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  • Using Definition of Done to Drive Agile Maturity

    - by Dylan Smith
    I’ve been an Agile Coach at a lot of different clients over the years, and I want to share an approach I use to help them adopt and mature over time. It’s important to realize that “Agile” is not a black/white yes/no thing. Teams can be varying degrees of agile. I think of this as their agile maturity level. When I coach teams I want them to start out being a little agile, and get more agile as they mature. The approach I teach them is to use the definition of done as a technique to continuously improve their agile maturity over time. We’re probably all familiar with the concept of “Done Done” that represents what *actually* being done a feature means. Not just when a developer says he’s done right after he writes that last line of code that makes the feature kind-of work. Done Done means the coding is done, it’s been tested, installers and deployment packages have been created, user manuals have been updated, architecture docs have been updated, etc. To enable teams to internalize the concept of “Done Done”, they usually get together and come up with their Definition of Done (DoD) that defines all the activities that need to be completed before a feature is considered Done Done. The Done Done technique typically is applied only to features (aka User Stories). What I do is extend this to apply to several concepts such as User Stories, Sprints, Releases (and sometimes Check-Ins). During project kick-off I’ll usually sit down with the team and go through an exercise of creating DoD’s for each of these concepts (Stories/Sprints/Releases). We’ll usually start by just brainstorming a bunch of activities that could end up in these various DoD’s. Here’s some examples: Code Reviews StyleCop FxCop User Manuals Updated Architecture Docs Updated Tested by QA Tested by UAT Installers Created Support Knowledge Base Updated Deployment Instructions (for Ops) written Automated Unit Tests Run Automated Integration Tests Run Then we start by arranging these activities into the place they occur today (e.g. Do you do UAT testing only once per release? every sprint? every feature?). If the team was previously Waterfall most of these activities probably end up in the Release DoD. An extremely mature agile team would probably have most of these activities in the DoD for the User Stories (because an extremely mature agile team will probably do continuous deployment and release every story). So what we need to do as a team, is work to move these activities from their current home (Release DoD) down into the Sprint DoD and eventually into the User Story DoD (and maybe into the lower-level Check-In DoD if we decide to use that). We don’t have to move them all down to User Story immediately, but as a team we figure out what we think we’re capable of moving down to the Sprint cycle, and Story cycle immediately, and that becomes our starting DoD’s. Over time the team makes an effort to continue moving activities down from Release->Sprint->Story as they become more agile and more mature. I try to encourage them to envision a world in which they deploy to production as each User Story is completed. They would need to be updating User Manuals, creating installers, doing UAT testing (typical Release cycle activities) on every single User Story. They may never actually reach that point, but they should envision that, and strive to keep driving the activities down closer to the User Story cycle s they mature. This is a great technique to give a team an easy-to-follow roadmap to mature their agile practices over time. Sure there’s other aspects to maturity outside of this, but it’s a great technique, that’s easy to visualize, to drive agility into the team. Just keep moving those activities (aka “gates”) down the board from Release->Sprint->Story. I’ll try to give an example of what a recent client of mine had for their DoD’s (this is from memory, so probably not 100% accurate): Release Create/Update deployment Instructions For Ops Instructional Videos Updated Run manual regression test suite UAT Testing In this case that meant deploying to an environment shared across the enterprise that mirrored production and asking other business groups to test their own apps to ensure we didn’t break anything outside our system Sprint Deploy to UAT Environment But not necessarily actually request UAT testing occur User Guides updated Sprint Features Video Created In this case we decided to create a video each sprint showing off the progress (video version of Sprint Demo) User Story Manual Test scripts developed and run Tested by BA Deployed in shared QA environment Using automated deployment process Peer Code Review Code Check-In Compiled (warning-free) Passes StyleCop Passes FxCop Create installer packages Run Automated Tests Run Automated Integration Tests PS – One of my clients had a great question when we went through this activity. They said that if a Sprint is by definition done when the end-date rolls around (time-boxed), isn’t a DoD on a sprint meaningless – it’s done on the end-date regardless of whether those other activities are complete or not? My answer is that while that statement is true – the sprint is done regardless when the end date rolls around – if the DoD activities haven’t been completed I would consider the Sprint a failure (similar to not completing what was committed/planned – failure may be too strong a word but you get the idea). In the Retrospective that will become an agenda item to discuss and understand why we weren’t able to complete the activities we agreed would need to be completed each Sprint.

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  • Take Two: Comparing JVMs on ARM/Linux

