virtual functions - Page 227

Windows 2008 Unknown Disks

- by Ailbe

I have a BL460c G7 blade server with OS Windows 2008 R2 SP1. This is a brand new C7000 enclosure, with FlexFabric interconnects. I got my FC switches setup and zoned properly to our Clariion CX4, and can see all the hosts that are assigned FCoE HBAs on both paths in both Navisphere and in HP Virtual Connect Manager. So I went ahead and created a storage group for a test server, assigned the appropriate host, assigned the LUN to the server. So far so good, log onto server and I can see 4 unknown disks.... No problem, I install MS MPIO, no luck, can't initialize the disks, and the multiple disks don't go away. Still no problem, I install PowerPath version 5.5 reboot. Now I see 3 disks. One is initialized and ready to go, but I still have 2 disks that I can't initialize, can't offline, can't delete. If I right click in storage manager and go to properties I can see that the MS MPIO tab, but I can't make a path active. I want to get rid of these phantom disks, but so far nothing is working and google searches are showing up some odd results, so obviously I'm not framing my question right. I thought I'd ask here real quick. Does anyone know a quick way to get rid of these unknown disks. Another question, do I need the MPIO feature installed if I have PowerPath installed? This is my first time installing Windows 2008 R2 in this fashion and I'm not sure if that feature is needed or not right now. So some more information to add to this. It seems I'm dealing with more of a Windows issue than anything else. I removed the LUN from the server, uninstalled PowerPath completely, removed the MPIO feature from the server, and rebooted twice. Now I am back to the original 4 Unknown Disks (plus the local Disk 0 containing the OS partition of course, which is working fine) I went to diskpart, I could see all 4 Unknown disks, I selected each disk, ran clean (just in case i'd somehow brought them online previously as GPT and didn't realize it) After a few minutes I was no longer able to see the disks when I ran list disk. However, the disks are still in Disk Management. When I try and offline the disks from Disk Management I get an error: Virtual Disk Manager - The system cannot find the file specified. Accompanied by an error in System Event Logs: Log Name: System Source: Virtual Disk Service Date: 6/25/2012 4:02:01 PM Event ID: 1 Task Category: None Level: Error Keywords: Classic User: N/A Computer: hostname.local Description: Unexpected failure. Error code: 2@02000018 Event Xml: 1 2 0 0x80000000000000 4239 System hostname.local 2@02000018 I feel sure there is a place I can go in the Registry to get rid of these, I just can't recall where and I am loathe to experiement. So to recap, there are currently no LUNS attached at all, I still have the phantom disks, and I'm getting The system cannot find the file specified from Virtual Disk Manager when I try to take them offline. Thanks!

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Linux not buffering block I/O when the device is not "in use" (i.e. mounted)

- by Radek Hladík

I am installing new server and I've found an interesting issue. The server is running Fedora 19 (3.11.7-200.fc19.x86_64 kernel) and is supposed to host a few KVM/Qemu virtual servers (mail server, file server, etc..). The HW is Intel(R) Xeon(R) CPU 5160 @ 3.00GHz with 16GB RAM. One of the most important features will be Samba server and we have decided to make it as virtual machine with almost direct access to the disks. So the real HDD is cached on SSD (via bcache) then raided with md and the final device is exported into the virtual machine via virtio. The virtual machine is again Fedora 19 with the same kernel. One important topic to find out is whether the virtualization layer will not introduce high overload into disk I/Os. So far I've been able to get up to 180MB/s in VM and up to 220MB/s on real HW (on the SSD disk). I am still not sure why the overhead is so big but it is more than the network can handle so I do not care so much. The interesting thing is that I've found that the disk reads are not buffered in the VM unless I create and mount FS on the disk or I use the disks somehow. Simply put: Lets do dd to read disk for the first time (the /dev/vdd is an old Raptor disk 70MB/s is its real speed): [root@localhost ~]# dd if=/dev/vdd of=/dev/null bs=256k count=10000 ; cat /proc/meminfo | grep Buffers 2621440000 bytes (2.6 GB) copied, 36.8038 s, 71.2 MB/s Buffers: 14444 kB Rereading the data shows that they are cached somewhere but not in buffers of the VM. Also the speed increased to "only" 500MB/s. The VM has 4GB of RAM (more that the test file) [root@localhost ~]# dd if=/dev/vdd of=/dev/null bs=256k count=10000 ; cat /proc/meminfo | grep Buffers 2621440000 bytes (2.6 GB) copied, 5.16016 s, 508 MB/s Buffers: 14444 kB [root@localhost ~]# dd if=/dev/vdd of=/dev/null bs=256k count=10000 ; cat /proc/meminfo | grep Buffers 2621440000 bytes (2.6 GB) copied, 5.05727 s, 518 MB/s Buffers: 14444 kB Now lets mount the FS on /dev/vdd and try the dd again: [root@localhost ~]# mount /dev/vdd /mnt/tmp [root@localhost ~]# dd if=/dev/vdd of=/dev/null bs=256k count=10000 ; cat /proc/meminfo | grep Buffers 2621440000 bytes (2.6 GB) copied, 4.68578 s, 559 MB/s Buffers: 2574592 kB [root@localhost ~]# dd if=/dev/vdd of=/dev/null bs=256k count=10000 ; cat /proc/meminfo | grep Buffers 2621440000 bytes (2.6 GB) copied, 1.50504 s, 1.7 GB/s Buffers: 2574592 kB While the first read was the same, all 2.6GB got buffered and the next read was at 1.7GB/s. And when I unmount the device: [root@localhost ~]# umount /mnt/tmp [root@localhost ~]# cat /proc/meminfo | grep Buffers Buffers: 14452 kB [root@localhost ~]# dd if=/dev/vdd of=/dev/null bs=256k count=10000 ; cat /proc/meminfo | grep Buffers 2621440000 bytes (2.6 GB) copied, 5.10499 s, 514 MB/s Buffers: 14468 kB The bcache was disabled while testing and the results are same on faster (newer) HDDs and on SSD (except for the initial read speed of course). To sum it up. When I read from the device via dd first time, it gets read from the disk. Next time I reread it gets cached in the host but not in the guest (thats actually the same issue, more on that later). When I mount the filesystem but try to read the device directly it gets cached in VM (via buffers). As soon as I stop "using" it, buffers are discarded and the device is not cached anymore in the VM. When I looked into buffers value on the host I realized that the situation is the same. The block I/O gets buffered only when the disk is in use, in this case it means "exported to a VM". On host, after all the measurement done: 3165552 buffers On the host, after the VM shutdown: 119176 buffers I know it is not important as the disks will be mounted all the time but I am curious and I would like to know why it is working like this.

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Postfix MySql Dovecot - SMTP Authentication Failure

- by borncamp

Hello I have a Postfix setup with Dovecot and MySql. The server is running Debian Squeeze. The MySql server is a slave that has data pushed to it from a primary (postfix) mail server(running a different os). The emails are stored on a replicated GlusterFS volume. I am able to check email using thunderbird over IMAP. However, SMTP requests fail. After turning on query logs for the MySql server I have noticed that no query statement is executed to retrieve the user information when an SMTP client tries to authenticate. I'd like to know what I'm doing wrong or what the next troubleshooting steps are. I'm about to pull my hair out. Below is some log and configuration data that I thought would be relevant. You're help is much obliged. The file /var/log/mail.log shows Oct 11 14:54:16 mailbox2 postfix/smtpd[25017]: connect from unknown[192.168.0.44] Oct 11 14:54:19 mailbox2 postfix/smtpd[25017]: warning: unknown[192.168.0.44]: SASL PLAIN authentication failed: Oct 11 14:54:25 mailbox2 postfix/smtpd[25017]: warning: unknown[192.168.0.44]: SASL LOGIN authentication failed: VXNlcm5hbWU6 Oct 11 14:55:48 mailbox2 postfix/smtpd[25017]: warning: unknown[192.168.0.44]: SASL PLAIN authentication failed: VXNlcm5hbWU6 Oct 11 14:55:54 mailbox2 postfix/smtpd[25017]: warning: unknown[192.168.0.44]: SASL LOGIN authentication failed: VXNlcm5hbWU6 Oct 11 14:55:57 mailbox2 postfix/smtpd[25017]: disconnect from unknown[192.168.0.44] This is my dovecot.conf file log_timestamp = "%Y-%m-%d %H:%M:%S " mail_location = maildir:/var/mail/virtual/%d/%n/ auth_mechanisms = plain login disable_plaintext_auth = no namespace { inbox = yes location = prefix = INBOX. separator = . type = private } passdb { args = /etc/dovecot/dovecot-mysql.conf driver = sql } protocols = imap pop3 service auth { unix_listener /var/spool/postfix/private/auth { group = postfix mode = 0660 user = postfix } unix_listener auth-master { mode = 0600 user = postfix } user = root } ssl_cert = </etc/ssl/certs/dovecot.pem ssl_key = </etc/ssl/private/dovecot.pem userdb { args = /etc/dovecot/dovecot-mysql.conf driver = sql } protocol lda { auth_socket_path = /var/run/dovecot/auth-master mail_plugins = sieve postmaster_address = [email protected] } protocol pop3 { pop3_uidl_format = %08Xu%08Xv } Here is my dovecot-mysql.conf file: connect = host=127.0.0.1 dbname=postfix user=postfix password=ffjM2MYAqQtAzRHX driver = mysql default_pass_scheme = MD5-CRYPT password_query = SELECT username AS user,password FROM mailbox WHERE username = '%u' AND active='1' user_query = SELECT CONCAT('/var/mail/virtual/', maildir) AS home, 1001 AS uid, 109 AS gid, CONCAT('*:messages=10000:bytes=',quota) as quota_rule, 'Trash:ignore' AS quota_rule2 FROM mailbox WHERE username = '%u' AND active='1' Here is my output from 'postconf -n': append_dot_mydomain = no biff = no bounce_template_file = /etc/postfix/bounce.cf broken_sasl_auth_clients = yes config_directory = /etc/postfix delay_warning_time = 0h dovecot_destination_recipient_limit = 1 inet_interfaces = all local_recipient_maps = $virtual_mailbox_maps local_transport = virtual mailbox_command = procmail -a "$EXTENSION" mailbox_size_limit = 0 maximal_queue_lifetime = 1d message_size_limit = 25600000 mydestination = mailbox2.cws.net, debian.local.cws.net, localhost.local.cws.net, localhost myhostname = mailbox2.cws.net mynetworks = 127.0.0.0/8 [::ffff:127.0.0.0]/104 [::1]/128 172.18.0.119 63.164.138.3 myorigin = /etc/mailname 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 $virtual_mailbox_limit_maps readme_directory = no recipient_delimiter = + relay_domains = relayhost = smtp_connect_timeout = 10 smtp_tls_session_cache_database = btree:${data_directory}/smtp_scache smtpd_banner = $myhostname ESMTP $mail_name (Debian/GNU) smtpd_client_message_rate_limit = 50 smtpd_client_recipient_rate_limit = 500 smtpd_client_restrictions = permit_sasl_authenticated, permit_mynetworks smtpd_delay_reject = yes smtpd_discard_ehlo_keyword_address_maps = hash:/etc/postfix/discard_ehlo smtpd_helo_required = yes smtpd_helo_restrictions = permit_mynetworks, reject_invalid_helo_hostname, permit smtpd_recipient_restrictions = permit_mynetworks,permit_sasl_authenticated,reject_unauth_destination smtpd_sasl_auth_enable = yes smtpd_sasl_authenticated_header = yes smtpd_sasl_path = private/auth smtpd_sasl_security_options = noanonymous smtpd_sasl_tls_security_options = $smtpd_sasl_security_options smtpd_sasl_type = dovecot smtpd_sender_restrictions = permit_mynetworks, reject_non_fqdn_sender, reject_unknown_sender_domain, permit smtpd_tls_cert_file = /etc/ssl/certs/ssl-cert-snakeoil.pem smtpd_tls_key_file = /etc/ssl/private/ssl-cert-snakeoil.key smtpd_tls_session_cache_database = btree:${data_directory}/smtpd_scache smtpd_use_tls = yes transport_maps = hash:/etc/postfix/transport virtual_alias_maps = proxy:mysql:/etc/postfix/sql/mysql_virtual_alias_maps.cf, proxy:mysql:/etc/postfix/sql/mysql_virtual_alias_domain_maps.cf, proxy:mysql:/etc/postfix/sql/mysql_virtual_alias_domain_catchall_maps.cf virtual_gid_maps = static:1001 virtual_mailbox_base = /var/mail/virtual/ virtual_mailbox_domains = proxy:mysql:/etc/postfix/sql/mysql_virtual_domains_maps.cf virtual_mailbox_maps = proxy:mysql:/etc/postfix/sql/mysql_virtual_mailbox_maps.cf, proxy:mysql:/etc/postfix/sql/mysql_virtual_alias_domain_mailbox_maps.cf virtual_transport = dovecot virtual_uid_maps = static:1001

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Problem installing build-essential and upgrading g++ on Ubuntu 8.04

- by ehsanul

I'm having some trouble with dependencies it seems, but myself don't really know how to resolve the issue. Here's the output: ~:) sudo apt-get install build-essential Reading package lists... Done Building dependency tree Reading state information... Done Some packages could not be installed. This may mean that you have requested an impossible situation or if you are using the unstable distribution that some required packages have not yet been created or been moved out of Incoming. Since you only requested a single operation it is extremely likely that the package is simply not installable and a bug report against that package should be filed. The following information may help to resolve the situation: The following packages have unmet dependencies: build-essential: Depends: g++ (>= 4:4.3.1) but 4:4.2.3-1ubuntu6 is to be installed E: Broken packages ~:) sudo apt-get install g++ Reading package lists... Done Building dependency tree Reading state information... Done Some packages could not be installed. This may mean that you have requested an impossible situation or if you are using the unstable distribution that some required packages have not yet been created or been moved out of Incoming. Since you only requested a single operation it is extremely likely that the package is simply not installable and a bug report against that package should be filed. The following information may help to resolve the situation: The following packages have unmet dependencies: g++: Depends: cpp (>= 4:4.3.1-1ubuntu2) but 4:4.2.3-1ubuntu6 is to be installed Depends: gcc (>= 4:4.3.1-1ubuntu2) but 4:4.2.3-1ubuntu6 is to be installed Depends: g++-4.3 (>= 4.3.1-1) but it is not going to be installed Depends: gcc-4.3 (>= 4.3.1-1) but it is not installable E: Broken packages ~:) Edit: I just tried aptitude instead of apt-get, as suggested. Doesn't work, had other problems: ~:) sudo aptitude install build-essential [sudo] password for ehsanul: Reading package lists... Done Building dependency tree Reading state information... Done Reading extended state information Initializing package states... Done Building tag database... Done The following packages are BROKEN: g++ g++-4.3 libstdc++6-4.3-dev The following packages have been automatically kept back: dpkg-dev fakeroot libdns35 libisc35 linux-libc-dev patch The following NEW packages will be automatically installed: libgmp3c2 libmpfr1ldbl The following packages have been kept back: adobe-flashplugin bind9-host dnsutils gvfs gvfs-backends gvfs-fuse libatm1 libbind9-30 libgvfscommon0 libisccc30 libisccfg30 liblwres30 libnautilus-extension1 linux-headers-2.6.24-24 linux-headers-2.6.24-24-generic linux-image-2.6.24-24-generic nautilus nautilus-data The following NEW packages will be installed: libgmp3c2 libmpfr1ldbl The following packages will be upgraded: build-essential The following partially installed packages will be configured: timidity 2 packages upgraded, 4 newly installed, 0 to remove and 24 not upgraded. Need to get 775kB/6265kB of archives. After unpacking 20.3MB will be used. The following packages have unmet dependencies: libstdc++6-4.3-dev: Depends: gcc-4.3-base (= 4.3.2-1ubuntu11) which is a virtual package. Depends: libstdc++6 (>= 4.3.2-1ubuntu11) but 4.2.4-1ubuntu4 is installed. g++-4.3: Depends: gcc-4.3-base (= 4.3.2-1ubuntu11) which is a virtual package. Depends: gcc-4.3 (= 4.3.2-1ubuntu11) which is a virtual package. Depends: libc6 (>= 2.8~20080505) but 2.7-10ubuntu4 is installed. g++: Depends: cpp (>= 4:4.3.1-1ubuntu2) but 4:4.2.3-1ubuntu6 is installed. Depends: gcc (>= 4:4.3.1-1ubuntu2) but 4:4.2.3-1ubuntu6 is installed. Depends: gcc-4.3 (>= 4.3.1-1) which is a virtual package. Resolving dependencies... The following actions will resolve these dependencies: Keep the following packages at their current version: build-essential [11.3ubuntu1 (hardy, now)] g++ [4:4.2.3-1ubuntu6 (hardy-updates, now)] g++-4.3 [Not Installed] libstdc++6-4.3-dev [Not Installed] Score is -9852 Accept this solution? [Y/n/q/?]

