 0377 7777 0667 6f6f .........www.goo 0x0030: 676c 6503 636f 6d00 001c 0001 c00c 0005 gle.com......... 0x0040: 0001 0001 517f 0008 0377 7777 016c c010 ....Q....www.l.. 0x0050: c030 0006 0001 0000 0258 0026 036e 7334 .0.......X.&.ns4 0x0060: c010 0964 6e73 2d61 646d 696e c010 0016 ...dns-admin.... 0x0070: 91f3 0000 0384 0000 0384 0000 0708 0000 ................ 0x0080: 003c .< I still don't know why the server's reply is adequate for ping but not for ssh/wget. If anyone has ideas, I'd be happy to hear them. For now, though, I can either refer to an outside DNS server or set up my own server on the new box. It's a workaround that seems like it should be unnecessary, but will allow me to proceed.

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  • LXC Container Networking

    - by digitaladdictions
    I just started to experiment with LXC containers. I was able to create a container and start it up but I cannot get dhcp to assign the container an IP address. If I assign a static address the container can ping the host IP but not outside the host IP. The host is CentOS 6.5 and the guest is Ubuntu 14.04LTS. I used the template downloaded by lxc-create -t download -n cn-01 command. If I am trying to get an IP address on the same subnet as the host I don't believe I should need the IP tables rule for masquerading but I added it anyways. Same with IP forwarding. I compiled LXC by hand from the following source https://linuxcontainers.org/downloads/lxc-1.0.4.tar.gz Host Operating System Version #> cat /etc/redhat-release CentOS release 6.5 (Final) #> uname -a Linux localhost.localdomain 2.6.32-431.20.3.el6.x86_64 #1 SMP Thu Jun 19 21:14:45 UTC 2014 x86_64 x86_64 x86_64 GNU/Linux Container Config #> cat /usr/local/var/lib/lxc/cn-01/config # Template used to create this container: /usr/local/share/lxc/templates/lxc-download # Parameters passed to the template: # For additional config options, please look at lxc.container.conf(5) # Distribution configuration lxc.include = /usr/local/share/lxc/config/ubuntu.common.conf lxc.arch = x86_64 # Container specific configuration lxc.rootfs = /usr/local/var/lib/lxc/cn-01/rootfs lxc.utsname = cn-01 # Network configuration lxc.network.type = veth lxc.network.flags = up lxc.network.link = br0 LXC default.confu 1500 qdisc pfifo_fast state UP qlen 1000 link/ether 00:0c:29:12:30:f2 brd ff:ff:ff:ff:f #> cat /usr/local/etc/lxc/default.conf lxc.network.type = veth lxc.network.link = br0 lxc.network.flags = up #> lxc-checkconfig Kernel configuration not found at /proc/config.gz; searching... Kernel configuration found at /boot/config-2.6.32-431.20.3.el6.x86_64 --- Namespaces --- Namespaces: enabled Utsname namespace: enabled Ipc namespace: enabled Pid namespace: enabled User namespace: enabled Network namespace: enabled Multiple /dev/pts instances: enabled --- Control groups --- Cgroup: enabled Cgroup namespace: enabled Cgroup device: enabled Cgroup sched: enabled Cgroup cpu account: enabled Cgroup memory controller: /usr/local/bin/lxc-checkconfig: line 103: [: too many arguments enabled Cgroup cpuset: enabled --- Misc --- Veth pair device: enabled Macvlan: enabled Vlan: enabled File capabilities: /usr/local/bin/lxc-checkconfig: line 118: [: -gt: unary operator expected Note : Before booting a new kernel, you can check its configuration usage : CONFIG=/path/to/config /usr/local/bin/lxc-checkconfig Network Config (HOST) #> cat /etc/sysconfig/network-scripts/ifcfg-br0 DEVICE=br0 TYPE=Bridge BOOTPROTO=dhcp ONBOOT=yes #> cat /etc/sysconfig/network-scripts/ifcfg-eth0 DEVICE=eth0 ONBOOT=yes TYPE=Ethernet IPV6INIT=no USERCTL=no BRIDGE=br0 #> cat /etc/networks default 0.0.0.0 loopback 127.0.0.0 link-local 169.254.0.0 #> ip a s 1: lo: <LOOPBACK,UP,LOWER_UP> mtu 16436 qdisc noqueue state UNKNOWN link/loopback 00:00:00:00:00:00 brd 00:00:00:00:00:00 inet 127.0.0.1/8 scope host lo inet6 ::1/128 scope host valid_lft forever preferred_lft forever 2: eth0: <BROADCAST,MULTICAST,UP,LOWER_UP> mtu 1500 qdisc pfifo_fast state UP qlen 1000 link/ether 00:0c:29:12:30:f2 brd ff:ff:ff:ff:ff:ff inet6 fe80::20c:29ff:fe12:30f2/64 scope link valid_lft forever preferred_lft forever 3: pan0: <BROADCAST,MULTICAST> mtu 1500 qdisc noop state DOWN link/ether 42:7e:43:b3:61:c5 brd ff:ff:ff:ff:ff:ff 4: br0: <BROADCAST,MULTICAST,UP,LOWER_UP> mtu 1500 qdisc noqueue state UNKNOWN link/ether 00:0c:29:12:30:f2 brd ff:ff:ff:ff:ff:ff inet 10.60.70.121/24 brd 10.60.70.255 scope global br0 inet6 fe80::20c:29ff:fe12:30f2/64 scope link valid_lft forever preferred_lft forever 12: vethT6BGL2: <BROADCAST,MULTICAST,UP,LOWER_UP> mtu 1500 qdisc pfifo_fast state UP qlen 1000 link/ether fe:a1:69:af:50:17 brd ff:ff:ff:ff:ff:ff inet6 fe80::fca1:69ff:feaf:5017/64 scope link valid_lft forever preferred_lft forever #> brctl show bridge name bridge id STP enabled interfaces br0 8000.000c291230f2 no eth0 vethT6BGL2 pan0 8000.000000000000 no #> cat /proc/sys/net/ipv4/ip_forward 1 # Generated by iptables-save v1.4.7 on Fri Jul 11 15:11:36 2014 *nat :PREROUTING ACCEPT [34:6287] :POSTROUTING ACCEPT [0:0] :OUTPUT ACCEPT [0:0] -A POSTROUTING -o eth0 -j MASQUERADE COMMIT # Completed on Fri Jul 11 15:11:36 2014 Network Config (Container) #> cat /etc/network/interfaces # This file describes the network interfaces available on your system # and how to activate them. For more information, see interfaces(5). # The loopback network interface auto lo iface lo inet loopback auto eth0 iface eth0 inet dhcp #> ip a s 11: eth0: <BROADCAST,MULTICAST,UP,LOWER_UP> mtu 1500 qdisc pfifo_fast state UP qlen 1000 link/ether 02:69:fb:42:ee:d7 brd ff:ff:ff:ff:ff:ff inet6 fe80::69:fbff:fe42:eed7/64 scope link valid_lft forever preferred_lft forever 13: lo: <LOOPBACK,UP,LOWER_UP> mtu 16436 qdisc noqueue state UNKNOWN link/loopback 00:00:00:00:00:00 brd 00:00:00:00:00:00 inet 127.0.0.1/8 scope host lo inet6 ::1/128 scope host valid_lft forever preferred_lft forever

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  • Data Profiling without SSIS

    Strangely enough for a predominantly SSIS blog, this post is all about how to perform data profiling without using SSIS. Whilst the Data Profiling Task is a worthy addition, there are a couple of limitations I’ve encountered of late. The first is that it requires SQL Server 2008, and not everyone is there yet. The second is that it can only target SQL Server 2005 and above. What about older systems, which are the ones that we probably need to investigate the most, or other vendor databases such as Oracle? With these limitations in mind I did some searching to find a quick and easy alternative to help me perform some data profiling for a project I was working on recently. I only had SQL Server 2005 available, and anyway most of my target source systems were Oracle, and of course I had short timescales. I looked at several options. Some never got beyond the download stage, they failed to install or just did not run, and others provided less than I could have produced myself by spending 2 minutes writing some basic SQL queries. In the end I settled on an open source product called DataCleaner. To quote from their website: DataCleaner is an Open Source application for profiling, validating and comparing data. These activities help you administer and monitor your data quality in order to ensure that your data is useful and applicable to your business situation. DataCleaner is the free alternative to software for master data management (MDM) methodologies, data warehousing (DW) projects, statistical research, preparation for extract-transform-load (ETL) activities and more. DataCleaner is developed in Java and licensed under LGPL. As quoted above it claims to support profiling, validating and comparing data, but I didn’t really get past the profiling functions, so won’t comment on the other two. The profiling whilst not prefect certainly saved some time compared to the limited alternatives. The ability to profile heterogeneous data sources is a big advantage over the SSIS option, and I found it overall quite easy to use and performance was good. I could see it struggling at times, but actually for what it does I was impressed. It had some data type niggles with Oracle, and some metrics seem a little strange, although thankfully they were easy to augment with some SQL queries to ensure a consistent picture. The report export options didn’t do it for me, but copy and paste with a bit of Excel magic was sufficient. One initial point for me personally is that I have had limited exposure to things of the Java persuasion and whilst I normally get by fine, sometimes the simplest things can throw me. For example installing a JDBC driver, why do I have to copy files to make it all work, has nobody ever heard of an MSI? In case there are other people out there like me who have become totally indoctrinated with the Microsoft software paradigm, I’ve written a quick start guide that details every step required. Steps 1- 5 are the key ones, the rest is really an excuse for some screenshots to show you the tool. Quick Start Guide Step 1  - Download Data Cleaner. The Microsoft Windows zipped exe option, and I chose the latest stable build, currently DataCleaner 1.5.3 (final). Extract the files to a suitable location. Step 2 - Download Java. If you try and run datacleaner.exe without Java it will warn you, and then open your default browser and take you to the Java download site. Follow the installation instructions from there, normally just click Download Java a couple of times and you’re done. Step 3 - Download Microsoft SQL Server JDBC Driver. You may have SQL Server installed, but you won’t have a JDBC driver. Version 3.0 is the latest as of April 2010. There is no real installer, we are in the Java world here, but run the exe you downloaded to extract the files. The default Unzip to folder is not much help, so try a fully qualified path such as C:\Program Files\Microsoft SQL Server JDBC Driver 3.0\ to ensure you can find the files afterwards. Step 4 - If you wish to use Windows Authentication to connect to your SQL Server then first we need to copy a file so that Data Cleaner can find it. Browse to the JDBC extract location from Step 3 and drill down to the file sqljdbc_auth.dll. You will have to choose the correct directory for your processor architecture. e.g. C:\Program Files\Microsoft SQL Server JDBC Driver 3.0\sqljdbc_3.0\enu\auth\x86\sqljdbc_auth.dll. Now copy this file to the Data Cleaner extract folder you chose in Step 1. An alternative method is to edit datacleaner.cmd in the data cleaner extract folder as detailed in this data cleaner wiki topic, but I find copying the file simpler. Step 5 – Now lets run Data Cleaner, just run datacleaner.exe from the extract folder you chose in Step 1. Step 6 – Complete or skip the registration screen, and ignore the task window for now. In the main window click settings. Step 7 – In the Settings dialog, select the Database drivers tab, then click Register database driver and select the Local JAR file option. Step 8 – Browse to the JDBC driver extract location from Step 3 and drill down to select sqljdbc4.jar. e.g. C:\Program Files\Microsoft SQL Server JDBC Driver 3.0\sqljdbc_3.0\enu\sqljdbc4.jar Step 9 – Select the Database driver class as com.microsoft.sqlserver.jdbc.SQLServerDriver, and then click the Test and Save database driver button. Step 10 - You should be back at the Settings dialog with a the list of drivers that includes SQL Server. Just click Save Settings to persist all your hard work. Step 11 – Now we can start to profile some data. In the main Data Cleaner window click New Task, and then Profile from the task window. Step 12 – In the Profile window click Open Database Step 13 – Now choose the SQL Server connection string option. Selecting a connection string gives us a template like jdbc:sqlserver://<hostname>:1433;databaseName=<database>, but obviously it requires some details to be entered for example  jdbc:sqlserver://localhost:1433;databaseName=SQLBits. This will connect to the database called SQLBits on my local machine. The port may also have to be changed if using such as when you have a multiple instances of SQL Server running. If using SQL Server Authentication enter a username and password as required and then click Connect to database. You can use Window Authentication, just add integratedSecurity=true to the end of your connection string. e.g jdbc:sqlserver://localhost:1433;databaseName=SQLBits;integratedSecurity=true.  If you didn’t complete Step 4 above you will need to do so now and restart Data Cleaner before it will work. Manually setting the connection string is fine, but creating a named connection makes more sense if you will be spending any length of time profiling a specific database. As highlighted in the left-hand screen-shot, at the bottom of the dialog it includes partial instructions on how to create named connections. In the folder shown C:\Users\<Username>\.datacleaner\1.5.3, open the datacleaner-config.xml file in your editor of choice add your own details. You’ll see a sample connection in the file already, just add yours following the same pattern. e.g. <!-- Darren's Named Connections --> <bean class="dk.eobjects.datacleaner.gui.model.NamedConnection"> <property name="name" value="SQLBits Local Connection" /> <property name="driverClass" value="com.microsoft.sqlserver.jdbc.SQLServerDriver" /> <property name="connectionString" value="jdbc:sqlserver://localhost:1433;databaseName=SQLBits;integratedSecurity=true" /> <property name="tableTypes"> <list> <value>TABLE</value> <value>VIEW</value> </list> </property> </bean> Step 14 – Once back at the Profile window, you should now see your schemas, tables and/or views listed down the left hand side. Browse this tree and double-click a table to select it for profiling. You can then click Add profile, and choose some profiling options, before finally clicking Run profiling. You can see below a sample output for three of the most common profiles, click the image for full size.   I hope this has given you a taster for DataCleaner, and should help you get up and running pretty quickly.