    - by user12608080
    Although the intent of the previous article, entitled Comparing JVMs on ARM/Linux, was to introduce and highlight the availability of the HotSpot server compiler (referred to as c2) for Java SE-Embedded ARM v7,  it seems, based on feedback, that everyone was more interested in the OpenJDK comparisons to Java SE-E.  In fact there were two main concerns: The fact that the previous article compared Java SE-E 7 against OpenJDK 6 might be construed as an unlevel playing field because version 7 is newer and therefore potentially more optimized. That the generic compiler settings chosen to build the OpenJDK implementations did not put those versions in a particularly favorable light. With those considerations in mind, we'll institute the following changes to this version of the benchmarking: In order to help alleviate an additional concern that there is some sort of benchmark bias, we'll use a different suite, called DaCapo.  Funded and supported by many prestigious organizations, DaCapo's aim is to benchmark real world applications.  Further information about DaCapo can be found at http://dacapobench.org. At the suggestion of Xerxes Ranby, who has been a great help through this entire exercise, a newer Linux distribution will be used to assure that the OpenJDK implementations were built with more optimal compiler settings.  The Linux distribution in this instance is Ubuntu 11.10 Oneiric Ocelot. Having experienced difficulties getting Ubuntu 11.10 to run on the original D2Plug ARMv7 platform, for these benchmarks, we'll switch to an embedded system that has a supported Ubuntu 11.10 release.  That platform is the Freescale i.MX53 Quick Start Board.  It has an ARMv7 Coretex-A8 processor running at 1GHz with 1GB RAM. We'll limit comparisons to 4 JVM implementations: Java SE-E 7 Update 2 c1 compiler (default) Java SE-E 6 Update 30 (c1 compiler is the only option) OpenJDK 6 IcedTea6 1.11pre 6b23~pre11-0ubuntu1.11.10.2 CACAO build 1.1.0pre2 OpenJDK 6 IcedTea6 1.11pre 6b23~pre11-0ubuntu1.11.10.2 JamVM build-1.6.0-devel Certain OpenJDK implementations were eliminated from this round of testing for the simple reason that their performance was not competitive.  The Java SE 7u2 c2 compiler was also removed because although quite respectable, it did not perform as well as the c1 compilers.  Recall that c2 works optimally in long-lived situations.  Many of these benchmarks completed in a relatively short period of time.  To get a feel for where c2 shines, take a look at the first chart in this blog. The first chart that follows includes performance of all benchmark runs on all platforms.  Later on we'll look more at individual tests.  In all runs, smaller means faster.  The DaCapo aficionado may notice that only 10 of the 14 DaCapo tests for this version were executed.  The reason for this is that these 10 tests represent the only ones successfully completed by all 4 JVMs.  Only the Java SE-E 6u30 could successfully run all of the tests.  Both OpenJDK instances not only failed to complete certain tests, but also experienced VM aborts too. One of the first observations that can be made between Java SE-E 6 and 7 is that, for all intents and purposes, they are on par with regards to performance.  While it is a fact that successive Java SE releases add additional optimizations, it is also true that Java SE 7 introduces additional complexity to the Java platform thus balancing out any potential performance gains at this point.  We are still early into Java SE 7.  We would expect further performance enhancements for Java SE-E 7 in future updates. In comparing Java SE-E to OpenJDK performance, among both OpenJDK VMs, Cacao results are respectable in 4 of the 10 tests.  The charts that follow show the individual results of those four tests.  Both Java SE-E versions do win every test and outperform Cacao in the range of 9% to 55%. For the remaining 6 tests, Java SE-E significantly outperforms Cacao in the range of 114% to 311% So it looks like OpenJDK results are mixed for this round of benchmarks.  In some cases, performance looks to have improved.  But in a majority of instances, OpenJDK still lags behind Java SE-Embedded considerably. Time to put on my asbestos suit.  Let the flames begin...

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  • Linksys WiFi usb dongle and linux woes

    - by MrStatic
    I have a Linksys WUSB54GC usb dongle and I have exhausted every thing I know about making this thing work in linux. I am using Fedora 13. Since it is not ready I can not view any networks. Any ideas would be great. tail of the system log Jun 2 20:14:35 localhost kernel: usb 1-7: new high speed USB device using ehci_hcd and address 8 Jun 2 20:14:35 localhost kernel: usb 1-7: New USB device found, idVendor=1737, idProduct=0077 Jun 2 20:14:35 localhost kernel: usb 1-7: New USB device strings: Mfr=1, Product=2, SerialNumber=3 Jun 2 20:14:35 localhost kernel: usb 1-7: Product: 802.11 g WLAN Jun 2 20:14:35 localhost kernel: usb 1-7: Manufacturer: Ralink Jun 2 20:14:35 localhost kernel: usb 1-7: SerialNumber: 1.0 Jun 2 20:14:35 localhost kernel: Registered led device: rt2800usb-phy3::radio Jun 2 20:14:35 localhost kernel: Registered led device: rt2800usb-phy3::assoc Jun 2 20:14:35 localhost kernel: Registered led device: rt2800usb-phy3::quality Jun 2 20:14:35 localhost NetworkManager[1367]: <info> found WiFi radio killswitch rfkill3 (at /sys/devices/pci0000:00/0000:00:1d.7/usb1/1-7/1-7:1.0/ieee80211/phy3/rfkill3) (driver <unknown>) Jun 2 20:14:35 localhost kernel: rt2800usb 1-7:1.0: firmware: requesting rt2870.bin Jun 2 20:14:35 localhost NetworkManager[1367]: <info> (wlan0): driver supports SSID scans (scan_capa 0x01). Jun 2 20:14:35 localhost NetworkManager[1367]: <info> (wlan0): new 802.11 WiFi device (driver: 'rt2800usb' ifindex: 6) Jun 2 20:14:35 localhost NetworkManager[1367]: <info> (wlan0): exported as /org/freedesktop/NetworkManager/Devices/4 Jun 2 20:14:35 localhost NetworkManager[1367]: <info> (wlan0): now managed Jun 2 20:14:35 localhost NetworkManager[1367]: <info> (wlan0): device state change: 1 -> 2 (reason 2) Jun 2 20:14:35 localhost NetworkManager[1367]: <info> (wlan0): bringing up device. Jun 2 20:14:35 localhost kernel: ADDRCONF(NETDEV_UP): wlan0: link is not ready Jun 2 20:14:35 localhost NetworkManager[1367]: <info> (wlan0): preparing device. Jun 2 20:14:35 localhost NetworkManager[1367]: <info> (wlan0): deactivating device (reason: 2). Jun 2 20:14:35 localhost NetworkManager[1367]: <info> (wlan0): supplicant interface state: starting -> ready Jun 2 20:14:35 localhost NetworkManager[1367]: <info> (wlan0): device state change: 2 -> 3 (reason 42) [root@localhost log]# iwconfig lo no wireless extensions. eth0 no wireless extensions. wlan0 IEEE 802.11bg Mode:Managed Access Point: Not-Associated Tx-Power=8 dBm Retry long limit:7 RTS thr:off Fragment thr:off Encryption key:off Power Management:on

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  • A Patent for Workload Management Based on Service Level Objectives