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Issue in nginx proxying to apache

- by Luis Masuelli

My current nginx configuration is as follows: specific configuration for (currently two) domains: server { listen 443 ssl; server_name studiotv.service.tebusco.lan phpmyadmin.service.tebusco.lan; ssl_certificate /home/administrador/nginx-confs/ssl/service.tebusco.lan.crt; ssl_certificate_key /home/administrador/nginx-confs/ssl/service.tebusco.lan.key; location / { proxy_pass http://127.0.0.1:8180; proxy_set_header Host $http_host:8180; } } default configuration for unmatched ssl connections: server { listen 443 default ssl; ssl_certificate /home/administrador/nginx-confs/ssl/service.tebusco.lan.crt; ssl_certificate_key /home/administrador/nginx-confs/ssl/service.tebusco.lan.key; location / { return 403; } } http configuration: server { listen 80; rewrite ^ https://$host$request_uri? permanent; } The intention is clear: Redirect http traffic to https. Proxy each https:// call from phpmyadmin.service.tebusco.lan and studiotv.service.tebusco.lan to apache2. This includes passing a host header, which is detected. Each unmatched ssl connection must return a 403 in nginx. Does not even reach apache2. In the apache2 side of the life, I have a default site, and a non-default site which will match studiotv.service.tebusco.lan: 000-default.conf file (available and enabled): <VirtualHost 127.0.0.1:8180> # The ServerName directive sets the request scheme, hostname and port that # the server uses to identify itself. This is used when creating # redirection URLs. In the context of virtual hosts, the ServerName # specifies what hostname must appear in the request's Host: header to # match this virtual host. For the default virtual host (this file) this # value is not decisive as it is used as a last resort host regardless. # However, you must set it for any further virtual host explicitly. ServerName localhost ServerAdmin webmaster@localhost DocumentRoot /var/www/html <Directory /var/www/html> Order deny,allow Require all granted </Directory> </VirtualHost> # vim: syntax=apache ts=4 sw=4 sts=4 sr noet studiotv.conf file (available and enabled): <VirtualHost *:8180> ServerName studiotv.service.tebusco.lan ServerAdmin [email protected] DocumentRoot /var/www/studiotv <Directory /var/www/studiotv/> Options -Indexes +FollowSymLinks AllowOverride None Order deny,allow Allow from all Require all granted </Directory> # Available loglevels: trace8, ..., trace1, debug, info, notice, warn, # error, crit, alert, emerg. # It is also possible to configure the loglevel for particular # modules, e.g. #LogLevel info ssl:warn # No usamos ${APACHE_LOG_DIR} sino en su lugar /var/log/<host> ErrorLog /var/log/apache2/studiotv/error.log CustomLog /var/log/apache2/studiotv/access.log combined </VirtualHost> # vim: syntax=apache ts=4 sw=4 sts=4 sr noet However, when I hit the browser with http://studiotv.service.tebusco.lan, the default php page is shown instead. Question: What am I missing? (apache 2.4.7, nginx 1.6.0, ubuntu server 14.04).

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Oracle Solaris: Zones on Shared Storage

- by Jeff Victor

Oracle Solaris 11.1 has several new features. At oracle.com you can find a detailed list. One of the significant new features, and the most significant new feature releated to Oracle Solaris Zones, is casually called "Zones on Shared Storage" or simply ZOSS (rhymes with "moss"). ZOSS offers much more flexibility because you can store Solaris Zones on shared storage (surprise!) so that you can perform quick and easy migration of a zone from one system to another. This blog entry describes and demonstrates the use of ZOSS. ZOSS provides complete support for a Solaris Zone that is stored on "shared storage." In this case, "shared storage" refers to fiber channel (FC) or iSCSI devices, although there is one lone exception that I will demonstrate soon. The primary intent is to enable you to store a zone on FC or iSCSI storage so that it can be migrated from one host computer to another much more easily and safely than in the past. With this blog entry, I wanted to make it easy for you to try this yourself. I couldn't assume that you have a SAN available - which is a good thing, because neither do I! What could I use, instead? [There he goes, foreshadowing again... -Ed.] Developing this entry reinforced the lesson that the solution to every lab problem is VirtualBox. Oracle VM VirtualBox (its formal name) helps here in a couple of important ways. It offers the ability to easily install multiple copies of Solaris as guests on top of any popular system (Microsoft Windows, MacOS, Solaris, Oracle Linux (and other Linuxes) etc.). It also offers the ability to create a separate virtual disk drive (VDI) that appears as a local hard disk to a guest. This virtual disk can be moved very easily from one guest to another. In other words, you can follow the steps below on a laptop or larger x86 system. Please note that the ability to use ZOSS to store a zone on a local disk is very useful for a lab environment, but not so useful for production. I do not suggest regularly moving disk drives among computers. In the method I describe below, that virtual hard disk will contain the zone that will be migrated among the (virtual) hosts. In production, you would use FC or iSCSI LUNs instead. The zonecfg(1M) man page details the syntax for each of the three types of devices. Why Migrate? Why is the migration of virtual servers important? Some of the most common reasons are: Moving a workload to a different computer so that the original computer can be turned off for extensive maintenance. Moving a workload to a larger system because the workload has outgrown its original system. If the workload runs in an environment (such as a Solaris Zone) that is stored on shared storage, you can restore the service of the workload on an alternate computer if the original computer has failed and will not reboot. You can simplify lifecycle management of a workload by developing it on a laptop, migrating it to a test platform when it's ready, and finally moving it to a production system. Concepts For ZOSS, the important new concept is named "rootzpool". You can read about it in the zonecfg(1M) man page, but here's the short version: it's the backing store (hard disk(s), or LUN(s)) that will be used to make a ZFS zpool - the zpool that will hold the zone. This zpool: contains the zone's Solaris content, i.e. the root file system does not contain any content not related to the zone can only be mounted by one Solaris instance at a time Method Overview Here is a brief list of the steps to create a zone on shared storage and migrate it. The next section shows the commands and output. You will need a host system with an x86 CPU (hopefully at least a couple of CPU cores), at least 2GB of RAM, and at least 25GB of free disk space. (The steps below will not actually use 25GB of disk space, but I don't want to lead you down a path that ends in a big sign that says "Your HDD is full. Good luck!") Configure the zone on both systems, specifying the rootzpool that both will use. The best way is to configure it on one system and then copy the output of "zonecfg export" to the other system to be used as input to zonecfg. This method reduces the chances of pilot error. (It is not necessary to configure the zone on both systems before creating it. You can configure this zone in multiple places, whenever you want, and migrate it to one of those places at any time - as long as those systems all have access to the shared storage.) Install the zone on one system, onto shared storage. Boot the zone. Provide system configuration information to the zone. (In the Real World(tm) you will usually automate this step.) Shutdown the zone. Detach the zone from the original system. Attach the zone to its new "home" system. Boot the zone. The zone can be used normally, and even migrated back, or to a different system. Details The rest of this shows the commands and output. The two hostnames are "sysA" and "sysB". Note that each Solaris guest might use a different device name for the VDI that they share. I used the device names shown below, but you must discover the device name(s) after booting each guest. In a production environment you would also discover the device name first and then configure the zone with that name. Fortunately, you can use the command "zpool import" or "format" to discover the device on the "new" host for the zone. The first steps create the VirtualBox guests and the shared disk drive. I describe the steps here without demonstrating them. Download VirtualBox and install it using a method normal for your host OS. You can read the complete instructions. Create two VirtualBox guests, each to run Solaris 11.1. Each will use its own VDI as its root disk. Install Solaris 11.1 in each guest.Install Solaris 11.1 in each guest. To install a Solaris 11.1 guest, you can either download a pre-built VirtualBox guest, and import it, or install Solaris 11.1 from the "text install" media. If you use the latter method, after booting you will not see a windowing system. To install the GUI and other important things, login and run "pkg install solaris-desktop" and take a break while it installs those important things. Life is usually easier if you install the VirtualBox Guest Additions because then you can copy and paste between the host and guests, etc. You can find the guest additions in the folder matching the version of VirtualBox you are using. You can also read the instructions for installing the guest additions. To create the zone's shared VDI in VirtualBox, you can open the storage configuration for one of the two guests, select the SATA controller, and click on the "Add Hard Disk" icon nearby. Choose "Create New Disk" and specify an appropriate path name for the file that will contain the VDI. The shared VDI must be at least 1.5 GB. Note that the guest must be stopped to do this. Add that VDI to the other guest - using its Storage configuration - so that each can access it while running. The steps start out the same, except that you choose "Choose Existing Disk" instead of "Create New Disk." Because the disk is configured on both of them, VirtualBox prevents you from running both guests at the same time. Identify device names of that VDI, in each of the guests. Solaris chooses the name based on existing devices. The names may be the same, or may be different from each other. This step is shown below as "Step 1." Assumptions In the example shown below, I make these assumptions. The guest that will own the zone at the beginning is named sysA. The guest that will own the zone after the first migration is named sysB. On sysA, the shared disk is named /dev/dsk/c7t2d0 On sysB, the shared disk is named /dev/dsk/c7t3d0 (Finally!) The Steps Step 1) Determine the name of the disk that will move back and forth between the systems. root@sysA:~# format Searching for disks...done AVAILABLE DISK SELECTIONS: 0. c7t0d0 /pci@0,0/pci8086,2829@d/disk@0,0 1. c7t2d0 /pci@0,0/pci8086,2829@d/disk@2,0 Specify disk (enter its number): ^D Step 2) The first thing to do is partition and label the disk. The magic needed to write an EFI label is not overly complicated. root@sysA:~# format -e c7t2d0 selecting c7t2d0 [disk formatted] FORMAT MENU: ... format fdisk No fdisk table exists. The default partition for the disk is: a 100% "SOLARIS System" partition Type "y" to accept the default partition, otherwise type "n" to edit the partition table. n SELECT ONE OF THE FOLLOWING: ... Enter Selection: 1 ... G=EFI_SYS 0=Exit? f SELECT ONE... ... 6 format label ... Specify Label type[1]: 1 Ready to label disk, continue? y format quit root@sysA:~# ls /dev/dsk/c7t2d0 /dev/dsk/c7t2d0 Step 3) Configure zone1 on sysA. root@sysA:~# zonecfg -z zone1 Use 'create' to begin configuring a new zone. zonecfg:zone1 create create: Using system default template 'SYSdefault' zonecfg:zone1 set zonename=zone1 zonecfg:zone1 set zonepath=/zones/zone1 zonecfg:zone1 add rootzpool zonecfg:zone1:rootzpool add storage dev:dsk/c7t2d0 zonecfg:zone1:rootzpool end zonecfg:zone1 exit root@sysA:~# oot@sysA:~# zonecfg -z zone1 info zonename: zone1 zonepath: /zones/zone1 brand: solaris autoboot: false bootargs: file-mac-profile: pool: limitpriv: scheduling-class: ip-type: exclusive hostid: fs-allowed: anet: ... rootzpool: storage: dev:dsk/c7t2d0 Step 4) Install the zone. This step takes the most time, but you can wander off for a snack or a few laps around the gym - or both! (Just not at the same time...) root@sysA:~# zoneadm -z zone1 install Created zone zpool: zone1_rpool Progress being logged to /var/log/zones/zoneadm.20121022T163634Z.zone1.install Image: Preparing at /zones/zone1/root. AI Manifest: /tmp/manifest.xml.RXaycg SC Profile: /usr/share/auto_install/sc_profiles/enable_sci.xml Zonename: zone1 Installation: Starting ... Creating IPS image Startup linked: 1/1 done Installing packages from: solaris origin: http://pkg.us.oracle.com/support/ DOWNLOAD PKGS FILES XFER (MB) SPEED Completed 183/183 33556/33556 222.2/222.2 2.8M/s PHASE ITEMS Installing new actions 46825/46825 Updating package state database Done Updating image state Done Creating fast lookup database Done Installation: Succeeded Note: Man pages can be obtained by installing pkg:/system/manual done. Done: Installation completed in 1696.847 seconds. Next Steps: Boot the zone, then log into the zone console (zlogin -C) to complete the configuration process. Log saved in non-global zone as /zones/zone1/root/var/log/zones/zoneadm.20121022T163634Z.zone1.install Step 5) Boot the Zone. root@sysA:~# zoneadm -z zone1 boot Step 6) Login to zone's console to complete the specification of system information. root@sysA:~# zlogin -C zone1 Answer the usual questions and wait for a login prompt. Then you can end the console session with the usual "~." incantation. Step 7) Shutdown the zone so it can be "moved." root@sysA:~# zoneadm -z zone1 shutdown Step 8) Detach the zone so that the original global zone can't use it. root@sysA:~# zoneadm list -cv ID NAME STATUS PATH BRAND IP 0 global running / solaris shared - zone1 installed /zones/zone1 solaris excl root@sysA:~# zpool list NAME SIZE ALLOC FREE CAP DEDUP HEALTH ALTROOT rpool 17.6G 11.2G 6.47G 63% 1.00x ONLINE - zone1_rpool 1.98G 484M 1.51G 23% 1.00x ONLINE - root@sysA:~# zoneadm -z zone1 detach Exported zone zpool: zone1_rpool Step 9) Review the result and shutdown sysA so that sysB can use the shared disk. root@sysA:~# zpool list NAME SIZE ALLOC FREE CAP DEDUP HEALTH ALTROOT rpool 17.6G 11.2G 6.47G 63% 1.00x ONLINE - root@sysA:~# zoneadm list -cv ID NAME STATUS PATH BRAND IP 0 global running / solaris shared - zone1 configured /zones/zone1 solaris excl root@sysA:~# init 0 Step 10) Now boot sysB and configure a zone with the parameters shown above in Step 1. (Again, the safest method is to use "zonecfg ... export" on sysA as described in section "Method Overview" above.) The one difference is the name of the rootzpool storage device, which was shown in the list of assumptions, and which you must determine by booting sysB and using the "format" or "zpool import" command. When that is done, you should see the output shown next. (I used the same zonename - "zone1" - in this example, but you can choose any valid zonename you want.) root@sysB:~# zoneadm list -cv ID NAME STATUS PATH BRAND IP 0 global running / solaris shared - zone1 configured /zones/zone1 solaris excl root@sysB:~# zonecfg -z zone1 info zonename: zone1 zonepath: /zones/zone1 brand: solaris autoboot: false bootargs: file-mac-profile: pool: limitpriv: scheduling-class: ip-type: exclusive hostid: fs-allowed: anet: linkname: net0 ... rootzpool: storage: dev:dsk/c7t3d0 Step 11) Attaching the zone automatically imports the zpool. root@sysB:~# zoneadm -z zone1 attach Imported zone zpool: zone1_rpool Progress being logged to /var/log/zones/zoneadm.20121022T184034Z.zone1.attach Installing: Using existing zone boot environment Zone BE root dataset: zone1_rpool/rpool/ROOT/solaris Cache: Using /var/pkg/publisher. Updating non-global zone: Linking to image /. Processing linked: 1/1 done Updating non-global zone: Auditing packages. No updates necessary for this image. Updating non-global zone: Zone updated. Result: Attach Succeeded. Log saved in non-global zone as /zones/zone1/root/var/log/zones/zoneadm.20121022T184034Z.zone1.attach root@sysB:~# zoneadm -z zone1 boot root@sysB:~# zlogin zone1 [Connected to zone 'zone1' pts/2] Oracle Corporation SunOS 5.11 11.1 September 2012 Step 12) Now let's migrate the zone back to sysA. Create a file in zone1 so we can verify it exists after we migrate the zone back, then begin migrating it back. root@zone1:~# ls /opt root@zone1:~# touch /opt/fileA root@zone1:~# ls -l /opt/fileA -rw-r--r-- 1 root root 0 Oct 22 14:47 /opt/fileA root@zone1:~# exit logout [Connection to zone 'zone1' pts/2 closed] root@sysB:~# zoneadm -z zone1 shutdown root@sysB:~# zoneadm -z zone1 detach Exported zone zpool: zone1_rpool root@sysB:~# init 0 Step 13) Back on sysA, check the status. Oracle Corporation SunOS 5.11 11.1 September 2012 root@sysA:~# zoneadm list -cv ID NAME STATUS PATH BRAND IP 0 global running / solaris shared - zone1 configured /zones/zone1 solaris excl root@sysA:~# zpool list NAME SIZE ALLOC FREE CAP DEDUP HEALTH ALTROOT rpool 17.6G 11.2G 6.47G 63% 1.00x ONLINE - Step 14) Re-attach the zone back to sysA. root@sysA:~# zoneadm -z zone1 attach Imported zone zpool: zone1_rpool Progress being logged to /var/log/zones/zoneadm.20121022T190441Z.zone1.attach Installing: Using existing zone boot environment Zone BE root dataset: zone1_rpool/rpool/ROOT/solaris Cache: Using /var/pkg/publisher. Updating non-global zone: Linking to image /. Processing linked: 1/1 done Updating non-global zone: Auditing packages. No updates necessary for this image. Updating non-global zone: Zone updated. Result: Attach Succeeded. Log saved in non-global zone as /zones/zone1/root/var/log/zones/zoneadm.20121022T190441Z.zone1.attach root@sysA:~# zpool list NAME SIZE ALLOC FREE CAP DEDUP HEALTH ALTROOT rpool 17.6G 11.2G 6.47G 63% 1.00x ONLINE - zone1_rpool 1.98G 491M 1.51G 24% 1.00x ONLINE - root@sysA:~# zoneadm -z zone1 boot root@sysA:~# zlogin zone1 [Connected to zone 'zone1' pts/2] Oracle Corporation SunOS 5.11 11.1 September 2012 root@zone1:~# zpool list NAME SIZE ALLOC FREE CAP DEDUP HEALTH ALTROOT rpool 1.98G 538M 1.46G 26% 1.00x ONLINE - Step 15) Check for the file created on sysB, earlier. root@zone1:~# ls -l /opt total 1 -rw-r--r-- 1 root root 0 Oct 22 14:47 fileA Next Steps Here is a brief list of some of the fun things you can try next. Add space to the zone by adding a second storage device to the rootzpool. Make sure that you add it to the configurations of both zones! Create a new zone, specifying two disks in the rootzpool when you first configure the zone. When you install that zone, or clone it from another zone, zoneadm uses those two disks to create a mirrored pool. (Three disks will result in a three-way mirror, etc.) Conclusion Hopefully you have seen the ease with which you can now move Solaris Zones from one system to another.