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  • Demystifying Silverlight Dependency Properties

    - by dwahlin
    I have the opportunity to teach a lot of people about Silverlight (amongst other technologies) and one of the topics that definitely confuses people initially is the concept of dependency properties. I confess that when I first heard about them my initial thought was “Why do we need a specialized type of property?” While you can certainly use standard CLR properties in Silverlight applications, Silverlight relies heavily on dependency properties for just about everything it does behind the scenes. In fact, dependency properties are an essential part of the data binding, template, style and animation functionality available in Silverlight. They simply back standard CLR properties. In this post I wanted to put together a (hopefully) simple explanation of dependency properties and why you should care about them if you’re currently working with Silverlight or looking to move to it.   What are Dependency Properties? XAML provides a great way to define layout controls, user input controls, shapes, colors and data binding expressions in a declarative manner. There’s a lot that goes on behind the scenes in order to make XAML work and an important part of that magic is the use of dependency properties. If you want to bind data to a property, style it, animate it or transform it in XAML then the property involved has to be a dependency property to work properly. If you’ve ever positioned a control in a Canvas using Canvas.Left or placed a control in a specific Grid row using Grid.Row then you’ve used an attached property which is a specialized type of dependency property. Dependency properties play a key role in XAML and the overall Silverlight framework. Any property that you bind, style, template, animate or transform must be a dependency property in Silverlight applications. You can programmatically bind values to controls and work with standard CLR properties, but if you want to use the built-in binding expressions available in XAML (one of my favorite features) or the Binding class available through code then dependency properties are a necessity. Dependency properties aren’t needed in every situation, but if you want to customize your application very much you’ll eventually end up needing them. For example, if you create a custom user control and want to expose a property that consumers can use to change the background color, you have to define it as a dependency property if you want bindings, styles and other features to be available for use. Now that the overall purpose of dependency properties has been discussed let’s take a look at how you can create them. Creating Dependency Properties When .NET first came out you had to write backing fields for each property that you defined as shown next: Brush _ScheduleBackground; public Brush ScheduleBackground { get { return _ScheduleBackground; } set { _ScheduleBackground = value; } } Although .NET 2.0 added auto-implemented properties (for example: public Brush ScheduleBackground { get; set; }) where the compiler would automatically generate the backing field used by get and set blocks, the concept is still the same as shown in the above code; a property acts as a wrapper around a field. Silverlight dependency properties replace the _ScheduleBackground field shown in the previous code and act as the backing store for a standard CLR property. The following code shows an example of defining a dependency property named ScheduleBackgroundProperty: public static readonly DependencyProperty ScheduleBackgroundProperty = DependencyProperty.Register("ScheduleBackground", typeof(Brush), typeof(Scheduler), null);   Looking through the code the first thing that may stand out is that the definition for ScheduleBackgroundProperty is marked as static and readonly and that the property appears to be of type DependencyProperty. This is a standard pattern that you’ll use when working with dependency properties. You’ll also notice that the property explicitly adds the word “Property” to the name which is another standard you’ll see followed. In addition to defining the property, the code also makes a call to the static DependencyProperty.Register method and passes the name of the property to register (ScheduleBackground in this case) as a string. The type of the property, the type of the class that owns the property and a null value (more on the null value later) are also passed. In this example a class named Scheduler acts as the owner. The code handles registering the property as a dependency property with the call to Register(), but there’s a little more work that has to be done to allow a value to be assigned to and retrieved from the dependency property. The following code shows the complete code that you’ll typically use when creating a dependency property. You can find code snippets that greatly simplify the process of creating dependency properties out on the web. The MVVM Light download available from http://mvvmlight.codeplex.com comes with built-in dependency properties snippets as well. public static readonly DependencyProperty ScheduleBackgroundProperty = DependencyProperty.Register("ScheduleBackground", typeof(Brush), typeof(Scheduler), null); public Brush ScheduleBackground { get { return (Brush)GetValue(ScheduleBackgroundProperty); } set { SetValue(ScheduleBackgroundProperty, value); } } The standard CLR property code shown above should look familiar since it simply wraps the dependency property. However, you’ll notice that the get and set blocks call GetValue and SetValue methods respectively to perform the appropriate operation on the dependency property. GetValue and SetValue are members of the DependencyObject class which is another key component of the Silverlight framework. Silverlight controls and classes (TextBox, UserControl, CompositeTransform, DataGrid, etc.) ultimately derive from DependencyObject in their inheritance hierarchy so that they can support dependency properties. Dependency properties defined in Silverlight controls and other classes tend to follow the pattern of registering the property by calling Register() and then wrapping the dependency property in a standard CLR property (as shown above). They have a standard property that wraps a registered dependency property and allows a value to be assigned and retrieved. If you need to expose a new property on a custom control that supports data binding expressions in XAML then you’ll follow this same pattern. Dependency properties are extremely useful once you understand why they’re needed and how they’re defined. Detecting Changes and Setting Defaults When working with dependency properties there will be times when you want to assign a default value or detect when a property changes so that you can keep the user interface in-sync with the property value. Silverlight’s DependencyProperty.Register() method provides a fourth parameter that accepts a PropertyMetadata object instance. PropertyMetadata can be used to hook a callback method to a dependency property. The callback method is called when the property value changes. PropertyMetadata can also be used to assign a default value to the dependency property. By assigning a value of null for the final parameter passed to Register() you’re telling the property that you don’t care about any changes and don’t have a default value to apply. Here are the different constructor overloads available on the PropertyMetadata class: PropertyMetadata Constructor Overload Description PropertyMetadata(Object) Used to assign a default value to a dependency property. PropertyMetadata(PropertyChangedCallback) Used to assign a property changed callback method. PropertyMetadata(Object, PropertyChangedCalback) Used to assign a default property value and a property changed callback.   There are many situations where you need to know when a dependency property changes or where you want to apply a default. Performing either task is easily accomplished by creating a new instance of the PropertyMetadata class and passing the appropriate values to its constructor. The following code shows an enhanced version of the initial dependency property code shown earlier that demonstrates these concepts: public Brush ScheduleBackground { get { return (Brush)GetValue(ScheduleBackgroundProperty); } set { SetValue(ScheduleBackgroundProperty, value); } } public static readonly DependencyProperty ScheduleBackgroundProperty = DependencyProperty.Register("ScheduleBackground", typeof(Brush), typeof(Scheduler), new PropertyMetadata(new SolidColorBrush(Colors.LightGray), ScheduleBackgroundChanged)); private static void ScheduleBackgroundChanged(DependencyObject d, DependencyPropertyChangedEventArgs e) { var scheduler = d as Scheduler; scheduler.Background = e.NewValue as Brush; } The code wires ScheduleBackgroundProperty to a property change callback method named ScheduleBackgroundChanged. What’s interesting is that this callback method is static (as is the dependency property) so it gets passed the instance of the object that owns the property that has changed (otherwise we wouldn’t be able to get to the object instance). In this example the dependency object is cast to a Scheduler object and its Background property is assigned to the new value of the dependency property. The code also handles assigning a default value of LightGray to the dependency property by creating a new instance of a SolidColorBrush. To Sum Up In this post you’ve seen the role of dependency properties and how they can be defined in code. They play a big role in XAML and the overall Silverlight framework. You can think of dependency properties as being replacements for fields that you’d normally use with standard CLR properties. In addition to a discussion on how dependency properties are created, you also saw how to use the PropertyMetadata class to define default dependency property values and hook a dependency property to a callback method. The most important thing to understand with dependency properties (especially if you’re new to Silverlight) is that they’re needed if you want a property to support data binding, animations, transformations and styles properly. Any time you create a property on a custom control or user control that has these types of requirements you’ll want to pick a dependency property over of a standard CLR property with a backing field. There’s more that can be covered with dependency properties including a related property called an attached property….more to come.