    - by jsavit
    I'm very pleased to announce that after a tiny :-) wait of about 5 years, my patent application for a workload manager was finally approved. Background Many operating systems have a resource manager which lets you control machine resources. For example, Solaris provides controls for CPU with several options: shares for proportional CPU allocation. If you have twice as many shares as me, and we are competing for CPU, you'll get about twice as many CPU cycles), dedicated CPU allocation in which a number of CPUs are exclusively dedicated to an application's use. You can say that a zone or project "owns" 8 CPUs on a 32 CPU machine, for example. And, capped CPU in which you specify the upper bound, or cap, of how much CPU an application gets. For example, you can throttle an application to 0.125 of a CPU. (This isn't meant to be an exhaustive list of Solaris RM controls.) Workload management Useful as that is (and tragic that some other operating systems have little resource management and isolation, and frighten people into running only 1 app per OS instance - and wastefully size every server for the peak workload it might experience) that's not really workload management. With resource management one controls the resources, and hope that's enough to meet application service objectives. In fact, we hold resource distribution constant, see if that was good enough, and adjust resource distribution if that didn't meet service level objectives. Here's an example of what happens today: Let's try 30% dedicated CPU. Not enough? Let's try 80% Oh, that's too much, and we're achieving much better response time than the objective, but other workloads are starving. Let's back that off and try again. It's not the process I object to - it's that we to often do this manually. Worse, we sometimes identify and adjust the wrong resource and fiddle with that to no useful result. Back in my days as a customer managing large systems, one of my users would call me up to beg for a "CPU boost": Me: "it won't make any difference - there's plenty of spare CPU to be had, and your application is completely I/O bound." User: "Please do it anyway." Me: "oh, all right, but it won't do you any good." (I did, because he was a friend, but it didn't help.) Prior art There are some operating environments that take a stab about workload management (rather than resource management) but I find them lacking. I know of one that uses synthetic "service units" composed of the sum of CPU, I/O and memory allocations multiplied by weighting factors. A workload is set to make a target rate of service units consumed per second. But this seems to be missing a key point: what is the relationship between artificial 'service units' and actually meeting a throughput or response time objective? What if I get plenty of one of the components (so am getting enough service units), but not enough of the resource whose needed to remove the bottleneck? Actual workload management That's not really the answer either. What is needed is to specify a workload's service levels in terms of externally visible metrics that are meaningful to a business, such as response times or transactions per second, and have the workload manager figure out which resources are not being adequately provided, and then adjust it as needed. If an application is not meeting its service level objectives and the reason is that it's not getting enough CPU cycles, adjust its CPU resource accordingly. If the reason is that the application isn't getting enough RAM to keep its working set in memory, then adjust its RAM assignment appropriately so it stops swapping. Simple idea, but that's a task we keep dumping on system administrators. In other words - don't hold the number of CPU shares constant and watch the achievement of service level vary. Instead, hold the service level constant, and dynamically adjust the number of CPU shares (or amount of other resources like RAM or I/O bandwidth) in order to meet the objective. Instrumenting non-instrumented applications There's one little problem here: how do I measure application performance in a way relating to a service level. I don't want to do it based on internal resources like number of CPU seconds it received per minute - We need to make resource decisions based on externally visible and meaningful measures of performance, not synthetic items or internal resource counters. If I have a way of marking the beginning and end of a transaction, I can then measure whether or not the application is meeting an objective based on it. If I can observe the delay factors for an application, I can see which resource shortages are slowing an application enough to keep it from meeting its objectives. I can then adjust resource allocations to relieve those shortages. Fortunately, Solaris provides facilities for both marking application progress and determining what factors cause application latency. The Solaris DTrace facility let's me introspect on application behavior: in particular I can see events like "receive a web hit" and "respond to that web hit" so I can get transaction rate and response time. DTrace (and tools like prstat) let me see where latency is being added to an application, so I know which resource to adjust. Summary After a delay of a mere few years, I am the proud creator of a patent (advice to anyone interested in going through the process: don't hold your breath!). The fundamental idea is fairly simple: instead of holding resource constant and suffering variable levels of success meeting service level objectives, properly characterise the service level objective in meaningful terms, instrument the application to see if it's meeting the objective, and then have a workload manager change resource allocations to remove delays preventing service level attainment. I've done it by hand for a long time - I think that's what a computer should do for me.