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How to maintain encapsulation with composition in C++?

- by iFreilicht

I am designing a class Master that is composed from multiple other classes, A, Base, C and D. These four classes have absolutely no use outside of Master and are meant to split up its functionality into manageable and logically divided packages. They also provide extensible functionality as in the case of Base, which can be inherited from by clients. But, how do I maintain encapsulation of Master with this design? So far, I've got two approaches, which are both far from perfect: 1. Replicate all accessors: Just write accessor-methods for all accessor-methods of all classes that Master is composed of. This leads to perfect encapsulation, because no implementation detail of Master is visible, but is extremely tedious and makes the class definition monstrous, which is exactly what the composition should prevent. Also, adding functionality to one of the composees (is that even a word?) would require to re-write all those methods in Master. An additional problem is that inheritors of Base could only alter, but not add functionality. 2. Use non-assignable, non-copyable member-accessors: Having a class accessor<T> that can not be copied, moved or assigned to, but overrides the operator-> to access an underlying shared_ptr, so that calls like Master->A()->niceFunction(); are made possible. My problem with this is that it kind of breaks encapsulation as I would now be unable to change my implementation of Master to use a different class for the functionality of niceFunction(). Still, it is the closest I've gotten without using the ugly first approach. It also fixes the inheritance issue quite nicely. A small side question would be if such a class already existed in std or boost. EDIT: Wall of code I will now post the code of the header files of the classes discussed. It may be a bit hard to understand, but I'll give my best in explaining all of it. 1. GameTree.h The foundation of it all. This basically is a doubly-linked tree, holding GameObject-instances, which we'll later get to. It also has it's own custom iterator GTIterator, but I left that out for brevity. WResult is an enum with the values SUCCESS and FAILED, but it's not really important. class GameTree { public: //Static methods for the root. Only one root is allowed to exist at a time! static void ConstructRoot(seed_type seed, unsigned int depth); inline static bool rootExists(){ return static_cast<bool>(rootObject_); } inline static weak_ptr<GameTree> root(){ return rootObject_; } //delta is in ms, this is used for velocity, collision and such void tick(unsigned int delta); //Interaction with the tree inline weak_ptr<GameTree> parent() const { return parent_; } inline unsigned int numChildren() const{ return static_cast<unsigned int>(children_.size()); } weak_ptr<GameTree> getChild(unsigned int index) const; template<typename GOType> weak_ptr<GameTree> addChild(seed_type seed, unsigned int depth = 9001){ GOType object{ new GOType(seed) }; return addChildObject(unique_ptr<GameTree>(new GameTree(std::move(object), depth))); } WResult moveTo(weak_ptr<GameTree> newParent); WResult erase(); //Iterators for for( : ) loop GTIterator& begin(){ return *(beginIter_ = std::move(make_unique<GTIterator>(children_.begin()))); } GTIterator& end(){ return *(endIter_ = std::move(make_unique<GTIterator>(children_.end()))); } //unloading should be used when objects are far away WResult unloadChildren(unsigned int newDepth = 0); WResult loadChildren(unsigned int newDepth = 1); inline const RenderObject& renderObject() const{ return gameObject_->renderObject(); } //Getter for the underlying GameObject (I have not tested the template version) weak_ptr<GameObject> gameObject(){ return gameObject_; } template<typename GOType> weak_ptr<GOType> gameObject(){ return dynamic_cast<weak_ptr<GOType>>(gameObject_); } weak_ptr<PhysicsObject> physicsObject() { return gameObject_->physicsObject(); } private: GameTree(const GameTree&); //copying is only allowed internally GameTree(shared_ptr<GameObject> object, unsigned int depth = 9001); //pointer to root static shared_ptr<GameTree> rootObject_; //internal management of a child weak_ptr<GameTree> addChildObject(shared_ptr<GameTree>); WResult removeChild(unsigned int index); //private members shared_ptr<GameObject> gameObject_; shared_ptr<GTIterator> beginIter_; shared_ptr<GTIterator> endIter_; //tree stuff vector<shared_ptr<GameTree>> children_; weak_ptr<GameTree> parent_; unsigned int selfIndex_; //used for deletion, this isn't necessary void initChildren(unsigned int depth); //constructs children }; 2. GameObject.h This is a bit hard to grasp, but GameObject basically works like this: When constructing a GameObject, you construct its basic attributes and a CResult-instance, which contains a vector<unique_ptr<Construction>>. The Construction-struct contains all information that is needed to construct a GameObject, which is a seed and a function-object that is applied at construction by a factory. This enables dynamic loading and unloading of GameObjects as done by GameTree. It also means that you have to define that factory if you inherit GameObject. This inheritance is also the reason why GameTree has a template-function gameObject<GOType>. GameObject can contain a RenderObject and a PhysicsObject, which we'll later get to. Anyway, here's the code. class GameObject; typedef unsigned long seed_type; //this declaration magic means that all GameObjectFactorys inherit from GameObjectFactory<GameObject> template<typename GOType> struct GameObjectFactory; template<> struct GameObjectFactory<GameObject>{ virtual unique_ptr<GameObject> construct(seed_type seed) const = 0; }; template<typename GOType> struct GameObjectFactory : GameObjectFactory<GameObject>{ GameObjectFactory() : GameObjectFactory<GameObject>(){} unique_ptr<GameObject> construct(seed_type seed) const{ return unique_ptr<GOType>(new GOType(seed)); } }; //same as with the factories. this is important for storing them in vectors template<typename GOType> struct Construction; template<> struct Construction<GameObject>{ virtual unique_ptr<GameObject> construct() const = 0; }; template<typename GOType> struct Construction : Construction<GameObject>{ Construction(seed_type seed, function<void(GOType*)> func = [](GOType* null){}) : Construction<GameObject>(), seed_(seed), func_(func) {} unique_ptr<GameObject> construct() const{ unique_ptr<GameObject> gameObject{ GOType::factory.construct(seed_) }; func_(dynamic_cast<GOType*>(gameObject.get())); return std::move(gameObject); } seed_type seed_; function<void(GOType*)> func_; }; typedef struct CResult { CResult() : constructions{} {} CResult(CResult && o) : constructions(std::move(o.constructions)) {} CResult& operator= (CResult& other){ if (this != &other){ for (unique_ptr<Construction<GameObject>>& child : other.constructions){ constructions.push_back(std::move(child)); } } return *this; } template<typename GOType> void push_back(seed_type seed, function<void(GOType*)> func = [](GOType* null){}){ constructions.push_back(make_unique<Construction<GOType>>(seed, func)); } vector<unique_ptr<Construction<GameObject>>> constructions; } CResult; //finally, the GameObject class GameObject { public: GameObject(seed_type seed); GameObject(const GameObject&); virtual void tick(unsigned int delta); inline Matrix4f trafoMatrix(){ return physicsObject_->transformationMatrix(); } //getter inline seed_type seed() const{ return seed_; } inline CResult& properties(){ return properties_; } inline const RenderObject& renderObject() const{ return *renderObject_; } inline weak_ptr<PhysicsObject> physicsObject() { return physicsObject_; } protected: virtual CResult construct_(seed_type seed) = 0; CResult properties_; shared_ptr<RenderObject> renderObject_; shared_ptr<PhysicsObject> physicsObject_; seed_type seed_; }; 3. PhysicsObject That's a bit easier. It is responsible for position, velocity and acceleration. It will also handle collisions in the future. It contains three Transformation objects, two of which are optional. I'm not going to include the accessors on the PhysicsObject class because I tried my first approach on it and it's just pure madness (way over 30 functions). Also missing: the named constructors that construct PhysicsObjects with different behaviour. class Transformation{ Vector3f translation_; Vector3f rotation_; Vector3f scaling_; public: Transformation() : translation_{ 0, 0, 0 }, rotation_{ 0, 0, 0 }, scaling_{ 1, 1, 1 } {}; Transformation(Vector3f translation, Vector3f rotation, Vector3f scaling); inline Vector3f translation(){ return translation_; } inline void translation(float x, float y, float z){ translation(Vector3f(x, y, z)); } inline void translation(Vector3f newTranslation){ translation_ = newTranslation; } inline void translate(float x, float y, float z){ translate(Vector3f(x, y, z)); } inline void translate(Vector3f summand){ translation_ += summand; } inline Vector3f rotation(){ return rotation_; } inline void rotation(float pitch, float yaw, float roll){ rotation(Vector3f(pitch, yaw, roll)); } inline void rotation(Vector3f newRotation){ rotation_ = newRotation; } inline void rotate(float pitch, float yaw, float roll){ rotate(Vector3f(pitch, yaw, roll)); } inline void rotate(Vector3f summand){ rotation_ += summand; } inline Vector3f scaling(){ return scaling_; } inline void scaling(float x, float y, float z){ scaling(Vector3f(x, y, z)); } inline void scaling(Vector3f newScaling){ scaling_ = newScaling; } inline void scale(float x, float y, float z){ scale(Vector3f(x, y, z)); } void scale(Vector3f factor){ scaling_(0) *= factor(0); scaling_(1) *= factor(1); scaling_(2) *= factor(2); } Matrix4f matrix(){ return WMatrix::Translation(translation_) * WMatrix::Rotation(rotation_) * WMatrix::Scale(scaling_); } }; class PhysicsObject; typedef void tickFunction(PhysicsObject& self, unsigned int delta); class PhysicsObject{ PhysicsObject(const Transformation& trafo) : transformation_(trafo), transformationVelocity_(nullptr), transformationAcceleration_(nullptr), tick_(nullptr) {} PhysicsObject(PhysicsObject&& other) : transformation_(other.transformation_), transformationVelocity_(std::move(other.transformationVelocity_)), transformationAcceleration_(std::move(other.transformationAcceleration_)), tick_(other.tick_) {} Transformation transformation_; unique_ptr<Transformation> transformationVelocity_; unique_ptr<Transformation> transformationAcceleration_; tickFunction* tick_; public: void tick(unsigned int delta){ tick_ ? tick_(*this, delta) : 0; } inline Matrix4f transformationMatrix(){ return transformation_.matrix(); } } 4. RenderObject RenderObject is a base class for different types of things that could be rendered, i.e. Meshes, Light Sources or Sprites. DISCLAIMER: I did not write this code, I'm working on this project with someone else. class RenderObject { public: RenderObject(float renderDistance); virtual ~RenderObject(); float renderDistance() const { return renderDistance_; } void setRenderDistance(float rD) { renderDistance_ = rD; } protected: float renderDistance_; }; struct NullRenderObject : public RenderObject{ NullRenderObject() : RenderObject(0.f){}; }; class Light : public RenderObject{ public: Light() : RenderObject(30.f){}; }; class Mesh : public RenderObject{ public: Mesh(unsigned int seed) : RenderObject(20.f) { meshID_ = 0; textureID_ = 0; if (seed == 1) meshID_ = Model::getMeshID("EM-208_heavy"); else meshID_ = Model::getMeshID("cube"); }; unsigned int getMeshID() const { return meshID_; } unsigned int getTextureID() const { return textureID_; } private: unsigned int meshID_; unsigned int textureID_; }; I guess this shows my issue quite nicely: You see a few accessors in GameObject which return weak_ptrs to access members of members, but that is not really what I want. Also please keep in mind that this is NOT, by any means, finished or production code! It is merely a prototype and there may be inconsistencies, unnecessary public parts of classes and such.