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  • Metro: Promises

    - by Stephen.Walther
    The goal of this blog entry is to describe the Promise class in the WinJS library. You can use promises whenever you need to perform an asynchronous operation such as retrieving data from a remote website or a file from the file system. Promises are used extensively in the WinJS library. Asynchronous Programming Some code executes immediately, some code requires time to complete or might never complete at all. For example, retrieving the value of a local variable is an immediate operation. Retrieving data from a remote website takes longer or might not complete at all. When an operation might take a long time to complete, you should write your code so that it executes asynchronously. Instead of waiting for an operation to complete, you should start the operation and then do something else until you receive a signal that the operation is complete. An analogy. Some telephone customer service lines require you to wait on hold – listening to really bad music – until a customer service representative is available. This is synchronous programming and very wasteful of your time. Some newer customer service lines enable you to enter your telephone number so the customer service representative can call you back when a customer representative becomes available. This approach is much less wasteful of your time because you can do useful things while waiting for the callback. There are several patterns that you can use to write code which executes asynchronously. The most popular pattern in JavaScript is the callback pattern. When you call a function which might take a long time to return a result, you pass a callback function to the function. For example, the following code (which uses jQuery) includes a function named getFlickrPhotos which returns photos from the Flickr website which match a set of tags (such as “dog” and “funny”): function getFlickrPhotos(tags, callback) { $.getJSON( "http://api.flickr.com/services/feeds/photos_public.gne?jsoncallback=?", { tags: tags, tagmode: "all", format: "json" }, function (data) { if (callback) { callback(data.items); } } ); } getFlickrPhotos("funny, dogs", function(data) { $.each(data, function(index, item) { console.log(item); }); }); The getFlickr() function includes a callback parameter. When you call the getFlickr() function, you pass a function to the callback parameter which gets executed when the getFlicker() function finishes retrieving the list of photos from the Flickr web service. In the code above, the callback function simply iterates through the results and writes each result to the console. Using callbacks is a natural way to perform asynchronous programming with JavaScript. Instead of waiting for an operation to complete, sitting there and listening to really bad music, you can get a callback when the operation is complete. Using Promises The CommonJS website defines a promise like this (http://wiki.commonjs.org/wiki/Promises): “Promises provide a well-defined interface for interacting with an object that represents the result of an action that is performed asynchronously, and may or may not be finished at any given point in time. By utilizing a standard interface, different components can return promises for asynchronous actions and consumers can utilize the promises in a predictable manner.” A promise provides a standard pattern for specifying callbacks. In the WinJS library, when you create a promise, you can specify three callbacks: a complete callback, a failure callback, and a progress callback. Promises are used extensively in the WinJS library. The methods in the animation library, the control library, and the binding library all use promises. For example, the xhr() method included in the WinJS base library returns a promise. The xhr() method wraps calls to the standard XmlHttpRequest object in a promise. The following code illustrates how you can use the xhr() method to perform an Ajax request which retrieves a file named Photos.txt: var options = { url: "/data/photos.txt" }; WinJS.xhr(options).then( function (xmlHttpRequest) { console.log("success"); var data = JSON.parse(xmlHttpRequest.responseText); console.log(data); }, function(xmlHttpRequest) { console.log("fail"); }, function(xmlHttpRequest) { console.log("progress"); } ) The WinJS.xhr() method returns a promise. The Promise class includes a then() method which accepts three callback functions: a complete callback, an error callback, and a progress callback: Promise.then(completeCallback, errorCallback, progressCallback) In the code above, three anonymous functions are passed to the then() method. The three callbacks simply write a message to the JavaScript Console. The complete callback also dumps all of the data retrieved from the photos.txt file. Creating Promises You can create your own promises by creating a new instance of the Promise class. The constructor for the Promise class requires a function which accepts three parameters: a complete, error, and progress function parameter. For example, the code below illustrates how you can create a method named wait10Seconds() which returns a promise. The progress function is called every second and the complete function is not called until 10 seconds have passed: (function () { "use strict"; var app = WinJS.Application; function wait10Seconds() { return new WinJS.Promise(function (complete, error, progress) { var seconds = 0; var intervalId = window.setInterval(function () { seconds++; progress(seconds); if (seconds > 9) { window.clearInterval(intervalId); complete(); } }, 1000); }); } app.onactivated = function (eventObject) { if (eventObject.detail.kind === Windows.ApplicationModel.Activation.ActivationKind.launch) { wait10Seconds().then( function () { console.log("complete") }, function () { console.log("error") }, function (seconds) { console.log("progress:" + seconds) } ); } } app.start(); })(); All of the work happens in the constructor function for the promise. The window.setInterval() method is used to execute code every second. Every second, the progress() callback method is called. If more than 10 seconds have passed then the complete() callback method is called and the clearInterval() method is called. When you execute the code above, you can see the output in the Visual Studio JavaScript Console. Creating a Timeout Promise In the previous section, we created a custom Promise which uses the window.setInterval() method to complete the promise after 10 seconds. We really did not need to create a custom promise because the Promise class already includes a static method for returning promises which complete after a certain interval. The code below illustrates how you can use the timeout() method. The timeout() method returns a promise which completes after a certain number of milliseconds. WinJS.Promise.timeout(3000).then( function(){console.log("complete")}, function(){console.log("error")}, function(){console.log("progress")} ); In the code above, the Promise completes after 3 seconds (3000 milliseconds). The Promise returned by the timeout() method does not support progress events. Therefore, the only message written to the console is the message “complete” after 10 seconds. Canceling Promises Some promises, but not all, support cancellation. When you cancel a promise, the promise’s error callback is executed. For example, the following code uses the WinJS.xhr() method to perform an Ajax request. However, immediately after the Ajax request is made, the request is cancelled. // Specify Ajax request options var options = { url: "/data/photos.txt" }; // Make the Ajax request var request = WinJS.xhr(options).then( function (xmlHttpRequest) { console.log("success"); }, function (xmlHttpRequest) { console.log("fail"); }, function (xmlHttpRequest) { console.log("progress"); } ); // Cancel the Ajax request request.cancel(); When you run the code above, the message “fail” is written to the Visual Studio JavaScript Console. Composing Promises You can build promises out of other promises. In other words, you can compose promises. There are two static methods of the Promise class which you can use to compose promises: the join() method and the any() method. When you join promises, a promise is complete when all of the joined promises are complete. When you use the any() method, a promise is complete when any of the promises complete. The following code illustrates how to use the join() method. A new promise is created out of two timeout promises. The new promise does not complete until both of the timeout promises complete: WinJS.Promise.join([WinJS.Promise.timeout(1000), WinJS.Promise.timeout(5000)]) .then(function () { console.log("complete"); }); The message “complete” will not be written to the JavaScript Console until both promises passed to the join() method completes. The message won’t be written for 5 seconds (5,000 milliseconds). The any() method completes when any promise passed to the any() method completes: WinJS.Promise.any([WinJS.Promise.timeout(1000), WinJS.Promise.timeout(5000)]) .then(function () { console.log("complete"); }); The code above writes the message “complete” to the JavaScript Console after 1 second (1,000 milliseconds). The message is written to the JavaScript console immediately after the first promise completes and before the second promise completes. Summary The goal of this blog entry was to describe WinJS promises. First, we discussed how promises enable you to easily write code which performs asynchronous actions. You learned how to use a promise when performing an Ajax request. Next, we discussed how you can create your own promises. You learned how to create a new promise by creating a constructor function with complete, error, and progress parameters. Finally, you learned about several advanced methods of promises. You learned how to use the timeout() method to create promises which complete after an interval of time. You also learned how to cancel promises and compose promises from other promises.

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  • Metro: Understanding Observables

    - by Stephen.Walther
    The goal of this blog entry is to describe how the Observer Pattern is implemented in the WinJS library. You learn how to create observable objects which trigger notifications automatically when their properties are changed. Observables enable you to keep your user interface and your application data in sync. For example, by taking advantage of observables, you can update your user interface automatically whenever the properties of a product change. Observables are the foundation of declarative binding in the WinJS library. The WinJS library is not the first JavaScript library to include support for observables. For example, both the KnockoutJS library and the Microsoft Ajax Library (now part of the Ajax Control Toolkit) support observables. Creating an Observable Imagine that I have created a product object like this: var product = { name: "Milk", description: "Something to drink", price: 12.33 }; Nothing very exciting about this product. It has three properties named name, description, and price. Now, imagine that I want to be notified automatically whenever any of these properties are changed. In that case, I can create an observable product from my product object like this: var observableProduct = WinJS.Binding.as(product); This line of code creates a new JavaScript object named observableProduct from the existing JavaScript object named product. This new object also has a name, description, and price property. However, unlike the properties of the original product object, the properties of the observable product object trigger notifications when the properties are changed. Each of the properties of the new observable product object has been changed into accessor properties which have both a getter and a setter. For example, the observable product price property looks something like this: price: { get: function () { return this.getProperty(“price”); } set: function (value) { this.setProperty(“price”, value); } } When you read the price property then the getProperty() method is called and when you set the price property then the setProperty() method is called. The getProperty() and setProperty() methods are methods of the observable product object. The observable product object supports the following methods and properties: · addProperty(name, value) – Adds a new property to an observable and notifies any listeners. · backingData – An object which represents the value of each property. · bind(name, action) – Enables you to execute a function when a property changes. · getProperty(name) – Returns the value of a property using the string name of the property. · notify(name, newValue, oldValue) – A private method which executes each function in the _listeners array. · removeProperty(name) – Removes a property and notifies any listeners. · setProperty(name, value) – Updates a property and notifies any listeners. · unbind(name, action) – Enables you to stop executing a function in response to a property change. · updateProperty(name, value) – Updates a property and notifies any listeners. So when you create an observable, you get a new object with the same properties as an existing object. However, when you modify the properties of an observable object, then you can notify any listeners of the observable that the value of a particular property has changed automatically. Imagine that you change the value of the price property like this: observableProduct.price = 2.99; In that case, the following sequence of events is triggered: 1. The price setter calls the setProperty(“price”, 2.99) method 2. The setProperty() method updates the value of the backingData.price property and calls the notify() method 3. The notify() method executes each function in the collection of listeners associated with the price property Creating Observable Listeners If you want to be notified when a property of an observable object is changed, then you need to register a listener. You register a listener by using the bind() method like this: (function () { "use strict"; var app = WinJS.Application; app.onactivated = function (eventObject) { if (eventObject.detail.kind === Windows.ApplicationModel.Activation.ActivationKind.launch) { // Simple product object var product = { name: "Milk", description: "Something to drink", price: 12.33 }; // Create observable product var observableProduct = WinJS.Binding.as(product); // Execute a function when price is changed observableProduct.bind("price", function (newValue) { console.log(newValue); }); // Change the price observableProduct.price = 2.99; } }; app.start(); })(); In the code above, the bind() method is used to associate the price property with a function. When the price property is changed, the function logs the new value of the price property to the Visual Studio JavaScript console. The price property is associated with the function using the following line of code: // Execute a function when price is changed observableProduct.bind("price", function (newValue) { console.log(newValue); }); Coalescing Notifications If you make multiple changes to a property – one change immediately following another – then separate notifications won’t be sent. Instead, any listeners are notified only once. The notifications are coalesced into a single notification. For example, in the following code, the product price property is updated three times. However, only one message is written to the JavaScript console. Only the last value assigned to the price property is written to the JavaScript Console window: // Simple product object var product = { name: "Milk", description: "Something to drink", price: 12.33 }; // Create observable product var observableProduct = WinJS.Binding.as(product); // Execute a function when price is changed observableProduct.bind("price", function (newValue) { console.log(newValue); }); // Change the price observableProduct.price = 3.99; observableProduct.price = 2.99; observableProduct.price = 1.99; Only the last value assigned to price, the value 1.99, appears in the console: If there is a time delay between changes to a property then changes result in different notifications. For example, the following code updates the price property every second: // Simple product object var product = { name: "Milk", description: "Something to drink", price: 12.33 }; // Create observable product var observableProduct = WinJS.Binding.as(product); // Execute a function when price is changed observableProduct.bind("price", function (newValue) { console.log(newValue); }); // Add 1 to price every second window.setInterval(function () { observableProduct.price += 1; }, 1000); In this case, separate notification messages are logged to the JavaScript Console window: If you need to prevent multiple notifications from being coalesced into one then you can take advantage of promises. I discussed WinJS promises in a previous blog entry: http://stephenwalther.com/blog/archive/2012/02/22/windows-web-applications-promises.aspx Because the updateProperty() method returns a promise, you can create different notifications for each change in a property by using the following code: // Change the price observableProduct.updateProperty("price", 3.99) .then(function () { observableProduct.updateProperty("price", 2.99) .then(function () { observableProduct.updateProperty("price", 1.99); }); }); In this case, even though the price is immediately changed from 3.99 to 2.99 to 1.99, separate notifications for each new value of the price property are sent. Bypassing Notifications Normally, if a property of an observable object has listeners and you change the property then the listeners are notified. However, there are certain situations in which you might want to bypass notification. In other words, you might need to change a property value silently without triggering any functions registered for notification. If you want to change a property without triggering notifications then you should change the property by using the backingData property. The following code illustrates how you can change the price property silently: // Simple product object var product = { name: "Milk", description: "Something to drink", price: 12.33 }; // Create observable product var observableProduct = WinJS.Binding.as(product); // Execute a function when price is changed observableProduct.bind("price", function (newValue) { console.log(newValue); }); // Change the price silently observableProduct.backingData.price = 5.99; console.log(observableProduct.price); // Writes 5.99 The price is changed to the value 5.99 by changing the value of backingData.price. Because the observableProduct.price property is not set directly, any listeners associated with the price property are not notified. When you change the value of a property by using the backingData property, the change in the property happens synchronously. However, when you change the value of an observable property directly, the change is always made asynchronously. Summary The goal of this blog entry was to describe observables. In particular, we discussed how to create observables from existing JavaScript objects and bind functions to observable properties. You also learned how notifications are coalesced (and ways to prevent this coalescing). Finally, we discussed how you can use the backingData property to update an observable property without triggering notifications. In the next blog entry, we’ll see how observables are used with declarative binding to display the values of properties in an HTML document.

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  • Adding an Admin user to an ASP.NET MVC 4 application using a single drop-in file