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  • Add user in CentOS 5

    - by Ron
    I created a new user in my CentOS web server with useradd. Added a password with passwd. But I can't log in with the user via SSH. I keep getting 'access denied'. I checked to make sure that the password was assigned and that the account is active. /var/log/secure shows the following error: Aug 13 03:41:40 server1 su: pam_unix(su:auth): authentication failure; logname= uid=500 euid=0 tty=pts/0 ruser=rwade rhost= user=root Please help, Thanks Thanks for the responses so far: I should add that it is a VPS on a remote computer, fresh out of the box. I can log in as the root user quite fine. I can also su to the new user, but I cannot log in as the new user. Here is my sshd_config file: # $OpenBSD: sshd_config,v 1.73 2005/12/06 22:38:28 reyk Exp $ # This is the sshd server system-wide configuration file. See # sshd_config(5) for more information. # This sshd was compiled with PATH=/usr/local/bin:/bin:/usr/bin # The strategy used for options in the default sshd_config shipped with # OpenSSH is to specify options with their default value where # possible, but leave them commented. Uncommented options change a # default value. #Port 22 #Protocol 2,1 Protocol 2 #AddressFamily any #ListenAddress 0.0.0.0 #ListenAddress :: # HostKey for protocol version 1 #HostKey /etc/ssh/ssh_host_key # HostKeys for protocol version 2 #HostKey /etc/ssh/ssh_host_rsa_key #HostKey /etc/ssh/ssh_host_dsa_key # Lifetime and size of ephemeral version 1 server key #KeyRegenerationInterval 1h #ServerKeyBits 768 # Logging # obsoletes QuietMode and FascistLogging #SyslogFacility AUTH SyslogFacility AUTHPRIV #LogLevel INFO # Authentication: #LoginGraceTime 2m #PermitRootLogin yes #StrictModes yes #MaxAuthTries 6 #RSAAuthentication yes #PubkeyAuthentication yes #AuthorizedKeysFile .ssh/authorized_keys # For this to work you will also need host keys in /etc/ssh/ssh_known_hosts #RhostsRSAAuthentication no # similar for protocol version 2 #HostbasedAuthentication no # Change to yes if you don't trust ~/.ssh/known_hosts for # RhostsRSAAuthentication and HostbasedAuthentication #IgnoreUserKnownHosts no # Don't read the user's ~/.rhosts and ~/.shosts files #IgnoreRhosts yes # To disable tunneled clear text passwords, change to no here! #PasswordAuthentication yes #PermitEmptyPasswords no PasswordAuthentication yes # Change to no to disable s/key passwords #ChallengeResponseAuthentication yes ChallengeResponseAuthentication no # Kerberos options #KerberosAuthentication no #KerberosOrLocalPasswd yes #KerberosTicketCleanup yes #KerberosGetAFSToken no # GSSAPI options #GSSAPIAuthentication no GSSAPIAuthentication yes #GSSAPICleanupCredentials yes GSSAPICleanupCredentials yes # Set this to 'yes' to enable PAM authentication, account processing, # and session processing. If this is enabled, PAM authentication will # be allowed through the ChallengeResponseAuthentication mechanism. # Depending on your PAM configuration, this may bypass the setting of # PasswordAuthentication, PermitEmptyPasswords, and # "PermitRootLogin without-password". If you just want the PAM account and # session checks to run without PAM authentication, then enable this but set # ChallengeResponseAuthentication=no #UsePAM no UsePAM yes # Accept locale-related environment variables AcceptEnv LANG LC_CTYPE LC_NUMERIC LC_TIME LC_COLLATE LC_MONETARY LC_MESSAGES AcceptEnv LC_PAPER LC_NAME LC_ADDRESS LC_TELEPHONE LC_MEASUREMENT AcceptEnv LC_IDENTIFICATION LC_ALL #AllowTcpForwarding yes #GatewayPorts no #X11Forwarding no X11Forwarding yes #X11DisplayOffset 10 #X11UseLocalhost yes #PrintMotd yes #PrintLastLog yes #TCPKeepAlive yes #UseLogin no #UsePrivilegeSeparation yes #PermitUserEnvironment no #Compression delayed #ClientAliveInterval 0 #ClientAliveCountMax 3 #ShowPatchLevel no #UseDNS yes #PidFile /var/run/sshd.pid #MaxStartups 10 #PermitTunnel no #ChrootDirectory none # no default banner path #Banner /some/path # override default of no subsystems Subsystem sftp /usr/libexec/openssh/sftp-server

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  • Exchange IMAP4 connector - Error Event ID 2006

    - by MikeB
    Hi, A couple of users in my organisation use IMAP4 to connect to Exchange 2007 (Update rollup 9 applied) because they prefer Thunderbird / Postbox clients. One of the users is generating errors in the Application Log as follows: An exception Microsoft.Exchange.Data.Storage.ConversionFailedException occurred while converting message Imap4Message 1523, user "*******", folder *********, subject: "******", date: "*******" into MIME format. Microsoft.Exchange.Data.Storage.ConversionFailedException: Message content has become corrupted. ---> System.ArgumentException: Value should be a valid content type in the form 'token/token' Parameter name: value at Microsoft.Exchange.Data.Mime.ContentTypeHeader.set_Value(String value) at Microsoft.Exchange.Data.Storage.MimeStreamWriter.WriteHeader(HeaderId type, String data) at Microsoft.Exchange.Data.Storage.ItemToMimeConverter.WriteMimeStreamAttachment(StreamAttachmentBase attachment, MimeFlags flags) --- End of inner exception stack trace --- at Microsoft.Exchange.Data.Storage.ItemToMimeConverter.WriteMimeStreamAttachment(StreamAttachmentBase attachment, MimeFlags flags) at Microsoft.Exchange.Data.Storage.ItemToMimeConverter.WriteMimeAttachment(MimePartInfo part, MimeFlags flags) at Microsoft.Exchange.Data.Storage.ItemToMimeConverter.WriteMimePart(MimePartInfo part, MimeFlags mimeFlags) at Microsoft.Exchange.Data.Storage.ItemToMimeConverter.WriteMimeParts(List`1 parts, MimeFlags mimeFlags) at Microsoft.Exchange.Data.Storage.ItemToMimeConverter.WriteMimePart(MimePartInfo part, MimeFlags mimeFlags) at Microsoft.Exchange.Data.Storage.ImapItemConverter.<>c__DisplayClass2.<WriteMimePart>b__0() at Microsoft.Exchange.Data.Storage.ConvertUtils.CallCts(Trace tracer, String methodName, String exceptionString, CtsCall ctsCall) at Microsoft.Exchange.Data.Storage.ImapItemConverter.WriteMimePart(ItemToMimeConverter converter, MimeStreamWriter writer, OutboundConversionOptions options, MimePartInfo partInfo, MimeFlags conversionFlags) at Microsoft.Exchange.Data.Storage.ImapItemConverter.GetBody(Stream outStream) at Microsoft.Exchange.Data.Storage.ImapItemConverter.GetBody(Stream outStream, UInt32[] indices) From my reading around it seems that the suggestion is to ask users to log in to Outlook / OWA and view the messages there. However, having logged in as the users myself, the messages cannot be found either through searching or by browsing the folder detailed in the log entry. The server returns the following error to the client: "The message could not be retrieved using the IMAP4 protocol. The message has not been deleted and may be accessible using either Microsoft Outlook or Microsoft Office Outlook Web Access. You can also try contacting the original sender of the message to find out about the contents of the message. Retrieval of this message will be retried when the server is updated with a fix that addresses the problem." Messages were transferred in to Exchange by copying them from the old Apple Xserve, accessed using IMAP. So my question, finally: 1. Is there any way to get the IMAP Exchange connector to rebuild its cache of messages since it doesn't seem to be pulling them directly from the MAPI store? 2. Alternatively, if there is no database, any ideas on why these messages don't appear in Outlook or OWA would be gratefully received. Many thanks, Mike