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IIS Strategies for Accessing Secured Network Resources

- by ErikE

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

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IIS Strategies for Accessing Secured Network Resources

- by Emtucifor

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

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Replacing jQuery.live() with jQuery.on()

- by Rick Strahl

jQuery 1.9 and 1.10 have introduced a host of changes, but for the most part these changes are mostly transparent to existing application usage of jQuery. After spending some time last week with a few of my projects and going through them with a specific eye for jQuery failures I found that for the most part there wasn't a big issue. The vast majority of code continues to run just fine with either 1.9 or 1.10 (which are supposed to be in sync but with 1.10 removing support for legacy Internet Explorer pre-9.0 versions). However, one particular change in the new versions has caused me quite a bit of update trouble, is the removal of the jQuery.live() function. This is my own fault I suppose - .live() has been deprecated for a while, but with 1.9 and later it was finally removed altogether from jQuery. In the past I had quite a bit of jQuery code that used .live() and it's one of the things that's holding back my upgrade process, although I'm slowly cleaning up my code and switching to the .on() function as the replacement. jQuery.live() jQuery.live() was introduced a long time ago to simplify handling events on matched elements that exist currently on the document and those that are are added in the future and also match the selector. jQuery uses event bubbling, special event binding, plus some magic using meta data attached to a parent level element to check and see if the original target event element matches the selected selected elements (for more info see Elijah Manor's comment below). An Example Assume a list of items like the following in HTML for example and further assume that the items in this list can be appended to at a later point. In this app there's a smallish initial list that loads to start, and as the user scrolls towards the end of the initial small list more items are loaded dynamically and added to the list.<div id="PostItemContainer" class="scrollbox"> <div class="postitem" data-id="4z6qhomm"> <div class="post-icon"></div> <div class="postitemheader"><a href="show/4z6qhomm" target="Content">1999 Buick Century For Sale!</a></div> <div class="postitemprice rightalign">$ 3,500 O.B.O.</div> <div class="smalltext leftalign">Jun. 07 @ 1:06am</div> <div class="post-byline">- Vehicles - Automobiles</div> </div> <div class="postitem" data-id="2jtvuu17"> <div class="postitemheader"><a href="show/2jtvuu17" target="Content">Toyota VAN 1987</a></div> <div class="postitemprice rightalign">$950</div> <div class="smalltext leftalign">Jun. 07 @ 12:29am</div> <div class="post-byline">- Vehicles - Automobiles</div> </div> … </div> With the jQuery.live() function you could easily select elements and hook up a click handler like this:$(".postitem").live("click", function() {...}); Simple and perfectly readable. The behavior of the .live handler generally was the same as the corresponding simple event handlers like .click(), except that you have to explicitly name the event instead of using one of the methods. Re-writing with jQuery.on() With .live() removed in 1.9 and later we have to re-write .live() code above with an alternative. The jQuery documentation points you at the .on() or .delegate() functions to update your code. jQuery.on() is a more generic event handler function, and it's what jQuery uses internally to map the high level event functions like .click(),.change() etc. that jQuery exposes. Using jQuery.on() however is not a one to one replacement of the .live() function. While .on() can handle events directly and use the same syntax as .live() did, you'll find if you simply switch out .live() with .on() that events on not-yet existing elements will not fire. IOW, the key feature of .live() is not working. You can use .on() to get the desired effect however, but you have to change the syntax to explicitly handle the event you're interested in on the container and then provide a filter selector to specify which elements you are actually interested in for handling the event for. Sounds more complicated than it is and it's easier to see with an example. For the list above hooking .postitem clicks, using jQuery.on() looks like this:$("#PostItemContainer").on("click", ".postitem", function() {...}); You specify a container that can handle the .click event and then provide a filter selector to find the child elements that trigger the the actual event. So here #PostItemContainer contains many .postitems, whose click events I want to handle. Any container will do including document, but I tend to use the container closest to the elements I actually want to handle the events on to minimize the event bubbling that occurs to capture the event. With this code I get the same behavior as with .live() and now as new .postitem elements are added the click events are always available. Sweet. Here's the full event signature for the .on() function: .on( events [, selector ] [, data ], handler(eventObject) ) Note that the selector is optional - if you omit it you essentially create a simple event handler that handles the event directly on the selected object. The filter/child selector required if you want life-like - uh, .live() like behavior to happen. While it's a bit more verbose than what .live() did, .on() provides the same functionality by being more explicit on what your parent container for trapping events is. .on() is good Practice even for ordinary static Element Lists As a side note, it's a good practice to use jQuery.on() or jQuery.delegate() for events in most cases anyway, using this 'container event trapping' syntax. That's because rather than requiring lots of event handlers on each of the child elements (.postitem in the sample above), there's just one event handler on the container, and only when clicked does jQuery drill down to find the matching filter element and tries to match it to the originating element. In the early days of jQuery I used manually build handlers that did this and manually drilled from the event object into the originalTarget to determine if it's a matching element. With later versions of jQuery the various event functions in jQuery essentially provide this functionality out of the box with functions like .on() and .delegate(). All of this is nothing new, but I thought I'd write this up because I have on a few occasions forgotten what exactly was needed to replace the many .live() function calls that litter my code - especially older code. This will be a nice reminder next time I have a memory blank on this topic. And maybe along the way I've helped one or two of you as well to clean up your .live() code…© Rick Strahl, West Wind Technologies, 2005-2013Posted in jQuery Tweet !function(d,s,id){var js,fjs=d.getElementsByTagName(s)[0];if(!d.getElementById(id)){js=d.createElement(s);js.id=id;js.src="//platform.twitter.com/widgets.js";fjs.parentNode.insertBefore(js,fjs);}}(document,"script","twitter-wjs"); (function() { var po = document.createElement('script'); po.type = 'text/javascript'; po.async = true; po.src = 'https://apis.google.com/js/plusone.js'; var s = document.getElementsByTagName('script')[0]; s.parentNode.insertBefore(po, s); })();

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Interesting articles and blogs on SPARC T4

- by mv

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

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Listing common SQL Code Smells.

- by Phil Factor

Once you’ve done a number of SQL Code-reviews, you’ll know those signs in the code that all might not be well. These ’Code Smells’ are coding styles that don’t directly cause a bug, but are indicators that all is not well with the code. . Kent Beck and Massimo Arnoldi seem to have coined the phrase in the "OnceAndOnlyOnce" page of www.C2.com, where Kent also said that code "wants to be simple". Bad Smells in Code was an essay by Kent Beck and Martin Fowler, published as Chapter 3 of the book ‘Refactoring: Improving the Design of Existing Code’ (ISBN 978-0201485677) Although there are generic code-smells, SQL has its own particular coding habits that will alert the programmer to the need to re-factor what has been written. See Exploring Smelly Code and Code Deodorants for Code Smells by Nick Harrison for a grounding in Code Smells in C# I’ve always been tempted by the idea of automating a preliminary code-review for SQL. It would be so useful to trawl through code and pick up the various problems, much like the classic ‘Lint’ did for C, and how the Code Metrics plug-in for .NET Reflector by Jonathan 'Peli' de Halleux is used for finding Code Smells in .NET code. The problem is that few of the standard procedural code smells are relevant to SQL, and we need an agreed list of code smells. Merrilll Aldrich made a grand start last year in his blog Top 10 T-SQL Code Smells.However, I'd like to make a start by discovering if there is a general opinion amongst Database developers what the most important SQL Smells are. One can be a bit defensive about code smells. I will cheerfully write very long stored procedures, even though they are frowned on. I’ll use dynamic SQL occasionally. You can only use them as an aid for your own judgment and it is fine to ‘sign them off’ as being appropriate in particular circumstances. Also, whole classes of ‘code smells’ may be irrelevant for a particular database. The use of proprietary SQL, for example, is only a ‘code smell’ if there is a chance that the database will have to be ported to another RDBMS. The use of dynamic SQL is a risk only with certain security models. As the saying goes, a CodeSmell is a hint of possible bad practice to a pragmatist, but a sure sign of bad practice to a purist. Plamen Ratchev’s wonderful article Ten Common SQL Programming Mistakes lists some of these ‘code smells’ along with out-and-out mistakes, but there are more. The use of nested transactions, for example, isn’t entirely incorrect, even though the database engine ignores all but the outermost: but it does flag up the possibility that the programmer thinks that nested transactions are supported. If anything requires some sort of general agreement, the definition of code smells is one. I’m therefore going to make this Blog ‘dynamic, in that, if anyone twitters a suggestion with a #SQLCodeSmells tag (or sends me a twitter) I’ll update the list here. If you add a comment to the blog with a suggestion of what should be added or removed, I’ll do my best to oblige. In other words, I’ll try to keep this blog up to date. The name against each 'smell' is the name of the person who Twittered me, commented about or who has written about the 'smell'. it does not imply that they were the first ever to think of the smell! Use of deprecated syntax such as *= (Dave Howard) Denormalisation that requires the shredding of the contents of columns. (Merrill Aldrich) Contrived interfaces Use of deprecated datatypes such as TEXT/NTEXT (Dave Howard) Datatype mis-matches in predicates that rely on implicit conversion.(Plamen Ratchev) Using Correlated subqueries instead of a join (Dave_Levy/ Plamen Ratchev) The use of Hints in queries, especially NOLOCK (Dave Howard /Mike Reigler) Few or No comments. Use of functions in a WHERE clause. (Anil Das) Overuse of scalar UDFs (Dave Howard, Plamen Ratchev) Excessive ‘overloading’ of routines. The use of Exec xp_cmdShell (Merrill Aldrich) Excessive use of brackets. (Dave Levy) Lack of the use of a semicolon to terminate statements Use of non-SARGable functions on indexed columns in predicates (Plamen Ratchev) Duplicated code, or strikingly similar code. Misuse of SELECT * (Plamen Ratchev) Overuse of Cursors (Everyone. Special mention to Dave Levy & Adrian Hills) Overuse of CLR routines when not necessary (Sam Stange) Same column name in different tables with different datatypes. (Ian Stirk) Use of ‘broken’ functions such as ‘ISNUMERIC’ without additional checks. Excessive use of the WHILE loop (Merrill Aldrich) INSERT ... EXEC (Merrill Aldrich) The use of stored procedures where a view is sufficient (Merrill Aldrich) Not using two-part object names (Merrill Aldrich) Using INSERT INTO without specifying the columns and their order (Merrill Aldrich) Full outer joins even when they are not needed. (Plamen Ratchev) Huge stored procedures (hundreds/thousands of lines). Stored procedures that can produce different columns, or order of columns in their results, depending on the inputs. Code that is never used. Complex and nested conditionals WHILE (not done) loops without an error exit. Variable name same as the Datatype Vague identifiers. Storing complex data or list in a character map, bitmap or XML field User procedures with sp_ prefix (Aaron Bertrand)Views that reference views that reference views that reference views (Aaron Bertrand) Inappropriate use of sql_variant (Neil Hambly) Errors with identity scope using SCOPE_IDENTITY @@IDENTITY or IDENT_CURRENT (Neil Hambly, Aaron Bertrand) Schemas that involve multiple dated copies of the same table instead of partitions (Matt Whitfield-Atlantis UK) Scalar UDFs that do data lookups (poor man's join) (Matt Whitfield-Atlantis UK) Code that allows SQL Injection (Mladen Prajdic) Tables without clustered indexes (Matt Whitfield-Atlantis UK) Use of "SELECT DISTINCT" to mask a join problem (Nick Harrison) Multiple stored procedures with nearly identical implementation. (Nick Harrison) Excessive column aliasing may point to a problem or it could be a mapping implementation. (Nick Harrison) Joining "too many" tables in a query. (Nick Harrison) Stored procedure returning more than one record set. (Nick Harrison) A NOT LIKE condition (Nick Harrison) excessive "OR" conditions. (Nick Harrison) User procedures with sp_ prefix (Aaron Bertrand) Views that reference views that reference views that reference views (Aaron Bertrand) sp_OACreate or anything related to it (Bill Fellows) Prefixing names with tbl_, vw_, fn_, and usp_ ('tibbling') (Jeremiah Peschka) Aliases that go a,b,c,d,e... (Dave Levy/Diane McNurlan) Overweight Queries (e.g. 4 inner joins, 8 left joins, 4 derived tables, 10 subqueries, 8 clustered GUIDs, 2 UDFs, 6 case statements = 1 query) (Robert L Davis) Order by 3,2 (Dave Levy) MultiStatement Table functions which are then filtered 'Sel * from Udf() where Udf.Col = Something' (Dave Ballantyne) running a SQL 2008 system in SQL 2000 compatibility mode(John Stafford)

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LightDM will not start after stopping it