    - by Jon Galloway
    I'm working on an ASP.NET MVC 4 tutorial and wanted to set it up so just dropping a file in App_Start would create a user named "Owner" and assign them to the "Administrator" role (more explanation at the end if you're interested). There are reasons why this wouldn't fit into most application scenarios: It's not efficient, as it checks for (and creates, if necessary) the user every time the app starts up The username, password, and role name are hardcoded in the app (although they could be pulled from config) Automatically creating an administrative account in code (without user interaction) could lead to obvious security issues if the user isn't informed However, with some modifications it might be more broadly useful - e.g. creating a test user with limited privileges, ensuring a required account isn't accidentally deleted, or - as in my case - setting up an account for demonstration or tutorial purposes. Challenge #1: Running on startup without requiring the user to install or configure anything I wanted to see if this could be done just by having the user drop a file into the App_Start folder and go. No copying code into Global.asax.cs, no installing addition NuGet packages, etc. That may not be the best approach - perhaps a NuGet package with a dependency on WebActivator would be better - but I wanted to see if this was possible and see if it offered the best experience. Fortunately ASP.NET 4 and later provide a PreApplicationStartMethod attribute which allows you to register a method which will run when the application starts up. You drop this attribute in your application and give it two parameters: a method name and the type that contains it. I created a static class named PreApplicationTasks with a static method named, then dropped this attribute in it: [assembly: PreApplicationStartMethod(typeof(PreApplicationTasks), "Initializer")] That's it. One small gotcha: the namespace can be a problem with assembly attributes. I decided my class didn't need a namespace. Challenge #2: Only one PreApplicationStartMethod per assembly In .NET 4, the PreApplicationStartMethod is marked as AllMultiple=false, so you can only have one PreApplicationStartMethod per assembly. This was fixed in .NET 4.5, as noted by Jon Skeet, so you can have as many PreApplicationStartMethods as you want (allowing you to keep your users waiting for the application to start indefinitely!). The WebActivator NuGet package solves the multiple instance problem if you're in .NET 4 - it registers as a PreApplicationStartMethod, then calls any methods you've indicated using [assembly: WebActivator.PreApplicationStartMethod(type, method)]. David Ebbo blogged about that here:  Light up your NuGets with startup code and WebActivator. In my scenario (bootstrapping a beginner level tutorial) I decided not to worry about this and stick with PreApplicationStartMethod. Challenge #3: PreApplicationStartMethod kicks in before configuration has been read This is by design, as Phil explains. It allows you to make changes that need to happen very early in the pipeline, well before Application_Start. That's fine in some cases, but it caused me problems when trying to add users, since the Membership Provider configuration hadn't yet been read - I got an exception stating that "Default Membership Provider could not be found." The solution here is to run code that requires configuration in a PostApplicationStart method. But how to do that? Challenge #4: Getting PostApplicationStartMethod without requiring WebActivator The WebActivator NuGet package, among other things, provides a PostApplicationStartMethod attribute. That's generally how I'd recommend running code that needs to happen after Application_Start: [assembly: WebActivator.PostApplicationStartMethod(typeof(TestLibrary.MyStartupCode), "CallMeAfterAppStart")] This works well, but I wanted to see if this would be possible without WebActivator. Hmm. Well, wait a minute - WebActivator works in .NET 4, so clearly it's registering and calling PostApplicationStartup tasks somehow. Off to the source code! Sure enough, there's even a handy comment in ActivationManager.cs which shows where PostApplicationStartup tasks are being registered: public static void Run() { if (!_hasInited) { RunPreStartMethods(); // Register our module to handle any Post Start methods. But outside of ASP.NET, just run them now if (HostingEnvironment.IsHosted) { Microsoft.Web.Infrastructure.DynamicModuleHelper.DynamicModuleUtility.RegisterModule(typeof(StartMethodCallingModule)); } else { RunPostStartMethods(); } _hasInited = true; } } Excellent. Hey, that DynamicModuleUtility seems familiar... Sure enough, K. Scott Allen mentioned it on his blog last year. This is really slick - a PreApplicationStartMethod can register a new HttpModule in code. Modules are run right after application startup, so that's a perfect time to do any startup stuff that requires configuration to be read. As K. Scott says, it's this easy: using System; using System.Web; using Microsoft.Web.Infrastructure.DynamicModuleHelper; [assembly:PreApplicationStartMethod(typeof(MyAppStart), "Start")] public class CoolModule : IHttpModule { // implementation not important // imagine something cool here } public static class MyAppStart { public static void Start() { DynamicModuleUtility.RegisterModule(typeof(CoolModule)); } } Challenge #5: Cooperating with SimpleMembership The ASP.NET MVC Internet template includes SimpleMembership. SimpleMembership is a big improvement over traditional ASP.NET Membership. For one thing, rather than forcing a database schema, it can work with your database schema. In the MVC 4 Internet template case, it uses Entity Framework Code First to define the user model. SimpleMembership bootstrap includes a call to InitializeDatabaseConnection, and I want to play nice with that. There's a new [InitializeSimpleMembership] attribute on the AccountController, which calls \Filters\InitializeSimpleMembershipAttribute.cs::OnActionExecuting(). That comment in that method that says "Ensure ASP.NET Simple Membership is initialized only once per app start" which sounds like good advice. I figured the best thing would be to call that directly: new Mvc4SampleApplication.Filters.InitializeSimpleMembershipAttribute().OnActionExecuting(null); I'm not 100% happy with this - in fact, it's my least favorite part of this solution. There are two problems - first, directly calling a method on a filter, while legal, seems odd. Worse, though, the Filter lives in the application's namespace, which means that this code no longer works well as a generic drop-in. The simplest workaround would be to duplicate the relevant SimpleMembership initialization code into my startup code, but I'd rather not. I'm interested in your suggestions here. Challenge #6: Module Init methods are called more than once When debugging, I noticed (and remembered) that the Init method may be called more than once per page request - it's run once per instance in the app pool, and an individual page request can cause multiple resource requests to the server. While SimpleMembership does have internal checks to prevent duplicate user or role entries, I'd rather not cause or handle those exceptions. So here's the standard single-use lock in the Module's init method: void IHttpModule.Init(HttpApplication context) { lock (lockObject) { if (!initialized) { //Do stuff } initialized = true; } } Putting it all together With all of that out of the way, here's the code I came up with: using Mvc4SampleApplication.Filters; using System.Web; using System.Web.Security; using WebMatrix.WebData; [assembly: PreApplicationStartMethod(typeof(PreApplicationTasks), "Initializer")] public static class PreApplicationTasks { public static void Initializer() { Microsoft.Web.Infrastructure.DynamicModuleHelper.DynamicModuleUtility .RegisterModule(typeof(UserInitializationModule)); } } public class UserInitializationModule : IHttpModule { private static bool initialized; private static object lockObject = new object(); private const string _username = "Owner"; private const string _password = "p@ssword123"; private const string _role = "Administrator"; void IHttpModule.Init(HttpApplication context) { lock (lockObject) { if (!initialized) { new InitializeSimpleMembershipAttribute().OnActionExecuting(null); if (!WebSecurity.UserExists(_username)) WebSecurity.CreateUserAndAccount(_username, _password); if (!Roles.RoleExists(_role)) Roles.CreateRole(_role); if (!Roles.IsUserInRole(_username, _role)) Roles.AddUserToRole(_username, _role); } initialized = true; } } void IHttpModule.Dispose() { } } The Verdict: Is this a good thing? Maybe. I think you'll agree that the journey was undoubtedly worthwhile, as it took us through some of the finer points of hooking into application startup, integrating with membership, and understanding why the WebActivator NuGet package is so useful Will I use this in the tutorial? I'm leaning towards no - I think a NuGet package with a dependency on WebActivator might work better: It's a little more clear what's going on Installing a NuGet package might be a little less error prone than copying a file A novice user could uninstall the package when complete It's a good introduction to NuGet, which is a good thing for beginners to see This code either requires either duplicating a little code from that filter or modifying the file to use the namespace Honestly I'm undecided at this point, but I'm glad that I can weigh the options. If you're interested: Why are you doing this? I'm updating the MVC Music Store tutorial to ASP.NET MVC 4, taking advantage of a lot of new ASP.NET MVC 4 features and trying to simplify areas that are giving people trouble. One change that addresses both needs us using the new OAuth support for membership as much as possible - it's a great new feature from an application perspective, and we get a fair amount of beginners struggling with setting up membership on a variety of database and development setups, which is a distraction from the focus of the tutorial - learning ASP.NET MVC. Side note: Thanks to some great help from Rick Anderson, we had a draft of the tutorial that was looking pretty good earlier this summer, but there were enough changes in ASP.NET MVC 4 all the way up to RTM that there's still some work to be done. It's high priority and should be out very soon. The one issue I ran into with OAuth is that we still need an Administrative user who can edit the store's inventory. I thought about a number of solutions for that - making the first user to register the admin, or the first user to use the username "Administrator" is assigned to the Administrator role - but they both ended up requiring extra code; also, I worried that people would use that code without understanding it or thinking about whether it was a good fit.

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  • How to tell if SPARC T4 crypto is being used?

    - by danx
    A question that often comes up when running applications on SPARC T4 systems is "How can I tell if hardware crypto accleration is being used?" To review, the SPARC T4 processor includes a crypto unit that supports several crypto instructions. For hardware crypto these include 11 AES instructions, 4 xmul* instructions (for AES GCM carryless multiply), mont for Montgomery multiply (optimizes RSA and DSA), and 5 des_* instructions (for DES3). For hardware hash algorithm optimization, the T4 has the md5, sha1, sha256, and sha512 instructions (the last two are used for SHA-224 an SHA-384). First off, it's easy to tell if the processor T4 crypto instructions—use the isainfo -v command and look for "sparcv9" and "aes" (and other hash and crypto algorithms) in the output: $ 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 These instructions are not-privileged, so are available for direct use in user-level applications and libraries (such as OpenSSL). Here is the "openssl speed -evp" command shown with the built-in t4 engine and with the pkcs11 engine. Both run the T4 AES instructions, but the t4 engine is faster than the pkcs11 engine because it has less overhead (especially for smaller packet sizes): t-4 $ /usr/bin/openssl version OpenSSL 1.0.0j 10 May 2012 t-4 $ /usr/bin/openssl engine (t4) SPARC T4 engine support (dynamic) Dynamic engine loading support (pkcs11) PKCS #11 engine support t-4 $ /usr/bin/openssl speed -evp aes-128-cbc # t4 engine used by default . . . The 'numbers' are in 1000s of bytes per second processed. type 16 bytes 64 bytes 256 bytes 1024 bytes 8192 bytes aes-128-cbc 487777.10k 816822.21k 986012.59k 1017029.97k 1053543.08k t-4 $ /usr/bin/openssl speed -engine pkcs11 -evp aes-128-cbc engine "pkcs11" set. . . . The 'numbers' are in 1000s of bytes per second processed. type 16 bytes 64 bytes 256 bytes 1024 bytes 8192 bytes aes-128-cbc 31703.58k 116636.39k 350672.81k 696170.50k 993599.49k Note: The "-evp" flag indicates use the OpenSSL "EnVeloPe" API, which gives more accurate results. That's because it tells OpenSSL to use the same API that external programs use when calling OpenSSL libcrypto functions, evp(3openssl). DTrace Shows if T4 Crypto Functions Are Used OK, good enough, the isainfo(1) command shows the instructions are present, but how does one know if they are being used? Chi-Chang Lin, who works on Oracle Solaris performance, wrote a Dtrace script to show if T4 instructions are being executed. To show the T4 instructions are being used, run the following Dtrace script. Look for functions named "t4" and "yf" in the output. The OpenSSL T4 engine uses functions named "t4" and the PKCS#11 engine uses functions named "yf". To demonstrate, I'll first run "openssl speed" with the built-in t4 engine then with the pkcs11 engine. The performance numbers are not valid due to dtrace probes slowing things down. t-4 # dtrace -Z -n ' pid$target::*yf*:entry,pid$target::*t4_*:entry{ @[probemod, probefunc] = count();}' \ -c "/usr/bin/openssl speed -evp aes-128-cbc" dtrace: description 'pid$target::*yf*:entry' matched 101 probes . . . dtrace: pid 2029 has exited libcrypto.so.1.0.0 ENGINE_load_t4 1 libcrypto.so.1.0.0 t4_DH 1 libcrypto.so.1.0.0 t4_DSA 1 libcrypto.so.1.0.0 t4_RSA 1 libcrypto.so.1.0.0 t4_destroy 1 libcrypto.so.1.0.0 t4_free_aes_ctr_NIDs 1 libcrypto.so.1.0.0 t4_init 1 libcrypto.so.1.0.0 t4_add_NID 3 libcrypto.so.1.0.0 t4_aes_expand128 5 libcrypto.so.1.0.0 t4_cipher_init_aes 5 libcrypto.so.1.0.0 t4_get_all_ciphers 6 libcrypto.so.1.0.0 t4_get_all_digests 59 libcrypto.so.1.0.0 t4_digest_final_sha1 65 libcrypto.so.1.0.0 t4_digest_init_sha1 65 libcrypto.so.1.0.0 t4_sha1_multiblock 126 libcrypto.so.1.0.0 t4_digest_update_sha1 261 libcrypto.so.1.0.0 t4_aes128_cbc_encrypt 1432979 libcrypto.so.1.0.0 t4_aes128_load_keys_for_encrypt 1432979 libcrypto.so.1.0.0 t4_cipher_do_aes_128_cbc 1432979 t-4 # dtrace -Z -n 'pid$target::*yf*:entry{ @[probemod, probefunc] = count();}   pid$target::*yf*:entry,pid$target::*t4_*:entry{ @[probemod, probefunc] = count();}' \ -c "/usr/bin/openssl speed -engine pkcs11 -evp aes-128-cbc" dtrace: description 'pid$target::*yf*:entry' matched 101 probes engine "pkcs11" set. . . . dtrace: pid 2033 has exited libcrypto.so.1.0.0 ENGINE_load_t4 1 libcrypto.so.1.0.0 t4_DH 1 libcrypto.so.1.0.0 t4_DSA 1 libcrypto.so.1.0.0 t4_RSA 1 libcrypto.so.1.0.0 t4_destroy 1 libcrypto.so.1.0.0 t4_free_aes_ctr_NIDs 1 libcrypto.so.1.0.0 t4_get_all_ciphers 1 libcrypto.so.1.0.0 t4_get_all_digests 1 libsoftcrypto.so.1 rijndael_key_setup_enc_yf 1 libsoftcrypto.so.1 yf_aes_expand128 1 libcrypto.so.1.0.0 t4_add_NID 3 libsoftcrypto.so.1 yf_aes128_cbc_encrypt 1542330 libsoftcrypto.so.1 yf_aes128_load_keys_for_encrypt 1542330 So, as shown above the OpenSSL built-in t4 engine executes t4_* functions (which are hand-coded assembly executing the T4 AES instructions) and the OpenSSL pkcs11 engine executes *yf* functions. Programmatic Use of OpenSSL T4 engine The OpenSSL t4 engine is used automatically with the /usr/bin/openssl command line. Chi-Chang Lin also points out that if you're calling the OpenSSL API (libcrypto.so) from a program, you must call ENGINE_load_built_engines(), otherwise the built-in t4 engine will not be loaded. You do not call ENGINE_set_default(). That's because "openssl speed -evp" test calls ENGINE_load_built_engines() even though the "-engine" option wasn't specified. OpenSSL T4 engine Availability The OpenSSL t4 engine is available with Solaris 11 and 11.1. For Solaris 10 08/11 (U10), you need to use the OpenSSL pkcs311 engine. The OpenSSL t4 engine is distributed only with the version of OpenSSL distributed with Solaris (and not third-party or self-compiled versions of OpenSSL). The OpenSSL engine implements the AES cipher for Solaris 11, released 11/2011. For Solaris 11.1, released 11/2012, the OpenSSL engine adds optimization for the MD5, SHA-1, and SHA-2 hash algorithms, and DES-3. Although the T4 processor has Camillia and Kasumi block cipher instructions, these are not implemented in the OpenSSL T4 engine. The following charts may help view availability of optimizations. The first chart shows what's available with Solaris CLIs and APIs, the second chart shows what's available in Solaris OpenSSL. Native Solaris Optimization for SPARC T4 This table is shows Solaris native CLI and API support. As such, they are all available with the OpenSSL pkcs11 engine. CLIs: "openssl -engine pkcs11", encrypt(1), decrypt(1), mac(1), digest(1), MD5sum(1), SHA1sum(1), SHA224sum(1), SHA256sum(1), SHA384sum(1), SHA512sum(1) APIs: PKCS#11 library libpkcs11(3LIB) (incluDES Openssl pkcs11 engine), libMD(3LIB), and Solaris kernel modules AlgorithmSolaris 1008/11 (U10)Solaris 11Solaris 11.1 AES-ECB, AES-CBC, AES-CTR, AES-CBC AES-CFB128 XXX DES3-ECB, DES3-CBC, DES2-ECB, DES2-CBC, DES-ECB, DES-CBC XXX bignum Montgomery multiply (RSA, DSA) XXX MD5, SHA-1, SHA-256, SHA-384, SHA-512 XXX SHA-224 X ARCFOUR (RC4) X Solaris OpenSSL T4 Engine Optimization This table is for the Solaris OpenSSL built-in t4 engine. Algorithms listed above are also available through the OpenSSL pkcs11 engine. CLI: openssl(1openssl) APIs: openssl(5), engine(3openssl), evp(3openssl), libcrypto crypto(3openssl) AlgorithmSolaris 11Solaris 11SRU2Solaris 11.1 AES-ECB, AES-CBC, AES-CTR, AES-CBC AES-CFB128 XXX DES3-ECB, DES3-CBC, DES-ECB, DES-CBC X bignum Montgomery multiply (RSA, DSA) X MD5, SHA-1, SHA-256, SHA-384, SHA-512 XX SHA-224 X Source Code Availability Solaris Most of the T4 assembly code that called the new T4 crypto instructions was written by Ferenc Rákóczi of the Solaris Security group, with assistance from others. You can download the Solaris source for this and other parts of Solaris as a few zip files at the Oracle Download website. The relevant source files are generally under directories usr/src/common/crypto/{aes,arcfour,des,md5,modes,sha1,sha2}}/sun4v/. and usr/src/common/bignum/sun4v/. Solaris 11 binary is available from the Oracle Solaris 11 download website. OpenSSL t4 engine The source for the OpenSSL t4 engine, which is based on the Solaris source above, is viewable through the OpenGrok source code browser in directory src/components/openssl/openssl-1.0.0/engines/t4 . You can download the source from the same website or through Mercurial source code management, hg(1). Conclusion Oracle Solaris with SPARC T4 provides a rich set of accelerated cryptographic and hash algorithms. Using the latest update, Solaris 11.1, provides the best set of optimized algorithms, but alternatives are often available, sometimes slightly slower, for releases back to Solaris 10 08/11 (U10). Reference See also these earlier blogs. SPARC T4 OpenSSL Engine by myself, Dan Anderson (2011), discusses the Openssl T4 engine and reviews the SPARC T4 processor for the Solaris 11 release. Exciting Crypto Advances with the T4 processor and Oracle Solaris 11 by Valerie Fenwick (2011) discusses crypto algorithms that were optimized for the T4 processor with the Solaris 11 FCS (11/11) and Solaris 10 08/11 (U10) release. T4 Crypto Cheat Sheet by Stefan Hinker (2012) discusses how to make T4 crypto optimization available to various consumers (such as SSH, Java, OpenSSL, Apache, etc.) High Performance Security For Oracle Database and Fusion Middleware Applications using SPARC T4 (PDF, 2012) discusses SPARC T4 and its usage to optimize application security. Configuring Oracle iPlanet WebServer / Oracle Traffic Director to use crypto accelerators on T4-1 servers by Meena Vyas (2012)