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  • Raid1 with active and spare partition

    - by Daniel Baron
    I am having the following problem with a RAID1 software raid partition on my Ubuntu system (10.04 LTS, 2.6.32-24-server in case it matters). One of my disks (sdb5) reported I/O errors and was therefore marked faulty in the array. The array was then degraded with one active device. Hence, I replaced the harddisk, cloned the partition table and added all new partitions to my raid arrays. After syncing all partitions ended up fine, having 2 active devices - except one of them. The partition which reported the faulty disk before, however, did not include the new partition as an active device but as a spare disk: md3 : active raid1 sdb5[2] sda5[1] 4881344 blocks [2/1] [_U] A detailed look reveals: root@server:~# mdadm --detail /dev/md3 [...] Number Major Minor RaidDevice State 2 8 21 0 spare rebuilding /dev/sdb5 1 8 5 1 active sync /dev/sda5 So here is the question: How do I tell my raid to turn the spare disk into an active one? And why has it been added as a spare device? Recreating or reassembling the array is not an option, because it is my root partition. And I can not find any hints to that subject in the Software Raid HOWTO. Any help would be appreciated. Current Solution I found a solution to my problem, but I am not sure that this is the actual way to do it. Having a closer look at my raid I found that sdb5 was always listed as a spare device: mdadm --examine /dev/sdb5 [...] Number Major Minor RaidDevice State this 2 8 21 2 spare /dev/sdb5 0 0 0 0 0 removed 1 1 8 5 1 active sync /dev/sda5 2 2 8 21 2 spare /dev/sdb5 so readding the device sdb5 to the array md3 always ended up in adding the device as a spare. Finally I just recreated the array mdadm --create /dev/md3 --level=1 -n2 -x0 /dev/sda5 /dev/sdb5 which worked. But the question remains open for me: Is there a better way to manipulate the summaries in the superblock and to tell the array to turn sdb5 from a spare disk to an active disk? I am still curious for an answer.

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  • MegaCli newly created disk doesn't appear under /dev/sdX

    - by Henry-Nicolas Tourneur
    After having successfully added 2 new disks in a new RAID virtual drive (background initialization done), I would have exepected it to appear under /dev/sdh but it's not there (so, unusable). The system is running a CentOS 5.2 64 bits, HAL and udev daemons are running, not records of any sdh apparition under the messsage log file or in dmesg, only MegaCli do see that virtual drive. Any idea ? Some data: [root@server ~]# ./MegaCli -LDInfo -LALL -a0 Adapter 0 -- Virtual Drive Information: Virtual Disk: 0 (target id: 0) Name: RAID Level: Primary-1, Secondary-0, RAID Level Qualifier-0 Size:139392MB State: Optimal Stripe Size: 64kB Number Of Drives:2 Span Depth:1 Default Cache Policy: WriteBack, ReadAheadNone, Direct, No Write Cache if Bad BBU Current Cache Policy: WriteBack, ReadAheadNone, Direct, No Write Cache if Bad BBU Access Policy: Read/Write Disk Cache Policy: Disk's Default Virtual Disk: 1 (target id: 1) Name: RAID Level: Primary-1, Secondary-0, RAID Level Qualifier-0 Size:285568MB State: Optimal Stripe Size: 64kB Number Of Drives:2 Span Depth:1 Default Cache Policy: WriteBack, ReadAheadNone, Direct, No Write Cache if Bad BBU Current Cache Policy: WriteBack, ReadAheadNone, Direct, No Write Cache if Bad BBU Access Policy: Read/Write Disk Cache Policy: Disk's Default [root@server ~]# ls -l /dev/disk/by-id/scsi-360* lrwxrwxrwx 1 root root 9 Nov 17 2010 /dev/disk/by-id/scsi-36001ec90f82fe100108ca0a704098d09 -> ../../sda lrwxrwxrwx 1 root root 10 Nov 17 2010 /dev/disk/by-id/scsi-36001ec90f82fe100108ca0a704098d09-part1 -> ../../sda1 lrwxrwxrwx 1 root root 10 Nov 17 2010 /dev/disk/by-id/scsi-36001ec90f82fe100108ca0a704098d09-part2 -> ../../sda2 lrwxrwxrwx 1 root root 9 Nov 17 2010 /dev/disk/by-id/scsi-36090a028e0fe07e78f94940c0000a0ee -> ../../sdf lrwxrwxrwx 1 root root 10 Nov 17 2010 /dev/disk/by-id/scsi-36090a028e0fe07e78f94940c0000a0ee-part1 -> ../../sdf1 lrwxrwxrwx 1 root root 9 Nov 17 2010 /dev/disk/by-id/scsi-36090a028e0fe972a3f91240a0000005f -> ../../sdb lrwxrwxrwx 1 root root 10 Nov 17 2010 /dev/disk/by-id/scsi-36090a028e0fe972a3f91240a0000005f-part1 -> ../../sdb1 lrwxrwxrwx 1 root root 9 Nov 17 2010 /dev/disk/by-id/scsi-36090a028e0fea7e18f94640c000020ec -> ../../sde lrwxrwxrwx 1 root root 10 Nov 17 2010 /dev/disk/by-id/scsi-36090a028e0fea7e18f94640c000020ec-part1 -> ../../sde1 lrwxrwxrwx 1 root root 9 Nov 17 2010 /dev/disk/by-id/scsi-36090a028e0feb7da8f94340c0000203d -> ../../sdd lrwxrwxrwx 1 root root 10 Nov 17 2010 /dev/disk/by-id/scsi-36090a028e0feb7da8f94340c0000203d-part1 -> ../../sdd1 lrwxrwxrwx 1 root root 9 Nov 17 2010 /dev/disk/by-id/scsi-36090a028e0fed7d78f94040c000080b7 -> ../../sdc lrwxrwxrwx 1 root root 10 Nov 17 2010 /dev/disk/by-id/scsi-36090a028e0fed7d78f94040c000080b7-part1 -> ../../sdc1 lrwxrwxrwx 1 root root 9 Nov 17 2010 /dev/disk/by-id/scsi-36090a05830145e58e0b9c479000010a1 -> ../../sdg lrwxrwxrwx 1 root root 10 Nov 17 2010 /dev/disk/by-id/scsi-36090a05830145e58e0b9c479000010a1-part1 -> ../../sdg1

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  • Why does writing a file to an NFS share send a COMMIT operation to the NFS server?