- by Sweeters

I am running Ubuntu 11.10 "Oneiric Ocelot", and in trying to install the nvidia CUDA developer drivers I switched to a virtual terminal (Ctrl-Alt-F5) and stopped lightdm (installation required that no X server instance be running) through sudo service lightdm stop. Re-starting lightdm with sudo service lightdm start did not work: A couple of * Starting [...] lines where displayed, but the process hanged. (I do not remember at which point, but I think it was * Starting System V runlevel compatibility. I manually rebooted my laptop, and ever since booting seems to hang, usually around the * Starting anac(h)ronistic cron [OK] log line (not consistently at that point, though). From that point on, I seem to be able to interact with my system only through a tty session (Ctrl-Alt-F1). I've tried purging and reinstalling both lightdm and gdm, as well as selecting both as the default display managers (through sudo dpkg-reconfigure [lightdm / gdm] or by manually editing /etc/X11/default-display-manager) through both apt-get and aptitude (that shouldn't make a difference anyway) after updating the packages, but the problem persists. Some of the responses I'm getting are the following: After running sudo dpkg-reconfigure lightdm (but not ... gdm) I get the following message: dpkg-maintscript-helper:warning: environment variable DPKG_MATINSCRIPT_NAME missing dpkg-maintscript-helper:warning: environment variable DPKG_MATINSCRIPT_PACKAGE missing After trying sudo service lightdm start or sudo start lightdm I get to see the boot loading screen again but nothing changes. If I go back to the tty shell I see lightdm start/running, process <num> but ps -e | grep lightdm gives no output. After trying sudo service gdm start or sudo starg gdm I get the gdm start/running, process <num> message, and gdm-binary is supposedly an active process, but all that happens is that the screen blinks a couple of times and nothing else. Other candidate solutions that I'd found on the web included running startx but when I try that I get an error output [...] Fatal server error: no screens found [...]. Moreover, I made sure that lightdm-gtk-greeter is installed but that did not help either. Please excuse my not including complete outputs/logs; I am writing this post from another computer and it's hard to manually copy the complete logs. Also, I've seen several posts that had to do with similar problems, but either there was no fix, or the one suggested did not work for me. In closing: Please help! I very much hope to avoid re-installing Ubuntu from scratch! :) Alex @mosi I did not manage to fix the NVIDIA kernel driver as per your instructions. I should perhaps mention that I'm on a Dell XPS15 laptop with an NVIDIA Optimus graphics card, and that I have bumblebee installed (which installs nvidia drivers during its installation, I believe). Issuing the mentioned commands I get the following: ~$uname -r 3.0.0-12-generic ~$lsmod | grep -i nvidia nvidia 11713772 0 ~$dmesg | grep -i nvidia [ 8.980041] nvidia: module license 'NVIDIA' taints kernel. [ 9.354860] nvidia 0000:01:00.0: power state changed by ACPI to D0 [ 9.354864] nvidia 0000:01:00.0: power state changed by ACPI to D0 [ 9.354868] nvidia 0000:01:00.0: enabling device (0006 -> 0007) [ 9.354873] nvidia 0000:01:00.0: PCI INT A -> GSI 16 (level, low) -> IRQ 16 [ 9.354879] nvidia 0000:01:00.0: setting latency timer to 64 [ 9.355052] NVRM: loading NVIDIA UNIX x86_64 Kernel Module 280.13 Wed Jul 27 16:53:56 PDT 2011 Also, running aptitude search nvidia gives me the following: p nvidia-173 - NVIDIA binary Xorg driver, kernel module a p nvidia-173-dev - NVIDIA binary Xorg driver development file p nvidia-173-updates - NVIDIA binary Xorg driver, kernel module a p nvidia-173-updates-dev - NVIDIA binary Xorg driver development file p nvidia-96 - NVIDIA binary Xorg driver, kernel module a p nvidia-96-dev - NVIDIA binary Xorg driver development file p nvidia-96-updates - NVIDIA binary Xorg driver, kernel module a p nvidia-96-updates-dev - NVIDIA binary Xorg driver development file p nvidia-cg-toolkit - Cg Toolkit - GPU Shader Authoring Language p nvidia-common - Find obsolete NVIDIA drivers i nvidia-current - NVIDIA binary Xorg driver, kernel module a p nvidia-current-dev - NVIDIA binary Xorg driver development file c nvidia-current-updates - NVIDIA binary Xorg driver, kernel module a p nvidia-current-updates-dev - NVIDIA binary Xorg driver development file i nvidia-settings - Tool of configuring the NVIDIA graphics dr p nvidia-settings-updates - Tool of configuring the NVIDIA graphics dr v nvidia-va-driver - v nvidia-va-driver - I've tried manually installing (sudo aptitude install <package>) packages nvidia-common and nvidia-settings-updates but to no avail. For example, sudo aptitude install nvidia-settings-updates returns the following log: Reading package lists... Building dependency tree... Reading state information... Reading extended state information... Initializing package states... Writing extended state information... No packages will be installed, upgraded, or removed. 0 packages upgraded, 0 newly installed, 0 to remove and 83 not upgraded. Need to get 0 B of archives. After unpacking 0 B will be used. Writing extended state information... Reading package lists... Building dependency tree... Reading state information... Reading extended state information... Initializing package states... Writing extended state information... The same happens with the Linux headers (i.e. I cannot seem to be able to install linux-headers-3.0.0-12-generic). The output of aptitude search linux-headers is as follows: v linux-headers - v linux-headers - v linux-headers-2.6 - i linux-headers-2.6.38-11 - Header files related to Linux kernel versi i linux-headers-2.6.38-11-generic - Linux kernel headers for version 2.6.38 on i A linux-headers-2.6.38-8 - Header files related to Linux kernel versi i A linux-headers-2.6.38-8-generic - Linux kernel headers for version 2.6.38 on v linux-headers-3 - v linux-headers-3.0 - v linux-headers-3.0 - i A linux-headers-3.0.0-12 - Header files related to Linux kernel versi p linux-headers-3.0.0-12-generic - Linux kernel headers for version 3.0.0 on p linux-headers-3.0.0-12-generic- - Linux kernel headers for version 3.0.0 on p linux-headers-3.0.0-12-server - Linux kernel headers for version 3.0.0 on p linux-headers-3.0.0-12-virtual - Linux kernel headers for version 3.0.0 on p linux-headers-generic - Generic Linux kernel headers p linux-headers-generic-pae - Generic Linux kernel headers v linux-headers-lbm - v linux-headers-lbm - v linux-headers-lbm-2.6 - v linux-headers-lbm-2.6 - p linux-headers-lbm-3.0.0-12-gene - Header files related to linux-backports-mo p linux-headers-lbm-3.0.0-12-gene - Header files related to linux-backports-mo p linux-headers-lbm-3.0.0-12-serv - Header files related to linux-backports-mo p linux-headers-server - Linux kernel headers on Server Equipment. p linux-headers-virtual - Linux kernel headers for virtual machines @heartsmagic I did try purging and reinstalling any nvidia driver packages, but it did not seem to make a difference, My xorg.conf file contains the following: # nvidia-xconfig: X configuration file generated by nvidia-xconfig # nvidia-xconfig: version 280.13 ([email protected]) Wed Jul 27 17:15:58 PDT 2011 Section "ServerLayout" Identifier "Layout0" Screen 0 "Screen0" 0 0 InputDevice "Keyboard0" "CoreKeyboard" InputDevice "Mouse0" "CorePointer" EndSection Section "Files" EndSection Section "InputDevice" # generated from default Identifier "Mouse0" Driver "mouse" Option "Protocol" "auto" Option "Device" "/dev/psaux" Option "Emulate3Buttons" "no" Option "ZAxisMapping" "4 5" EndSection Section "InputDevice" # generated from default Identifier "Keyboard0" Driver "kbd" EndSection Section "Monitor" Identifier "Monitor0" VendorName "Unknown" ModelName "Unknown" HorizSync 28.0 - 33.0 VertRefresh 43.0 - 72.0 Option "DPMS" EndSection Section "Device" Identifier "Device0" Driver "nvidia" VendorName "NVIDIA Corporation" EndSection Section "Screen" Identifier "Screen0" Device "Device0" Monitor "Monitor0" DefaultDepth 24 SubSection "Display" Depth 24 EndSubSection EndSection

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Networking in VirtualBox

- by Fat Bloke

Networking in VirtualBox is extremely powerful, but can also be a bit daunting, so here's a quick overview of the different ways you can setup networking in VirtualBox, with a few pointers as to which configurations should be used and when. VirtualBox allows you to configure up to 8 virtual NICs (Network Interface Controllers) for each guest vm (although only 4 are exposed in the GUI) and for each of these NICs you can configure: Which virtualized NIC-type is exposed to the Guest. Examples include: Intel PRO/1000 MT Server (82545EM), AMD PCNet FAST III (Am79C973, the default) or a Paravirtualized network adapter (virtio-net). How the NIC operates with respect to your Host's physical networking. The main modes are: Network Address Translation (NAT) Bridged networking Internal networking Host-only networking NAT with Port-forwarding The choice of NIC-type comes down to whether the guest has drivers for that NIC. VirtualBox, suggests a NIC based on the guest OS-type that you specify during creation of the vm, and you rarely need to modify this. But the choice of networking mode depends on how you want to use your vm (client or server) and whether you want other machines on your network to see it. So let's look at each mode in a bit more detail... Network Address Translation (NAT) This is the default mode for new vm's and works great in most situations when the Guest is a "client" type of vm. (i.e. most network connections are outbound). Here's how it works: When the guest OS boots, it typically uses DHCP to get an IP address. VirtualBox will field this DHCP request and tell the guest OS its assigned IP address and the gateway address for routing outbound connections. In this mode, every vm is assigned the same IP address (10.0.2.15) because each vm thinks they are on their own isolated network. And when they send their traffic via the gateway (10.0.2.2) VirtualBox rewrites the packets to make them appear as though they originated from the Host, rather than the Guest (running inside the Host). This means that the Guest will work even as the Host moves from network to network (e.g. laptop moving between locations), and from wireless to wired connections too. However, how does another computer initiate a connection into a Guest? e.g. connecting to a web server running in the Guest. This is not (normally) possible using NAT mode as there is no route into the Guest OS. So for vm's running servers we need a different networking mode.... Bridged Networking Bridged Networking is used when you want your vm to be a full network citizen, i.e. to be an equal to your host machine on the network. In this mode, a virtual NIC is "bridged" to a physical NIC on your host, like this: The effect of this is that each VM has access to the physical network in the same way as your host. It can access any service on the network such as external DHCP services, name lookup services, and routing information just as the host does. Logically, the network looks like this: The downside of this mode is that if you run many vm's you can quickly run out of IP addresses or your network administrator gets fed up with you asking for statically assigned IP addresses. Secondly, if your host has multiple physical NICs (e.g. Wireless and Wired) you must reconfigure the bridge when your host jumps networks. Hmm, so what if you want to run servers in vm's but don't want to involve your network administrator? Maybe one of the next 2 modes is for you... Internal Networking When you configure one or more vm's to sit on an Internal network, VirtualBox ensures that all traffic on that network stays within the host and is only visible to vm's on that virtual network. Configuration looks like this: The internal network ( in this example "intnet" ) is a totally isolated network and so is very "quiet". This is good for testing when you need a separate, clean network, and you can create sophisticated internal networks with vm's that provide their own services to the internal network. (e.g. Active Directory, DHCP, etc). Note that not even the Host is a member of the internal network, but this mode allows vm's to function even when the Host is not connected to a network (e.g. on a plane). Note that in this mode, VirtualBox provides no "convenience" services such as DHCP, so your machines must be statically configured or one of the vm's needs to provide a DHCP/Name service. Multiple internal networks are possible and you can configure vm's to have multiple NICs to sit across internal and other network modes and thereby provide routes if needed. But all this sounds tricky. What if you want an Internal Network that the host participates on with VirtualBox providing IP addresses to the Guests? Ah, then for this, you might want to consider Host-only Networking... Host-only Networking Host-only Networking is like Internal Networking in that you indicate which network the Guest sits on, in this case, "vboxnet0": All vm's sitting on this "vboxnet0" network will see each other, and additionally, the host can see these vm's too. However, other external machines cannot see Guests on this network, hence the name "Host-only". Logically, the network looks like this: This looks very similar to Internal Networking but the host is now on "vboxnet0" and can provide DHCP services. To configure how a Host-only network behaves, look in the VirtualBox Manager...Preferences...Network dialog: Port-Forwarding with NAT Networking Now you may think that we've provided enough modes here to handle every eventuality but here's just one more... What if you cart around a mobile-demo or dev environment on, say, a laptop and you have one or more vm's that you need other machines to connect into? And you are continually hopping onto different (customer?) networks. In this scenario: NAT - won't work because external machines need to connect in. Bridged - possibly an option, but does your customer want you eating IP addresses and can your software cope with changing networks? Internal - we need the vm(s) to be visible on the network, so this is no good. Host-only - same problem as above, we want external machines to connect in to the vm's. Enter Port-forwarding to save the day! Configure your vm's to use NAT networking; Add Port Forwarding rules; External machines connect to "host":"port number" and connections are forwarded by VirtualBox to the guest:port number specified. For example, if your vm runs a web server on port 80, you could set up rules like this: ...which reads: "any connections on port 8080 on the Host will be forwarded onto this vm's port 80". This provides a mobile demo system which won't need re-configuring every time you open your laptop lid. Summary VirtualBox has a very powerful set of options allowing you to set up almost any configuration your heart desires. For more information, check out the VirtualBox User Manual on Virtual Networking. -FB

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cannot connect with huawei e173 after upgrade to 12.10 using network manager

- by user104195

Since upgrade from 12.04 to 12.10 I can't connect to internet using mobile broadband modem Huawei e173. It worked earlier without problems and now it seems to be properly recognized (at least its connections appear in network manager applet), and after selecting connection manually it starts connection procedure. After about 20 seconds it returns to state disconnected. After browsing internet I've found that running network manager with: NM_PPP_DEBUG=1 /usr/sbin/NetworkManager --no-daemon After inserting modem I get: NetworkManager[507]: <warn> (ttyUSB2): failed to look up interface index NetworkManager[507]: <info> (ttyUSB2): new GSM/UMTS device (driver: 'option1' ifindex: 0) NetworkManager[507]: <info> (ttyUSB2): exported as /org/freedesktop/NetworkManager/Devices/2 NetworkManager[507]: <info> (ttyUSB2): now managed NetworkManager[507]: <info> (ttyUSB2): device state change: unmanaged -> unavailable (reason 'managed') [10 20 2] NetworkManager[507]: <info> (ttyUSB2): deactivating device (reason 'managed') [2] NetworkManager[507]: <info> (ttyUSB2): device state change: unavailable -> disconnected (reason 'none') [20 30 0] where 'failed to look up interface index' seems to be suspicious. After starting connecting: NetworkManager[507]: <info> Activation (ttyUSB2) starting connection 'Plus - Dostep standardowy' NetworkManager[507]: <info> (ttyUSB2): device state change: disconnected -> prepare (reason 'none') [30 40 0] NetworkManager[507]: <info> Activation (ttyUSB2) Stage 1 of 5 (Device Prepare) scheduled... NetworkManager[507]: <info> Activation (ttyUSB2) Stage 1 of 5 (Device Prepare) started... NetworkManager[507]: <info> (ttyUSB2): device state change: prepare -> need-auth (reason 'none') [40 60 0] NetworkManager[507]: <info> Activation (ttyUSB2) Stage 1 of 5 (Device Prepare) complete. NetworkManager[507]: <info> Activation (ttyUSB2) Stage 1 of 5 (Device Prepare) scheduled... NetworkManager[507]: <info> Activation (ttyUSB2) Stage 1 of 5 (Device Prepare) started... NetworkManager[507]: <info> (ttyUSB2): device state change: need-auth -> prepare (reason 'none') [60 40 0] NetworkManager[507]: <info> Activation (ttyUSB2) Stage 1 of 5 (Device Prepare) complete. NetworkManager[507]: <info> WWAN now enabled by management service NetworkManager[507]: <info> Activation (ttyUSB2) Stage 2 of 5 (Device Configure) scheduled... NetworkManager[507]: <info> Activation (ttyUSB2) Stage 2 of 5 (Device Configure) starting... NetworkManager[507]: <info> (ttyUSB2): device state change: prepare -> config (reason 'none') [40 50 0] NetworkManager[507]: <info> Activation (ttyUSB2) Stage 2 of 5 (Device Configure) successful. NetworkManager[507]: <info> Activation (ttyUSB2) Stage 3 of 5 (IP Configure Start) scheduled. NetworkManager[507]: <info> Activation (ttyUSB2) Stage 2 of 5 (Device Configure) complete. NetworkManager[507]: <info> Activation (ttyUSB2) Stage 3 of 5 (IP Configure Start) started... NetworkManager[507]: <info> (ttyUSB2): device state change: config -> ip-config (reason 'none') [50 70 0] NetworkManager[507]: <info> starting PPP connection NetworkManager[507]: <info> pppd started with pid 663 NetworkManager[507]: <info> Activation (ttyUSB2) Stage 4 of 5 (IPv6 Configure Timeout) scheduled... NetworkManager[507]: <info> Activation (ttyUSB2) Stage 3 of 5 (IP Configure Start) complete. NetworkManager[507]: <info> Activation (ttyUSB2) Stage 4 of 5 (IPv6 Configure Timeout) started... NetworkManager[507]: <info> Activation (ttyUSB2) Stage 4 of 5 (IPv6 Configure Timeout) complete. Plugin /usr/lib/pppd/2.4.5/nm-pppd-plugin.so loaded. ** Message: nm-ppp-plugin: (plugin_init): initializing ** Message: nm-ppp-plugin: (nm_phasechange): status 3 / phase 'serial connection' Removed stale lock on ttyUSB2 (pid 32146) using channel 23 NetworkManager[507]: SCPlugin-Ifupdown: devices added (path: /sys/devices/virtual/net/ppp0, iface: ppp0) NetworkManager[507]: SCPlugin-Ifupdown: device added (path: /sys/devices/virtual/net/ppp0, iface: ppp0): no ifupdown configuration found. NetworkManager[507]: <warn> /sys/devices/virtual/net/ppp0: couldn't determine device driver; ignoring... Using interface ppp0 Connect: ppp0 <--> /dev/ttyUSB2 ** Message: nm-ppp-plugin: (nm_phasechange): status 5 / phase 'establish' sent [LCP ConfReq id=0x1 <asyncmap 0x0> <magic 0x64b4024a> <pcomp> <accomp>] sent [LCP ConfReq id=0x1 <asyncmap 0x0> <magic 0x64b4024a> <pcomp> <accomp>] sent [LCP ConfReq id=0x1 <asyncmap 0x0> <magic 0x64b4024a> <pcomp> <accomp>] sent [LCP ConfReq id=0x1 <asyncmap 0x0> <magic 0x64b4024a> <pcomp> <accomp>] sent [LCP ConfReq id=0x1 <asyncmap 0x0> <magic 0x64b4024a> <pcomp> <accomp>] sent [LCP ConfReq id=0x1 <asyncmap 0x0> <magic 0x64b4024a> <pcomp> <accomp>] sent [LCP ConfReq id=0x1 <asyncmap 0x0> <magic 0x64b4024a> <pcomp> <accomp>] NetworkManager[507]: <warn> pppd timed out or didn't initialize our dbus module NetworkManager[507]: <info> Activation (ttyUSB2) Stage 4 of 5 (IPv4 Configure Timeout) scheduled... NetworkManager[507]: <info> Activation (ttyUSB2) Stage 4 of 5 (IPv4 Configure Timeout) started... NetworkManager[507]: <info> (ttyUSB2): device state change: ip-config -> failed (reason 'ip-config-unavailable') [70 120 5] NetworkManager[507]: <warn> Activation (ttyUSB2) failed for connection 'Plus - Dostep standardowy' NetworkManager[507]: <info> Activation (ttyUSB2) Stage 4 of 5 (IPv4 Configure Timeout) complete. NetworkManager[507]: <info> (ttyUSB2): device state change: failed -> disconnected (reason 'none') [120 30 0] NetworkManager[507]: <info> (ttyUSB2): deactivating device (reason 'none') [0] Terminating on signal 15 ** Message: nm-ppp-plugin: (nm_phasechange): status 10 / phase 'terminate' sent [LCP TermReq id=0x2 "User request"] NetworkManager[507]: SCPlugin-Ifupdown: devices removed (path: /sys/devices/virtual/net/ppp0, iface: ppp0) where repeated: sent [LCP ConfReq id=0x1 <asyncmap 0x0> <magic 0x64b4024a> <pcomp> <accomp>] last for about 20 seconds. I've tried to downgrade network manager but failed due to many dependencies. Can anyone point me to solution or tell what should I do to further investigate the problem?