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  • Using a portable USB monitor in Ubuntu 13.04 (AOC e1649Fwu - DisplayLink)

    Having access to a little bit of IT hardware extravaganza isn't that easy here in Mauritius for exactly two reasons - either it is simply not available or it is expensive like nowhere. Well, by chance I came across an advert by a local hardware supplier and their offer of the week caught my attention - a portable USB monitor. Sounds cool, and the specs are okay as well. It's completely driven via USB 2.0, has a light weight, the dimensions would fit into my laptop bag and the resolution of 1366 x 768 pixels is okay for a second screen. Long story, short ending: I called them and only got to understand that they are out of stock - how convenient! Well, as usual I left some contact details and got the regular 'We call you back' answer. Surprisingly, I didn't receive a phone call as promised and after starting to complain via social media networks they finally came back to me with new units available - and *drum-roll* still the same price tag as promoted (and free delivery on top as one of their employees lives in Flic en Flac). Guess, it was a no-brainer to get at least one unit to fool around with. In worst case it might end up as image frame on the shelf or so... The usual suspects... Ubuntu first! Of course, the packing mentions only Windows or Mac OS as supported operating systems and without hesitation at all, I hooked up the device on my main machine running on Ubuntu 13.04. Result: Blackout... Hm, actually not the situation I was looking for but okay can't be too difficult to get this piece of hardware up and running. Following the output of syslogd (or dmesg if you prefer) the device has been recognised successfully but we got stuck in the initialisation phase. Oct 12 08:17:23 iospc2 kernel: [69818.689137] usb 2-4: new high-speed USB device number 5 using ehci-pciOct 12 08:17:23 iospc2 kernel: [69818.800306] usb 2-4: device descriptor read/64, error -32Oct 12 08:17:24 iospc2 kernel: [69819.043620] usb 2-4: New USB device found, idVendor=17e9, idProduct=4107Oct 12 08:17:24 iospc2 kernel: [69819.043630] usb 2-4: New USB device strings: Mfr=1, Product=2, SerialNumber=3Oct 12 08:17:24 iospc2 kernel: [69819.043636] usb 2-4: Product: e1649FwuOct 12 08:17:24 iospc2 kernel: [69819.043642] usb 2-4: Manufacturer: DisplayLinkOct 12 08:17:24 iospc2 kernel: [69819.043647] usb 2-4: SerialNumber: FJBD7HA000778Oct 12 08:17:24 iospc2 kernel: [69819.046073] hid-generic 0003:17E9:4107.0008: hiddev0,hidraw5: USB HID v1.10 Device [DisplayLink e1649Fwu] on usb-0000:00:1d.7-4/input1Oct 12 08:17:24 iospc2 mtp-probe: checking bus 2, device 5: "/sys/devices/pci0000:00/0000:00:1d.7/usb2/2-4"Oct 12 08:17:24 iospc2 mtp-probe: bus: 2, device: 5 was not an MTP deviceOct 12 08:17:30 iospc2 kernel: [69825.411220] [drm] vendor descriptor length:17 data:17 5f 01 00 15 05 00 01 03 00 04Oct 12 08:17:30 iospc2 kernel: [69825.498778] udl 2-4:1.0: fb1: udldrmfb frame buffer deviceOct 12 08:17:30 iospc2 kernel: [69825.498786] [drm] Initialized udl 0.0.1 20120220 on minor 1Oct 12 08:17:30 iospc2 kernel: [69825.498909] usbcore: registered new interface driver udl The device has been recognised as USB device without any question and it is listed properly: # lsusb...Bus 002 Device 005: ID 17e9:4107 DisplayLink ... A quick and dirty research on the net gave me some hints towards the udlfb framebuffer device for USB DisplayLink devices. By default this kernel module is blacklisted $ less /etc/modprobe.d/blacklist-framebuffer.conf | grep udl#blacklist udlblacklist udlfb and it is recommended to load it manually. So, unloading the whole udl stack and giving udlfb a shot: Oct 12 08:22:31 iospc2 kernel: [70126.642809] usbcore: registered new interface driver udlfb But still no reaction on the external display which supposedly should have been on and green. Display okay? Test run on Windows Just to be on the safe side and to exclude any hardware related defects or whatsoever - you never know what happened during delivery. I moved the display to a new position on the opposite side of my laptop, installed the display drivers first in Windows Vista (I know, I know...) as recommended in the manual, and then finally hooked it up on that machine. Tada! Display has been recognised correctly and I have a proper choice between cloning and extending my desktop. Testing whether the display is working properly - using Windows Vista Okay, good to know that there is nothing wrong on the hardware side just software... Back to Ubuntu - Kernel too old Some more research on Google and various hits recommend that the original displaylink driver has been merged into the recent kernel development and one should manually upgrade the kernel image (and both header) packages for Ubuntu. At least kernel 3.9 or higher would be necessary, and so I went out to this URL: http://kernel.ubuntu.com/~kernel-ppa/mainline/ and I downloaded all the good stuff from the v3.9-raring directory. The installation itself is easy going via dpkg: $ sudo dpkg -i linux-image-3.9.0-030900-generic_3.9.0-030900.201304291257_amd64.deb$ sudo dpkg -i linux-headers-3.9.0-030900_3.9.0-030900.201304291257_all.deb$ sudo dpkg -i linux-headers-3.9.0-030900-generic_3.9.0-030900.201304291257_amd64.deb As with any kernel upgrades it is necessary to restart the system in order to use the new one. Said and done: $ uname -r3.9.0-030900-generic And now connecting the external display gives me the following output in /var/log/syslog: Oct 12 17:51:36 iospc2 kernel: [ 2314.984293] usb 2-4: new high-speed USB device number 6 using ehci-pciOct 12 17:51:36 iospc2 kernel: [ 2315.096257] usb 2-4: device descriptor read/64, error -32Oct 12 17:51:36 iospc2 kernel: [ 2315.337105] usb 2-4: New USB device found, idVendor=17e9, idProduct=4107Oct 12 17:51:36 iospc2 kernel: [ 2315.337115] usb 2-4: New USB device strings: Mfr=1, Product=2, SerialNumber=3Oct 12 17:51:36 iospc2 kernel: [ 2315.337122] usb 2-4: Product: e1649FwuOct 12 17:51:36 iospc2 kernel: [ 2315.337127] usb 2-4: Manufacturer: DisplayLinkOct 12 17:51:36 iospc2 kernel: [ 2315.337132] usb 2-4: SerialNumber: FJBD7HA000778Oct 12 17:51:36 iospc2 kernel: [ 2315.338292] udlfb: DisplayLink e1649Fwu - serial #FJBD7HA000778Oct 12 17:51:36 iospc2 kernel: [ 2315.338299] udlfb: vid_17e9&pid_4107&rev_0129 driver's dlfb_data struct at ffff880117e59000Oct 12 17:51:36 iospc2 kernel: [ 2315.338303] udlfb: console enable=1Oct 12 17:51:36 iospc2 kernel: [ 2315.338306] udlfb: fb_defio enable=1Oct 12 17:51:36 iospc2 kernel: [ 2315.338309] udlfb: shadow enable=1Oct 12 17:51:36 iospc2 kernel: [ 2315.338468] udlfb: vendor descriptor length:17 data:17 5f 01 0015 05 00 01 03 00 04Oct 12 17:51:36 iospc2 kernel: [ 2315.338473] udlfb: DL chip limited to 1500000 pixel modesOct 12 17:51:36 iospc2 kernel: [ 2315.338565] udlfb: allocated 4 65024 byte urbsOct 12 17:51:36 iospc2 kernel: [ 2315.343592] hid-generic 0003:17E9:4107.0009: hiddev0,hidraw5: USB HID v1.10 Device [DisplayLink e1649Fwu] on usb-0000:00:1d.7-4/input1Oct 12 17:51:36 iospc2 mtp-probe: checking bus 2, device 6: "/sys/devices/pci0000:00/0000:00:1d.7/usb2/2-4"Oct 12 17:51:36 iospc2 mtp-probe: bus: 2, device: 6 was not an MTP deviceOct 12 17:51:36 iospc2 kernel: [ 2315.426583] udlfb: 1366x768 @ 59 Hz valid modeOct 12 17:51:36 iospc2 kernel: [ 2315.426589] udlfb: Reallocating framebuffer. Addresses will change!Oct 12 17:51:36 iospc2 kernel: [ 2315.428338] udlfb: 1366x768 @ 59 Hz valid modeOct 12 17:51:36 iospc2 kernel: [ 2315.428343] udlfb: set_par mode 1366x768Oct 12 17:51:36 iospc2 kernel: [ 2315.430620] udlfb: DisplayLink USB device /dev/fb1 attached. 1366x768 resolution. Using 4104K framebuffer memory Okay, that's looks more promising but still only blackout on the external screen... And yes, due to my previous modifications I swapped the blacklisted kernel modules: $ less /etc/modprobe.d/blacklist-framebuffer.conf | grep udlblacklist udl#blacklist udlfb Silly me! Okay, back to the original situation in which udl is allowed and udlfb blacklisted. Now, the logging looks similar to this and the screen shows those maroon-brown and azure-blue horizontal bars as described on other online resources. Oct 15 21:27:23 iospc2 kernel: [80934.308238] usb 2-4: new high-speed USB device number 5 using ehci-pciOct 15 21:27:23 iospc2 kernel: [80934.420244] usb 2-4: device descriptor read/64, error -32Oct 15 21:27:24 iospc2 kernel: [80934.660822] usb 2-4: New USB device found, idVendor=17e9, idProduct=4107Oct 15 21:27:24 iospc2 kernel: [80934.660832] usb 2-4: New USB device strings: Mfr=1, Product=2, SerialNumber=3Oct 15 21:27:24 iospc2 kernel: [80934.660838] usb 2-4: Product: e1649FwuOct 15 21:27:24 iospc2 kernel: [80934.660844] usb 2-4: Manufacturer: DisplayLinkOct 15 21:27:24 iospc2 kernel: [80934.660850] usb 2-4: SerialNumber: FJBD7HA000778Oct 15 21:27:24 iospc2 kernel: [80934.663391] hid-generic 0003:17E9:4107.0008: hiddev0,hidraw5: USB HID v1.10 Device [DisplayLink e1649Fwu] on usb-0000:00:1d.7-4/input1Oct 15 21:27:24 iospc2 mtp-probe: checking bus 2, device 5: "/sys/devices/pci0000:00/0000:00:1d.7/usb2/2-4"Oct 15 21:27:24 iospc2 mtp-probe: bus: 2, device: 5 was not an MTP deviceOct 15 21:27:25 iospc2 kernel: [80935.742407] [drm] vendor descriptor length:17 data:17 5f 01 00 15 05 00 01 03 00 04Oct 15 21:27:25 iospc2 kernel: [80935.834403] udl 2-4:1.0: fb1: udldrmfb frame buffer deviceOct 15 21:27:25 iospc2 kernel: [80935.834416] [drm] Initialized udl 0.0.1 20120220 on minor 1Oct 15 21:27:25 iospc2 kernel: [80935.836389] usbcore: registered new interface driver udlOct 15 21:27:25 iospc2 kernel: [80936.021458] [drm] write mode info 153 Next, it's time to enable the display for our needs... This can be done either via UI or console, just as you'd prefer it. Adding the external USB display under Linux isn't an issue after all... Settings Manager => Display Personally, I like the console. With the help of xrandr we get the screen identifier first $ xrandrScreen 0: minimum 320 x 200, current 3200 x 1080, maximum 32767 x 32767LVDS1 connected 1280x800+0+0 (normal left inverted right x axis y axis) 331mm x 207mm...DVI-0 connected 1366x768+0+0 (normal left inverted right x axis y axis) 344mm x 193mm   1366x768       60.0*+ and then give it the usual shot with auto-configuration. Let the system decide what's best for your hardware... $ xrandr --output DVI-0 --off$ xrandr --output DVI-0 --auto And there we go... Cloned output of main display: New kernel, new display... The external USB display works out-of-the-box with a Linux kernel > 3.9.0. Despite of a good number of resources it is absolutely not necessary to create a Device or Screen section in one of Xorg.conf files. This information belongs to the past and is not valid on kernel 3.9 or higher. Same hardware but Windows 8 Of course, I wanted to know how the latest incarnation from Redmond would handle the new hardware... Flawless! Most interesting aspect here: I did not use the driver installation medium on purpose. And I was right... not too long afterwards a dialog with the EULA of DisplayLink appeared on the main screen. And after confirmation of same it took some more seconds and the external USB monitor was ready to rumble. Well, and not only that one... but see for yourself. This time Windows 8 was the easiest solution after all. Resume I can highly recommend this type of hardware to anyone asking me. Although, it's dimensions are 15.6" it is actually lighter than my Samsung Galaxy Tab 10.1 and it still fits into my laptop bag without any issues. From now on... no more single screen while developing software on the road!