    - by Antonis Christofides
    I have a Debian squeeze (2.6.32-5-amd64) which is at the same time a NFS4 server and client (it mounts itself through NFS4). The local directory that leads directly to disk is /nfs4exports/mydir, whereas /nfs4mounts/mydir is the same thing mounted through NFS, using the machine's external IP address. Here is the line from fstab: 192.168.1.75:/mydir /nfs4mounts/mydir nfs4 soft 0 0 I have an application that writes many small files. If I write directly to /nfs4exports/mydir, it writes thousands of files per second; but if I write to /nfs4mounts/mydir, it writes 4 files per second or so. I can greatly increase speed if I add async to /etc/exports. (Writing a single large file to the NFS-mounted directory goes at more than 100 MB/s.) I examine the server statistics and I see that whenever a file is written, it is "committed" (this also happens with NFSv3): root@debianvboxtest:~# mount -t nfs4 192.168.1.75:/mydir /mnt root@debianvboxtest:~# nfsstat|grep -A 2 'nfs v4 operations' Server nfs v4 operations: op0-unused op1-unused op2-future access close commit 0 0% 0 0% 0 0% 10 4% 1 0% 1 0% root@debianvboxtest:~# echo 'hello' >/mnt/test1056 root@debianvboxtest:~# nfsstat|grep -A 2 'nfs v4 operations' Server nfs v4 operations: op0-unused op1-unused op2-future access close commit 0 0% 0 0% 0 0% 11 4% 2 0% 2 0% Now in the RFC, I read this: The COMMIT operation is similar in operation and semantics to the POSIX fsync(2) system call that synchronizes a file's state with the disk (file data and metadata is flushed to disk or stable storage). COMMIT performs the same operation for a client, flushing any unsynchronized data and metadata on the server to the server's disk or stable storage for the specified file. I don't understand why the client commits. I don't think that the "echo" shell built-in command runs fsync; if echo wrote to a local file and then the machine went down, the file might be lost. In contrast, the NFS client appears to be sending a COMMIT upon completion of the echo. Why? I am reluctant to use the async NFS server option, because it would apparently ignore COMMIT. I feel as if I had a local filesystem and I had to choose between syncing every file upon close and ignoring fsync altogether. What have I understood wrong?

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  • How to setup ssh's umask for all type of connections

    - by Unode
    I've been searching for a way to setup OpenSSH's umask to 0027 in a consistent way across all connection types. By connection types I'm referring to: sftp scp ssh hostname ssh hostname program The difference between 3. and 4. is that the former starts a shell which usually reads the /etc/profile information while the latter doesn't. In addition by reading this post I've became aware of the -u option that is present in newer versions of OpenSSH. However this doesn't work. I must also add that /etc/profile now includes umask 0027. Going point by point: sftp - Setting -u 0027 in sshd_config as mentioned here, is not enough. If I don't set this parameter, sftp uses by default umask 0022. This means that if I have the two files: -rwxrwxrwx 1 user user 0 2011-01-29 02:04 execute -rw-rw-rw- 1 user user 0 2011-01-29 02:04 read-write When I use sftp to put them in the destination machine I actually get: -rwxr-xr-x 1 user user 0 2011-01-29 02:04 execute -rw-r--r-- 1 user user 0 2011-01-29 02:04 read-write However when I set -u 0027 on sshd_config of the destination machine I actually get: -rwxr--r-- 1 user user 0 2011-01-29 02:04 execute -rw-r--r-- 1 user user 0 2011-01-29 02:04 read-write which is not expected, since it should actually be: -rwxr-x--- 1 user user 0 2011-01-29 02:04 execute -rw-r----- 1 user user 0 2011-01-29 02:04 read-write Anyone understands why this happens? scp - Independently of what is setup for sftp, permissions are always umask 0022. I currently have no idea how to alter this. ssh hostname - no problem here since the shell reads /etc/profile by default which means umask 0027 in the current setup. ssh hostname program - same situation as scp. In sum, setting umask on sftp alters the result but not as it should, ssh hostname works as expected reading /etc/profile and both scp and ssh hostname program seem to have umask 0022 hardcoded somewhere. Any insight on any of the above points is welcome. EDIT: I would like to avoid patches that require manually compiling openssh. The system is running Ubuntu Server 10.04.01 (lucid) LTS with openssh packages from maverick. Answer: As indicated by poige, using pam_umask did the trick. The exact changes were: Lines added to /etc/pam.d/sshd: # Setting UMASK for all ssh based connections (ssh, sftp, scp) session optional pam_umask.so umask=0027 Also, in order to affect all login shells regardless of if they source /etc/profile or not, the same lines were also added to /etc/pam.d/login. EDIT: After some of the comments I retested this issue. At least in Ubuntu (where I tested) it seems that if the user has a different umask set in their shell's init files (.bashrc, .zshrc,...), the PAM umask is ignored and the user defined umask used instead. Changes in /etc/profile did't affect the outcome unless the user explicitly sources those changes in the init files. It is unclear at this point if this behavior happens in all distros.