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How to solve JAVA menubar, layout and panel problem ?

- by Berkay

i'm not a java guy however i start implementing some security tools with java, in these days i'm creating a gui for my security tools : Here is the how my Gui looks: menuBar = new JMenuBar(); // construct menu bar // hash functions fileMenu = new JMenu("HASH FUNCTIONS"); // define file menu fileMenu.setMnemonic('H'); // shortcut hashes = new JMenuItem("Md5&Sha1"); // define file menu options hashes.setMnemonic('M'); hashes.addActionListener(this); fileMenu.add(hashes); menuBar.add(fileMenu); // add file menu to menu bar // symmetric encryption asMenu = new JMenu("SYMMETRIC ENCRYPTION"); // define format menu asMenu.setMnemonic('S'); // shortcut desItem = new JMenuItem("DES"); // define format menu options desItem.setMnemonic('D'); desItem.addActionListener(this); asMenu.add(desItem); ... ... menuBar.add(helpMenu); // add help menu to menu bar setJMenuBar(menuBar); // put menu bar on application textColor = Color.RED; when from the Menu desitem is selected, desvar is just for not showing the panel multiple times, it calls Panels () else if(e.getSource() == desItem && desvar ==1 ) { // make other panels unvisible. if(hashvar!=1) MyPanel.setVisible(false);//hash functions if(rsavar!=1) MyPanel3.setVisible(false);//rsa function if (dhvar!=1) MyPanel4.setVisible(false);//dh diffie hellman hashvar=1; rsavar=1; dhvar=1; ++desvar; desPanel=true; Panels(); } and in Panels() Method: if (hashPanel){ MyPanel.add("West",radioSHA1); MyPanel.add("West",radioMD5); MyPanel.add("Center", inputField); MyPanel.add("East",SubmitButton); MyPanel.add("South",resultHash); add(MyPanel); validate(); hashPanel=false; } i have many panels for example : -hash functions=mypanel1 , des=mypanel2, rsa functions= mypanel3 , dh= mypanel4 However in may other panals such as rsa function: i have to use some layout properties of the java:in this panel i selected to use gridbaglayout if (dhPanel){ System.out.println("rsa panel burda misin"); MyPanel3.add(pqLabel); MyPanel3.add(pLabel); MyPanel3.add(pTextArea); MyPanel3.add(qLabel); ... ... add(MyPanel3); generate_pqButton.addActionListener(this); calculate_nButton.addActionListener(this); ... ... GridBagLayout layout = new GridBagLayout(); GridBagConstraints gbc = new GridBagConstraints(); setLayout(layout); // x, y, w, h, wx, wy gbc.fill = GridBagConstraints.NONE; add (bit_length_label, gbc, 0, 0, 1, 1, 0, 10); add (p, gbc, 0, 1, 1, 1, 0, 10); add (g, gbc, 0, 2, 1, 1, 0, 10); add (a, gbc, 1, 3, 1, 1, 100, 10); add (x, gbc, 0, 4, 1, 1, 0, 10); add (gx, gbc, 0, 5, 1, 1, 0, 10); add (gxy, gbc, 0, 6, 1, 1, 0, 10); add (b, gbc, 1, 7, 1, 1, 100, 10); add (y, gbc, 0, 8, 1, 1, 0, 10); add (gy, gbc, 0, 9, 1, 1, 0, 10); add (gyx, gbc, 0, 10, 1, 1, 0, 10); add (sk, gbc, 1, 11, 1, 1, 100, 10); add (key, gbc, 0, 12, 1, 1, 0, 10); add (status, gbc, 1, 13, 1, 1, 100, 10); add (dhstart, gbc, 1, 14, 1, 1, 100, 10); add (bit_length_value, gbc, 1, 0, 1, 1, 100, 10); add (p_value, gbc, 1, 1, 1, 1, 100, 10); add (g_value, gbc, 1, 2, 1, 1, 100, 10); add (x_value, gbc, 1, 4, 1, 1, 100, 10); add (gx_value, gbc, 1, 5, 1, 1, 100, 10); add (gxy_value, gbc, 1, 6, 1, 1, 100, 10); add (y_value, gbc, 1, 8, 1, 1, 100, 10); add (gy_value, gbc, 1, 9, 1, 1, 100, 10); add (gyx_value, gbc, 1, 10, 1, 1, 100, 10); validate(); repaint(); rsaPanel=false; } Everthing seems okey however when i swith from one menu to another sometimes components are seen or apper in wrong places or mixed. where i'm doing wrong?

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Prefer extension methods for encapsulation and reusability?

- by tzaman

edit4: wikified, since this seems to have morphed more into a discussion than a specific question. In C++ programming, it's generally considered good practice to "prefer non-member non-friend functions" instead of instance methods. This has been recommended by Scott Meyers in this classic Dr. Dobbs article, and repeated by Herb Sutter and Andrei Alexandrescu in C++ Coding Standards (item 44); the general argument being that if a function can do its job solely by relying on the public interface exposed by the class, it actually increases encapsulation to have it be external. While this confuses the "packaging" of the class to some extent, the benefits are generally considered worth it. Now, ever since I've started programming in C#, I've had a feeling that here is the ultimate expression of the concept that they're trying to achieve with "non-member, non-friend functions that are part of a class interface". C# adds two crucial components to the mix - the first being interfaces, and the second extension methods: Interfaces allow a class to formally specify their public contract, the methods and properties that they're exposing to the world. Any other class can choose to implement the same interface and fulfill that same contract. Extension methods can be defined on an interface, providing any functionality that can be implemented via the interface to all implementers automatically. And best of all, because of the "instance syntax" sugar and IDE support, they can be called the same way as any other instance method, eliminating the cognitive overhead! So you get the encapsulation benefits of "non-member, non-friend" functions with the convenience of members. Seems like the best of both worlds to me; the .NET library itself providing a shining example in LINQ. However, everywhere I look I see people warning against extension method overuse; even the MSDN page itself states: In general, we recommend that you implement extension methods sparingly and only when you have to. (edit: Even in the current .NET library, I can see places where it would've been useful to have extensions instead of instance methods - for example, all of the utility functions of List<T> (Sort, BinarySearch, FindIndex, etc.) would be incredibly useful if they were lifted up to IList<T> - getting free bonus functionality like that adds a lot more benefit to implementing the interface.) So what's the verdict? Are extension methods the acme of encapsulation and code reuse, or am I just deluding myself? (edit2: In response to Tomas - while C# did start out with Java's (overly, imo) OO mentality, it seems to be embracing more multi-paradigm programming with every new release; the main thrust of this question is whether using extension methods to drive a style change (towards more generic / functional C#) is useful or worthwhile..) edit3: overridable extension methods The only real problem identified so far with this approach, is that you can't specialize extension methods if you need to. I've been thinking about the issue, and I think I've come up with a solution. Suppose I have an interface MyInterface, which I want to extend - I define my extension methods in a MyExtension static class, and pair it with another interface, call it MyExtensionOverrider. MyExtension methods are defined according to this pattern: public static int MyMethod(this MyInterface obj, int arg, bool attemptCast=true) { if (attemptCast && obj is MyExtensionOverrider) { return ((MyExtensionOverrider)obj).MyMethod(arg); } // regular implementation here } The override interface mirrors all of the methods defined in MyExtension, except without the this or attemptCast parameters: public interface MyExtensionOverrider { int MyMethod(int arg); string MyOtherMethod(); } Now, any class can implement the interface and get the default extension functionality: public class MyClass : MyInterface { ... } Anyone that wants to override it with specific implementations can additionally implement the override interface: public class MySpecializedClass : MyInterface, MyExtensionOverrider { public int MyMethod(int arg) { //specialized implementation for one method } public string MyOtherMethod() { // fallback to default for others MyExtension.MyOtherMethod(this, attemptCast: false); } } And there we go: extension methods provided on an interface, with the option of complete extensibility if needed. Fully general too, the interface itself doesn't need to know about the extension / override, and multiple extension / override pairs can be implemented without interfering with each other. I can see three problems with this approach - It's a little bit fragile - the extension methods and override interface have to be kept synchronized manually. It's a little bit ugly - implementing the override interface involves boilerplate for every function you don't want to specialize. It's a little bit slow - there's an extra bool comparison and cast attempt added to the mainline of every method. Still, all those notwithstanding, I think this is the best we can get until there's language support for interface functions. Thoughts?

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Cloud Computing = Elasticity * Availability

- by Herve Roggero

What is cloud computing? Is hosting the same thing as cloud computing? Are you running a cloud if you already use virtual machines? What is the difference between Infrastructure as a Service (IaaS) and a cloud provider? And the list goes on… these questions keep coming up and all try to fundamentally explain what “cloud” means relative to other concepts. At the risk of over simplification, answering these questions becomes simpler once you understand the primary foundations of cloud computing: Elasticity and Availability. Elasticity The basic value proposition of cloud computing is to pay as you go, and to pay for what you use. This implies that an application can expand and contract on demand, across all its tiers (presentation layer, services, database, security…). This also implies that application components can grow independently from each other. So if you need more storage for your database, you should be able to grow that tier without affecting, reconfiguring or changing the other tiers. Basically, cloud applications behave like a sponge; when you add water to a sponge, it grows in size; in the application world, the more customers you add, the more it grows. Pure IaaS providers will provide certain benefits, specifically in terms of operating costs, but an IaaS provider will not help you in making your applications elastic; neither will Virtual Machines. The smallest elasticity unit of an IaaS provider and a Virtual Machine environment is a server (physical or virtual). While adding servers in a datacenter helps in achieving scale, it is hardly enough. The application has yet to use this hardware. If the process of adding computing resources is not transparent to the application, the application is not elastic. As you can see from the above description, designing for the cloud is not about more servers; it is about designing an application for elasticity regardless of the underlying server farm. Availability The fact of the matter is that making applications highly available is hard. It requires highly specialized tools and trained staff. On top of it, it's expensive. Many companies are required to run multiple data centers due to high availability requirements. In some organizations, some data centers are simply on standby, waiting to be used in a case of a failover. Other organizations are able to achieve a certain level of success with active/active data centers, in which all available data centers serve incoming user requests. While achieving high availability for services is relatively simple, establishing a highly available database farm is far more complex. In fact it is so complex that many companies establish yearly tests to validate failover procedures. To a certain degree certain IaaS provides can assist with complex disaster recovery planning and setting up data centers that can achieve successful failover. However the burden is still on the corporation to manage and maintain such an environment, including regular hardware and software upgrades. Cloud computing on the other hand removes most of the disaster recovery requirements by hiding many of the underlying complexities. Cloud Providers A cloud provider is an infrastructure provider offering additional tools to achieve application elasticity and availability that are not usually available on-premise. For example Microsoft Azure provides a simple configuration screen that makes it possible to run 1 or 100 web sites by clicking a button or two on a screen (simplifying provisioning), and soon SQL Azure will offer Data Federation to allow database sharding (which allows you to scale the database tier seamlessly and automatically). Other cloud providers offer certain features that are not available on-premise as well, such as the Amazon SC3 (Simple Storage Service) which gives you virtually unlimited storage capabilities for simple data stores, which is somewhat equivalent to the Microsoft Azure Table offering (offering a server-independent data storage model). Unlike IaaS providers, cloud providers give you the necessary tools to adopt elasticity as part of your application architecture. Some cloud providers offer built-in high availability that get you out of the business of configuring clustered solutions, or running multiple data centers. Some cloud providers will give you more control (which puts some of that burden back on the customers' shoulder) and others will tend to make high availability totally transparent. For example, SQL Azure provides high availability automatically which would be very difficult to achieve (and very costly) on premise. Keep in mind that each cloud provider has its strengths and weaknesses; some are better at achieving transparent scalability and server independence than others. Not for Everyone Note however that it is up to you to leverage the elasticity capabilities of a cloud provider, as discussed previously; if you build a website that does not need to scale, for which elasticity is not important, then you can use a traditional host provider unless you also need high availability. Leveraging the technologies of cloud providers can be difficult and can become a journey for companies that build their solutions in a scale up fashion. Cloud computing promises to address cost containment and scalability of applications with built-in high availability. If your application does not need to scale or you do not need high availability, then cloud computing may not be for you. In fact, you may pay a premium to run your applications with cloud providers due to the underlying technologies built specifically for scalability and availability requirements. And as such, the cloud is not for everyone. Consistent Customer Experience, Predictable Cost With all its complexities, buzz and foggy definition, cloud computing boils down to a simple objective: consistent customer experience at a predictable cost. The objective of a cloud solution is to provide the same user experience to your last customer than the first, while keeping your operating costs directly proportional to the number of customers you have. Making your applications elastic and highly available across all its tiers, with as much automation as possible, achieves the first objective of a consistent customer experience. And the ability to expand and contract the infrastructure footprint of your application dynamically achieves the cost containment objectives. Herve Roggero is a SQL Azure MVP and co-author of Pro SQL Azure (APress). He is the co-founder of Blue Syntax Consulting (www.bluesyntax.net), a company focusing on cloud computing technologies helping customers understand and adopt cloud computing technologies. For more information contact herve at hroggero @ bluesyntax.net .