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  • Introduction to LinqPad Driver for StreamInsight 2.1

    - by Roman Schindlauer
    We are announcing the availability of the LinqPad driver for StreamInsight 2.1. The purpose of this blog post is to offer a quick introduction into the new features that we added to the StreamInsight LinqPad driver. We’ll show you how to connect to a remote server, how to inspect the entities present of that server, how to compose on top of them and how to manage their lifetime. Installing the driver Info on how to install the driver can be found in an earlier blog post here. Establishing connections As you click on the “Add Connection” link in the left pane you will notice that now it’s possible to build the data context automatically. The new driver appears as an option in the upper list, and if you pick it you will open a connection dialog that lets you connect to a remote StreamInsight server. The connection dialog lets you specify the address of the remote server. You will notice that it’s possible to pick up the binding information from the configuration file of the LinqPad application (which is normally in the same folder as LinqPad.exe and is called LinqPad.exe.config). In order for the context to be generated you need to pick an application from the server. The control is editable hence you can create a new application if you don’t want to make changes to an existing application. If you choose a new application name you will be prompted for confirmation before this gets created. Once you click OK the connection is created and you can start issuing queries against the remote server. If there’s any connectivity error the connection is marked with a red X and you can see the error message informing you what went wrong (i.e., the remote server could not be reached etc.). The context for remote servers Let’s take a look at what happens after we are connected successfully. Every LinqPad query runs inside a context – think of it as a class that wraps all the code that you’re writing. If you’re connecting to a live server the context will contain the following: The application object itself. All entities present in this application (sources, sinks, subjects and processes). The picture below shows a snapshot of the left pane of LinqPad after a successful connection. Every entity on the server has a different icon which will allow users to figure out its purpose. You will also notice that some entities have a string in parentheses following the name. It should be interpreted as such: the first name is the name of the property of the context class and the second name is the name of the entity as it exists on the server. Not all valid entity names are valid identifier names so in cases where we had to make a transformation you see both. Note also that as you hover over the entities you get IntelliSense with their types – more on that later. Remoting is not supported As you play with the entities exposed by the context you will notice that you can’t read and write directly to/from them. If for instance you’re trying to dump the content of an entity you will get an error message telling you that in the current version remoting is not supported. This is because the entity lives on the remote server and dumping its content means reading the events produced by this entity into the local process. ObservableSource.Dump(); Will yield the following error: Reading from a remote 'System.Reactive.Linq.IQbservable`1[System.Int32]' is not supported. Use the 'Microsoft.ComplexEventProcessing.Linq.RemoteProvider.Bind' method to read from the source using a remote observer. This basically tells you that you can call the Bind() method to direct the output of this source to a sink that has to be defined on the remote machine as well. You can’t bring the results to the LinqPad window unless you write code specifically for that. Compose queries You may ask – what's the purpose of all that? After all the same information is present in the EventFlowDebugger, why bother with showing it in LinqPad? First of all, What gets exposed in LinqPad is not what you see in the debugger. In LinqPad we have a property on the context class for every entity that lives on the server. Because LinqPad offers IntelliSense we in fact have much more information about the entity, and more importantly we can compose with that entity very easily. For example, let’s say that this code creates an entity: using (var server = Server.Connect(...)) {     var a = server.CreateApplication("WhiteFish");     var src = a         .DefineObservable<int>(() => Observable.Range(0, 3))         .Deploy("ObservableSource"); If later we want to compose with the source we have to fetch it and then we can bind something to     a.GetObservable<int>("ObservableSource)").Bind(... This means that we had to know a bunch of things about this: that it’s a source, that it’s an observable, it produces a result with payload Int32 and it’s named “ObservableSource”. Only the second and last bits of information are present in the debugger, by the way. As you type in the query window you see that all the entities are present, you get IntelliSense support for them and it’s much easier to make sense of what’s available. Let’s look at a scenario where composition is plausible. With the new programming model it’s possible to create “cold” sources that are parameterized. There was a way to accomplish that even in the previous version by passing parameters to the adapters, but this time it’s much more elegant because the expression declares what parameters are required. Say that we hover the mouse over the ThrottledSource source – we will see that its type is Func<int, int, IQbservable<int>> - this in effect means that we need to pass two int parameters before we can get a source that produces events, and the type for those events is int – in the particular case of my example I had the source produce a range of integers and the two parameters were the start and end of the range. So we see how a developer can create a source that is not running yet. Then someone else (e.g. an administrator) can pass whatever parameters appropriate and run the process. Proxy Types Here’s an interesting scenario – what if someone created a source on a server but they forgot to tell you what type they used. Worse yet, they might have used an anonymous type and even though they can refer to it by name you can’t figure out how to use that type. Let’s walk through an example that shows how you can compose against types you don’t need to have the definition of. This is how we can create a source that returns an anonymous type: Application.DefineObservable(() => Observable.Range(1, 10).Select(i => new { I = i })).Deploy("O1"); Now if we refresh the connection we can see the new source named O1 appear in the list. But what’s more important is that we now have a type to work with. So we can compose a query that refers to the anonymous type. var threshold = new StreamInsightDynamicDriver.TypeProxies.AnonymousType1_0<int>(5); var filter = from i in O1              where i > threshold              select i; filter.Deploy("O2"); You will notice that the anonymous type defined with this statement: new { I = i } can now be manipulated by a client that does not have access to it because the LinqPad driver has generated another type in its stead, named StreamInsightDynamicDriver.TypeProxies.AnonymousType1_0. This type has all the properties and fields of the type defined on the server, except in this case we can instantiate values and use it to compose more queries. It is worth noting that the same thing works for types that are not anonymous – the test is if the LinqPad driver can resolve the type or not. If it’s not possible then a new type will be generated that approximates the type that exists on the server. Control metadata In addition to composing processes on top of the existing entities we can do other useful things. We can delete them – nothing new here as we simply access the entities through the Entities collection of the application class. Here is where having their real name in parentheses comes handy. There’s another way to find out what’s behind a property – dump its expression. The first line in the output tells us what’s the name of the entity used to build this property in the context. Runtime information So let’s create a process to see what happens. We can bind a source to a sink and run the resulting process. If you right click on the connection you can refresh it and see the process present in the list of entities. Then you can drag the process to the query window and see that you can have access to process object in the Processes collection of the application. You can then manipulate the process (delete it, read its diagnostic view etc.). Regards, The StreamInsight Team

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  • Getting Started with TypeScript – Classes, Static Types and Interfaces