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  • =?UTF-8?B??= in Emails sent via php mail problem

    - by Camran
    I have a website, and in the "Contact" section I have a form which users may fill in to contact me. The form is a simple form which action is a php page. The php code: $to = "[email protected]"; $name=$_POST['name']; // sender name $email=$_POST['email']; // sender email $tel= $_POST['tel']; // sender tel $subject=$_POST['subject']; // subject CHOSEN FROM DROPLIST, ALL TESTED $text=$_POST['text']; // Message from sender $text.="\n\nTel:".$tel; // Added to message to show me the telephone nr to the sender at bottom of message $headers="MIME-Version: 1.0"."\n"; $headers.="Content-type: text/plain; charset=UTF-8"."\n"; $headers.="From: $name <$email>"."\n"; mail($to, '=?UTF-8?B?'.base64_encode($subject).'?=', $text, $headers, '[email protected]'); Could somebody please tell me why this works most of the time, but sometimes I receive email whith no text and the subject line showing =?UTF-8?B??= I use outlook express, and I have read this http://stackoverflow.com/questions/454833/system-net-mail-and-utf-8bxxxxx-headers but it didn't help. The problem is not in Outlook, because when I log in to the actual mailprogram where I fetch the POP3 emails from, the email looks the same. When I right click in Outlook and chose "message source" then there is no "From" information. Ex, a good message should look like this: Subject: =?UTF-8?B?w5Z2cmlndA==?= MIME-Version: 1.0 Content-type: text/plain; charset=UTF-8 From: John Doe However, the ones with problem looks like this: Subject: =?UTF-8?B??= MIME-Version: 1.0 Content-type: text/plain; charset=UTF-8 From: As if the information has been lost somewhere. You should know also that I have a VPS, which I manage myself. I use postfix as an emailserver, if thats got anything to do with it. But then again, why does it work sometimes? Also another thing that I have noticed is that sometimes special characters are not shown correctly (by both Outlook and the webmail). For instance, the name "Björkman" in swedish is shown like Björkman, but again, only sometimes. I hope anybody knows something about this problem, because it is very hard to track down for me atleast. If you need more input let me know.

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  • startup Error for Zend Server CE

    - by Jamison
    Hello! I've got a strange startup error for Zend Server CE - it's probably easy to fix, but I don't have much experience with Zend Server! I'm running the latest OSX 10.6.6 and the latest Zend Server CE for Mac. When I run the "start" command from the command line, here is what I get: /usr/local/zend/bin/apachectl start [OK] spawn-fcgi: child spawned successfully: PID: 4206 /usr/local/zend/bin/shell_functions.rc: line 133: 4210 Bus error $WATCHDOG -i $BINARY 1>&3 2>&4 /usr/local/zend/bin/shell_functions.rc: line 133: 4211 Bus error $WATCHDOG -u $WD_UID -g $WD_GID -s $BINARY 1>&3 2>&4 Starting Zend Server GUI [Lighttpd] [FAILED] /usr/local/zend/bin/lighttpdctl.sh: line 46: 4212 Bus error $WATCHDOG -i $BINARY Starting MySQL SUCCESS! /usr/local/zend/bin/shell_functions.rc: line 133: 4304 Bus error $WATCHDOG -i $BINARY 1>&3 2>&4 /usr/local/zend/bin/shell_functions.rc: line 133: 4425 Bus error $WATCHDOG -u $WD_UID -g $WD_GID -s $BINARY 1>&3 2>&4 Starting Java bridge [FAILED] /usr/local/zend/bin/java_bridge.sh: line 39: 4426 Bus error $WATCHDOG -i $BINARY Zend Server started... The challenge is that ZEND SERVER wont open the GUI with this error, and seemingly I can click on Zend Server in the Applications folder and it opens for a second and immediately closes. I've made sure that Web Sharing is turned off to avoid conflicts, and I've run Disk Utility from my recovery disk to make sure there are no file system errors. Here is what the lines that are referenced in the errors have in terms of code: shell_functions.rc: (starting on line 132 - the error message says line 133...): launch() { if [ -z "$DEBUG" ]; then exec 3>/dev/null 4>&3 else exec 3>&1 4>&2 fi $WATCHDOG -i $BINARY 1>&3 2>&4 RET=$? if [ $RET -eq 0 ];then $ECHO_CMD "$BINARY watchdog is up and running.. ${OK_COLOR}[OK]${T_RESET}" return $RET else #$WATCHDOG -u $WD_UID -g $WD_GID -s $BINARY >> "$PREFIX/logs/watchdog_$BINARY.log" 2>&1 $WATCHDOG -u $WD_UID -g $WD_GID -s $BINARY 1>&3 2>&4 report $? "Starting" fi } _kill() { $WATCHDOG -i $BINARY > /dev/null 2>&1 if [ $? -eq 1 ];then $ECHO_CMD "$BINARY is not running" else $WATCHDOG -t $BINARY > /dev/null 2>&1 report $? "Stopping" fi } lighttpdctl.sh: (starting on line 45 - the error message says line 46...): status() { $WATCHDOG -i $BINARY } case "$1" in start) start status ;; stop) stop ;; restart) stop sleep 1 start ;; status) status ;; *) usage exit 1 esac exit $? java_bridge.sh: (starting on line 38 - the error message says line 39...): status() { $WATCHDOG -i $BINARY } Question: "Watchdog" is library in this zend BIN folder - it seems to handle error reporting? all the errors in my start command seem to deal with this Watchdog thing, but I don't know what to do about it... Thanks!

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  • Part 8: How to name EBS Customizations