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Announcing: Improvements to the Windows Azure Portal

- by ScottGu

Earlier today we released a number of enhancements to the new Windows Azure Management Portal. These new capabilities include: Service Bus Management and Monitoring Support for Managing Co-administrators Import/Export support for SQL Databases Virtual Machine Experience Enhancements Improved Cloud Service Status Notifications Media Services Monitoring Support Storage Container Creation and Access Control Support All of these improvements are now live in production and available to start using immediately. Below are more details on them: Service Bus Management and Monitoring The new Windows Azure Management Portal now supports Service Bus management and monitoring. Service Bus provides rich messaging infrastructure that can sit between applications (or between cloud and on-premise environments) and allow them to communicate in a loosely coupled way for improved scale and resiliency. With the new Service Bus experience, you can now create and manage Service Bus Namespaces, Queues, Topics, Relays and Subscriptions. You can also get rich monitoring for Service Bus Queues, Topics and Subscriptions. To create a Service Bus namespace, you can now select the “Service Bus” tab in the Windows Azure portal and then simply select the CREATE command: Doing so will bring up a new “Create a Namespace” dialog that allows you to name and create a new Service Bus Namespace: Once created, you can obtain security credentials associated with the Namespace via the ACCESS KEY command. This gives you the ability to obtain the connection string associated with the service namespace. You can copy and paste these values into any application that requires these credentials: It is also now easy to create Service Bus Queues and Topics via the NEW experience in the portal drawer. Simply click the NEW command and navigate to the “App Services” category to create a new Service Bus entity: Once you provision a new Queue or Topic it can be managed in the portal. Clicking on a namespace will display all queues and topics within it: Clicking on an item in the list will allow you to drill down into a dashboard view that allows you to monitor the activity and traffic within it, as well as perform operations on it. For example, below is a view of an “orders” queue – note how we now surface both the incoming and outgoing message flow rate, as well as the total queue length and queue size: To monitor pub/sub subscriptions you can use the ADD METRICS command within a topic and select a specific subscription to monitor. Support for Managing Co-Administrators You can now add co-administrators for your Windows Azure subscription using the new Windows Azure Portal. This allows you to share management of your Windows Azure services with other users. Subscription co-administrators share the same administrative rights and permissions that service administrator have - except a co-administrator cannot change or view billing details about the account, nor remove the service administrator from a subscription. In the SETTINGS section, click on the ADMINISTRATORS tab, and select the ADD button to add a co-administrator to your subscription: To add a co-administrator, you specify the email address for a Microsoft account (formerly Windows Live ID) or an organizational account, and choose the subscription you want to add them to: You can later update the subscriptions that the co-administrator has access to by clicking on the EDIT button, and then select or deselect the subscriptions to which they belong. Import/Export Support for SQL Databases The Windows Azure administration portal now supports importing and exporting SQL Databases to/from Blob Storage. Databases can be imported/exported to blob storage using the same BACPAC file format that is supported with SQL Server 2012. Among other benefits, this makes it easy to copy and migrate databases between on-premise and cloud environments. SQL Databases now have an EXPORT command in the bottom drawer that when pressed will prompt you to save your database to a Windows Azure storage container: The UI allows you to choose an existing storage account or create a new one, as well as the name of the BACPAC file to persist in blob storage: You can also now import and create a new SQL Database by using the NEW command. This will prompt you to select the storage container and file to import the database from: The Windows Azure Portal enables you to monitor the progress of import and export operations. If you choose to log out of the portal, you can come back later and check on the status of all of the operations in the new history tab of the SQL Database server – this shows your entire import and export history and the status (success/fail) of each: Enhancements to the Virtual Machine Experience One of the common pain-points we have heard from customers using the preview of our new Virtual Machine support has been the inability to delete the associated VHDs when a VM instance (or VM drive) gets deleted. Prior to today’s release the VHDs would continue to be in your storage account and accumulate storage charges. You can now navigate to the Disks tab within the Virtual Machine extension, select a VM disk to delete, and click the DELETE DISK command: When you click the DELETE DISK button you have the option to delete the disk + associated .VHD file (completely clearing it from storage). Alternatively you can delete the disk but still retain a .VHD copy of it in storage. Improved Cloud Service Status Notifications The Windows Azure portal now exposes more information of the health status of role instances. If any of the instances are in a non-running state, the status at the top of the dashboard will summarize the status (and update automatically as the role health changes): Clicking the instance hyperlink within this status summary view will navigate you to a detailed role instance view, and allow you to get more detailed health status of each of the instances. The portal has been updated to provide more specific status information within this detailed view – giving you better visibility into the health of your app: Monitoring Support for Media Services Windows Azure Media Services allows you to create media processing jobs (for example: encoding media files) in your Windows Azure Media Services account. In the Windows Azure Portal, you can now monitor the number of encoding jobs that are queued up for processing as well as active, failed and queued tasks for encoding jobs. On your media services account dashboard, you can visualize the monitoring data for last 6 hours, 24 hours or 7 days. Storage Container Creation and Access Control Support You can now create Windows Azure Storage storage containers from within the Windows Azure Portal. After selecting a storage account, you can navigate to the CONTAINERS tab and click the ADD CONTAINER command: This will display a dialog that lets you name the new container and control access to it: You can also update the access setting as well as container metadata of existing containers by selecting one and then using the new EDIT CONTAINER command: This will then bring up the edit container dialog that allows you to change and save its settings: In addition to creating and editing containers, you can click on them within the portal to drill-in and view blobs within them. Summary The above features are all now live in production and available to use immediately. If you don’t already have a Windows Azure account, you can sign-up for a free trial and start using them today. Visit the Windows Azure Developer Center to learn more about how to build apps with it. We’ll have even more new features and enhancements coming later this month – including support for the recent Windows Server 2012 and .NET 4.5 releases (we will enable new web and worker role images with Windows Server 2012 and .NET 4.5, and support .NET 4.5 with Websites). Keep an eye out on my blog for details as these new features become available. Hope this helps, Scott P.S. In addition to blogging, I am also now using Twitter for quick updates and to share links. Follow me at: twitter.com/scottgu

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RiverTrail - JavaScript GPPGU Data Parallelism

- by JoshReuben

Where is WebCL ? The Khronos WebCL working group is working on a JavaScript binding to the OpenCL standard so that HTML 5 compliant browsers can host GPGPU web apps – e.g. for image processing or physics for WebGL games - http://www.khronos.org/webcl/ . While Nokia & Samsung have some protype WebCL APIs, Intel has one-upped them with a higher level of abstraction: RiverTrail. Intro to RiverTrail Intel Labs JavaScript RiverTrail provides GPU accelerated SIMD data-parallelism in web applications via a familiar JavaScript programming paradigm. It extends JavaScript with simple deterministic data-parallel constructs that are translated at runtime into a low-level hardware abstraction layer. With its high-level JS API, programmers do not have to learn a new language or explicitly manage threads, orchestrate shared data synchronization or scheduling. It has been proposed as a draft specification to ECMA a (known as ECMA strawman). RiverTrail runs in all popular browsers (except I.E. of course). To get started, download a prebuilt version https://github.com/downloads/RiverTrail/RiverTrail/rivertrail-0.17.xpi , install Intel's OpenCL SDK http://www.intel.com/go/opencl and try out the interactive River Trail shell http://rivertrail.github.com/interactive For a video overview, see http://www.youtube.com/watch?v=jueg6zB5XaM . ParallelArray the ParallelArray type is the central component of this API & is a JS object that contains ordered collections of scalars – i.e. multidimensional uniform arrays. A shape property describes the dimensionality and size– e.g. a 2D RGBA image will have shape [height, width, 4]. ParallelArrays are immutable & fluent – they are manipulated by invoking methods on them which produce new ParallelArray objects. ParallelArray supports several constructors over arrays, functions & even the canvas. // Create an empty Parallel Array var pa = new ParallelArray(); // pa0 = <> // Create a ParallelArray out of a nested JS array. // Note that the inner arrays are also ParallelArrays var pa = new ParallelArray([ [0,1], [2,3], [4,5] ]); // pa1 = <<0,1>, <2,3>, <4.5>> // Create a two-dimensional ParallelArray with shape [3, 2] using the comprehension constructor var pa = new ParallelArray([3, 2], function(iv){return iv[0] * iv[1];}); // pa7 = <<0,0>, <0,1>, <0,2>> // Create a ParallelArray from canvas. This creates a PA with shape [w, h, 4], var pa = new ParallelArray(canvas); // pa8 = CanvasPixelArray ParallelArray exposes fluent API functions that take an elemental JS function for data manipulation: map, combine, scan, filter, and scatter that return a new ParallelArray. Other functions are scalar - reduce returns a scalar value & get returns the value located at a given index. The onus is on the developer to ensure that the elemental function does not defeat data parallelization optimization (avoid global var manipulation, recursion). For reduce & scan, order is not guaranteed - the onus is on the dev to provide an elemental function that is commutative and associative so that scan will be deterministic – E.g. Sum is associative, but Avg is not. map Applies a provided elemental function to each element of the source array and stores the result in the corresponding position in the result array. The map method is shape preserving & index free - can not inspect neighboring values. // Adding one to each element. var source = new ParallelArray([1,2,3,4,5]); var plusOne = source.map(function inc(v) { return v+1; }); //<2,3,4,5,6> combine Combine is similar to map, except an index is provided. This allows elemental functions to access elements from the source array relative to the one at the current index position. While the map method operates on the outermost dimension only, combine, can choose how deep to traverse - it provides a depth argument to specify the number of dimensions it iterates over. The elemental function of combine accesses the source array & the current index within it - element is computed by calling the get method of the source ParallelArray object with index i as argument. It requires more code but is more expressive. var source = new ParallelArray([1,2,3,4,5]); var plusOne = source.combine(function inc(i) { return this.get(i)+1; }); reduce reduces the elements from an array to a single scalar result – e.g. Sum. // Calculate the sum of the elements var source = new ParallelArray([1,2,3,4,5]); var sum = source.reduce(function plus(a,b) { return a+b; }); scan Like reduce, but stores the intermediate results – return a ParallelArray whose ith elements is the results of using the elemental function to reduce the elements between 0 and I in the original ParallelArray. // do a partial sum var source = new ParallelArray([1,2,3,4,5]); var psum = source.scan(function plus(a,b) { return a+b; }); //<1, 3, 6, 10, 15> scatter a reordering function - specify for a certain source index where it should be stored in the result array. An optional conflict function can prevent an exception if two source values are assigned the same position of the result: var source = new ParallelArray([1,2,3,4,5]); var reorder = source.scatter([4,0,3,1,2]); // <2, 4, 5, 3, 1> // if there is a conflict use the max. use 33 as a default value. var reorder = source.scatter([4,0,3,4,2], 33, function max(a, b) {return a>b?a:b; }); //<2, 33, 5, 3, 4> filter // filter out values that are not even var source = new ParallelArray([1,2,3,4,5]); var even = source.filter(function even(iv) { return (this.get(iv) % 2) == 0; }); // <2,4> Flatten used to collapse the outer dimensions of an array into a single dimension. pa = new ParallelArray([ [1,2], [3,4] ]); // <<1,2>,<3,4>> pa.flatten(); // <1,2,3,4> Partition used to restore the original shape of the array. var pa = new ParallelArray([1,2,3,4]); // <1,2,3,4> pa.partition(2); // <<1,2>,<3,4>> Get return value found at the indices or undefined if no such value exists. var pa = new ParallelArray([0,1,2,3,4], [10,11,12,13,14], [20,21,22,23,24]) pa.get([1,1]); // 11 pa.get([1]); // <10,11,12,13,14>

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How to implement Cocoa copyWithZone on derived object in MonoMac C#?

- by Justin Aquadro

I'm currently porting a small Winforms-based .NET application to use a native Mac front-end with MonoMac. The application has a TreeControl with icons and text, which does not exist out of the box in Cocoa. So far, I've ported almost all of the ImageAndTextCell code in Apple's DragNDrop example: https://developer.apple.com/library/mac/#samplecode/DragNDropOutlineView/Listings/ImageAndTextCell_m.html#//apple_ref/doc/uid/DTS40008831-ImageAndTextCell_m-DontLinkElementID_6, which is assigned to an NSOutlineView as a custom cell. It seems to be working almost perfectly, except that I have not figured out how to properly port the copyWithZone method. Unfortunately, this means the internal copies that NSOutlineView is making do not have the image field, and it leads to the images briefly vanishing during expand and collapse operations. The objective-c code in question is: - (id)copyWithZone:(NSZone *)zone { ImageAndTextCell *cell = (ImageAndTextCell *)[super copyWithZone:zone]; // The image ivar will be directly copied; we need to retain or copy it. cell->image = [image retain]; return cell; } The first line is what's tripping me up, as MonoMac does not expose a copyWithZone method, and I don't know how to otherwise call it. Update Based on current answers and additional research and testing, I've come up with a variety of models for copying an object. static List<ImageAndTextCell> _refPool = new List<ImageAndTextCell>(); // Method 1 static IntPtr selRetain = Selector.GetHandle ("retain"); [Export("copyWithZone:")] public virtual NSObject CopyWithZone(IntPtr zone) { ImageAndTextCell cell = new ImageAndTextCell() { Title = Title, Image = Image, }; Messaging.void_objc_msgSend (cell.Handle, selRetain); return cell; } // Method 2 [Export("copyWithZone:")] public virtual NSObject CopyWithZone(IntPtr zone) { ImageAndTextCell cell = new ImageAndTextCell() { Title = Title, Image = Image, }; _refPool.Add(cell); return cell; } [Export("dealloc")] public void Dealloc () { _refPool.Remove(this); this.Dispose(); } // Method 3 static IntPtr selRetain = Selector.GetHandle ("retain"); [Export("copyWithZone:")] public virtual NSObject CopyWithZone(IntPtr zone) { ImageAndTextCell cell = new ImageAndTextCell() { Title = Title, Image = Image, }; _refPool.Add(cell); Messaging.void_objc_msgSend (cell.Handle, selRetain); return cell; } // Method 4 static IntPtr selRetain = Selector.GetHandle ("retain"); static IntPtr selRetainCount = Selector.GetHandle("retainCount"); [Export("copyWithZone:")] public virtual NSObject CopyWithZone (IntPtr zone) { ImageAndTextCell cell = new ImageAndTextCell () { Title = Title, Image = Image, }; _refPool.Add (cell); Messaging.void_objc_msgSend (cell.Handle, selRetain); return cell; } public void PeriodicCleanup () { List<ImageAndTextCell> markedForDelete = new List<ImageAndTextCell> (); foreach (ImageAndTextCell cell in _refPool) { uint count = Messaging.UInt32_objc_msgSend (cell.Handle, selRetainCount); if (count == 1) markedForDelete.Add (cell); } foreach (ImageAndTextCell cell in markedForDelete) { _refPool.Remove (cell); cell.Dispose (); } } // Method 5 static IntPtr selCopyWithZone = Selector.GetHandle("copyWithZone:"); [Export("copyWithZone:")] public virtual NSObject CopyWithZone(IntPtr zone) { IntPtr copyHandle = Messaging.IntPtr_objc_msgSendSuper_IntPtr(SuperHandle, selCopyWithZone, zone); ImageAndTextCell cell = new ImageAndTextCell(copyHandle) { Image = Image, }; _refPool.Add(cell); return cell; } Method 1: Increases the retain count of the unmanaged object. The unmanaged object will persist persist forever (I think? dealloc never called), and the managed object will be harvested early. Seems to be lose-lose all-around, but runs in practice. Method 2: Saves a reference of the managed object. The unmanaged object is left alone, and dealloc appears to be invoked at a reasonable time by the caller. At this point the managed object is released and disposed. This seems reasonable, but on the downside the base type's dealloc won't be run (I think?) Method 3: Increases the retain count and saves a reference. Unmanaged and managed objects leak forever. Method 4: Extends Method 3 by adding a cleanup function that is run periodically (e.g. during Init of each new ImageAndTextCell object). The cleanup function checks the retain counts of the stored objects. A retain count of 1 means the caller has released it, so we should as well. Should eliminate leaking in theory. Method 5: Attempt to invoke the copyWithZone method on the base type, and then construct a new ImageAndTextView object with the resulting handle. Seems to do the right thing (the base data is cloned). Internally, NSObject bumps the retain count on objects constructed like this, so we also use the PeriodicCleanup function to release these objects when they're no longer used. Based on the above, I believe Method 5 is the best approach since it should be the only one that results in a truly correct copy of the base type data, but I don't know if the approach is inherently dangerous (I am also making some assumptions about the underlying implementation of NSObject). So far nothing bad has happened "yet", but if anyone is able to vet my analysis then I would be more confident going forward.