    - by dwahlin
    I had the opportunity to speak on different JavaScript topics at DevConnections in Las Vegas this fall and heard a lot of interesting comments about JavaScript as I talked with people. The most frequent comment I heard from people was, “I guess it’s time to start learning JavaScript”. Yep – if you don’t already know JavaScript then it’s time to learn it. As HTML5 becomes more and more popular the amount of JavaScript code written will definitely increase. After all, many of the HTML5 features available in browsers have little to do with “tags” and more to do with JavaScript (web workers, web sockets, canvas, local storage, etc.). As the amount of JavaScript code being used in applications increases, it’s more important than ever to structure the code in a way that’s maintainable and easy to debug. While JavaScript patterns can certainly be used (check out my previous posts on the subject or my course on Pluralsight.com), several alternatives have come onto the scene such as CoffeeScript, Dart and TypeScript. In this post I’ll describe some of the features TypeScript offers and the benefits that they can potentially offer enterprise-scale JavaScript applications. It’s important to note that while TypeScript has several great features, it’s definitely not for everyone or every project especially given how new it is. The goal of this post isn’t to convince you to use TypeScript instead of standard JavaScript….I’m a big fan of JavaScript. Instead, I’ll present several TypeScript features and let you make the decision as to whether TypeScript is a good fit for your applications. TypeScript Overview Here’s the official definition of TypeScript from the http://typescriptlang.org site: “TypeScript is a language for application-scale JavaScript development. TypeScript is a typed superset of JavaScript that compiles to plain JavaScript. Any browser. Any host. Any OS. Open Source.” TypeScript was created by Anders Hejlsberg (the creator of the C# language) and his team at Microsoft. To sum it up, TypeScript is a new language that can be compiled to JavaScript much like alternatives such as CoffeeScript or Dart. It isn’t a stand-alone language that’s completely separate from JavaScript’s roots though. It’s a superset of JavaScript which means that standard JavaScript code can be placed in a TypeScript file (a file with a .ts extension) and used directly. That’s a very important point/feature of the language since it means you can use existing code and frameworks with TypeScript without having to do major code conversions to make it all work. Once a TypeScript file is saved it can be compiled to JavaScript using TypeScript’s tsc.exe compiler tool or by using a variety of editors/tools. TypeScript offers several key features. First, it provides built-in type support meaning that you define variables and function parameters as being “string”, “number”, “bool”, and more to avoid incorrect types being assigned to variables or passed to functions. Second, TypeScript provides a way to write modular code by directly supporting class and module definitions and it even provides support for custom interfaces that can be used to drive consistency. Finally, TypeScript integrates with several different tools such as Visual Studio, Sublime Text, Emacs, and Vi to provide syntax highlighting, code help, build support, and more depending on the editor. Find out more about editor support at http://www.typescriptlang.org/#Download. TypeScript can also be used with existing JavaScript frameworks such as Node.js, jQuery, and others and even catch type issues and provide enhanced code help. Special “declaration” files that have a d.ts extension are available for Node.js, jQuery, and other libraries out-of-the-box. Visit http://typescript.codeplex.com/SourceControl/changeset/view/fe3bc0bfce1f#samples%2fjquery%2fjquery.d.ts for an example of a jQuery TypeScript declaration file that can be used with tools such as Visual Studio 2012 to provide additional code help and ensure that a string isn’t passed to a parameter that expects a number. Although declaration files certainly aren’t required, TypeScript’s support for declaration files makes it easier to catch issues upfront while working with existing libraries such as jQuery. In the future I expect TypeScript declaration files will be released for different HTML5 APIs such as canvas, local storage, and others as well as some of the more popular JavaScript libraries and frameworks. Getting Started with TypeScript To get started learning TypeScript visit the TypeScript Playground available at http://www.typescriptlang.org. Using the playground editor you can experiment with TypeScript code, get code help as you type, and see the JavaScript that TypeScript generates once it’s compiled. Here’s an example of the TypeScript playground in action:   One of the first things that may stand out to you about the code shown above is that classes can be defined in TypeScript. This makes it easy to group related variables and functions into a container which helps tremendously with re-use and maintainability especially in enterprise-scale JavaScript applications. While you can certainly simulate classes using JavaScript patterns (note that ECMAScript 6 will support classes directly), TypeScript makes it quite easy especially if you come from an object-oriented programming background. An example of the Greeter class shown in the TypeScript Playground is shown next: class Greeter { greeting: string; constructor (message: string) { this.greeting = message; } greet() { return "Hello, " + this.greeting; } } Looking through the code you’ll notice that static types can be defined on variables and parameters such as greeting: string, that constructors can be defined, and that functions can be defined such as greet(). The ability to define static types is a key feature of TypeScript (and where its name comes from) that can help identify bugs upfront before even running the code. Many types are supported including primitive types like string, number, bool, undefined, and null as well as object literals and more complex types such as HTMLInputElement (for an <input> tag). Custom types can be defined as well. The JavaScript output by compiling the TypeScript Greeter class (using an editor like Visual Studio, Sublime Text, or the tsc.exe compiler) is shown next: var Greeter = (function () { function Greeter(message) { this.greeting = message; } Greeter.prototype.greet = function () { return "Hello, " + this.greeting; }; return Greeter; })(); Notice that the code is using JavaScript prototyping and closures to simulate a Greeter class in JavaScript. The body of the code is wrapped with a self-invoking function to take the variables and functions out of the global JavaScript scope. This is important feature that helps avoid naming collisions between variables and functions. In cases where you’d like to wrap a class in a naming container (similar to a namespace in C# or a package in Java) you can use TypeScript’s module keyword. The following code shows an example of wrapping an AcmeCorp module around the Greeter class. In order to create a new instance of Greeter the module name must now be used. This can help avoid naming collisions that may occur with the Greeter class.   module AcmeCorp { export class Greeter { greeting: string; constructor (message: string) { this.greeting = message; } greet() { return "Hello, " + this.greeting; } } } var greeter = new AcmeCorp.Greeter("world"); In addition to being able to define custom classes and modules in TypeScript, you can also take advantage of inheritance by using TypeScript’s extends keyword. The following code shows an example of using inheritance to define two report objects:   class Report { name: string; constructor (name: string) { this.name = name; } print() { alert("Report: " + this.name); } } class FinanceReport extends Report { constructor (name: string) { super(name); } print() { alert("Finance Report: " + this.name); } getLineItems() { alert("5 line items"); } } var report = new FinanceReport("Month's Sales"); report.print(); report.getLineItems();   In this example a base Report class is defined that has a variable (name), a constructor that accepts a name parameter of type string, and a function named print(). The FinanceReport class inherits from Report by using TypeScript’s extends keyword. As a result, it automatically has access to the print() function in the base class. In this example the FinanceReport overrides the base class’s print() method and adds its own. The FinanceReport class also forwards the name value it receives in the constructor to the base class using the super() call. TypeScript also supports the creation of custom interfaces when you need to provide consistency across a set of objects. The following code shows an example of an interface named Thing (from the TypeScript samples) and a class named Plane that implements the interface to drive consistency across the app. Notice that the Plane class includes intersect and normal as a result of implementing the interface.   interface Thing { intersect: (ray: Ray) => Intersection; normal: (pos: Vector) => Vector; surface: Surface; } class Plane implements Thing { normal: (pos: Vector) =>Vector; intersect: (ray: Ray) =>Intersection; constructor (norm: Vector, offset: number, public surface: Surface) { this.normal = function (pos: Vector) { return norm; } this.intersect = function (ray: Ray): Intersection { var denom = Vector.dot(norm, ray.dir); if (denom > 0) { return null; } else { var dist = (Vector.dot(norm, ray.start) + offset) / (-denom); return { thing: this, ray: ray, dist: dist }; } } } }   At first glance it doesn’t appear that the surface member is implemented in Plane but it’s actually included automatically due to the public surface: Surface parameter in the constructor. Adding public varName: Type to a constructor automatically adds a typed variable into the class without having to explicitly write the code as with normal and intersect. TypeScript has additional language features but defining static types and creating classes, modules, and interfaces are some of the key features it offers. So is TypeScript right for you and your applications? That’s a not a question that I or anyone else can answer for you. You’ll need to give it a spin to see what you think. In future posts I’ll discuss additional details about TypeScript and how it can be used with enterprise-scale JavaScript applications. In the meantime, I’m in the process of working with John Papa on a new Typescript course for Pluralsight that we hope to have out in December of 2012.

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  • Introduction to LinqPad Driver for StreamInsight 2.1

    - by Roman Schindlauer
    We are announcing the availability of the LinqPad driver for StreamInsight 2.1. The purpose of this blog post is to offer a quick introduction into the new features that we added to the StreamInsight LinqPad driver. We’ll show you how to connect to a remote server, how to inspect the entities present of that server, how to compose on top of them and how to manage their lifetime. Installing the driver Info on how to install the driver can be found in an earlier blog post here. Establishing connections As you click on the “Add Connection” link in the left pane you will notice that now it’s possible to build the data context automatically. The new driver appears as an option in the upper list, and if you pick it you will open a connection dialog that lets you connect to a remote StreamInsight server. The connection dialog lets you specify the address of the remote server. You will notice that it’s possible to pick up the binding information from the configuration file of the LinqPad application (which is normally in the same folder as LinqPad.exe and is called LinqPad.exe.config). In order for the context to be generated you need to pick an application from the server. The control is editable hence you can create a new application if you don’t want to make changes to an existing application. If you choose a new application name you will be prompted for confirmation before this gets created. Once you click OK the connection is created and you can start issuing queries against the remote server. If there’s any connectivity error the connection is marked with a red X and you can see the error message informing you what went wrong (i.e., the remote server could not be reached etc.). The context for remote servers Let’s take a look at what happens after we are connected successfully. Every LinqPad query runs inside a context – think of it as a class that wraps all the code that you’re writing. If you’re connecting to a live server the context will contain the following: The application object itself. All entities present in this application (sources, sinks, subjects and processes). The picture below shows a snapshot of the left pane of LinqPad after a successful connection. Every entity on the server has a different icon which will allow users to figure out its purpose. You will also notice that some entities have a string in parentheses following the name. It should be interpreted as such: the first name is the name of the property of the context class and the second name is the name of the entity as it exists on the server. Not all valid entity names are valid identifier names so in cases where we had to make a transformation you see both. Note also that as you hover over the entities you get IntelliSense with their types – more on that later. Remoting is not supported As you play with the entities exposed by the context you will notice that you can’t read and write directly to/from them. If for instance you’re trying to dump the content of an entity you will get an error message telling you that in the current version remoting is not supported. This is because the entity lives on the remote server and dumping its content means reading the events produced by this entity into the local process. ObservableSource.Dump(); Will yield the following error: Reading from a remote 'System.Reactive.Linq.IQbservable`1[System.Int32]' is not supported. Use the 'Microsoft.ComplexEventProcessing.Linq.RemoteProvider.Bind' method to read from the source using a remote observer. This basically tells you that you can call the Bind() method to direct the output of this source to a sink that has to be defined on the remote machine as well. You can’t bring the results to the LinqPad window unless you write code specifically for that. Compose queries You may ask – what's the purpose of all that? After all the same information is present in the EventFlowDebugger, why bother with showing it in LinqPad? First of all, What gets exposed in LinqPad is not what you see in the debugger. In LinqPad we have a property on the context class for every entity that lives on the server. Because LinqPad offers IntelliSense we in fact have much more information about the entity, and more importantly we can compose with that entity very easily. For example, let’s say that this code creates an entity: using (var server = Server.Connect(...)) {     var a = server.CreateApplication("WhiteFish");     var src = a         .DefineObservable<int>(() => Observable.Range(0, 3))         .Deploy("ObservableSource"); If later we want to compose with the source we have to fetch it and then we can bind something to     a.GetObservable<int>("ObservableSource)").Bind(... This means that we had to know a bunch of things about this: that it’s a source, that it’s an observable, it produces a result with payload Int32 and it’s named “ObservableSource”. Only the second and last bits of information are present in the debugger, by the way. As you type in the query window you see that all the entities are present, you get IntelliSense support for them and it’s much easier to make sense of what’s available. Let’s look at a scenario where composition is plausible. With the new programming model it’s possible to create “cold” sources that are parameterized. There was a way to accomplish that even in the previous version by passing parameters to the adapters, but this time it’s much more elegant because the expression declares what parameters are required. Say that we hover the mouse over the ThrottledSource source – we will see that its type is Func<int, int, IQbservable<int>> - this in effect means that we need to pass two int parameters before we can get a source that produces events, and the type for those events is int – in the particular case of my example I had the source produce a range of integers and the two parameters were the start and end of the range. So we see how a developer can create a source that is not running yet. Then someone else (e.g. an administrator) can pass whatever parameters appropriate and run the process. Proxy Types Here’s an interesting scenario – what if someone created a source on a server but they forgot to tell you what type they used. Worse yet, they might have used an anonymous type and even though they can refer to it by name you can’t figure out how to use that type. Let’s walk through an example that shows how you can compose against types you don’t need to have the definition of. This is how we can create a source that returns an anonymous type: Application.DefineObservable(() => Observable.Range(1, 10).Select(i => new { I = i })).Deploy("O1"); Now if we refresh the connection we can see the new source named O1 appear in the list. But what’s more important is that we now have a type to work with. So we can compose a query that refers to the anonymous type. var threshold = new StreamInsightDynamicDriver.TypeProxies.AnonymousType1_0<int>(5); var filter = from i in O1              where i > threshold              select i; filter.Deploy("O2"); You will notice that the anonymous type defined with this statement: new { I = i } can now be manipulated by a client that does not have access to it because the LinqPad driver has generated another type in its stead, named StreamInsightDynamicDriver.TypeProxies.AnonymousType1_0. This type has all the properties and fields of the type defined on the server, except in this case we can instantiate values and use it to compose more queries. It is worth noting that the same thing works for types that are not anonymous – the test is if the LinqPad driver can resolve the type or not. If it’s not possible then a new type will be generated that approximates the type that exists on the server. Control metadata In addition to composing processes on top of the existing entities we can do other useful things. We can delete them – nothing new here as we simply access the entities through the Entities collection of the application class. Here is where having their real name in parentheses comes handy. There’s another way to find out what’s behind a property – dump its expression. The first line in the output tells us what’s the name of the entity used to build this property in the context. Runtime information So let’s create a process to see what happens. We can bind a source to a sink and run the resulting process. If you right click on the connection you can refresh it and see the process present in the list of entities. Then you can drag the process to the query window and see that you can have access to process object in the Processes collection of the application. You can then manipulate the process (delete it, read its diagnostic view etc.). Regards, The StreamInsight Team