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Windows Azure – Write, Run or Use Software

- by BuckWoody

Windows Azure is a platform that has you covered, whether you need to write software, run software that is already written, or Install and use “canned” software whether you or someone else wrote it. Like any platform, it’s a set of tools you can use where it makes sense to solve a problem. The primary location for Windows Azure information is located at http://windowsazure.com. You can find everything there from the development kits for writing software to pricing, licensing and tutorials on all of that. I have a few links here for learning to use Windows Azure – although it’s best if you focus not on the tools, but what you want to solve. I’ve got it broken down here into various sections, so you can quickly locate things you want to know. I’ll include resources here from Microsoft and elsewhere – I use these same resources in the Architectural Design Sessions (ADS) I do with my clients worldwide. Write Software Also called “Platform as a Service” (PaaS), Windows Azure has lots of components you can use together or separately that allow you to write software in .NET or various Open Source languages to work completely online, or in partnership with code you have on-premises or both – even if you’re using other cloud providers. Keep in mind that all of the features you see here can be used together, or independently. For instance, you might only use a Web Site, or use Storage, but you can use both together. You can access all of these components through standard REST API calls, or using our Software Development Kit’s API’s, which are a lot easier. In any case, you simply use Visual Studio, Eclipse, Cloud9 IDE, or even a text editor to write your code from a Mac, PC or Linux. Components you can use: Azure Web Sites: Windows Azure Web Sites allow you to quickly write an deploy websites, without setting a Virtual Machine, installing a web server or configuring complex settings. They work alone, with other Windows Azure Web Sites, or with other parts of Windows Azure. Web and Worker Roles: Windows Azure Web Roles give you a full stateless computing instance with Internet Information Services (IIS) installed and configured. Windows Azure Worker Roles give you a full stateless computing instance without Information Services (IIS) installed, often used in a "Services" mode. Scale-out is achieved either manually or programmatically under your control. Storage: Windows Azure Storage types include Blobs to store raw binary data, Tables to use key/value pair data (like NoSQL data structures), Queues that allow interaction between stateless roles, and a relational SQL Server database. Other Services: Windows Azure has many other services such as a security mechanism, a Cache (memcacheD compliant), a Service Bus, a Traffic Manager and more. Once again, these features can be used with a Windows Azure project, or alone based on your needs. Various Languages: Windows Azure supports the .NET stack of languages, as well as many Open-Source languages like Java, Python, PHP, Ruby, NodeJS, C++ and more. Use Software Also called “Software as a Service” (SaaS) this often means consumer or business-level software like Hotmail or Office 365. In other words, you simply log on, use the software, and log off – there’s nothing to install, and little to even configure. For the Information Technology professional, however, It’s not quite the same. We want software that provides services, but in a platform. That means we want things like Hadoop or other software we don’t want to have to install and configure. Components you can use: Kits: Various software “kits” or packages are supported with just a few clicks, such as Umbraco, Wordpress, and others. Windows Azure Media Services: Windows Azure Media Services is a suite of services that allows you to upload media for encoding, processing and even streaming – or even one or more of those functions. We can add DRM and even commercials to your media if you like. Windows Azure Media Services is used to stream large events all the way down to small training videos. High Performance Computing and “Big Data”: Windows Azure allows you to scale to huge workloads using a few clicks to deploy Hadoop Clusters or the High Performance Computing (HPC) nodes, accepting HPC Jobs, Pig and Hive Jobs, and even interfacing with Microsoft Excel. Windows Azure Marketplace: Windows Azure Marketplace offers data and programs you can quickly implement and use – some free, some for-fee. Run Software Also known as “Infrastructure as a Service” (IaaS), this offering allows you to build or simply choose a Virtual Machine to run server-based software. Components you can use: Persistent Virtual Machines: You can choose to install Windows Server, Windows Server with Active Directory, with SQL Server, or even SharePoint from a pre-configured gallery. You can configure your own server images with standard Hyper-V technology and load them yourselves – and even bring them back when you’re done. As a new offering, we also even allow you to select various distributions of Linux – a first for Microsoft. Windows Azure Connect: You can connect your on-premises networks to Windows Azure Instances. Storage: Windows Azure Storage can be used as a remote backup, a hybrid storage location and more using software or even hardware appliances. Decision Matrix With all of these options, you can use Windows Azure to solve just about any computing problem. It’s often hard to know when to use something on-premises, in the cloud, and what kind of service to use. I’ve used a decision matrix in the last couple of years to take a particular problem and choose the proper technology to solve it. It’s all about options – there is no “silver bullet”, whether that’s Windows Azure or any other set of functions. I take the problem, decide which particular component I want to own and control – and choose the column that has that box darkened. For instance, if I have to control the wiring for a solution (a requirement in some military and government installations), that means the “Networking” component needs to be dark, and so I select the “On Premises” column for that particular solution. If I just need the solution provided and I want no control at all, I can look as “Software as a Service” solutions. Security, Pricing, and Other Info Security: Security is one of the first questions you should ask in any distributed computing environment. We have certification info, coding guidelines and more, even a general “Request for Information” RFI Response already created for you. Pricing: Are there licenses? How much does this cost? Is there a way to estimate the costs in this new environment? New Features: Many new features were added to Windows Azure - a good roundup of those changes can be found here. Support: Software Support on Virtual Machines, general support.

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Dynamically register constructor methods in an AbstractFactory at compile time using C++ templates

- by Horacio

When implementing a MessageFactory class to instatiate Message objects I used something like: class MessageFactory { public: static Message *create(int type) { switch(type) { case PING_MSG: return new PingMessage(); case PONG_MSG: return new PongMessage(); .... } } This works ok but every time I add a new message I have to add a new XXX_MSG and modify the switch statement. After some research I found a way to dynamically update the MessageFactory at compile time so I can add as many messages as I want without need to modify the MessageFactory itself. This allows for cleaner and easier to maintain code as I do not need to modify three different places to add/remove message classes: #include <stdio.h> #include <stdlib.h> #include <string.h> #include <inttypes.h> class Message { protected: inline Message() {}; public: inline virtual ~Message() { } inline int getMessageType() const { return m_type; } virtual void say() = 0; protected: uint16_t m_type; }; template<int TYPE, typename IMPL> class MessageTmpl: public Message { enum { _MESSAGE_ID = TYPE }; public: static Message* Create() { return new IMPL(); } static const uint16_t MESSAGE_ID; // for registration protected: MessageTmpl() { m_type = MESSAGE_ID; } //use parameter to instanciate template }; typedef Message* (*t_pfFactory)(); class MessageFactory· { public: static uint16_t Register(uint16_t msgid, t_pfFactory factoryMethod) { printf("Registering constructor for msg id %d\n", msgid); m_List[msgid] = factoryMethod; return msgid; } static Message *Create(uint16_t msgid) { return m_List[msgid](); } static t_pfFactory m_List[65536]; }; template <int TYPE, typename IMPL> const uint16_t MessageTmpl<TYPE, IMPL >::MESSAGE_ID = MessageFactory::Register( MessageTmpl<TYPE, IMPL >::_MESSAGE_ID, &MessageTmpl<TYPE, IMPL >::Create); class PingMessage: public MessageTmpl < 10, PingMessage > {· public: PingMessage() {} virtual void say() { printf("Ping\n"); } }; class PongMessage: public MessageTmpl < 11, PongMessage > {· public: PongMessage() {} virtual void say() { printf("Pong\n"); } }; t_pfFactory MessageFactory::m_List[65536]; int main(int argc, char **argv) { Message *msg1; Message *msg2; msg1 = MessageFactory::Create(10); msg1->say(); msg2 = MessageFactory::Create(11); msg2->say(); delete msg1; delete msg2; return 0; } The template here does the magic by registering into the MessageFactory class, all new Message classes (e.g. PingMessage and PongMessage) that subclass from MessageTmpl. This works great and simplifies code maintenance but I still have some questions about this technique: Is this a known technique/pattern? what is the name? I want to search more info about it. I want to make the array for storing new constructors MessageFactory::m_List[65536] a std::map but doing so causes the program to segfault even before reaching main(). Creating an array of 65536 elements is overkill but I have not found a way to make this a dynamic container. For all message classes that are subclasses of MessageTmpl I have to implement the constructor. If not it won't register in the MessageFactory. For example commenting the constructor of the PongMessage: class PongMessage: public MessageTmpl < 11, PongMessage > { public: //PongMessage() {} /* HERE */ virtual void say() { printf("Pong\n"); } }; would result in the PongMessage class not being registered by the MessageFactory and the program would segfault in the MessageFactory::Create(11) line. The question is why the class won't register? Having to add the empty implementation of the 100+ messages I need feels inefficient and unnecessary.

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compile time if && return string reference optimization

- by Truncheon

Hi. I'm writing a series classes that inherit from a base class using virtual. They are INT, FLOAT and STRING objects that I want to use in a scripting language. I'm trying to implement weak typing, but I don't want STRING objects to return copies of themselves when used in the following way (instead I would prefer to have a reference returned which can be used in copying): a = "hello "; b = "world"; c = a + b; I have written the following code as a mock example: #include <iostream> #include <string> #include <cstdio> #include <cstdlib> std::string dummy("<int object cannot return string reference>"); struct BaseImpl { virtual bool is_string() = 0; virtual int get_int() = 0; virtual std::string get_string_copy() = 0; virtual std::string const& get_string_ref() = 0; }; struct INT : BaseImpl { int value; INT(int i = 0) : value(i) { std::cout << "constructor called\n"; } INT(BaseImpl& that) : value(that.get_int()) { std::cout << "copy constructor called\n"; } bool is_string() { return false; } int get_int() { return value; } std::string get_string_copy() { char buf[33]; sprintf(buf, "%i", value); return buf; } std::string const& get_string_ref() { return dummy; } }; struct STRING : BaseImpl { std::string value; STRING(std::string s = "") : value(s) { std::cout << "constructor called\n"; } STRING(BaseImpl& that) { if (that.is_string()) value = that.get_string_ref(); else value = that.get_string_copy(); std::cout << "copy constructor called\n"; } bool is_string() { return true; } int get_int() { return atoi(value.c_str()); } std::string get_string_copy() { return value; } std::string const& get_string_ref() { return value; } }; struct Base { BaseImpl* impl; Base(BaseImpl* p = 0) : impl(p) {} ~Base() { delete impl; } }; int main() { Base b1(new INT(1)); Base b2(new STRING("Hello world")); Base b3(new INT(*b1.impl)); Base b4(new STRING(*b2.impl)); std::cout << "\n"; std::cout << b1.impl->get_int() << "\n"; std::cout << b2.impl->get_int() << "\n"; std::cout << b3.impl->get_int() << "\n"; std::cout << b4.impl->get_int() << "\n"; std::cout << "\n"; std::cout << b1.impl->get_string_ref() << "\n"; std::cout << b2.impl->get_string_ref() << "\n"; std::cout << b3.impl->get_string_ref() << "\n"; std::cout << b4.impl->get_string_ref() << "\n"; std::cout << "\n"; std::cout << b1.impl->get_string_copy() << "\n"; std::cout << b2.impl->get_string_copy() << "\n"; std::cout << b3.impl->get_string_copy() << "\n"; std::cout << b4.impl->get_string_copy() << "\n"; return 0; } It was necessary to add an if check in the STRING class to determine whether its safe to request a reference instead of a copy: Script code: a = "test"; b = a; c = 1; d = "" + c; /* not safe to request reference by standard */ C++ code: STRING(BaseImpl& that) { if (that.is_string()) value = that.get_string_ref(); else value = that.get_string_copy(); std::cout << "copy constructor called\n"; } If was hoping there's a way of moving that if check into compile time, rather than run time.

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Who could ask for more with LESS CSS? (Part 2 of 3–Setup)

- by ToStringTheory

Welcome to part two in my series covering the LESS CSS language. In the first post, I covered the two major CSS precompiled languages - LESS and SASS to a small extent, iterating over some of the features that you could expect to find in them. In this post, I will go a little further in depth into the setup and execution of using the LESS framework. Introduction It really doesn’t take too much to get LESS working in your project. The basic workflow will be including the necessary translator in your project, defining bundles for the LESS files, add the necessary code to your layouts.cshtml file, and finally add in all your necessary styles to the LESS files! Lets get started… New Project Just like all great experiments in Visual Studio, start up a File > New Project, and create a new MVC 4 Web Application. The Base Package After you have the new project spun up, use the Nuget Package Manager to install the Bundle Transformer: LESS package. This will take care of installing the main translator that we will be using for LESS code (dotless which is another Nuget package), as well as the core framework for the Bundle Transformer library. The installation will come up with some instructions in a readme file on how to modify your web.config to handle all your *.less requests through the Bundle Transformer, which passes the translating onto dotless. Where To Put These LESS Files?! This step isn’t really a requirement, however I find that I don’t like how ASP.Net MVC just has a content directory where they store CSS, content images, css images…. In my project, I went ahead and created a new directory just for styles – LESS files, CSS files, and images that are only referenced in LESS or CSS. Ignore the MVC directory as this was my testbed for another project I was working on at the same time. As you can see here, I have: A top level directory for images which contains only images used in a page A top level directory for scripts A top level directory for Styles A few directories for plugins I am using (Colrizr, JQueryUI, Farbtastic) Multiple *.less files for different functions (I’ll go over these in a minute) I find that this layout offers the best separation of content types. Bring Out Your Bundles! The next thing that we need to do is add in the necessary code for the bundling of these LESS files. Go ahead and open your BundleConfig.cs file, usually located in the /App_Start/ folder of the project. As you will see in a minute, instead of using the method Microsoft does in the base MVC 4 project, I change things up a bit. Define Constants The first thing I do is define constants for each of the virtual paths that will be used in the bundler: The main reason is that I hate magic strings in my program, so the fact that you first defined a virtual path in the BundleConfig file, and then used that path in the _Layout.cshtml file really irked me. Add Bundles to the BundleCollection Next, I am going to define the bundles for my styles in my AddStyleBundles method: That is all it takes to get all of my styles in play with LESS. The CssTransformer and NullOrderer types come from the Bundle Transformer we grabbed earlier. If we didn’t use that package, we would have to write our own function (not too hard, but why do it if it’s been done). I use the site.less file as my main hub for LESS - I will cover that more in the next section. Add Bundles To Layout.cshtml File With the constants in the BundleConfig file, instead of having to use the same magic string I defined for the bundle virtual path, I am able to do this: Notice here that besides the RenderSection magic strings (something I am working on in another side project), all of the bundles are now based on const strings. If I need to change the virtual path, I only have to do it in one place. Nifty! Get Started! We are now ready to roll! As I said in the previous section, I use the site.less file as a central hub for my styles: As seen here, I have a reset.css file which is a simple CSS reset. Next, I have created a file for managing all my color variables – colors.less: Here, you can see some of the standards I started to use, in this case for color variables. I define all color variables with the @col prefix. Currently, I am going for verbose variable names. The next file imported is my font.less file that defines the typeface information for the site: Simple enough. A couple of imports for fonts from Google, and then declaring variables for use throughout LESS. I also set up the heading sizes, margins, etc.. You can also see my current standardization for font declaration strings – @font. Next, I pull in a mixins.less file that I grabbed from the Twitter Bootstrap library that gives some useful parameterized mixins for use such as border-radius, gradient, box-shadow, etc… The common.less file is a file that just contains items that I will be defining that can be used across all my LESS files. Kind of like my own mixins or font-helpers: Finally I have my layout.less file that contains all of my definitions for general site layout – width, main/sidebar widths, footer layout, etc: That’s it! For the rest of my one off definitions/corrections, I am currently putting them into the site.less file beneath my original imports Note Probably my favorite side effect of using the LESS handler/translator while bundling is that it also does a CSS checkup when rendering… See, when your web.config is set to debug, bundling will output the url to the direct less file, not the bundle, and the http handler intercepts the call, compiles the less, and returns the result. If there is an error in your LESS code, the CSS file can be returned empty, or may have the error output as a comment on the first couple lines. If you have the web.config set to not debug, then if there is an error in your code, you will end up with the usual ASP.Net exception page (unless you catch the exception of course), with information regarding the failure of the conversion, such as brace mismatch, undefined variable, etc… I find it nifty. Conclusion This is really just the beginning. LESS is very powerful and exciting! My next post will show an actual working example of why LESS is so powerful with its functions and variables… At least I hope it will! As for now, if you have any questions, comments, or suggestions on my current practice, I would love to hear them! Feel free to drop a comment or shoot me an email using the contact page. In the mean time, I plan on posting the final post in this series tomorrow or the day after, with my side project, as well as a whole base ASP.Net MVC4 templated project with LESS added in it so that you can check out the layout I have in this post. Until next time…

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