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  • Looking into ASP.Net MVC 4.0 Mobile Development - part 1

    - by nikolaosk
    In this post I will be looking how ASP.Net MVC 4.0 helps us to create web solutions that target mobile devices.We all experience the magic that is the World Wide Web through mobile devices. Millions of people around the world, use tablets and smartphones to view the contents of websites,e-shops and portals.ASP.Net MVC 4.0 includes a new mobile project template and the ability to render a different set of views for different types of devices.There is a new feature that is called browser overriding which allows us to control exactly what a user is going to see from your web application regardless of what type of device he is using.In order to follow along this post you must have Visual Studio 2012 and .Net Framework 4.5 installed in your machine.Download and install VS 2012 using this link.My machine runs on Windows 8 and Visual Studio 2012 works just fine.It will work fine in Windows 7 as well so do not worry if you do not have the latest Microsoft operating system.1) Launch VS 2012 and create a new Web Forms application by going to File - >New Project - > ASP.Net MVC 4 Web Application and then click OKHave a look at the picture below  2) From the available templates select Mobile Application and then click OK.Have a look at the picture below 3) When I run the application I get the mobile view of the page. I would like to show you what a typical ASP.Net MVC 4.0 application looks like. So I will create a new simple ASP.Net MVC 4.0 Web Application. When I run the application I get the normal page view.Have a look at the picture below.On the left is the mobile view and on the right the normal view. As you can see we have more or less the same content in our mobile application (log in,register) compared with the normal ASP.Net MVC 4.0 application but it is optimised for mobile devices. 4) Let me explain how and when the mobile view is selected and finally rendered.There is a feature in MVC 4.0 that is called Display Modes and with this feature the runtime will select a view.If we have 2 views e.g contact.mobile.cshtml and contact.cshtml in our application the Controller at some point will instruct the runtime to select and render a view named contact.The runtime will look at the browser making the request and will determine if it is a mobile browser or a desktop browser. So if there is a request from my IPhone Safari browser for a particular site, if there is a mobile view the MVC 4.0 will select it and render it. If there is not a mobile view, the normal view will be rendered.5) In the  ASP.Net MVC 4.0 (Internet application) I created earlier (not the first project which was a mobile one) I can run it once more and see how it looks on the browser. If I want to view it with a mobile browser I must download one emulator like Opera Mobile.You can download Opera Mobile hereWhen I run the application I get the same view in both the desktop and the mobile browser. That was to be expected. Have a look at the picture below 6) Then I create another version of the _Layout.mobile.cshtml view in the Shared folder.I simply copy and paste the _Layout.cshtml  into the same folder and then rename it to _Layout.mobile.cshtml and then just alter the contents of the _Layout.mobile.cshtml.When I run again the application I get a different view on the desktop browser and a different one on the Opera mobile browser.Have a look at the picture below ?he Controller will instruct the ASP.Net runtime to select and render a view named _Layout.mobile.cshtml when the request will come from a mobile browser.?he runtime knows that a browser is a mobile one through the ASP.Net browser capability provider. Hope it helps!!!

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  • JavaScript: this

    - by bdukes
    JavaScript is a language steeped in juxtaposition.  It was made to “look like Java,” yet is dynamic and classless.  From this origin, we get the new operator and the this keyword.  You are probably used to this referring to the current instance of a class, so what could it mean in a language without classes? In JavaScript, this refers to the object off of which a function is referenced when it is invoked (unless it is invoked via call or apply). What this means is that this is not bound to your function, and can change depending on how your function is invoked. It also means that this changes when declaring a function inside another function (i.e. each function has its own this), such as when writing a callback. Let's see some of this in action: var obj = { count: 0, increment: function () { this.count += 1; }, logAfterTimeout = function () { setTimeout(function () { console.log(this.count); }, 1); } }; obj.increment(); console.log(obj.count); // 1 var increment = obj.increment; window.count = 'global count value: '; increment(); console.log(obj.count); // 1 console.log(window.count); // global count value: 1 var newObj = {count:50}; increment.call(newObj); console.log(newObj.count); // 51 obj.logAfterTimeout();// global count value: 1 obj.logAfterTimeout = function () { var proxiedFunction = $.proxy(function () { console.log(this.count); }, this); setTimeout(proxiedFunction, 1); }; obj.logAfterTimeout(); // 1 obj.logAfterTimeout = function () { var that = this; setTimeout(function () { console.log(that.count); }, 1); }; obj.logAfterTimeout(); // 1 The last couple of examples here demonstrate some methods for making sure you get the values you expect.  The first time logAfterTimeout is redefined, we use jQuery.proxy to create a new function which has its this permanently set to the passed in value (in this case, the current this).  The second time logAfterTimeout is redefined, we save the value of this in a variable (named that in this case, also often named self) and use the new variable in place of this. Now, all of this is to clarify what’s going on when you use this.  However, it’s pretty easy to avoid using this altogether in your code (especially in the way I’ve demonstrated above).  Instead of using this.count all over the place, it would have been much easier if I’d made count a variable instead of a property, and then I wouldn’t have to use this to refer to it.  var obj = (function () { var count = 0; return { increment: function () { count += 1; }, logAfterTimeout = function () { setTimeout(function () { console.log(count); }, 1); }, getCount: function () { return count; } }; }()); If you’re writing your code in this way, the main place you’ll run into issues with this is when handling DOM events (where this is the element on which the event occurred).  In that case, just be careful when using a callback within that event handler, that you’re not expecting this to still refer to the element (and use proxy or that/self if you need to refer to it). Finally, as demonstrated in the example, you can use call or apply on a function to set its this value.  This isn’t often needed, but you may also want to know that you can use apply to pass in an array of arguments to a function (e.g. console.log.apply(console, [1, 2, 3, 4])).

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  • Ubuntu 10.04 & IBM DS3524 with FC multipath, inactive path is [failed][faulty] instead of [active][ghost]

    - by Graeme Donaldson
    OK, this is my setup: FC Switches IBM/Brocade, Switch1 and Switch2, independent fabrics. Server IBM x3650 M2, 2x QLogic QLE2460, 1 connected to each FC Switch. Storage IBM DS3524, 2x controllers with 4x FC ports each, but only 2x connected on each. +-----------------------------------------------------------------------+ | HBA1 Server HBA2 | +-----------------------------------------------------------------------+ | | | | | | +-----------------------------+ +------------------------------+ | Switch1 | | Switch2 | +-----------------------------+ +------------------------------+ | | | | | | | | | | | | | | | | | | | | +-----------------------------------+-----------------------------------+ | Contr A, port 3 | Contr A, port 4 | Contr B, port 3 | Contr B, port 4 | +-----------------------------------+-----------------------------------+ | Storage | +-----------------------------------------------------------------------+ My /etc/multipath.conf is from the IBM redbook for the DS3500, except I use a different setting for prio_callout, IBM uses /sbin/mpath_prio_tpc, but according to http://changelogs.ubuntu.com/changelogs/pool/main/m/multipath-tools/multipath-tools_0.4.8-7ubuntu2/changelog, this was renamed to /sbin/mpath_prio_rdac, which I'm using. devices { device { #ds3500 vendor "IBM" product "1746 FAStT" hardware_handler "1 rdac" path_checker rdac failback 0 path_grouping_policy multibus prio_callout "/sbin/mpath_prio_rdac /dev/%n" } } multipaths { multipath { wwid xxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxx alias array07 path_grouping_policy multibus path_checker readsector0 path_selector "round-robin 0" failback "5" rr_weight priorities no_path_retry "5" } } The output of multipath -ll with controller A as the preferred path: root@db06:~# multipath -ll sdg: checker msg is "directio checker reports path is down" sdh: checker msg is "directio checker reports path is down" array07 (xxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxx) dm-2 IBM ,1746 FASt [size=4.9T][features=1 queue_if_no_path][hwhandler=0] \_ round-robin 0 [prio=2][active] \_ 5:0:1:0 sdd 8:48 [active][ready] \_ 5:0:2:0 sde 8:64 [active][ready] \_ 6:0:1:0 sdg 8:96 [failed][faulty] \_ 6:0:2:0 sdh 8:112 [failed][faulty] If I change the preferred path using IBM DS Storage Manager to Controller B, the output swaps accordingly: root@db06:~# multipath -ll sdd: checker msg is "directio checker reports path is down" sde: checker msg is "directio checker reports path is down" array07 (xxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxx) dm-2 IBM ,1746 FASt [size=4.9T][features=1 queue_if_no_path][hwhandler=0] \_ round-robin 0 [prio=2][active] \_ 5:0:1:0 sdd 8:48 [failed][faulty] \_ 5:0:2:0 sde 8:64 [failed][faulty] \_ 6:0:1:0 sdg 8:96 [active][ready] \_ 6:0:2:0 sdh 8:112 [active][ready] According to IBM, the inactive path should be "[active][ghost]", not "[failed][faulty]". Despite this, I don't seem to have any I/O issues, but my syslog is being spammed with this every 5 seconds: Jun 1 15:30:09 db06 multipathd: sdg: directio checker reports path is down Jun 1 15:30:09 db06 kernel: [ 2350.282065] sd 6:0:2:0: [sdh] Result: hostbyte=DID_OK driverbyte=DRIVER_SENSE Jun 1 15:30:09 db06 kernel: [ 2350.282071] sd 6:0:2:0: [sdh] Sense Key : Illegal Request [current] Jun 1 15:30:09 db06 kernel: [ 2350.282076] sd 6:0:2:0: [sdh] <<vendor>> ASC=0x94 ASCQ=0x1ASC=0x94 ASCQ=0x1 Jun 1 15:30:09 db06 kernel: [ 2350.282083] sd 6:0:2:0: [sdh] CDB: Read(10): 28 00 00 00 00 00 00 00 08 00 Jun 1 15:30:09 db06 kernel: [ 2350.282092] end_request: I/O error, dev sdh, sector 0 Jun 1 15:30:10 db06 multipathd: sdh: directio checker reports path is down Jun 1 15:30:14 db06 kernel: [ 2355.312270] sd 6:0:1:0: [sdg] Result: hostbyte=DID_OK driverbyte=DRIVER_SENSE Jun 1 15:30:14 db06 kernel: [ 2355.312277] sd 6:0:1:0: [sdg] Sense Key : Illegal Request [current] Jun 1 15:30:14 db06 kernel: [ 2355.312282] sd 6:0:1:0: [sdg] <<vendor>> ASC=0x94 ASCQ=0x1ASC=0x94 ASCQ=0x1 Jun 1 15:30:14 db06 kernel: [ 2355.312290] sd 6:0:1:0: [sdg] CDB: Read(10): 28 00 00 00 00 00 00 00 08 00 Jun 1 15:30:14 db06 kernel: [ 2355.312299] end_request: I/O error, dev sdg, sector 0 Does anyone know how I can get the inactive path to show "[active][ghost]" instead of "[failed][faulty]"? I assume that once I can get that right then the spam in my syslog will end as well. One final thing worth mentioning is that the IBM redbook doc targets SLES 11 so I'm assuming there's something a little different under Ubuntu that I just haven't figured out yet. Update: As suggested by Mitch, I've tried removing /etc/multipath.conf, and now the output of multipath -ll looks like this: root@db06:~# multipath -ll sdg: checker msg is "directio checker reports path is down" sdh: checker msg is "directio checker reports path is down" xxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxdm-1 IBM ,1746 FASt [size=4.9T][features=0][hwhandler=0] \_ round-robin 0 [prio=1][active] \_ 5:0:2:0 sde 8:64 [active][ready] \_ round-robin 0 [prio=1][enabled] \_ 5:0:1:0 sdd 8:48 [active][ready] \_ round-robin 0 [prio=0][enabled] \_ 6:0:1:0 sdg 8:96 [failed][faulty] \_ round-robin 0 [prio=0][enabled] \_ 6:0:2:0 sdh 8:112 [failed][faulty] So its more or less the same, with the same message in the syslog every 5 minutes as before, but the grouping has changed.

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