    - by volker.eckardt(at)oracle.com
    You might wonder why I am discussing this here. The reason is simple: nearly every project has a bit different naming conventions, which makes a the life always a bit complicated (for developers, but also setup responsible, and also for consultants).  Although we always create a document to describe the technical object naming conventions, I have rarely seen a dedicated document  with functional naming conventions. To be precisely, from my stand point, there should always be one global naming definition for an implementation! Let me discuss some related questions: What is the best convention for the customization reference? How to name database objects (tables, packages etc.)? How to name functional objects like Value Sets, Concurrent Programs, etc. How to separate customizations from standard objects best? What is the best convention for the customization reference? The customization reference is the key you use to reference your customization from other lists, from the project plan etc. Usually it is something like XXHU_CONV_22 (HU=customer abbreviation, CONV=Conversion object #22) or XXFA_DEPRN_RPT_02 (FA=Fixed Assets, DEPRN=Short object group, here depreciation, RPT=Report, 02=2nd report in this area) As this is just a reference (not an object name yet), I would prefer the second option. XX=Customization, FA=Main EBS Module linked (you may have sometimes more, but FA is the main) DEPRN_RPT=Short name to specify the customization 02=a unique number Important here is that the HU isn’t used, because XX is enough to mark a custom object, and the 3rd+4th char can be used by the EBS module short name. How to name database objects (tables, packages etc.)? I was leading different developer teams, and I know that one common way is it to take the Customization reference and add more chars behind to classify the object (like _V for view and _T1 for triggers etc.). The only concern I have with this approach is the reusability. If you name your view XXFA_DEPRN_RPT_02_V, no one will by choice reuse this nice view, as it seams to be specific for this CEMLI. My suggestion is rather to name the view XXFA_DEPRN_PERIODS_V and allow herewith reusability for other CEMLIs (although the view will be deployed primarily with CEMLI package XXFA_DEPRN_RPT_02). (check also one of the following Blogs where I will talk about deployment.) How to name Value Sets, Concurrent Programs, etc. For Value Sets I would go with the same convention as for database objects, starting with XX<Module> …. For Concurrent Programs the situation is a bit different. This “object” is seen and used by a lot of users, and they will search for. In many projects it is common to start again with the company short name, or with XX. My proposal would differ. If you have created your own report and you name it “XX: Invoice Report”, the user has to remember that this report does not start with “I”, it starts with X. Would you like typing an X if you are looking for an Invoice report? No, you wouldn’t! So my advise would be to name it:   “Invoice Report (XXAP)”. Still we know it is custom (because of the XXAP), but the end user will type the key “i” to get it (and will see similar reports starting also with “i”). I hope that the general schema behind has now become obvious. How to separate customizations from standard objects best? I would not have this section here if the naming would not play an important role. Unfortunately, we can not always link a custom application to our own object, therefore the naming is really important. In the file system structure we use our $XXyy_TOP, in JAVA_TOP it is perhaps also “xx” in front. But in the database itself? Although there are different concepts in place, still many implementations are using the standard “apps” approach, means custom objects are stored in the apps schema (which should not cause any trouble). Final advise: review the naming conventions regularly, once a month. You may have to add more! And, publish them! To summarize: Technical and functional customized objects should always follow a naming convention. This naming convention should be project wide, and only one place shall be used to maintain (like in a Wiki). If the name is for the end user, rather put a customization identifier at the end; if it is an internal name, start with XX…

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  • Migrating Virtual Iron guest to Oracle VM 3.x

    - by scoter
    As stated on the official site, Oracle in 2009, acquired a provider of server virtualization management software named Virtual Iron; you can find all the acquisition details at this link. Into the FAQ on the official site you can also view that, for the future, Oracle plans to fully integrate Virtual Iron technology into Oracle VM products, and any enhancements will be delivered as a part of the combined solution; this is what is going on with Oracle VM 3.x. So, customers started asking us to migrate Virtual Iron guests to Oracle VM. IMPORTANT: This procedure needs a dedicated OVM-Server with no-guests running on top; be careful while execute this procedure on production environments. In these little steps you will find how-to migrate, as fast as possible, your guests between VI ( Virtual Iron ) and Oracle VM; keep in mind that OracleVM has a built-in P2V utility ( Official Documentation )  that you can use to migrate guests between VI and Oracle VM. Concepts: VI repositories.  On VI we have the same "repository" concept as in Oracle VM; the difference between these two products is that VI use a raw-lun as repository ( instead of using ocfs2 and its capabilities, like ref-links ). The VI "raw-lun" repository, with a pure operating-system perspective, may be presented as in this picture: Infact on this "raw-lun" VI create an LVM2 volume-group. The VI "raw-lun" repository, with an hypervisor perspective, may be presented as in this picture: So, the relationships are: LVM2-Volume-Group <-> VI Repository LVM2-Logical-Volume <-> VI guest virtual-disk The first step is to present the VI repository ( raw-lun ) to your dedicated OVM-Server. Prepare dedicated OVM-Server On the OVM-Server ( OVS ) you need to discover new lun and, after that, discover volume-group and logical-volumes containted in VI repository; due to default OVS configuration you need to edit lvm2 configuration file: /etc/lvm/lvm.conf     # By default for OVS we restrict every block device:     # filter = [ "r/.*/" ] and comment the line starting with "filter" as above. Now you have to discover the raw-lun presented and, next, activate volume-group and logical-volumes: #!/bin/bash for HOST in `ls /sys/class/scsi_host`;do echo '- - -' > /sys/class/scsi_host/$HOST/scan; done CPATH=`pwd` cd /dev for DEVICE in `ls sd[a-z] sd?[a-z]`;do echo '1' > /sys/block/$DEVICE/device/rescan; done cd $CPATH cd /dev/mapper for PARTITION in `ls *[a-z] *?[a-z]`;do partprobe /dev/mapper/$PARTITION; done cd $CPATH vgchange -a yAfter that you will see a new device:[root@ovs01 ~]# cd /dev/6000F4B00000000000210135bef64994[root@ovs01 6000F4B00000000000210135bef64994]# ls -l 6000F4B0000000000061013* lrwxrwxrwx 1 root root 77 Oct 29 10:50 6000F4B00000000000610135c3a0b8cb -> /dev/mapper/6000F4B00000000000210135bef64994-6000F4B00000000000610135c3a0b8cb By your OVM-Manager create a guest server with the same definition as on VI:same core number as VI source guestsame memory as VI source guestsame number of disks as VI source guest ( you can create OVS virtual disk with a small size of 1GB because the "clone" will, eventually, extend the size of your new virtual disks )Summarizing:source-virtual-disk path ( VI ):/dev/mapper/6000F4B00000000000210135bef64994-6000F4B00000000000610135c3a0b8cbdest-virtual-disk path ( OVS ):/OVS/Repositories/0004fb00000300006cfeb81c12f12f00/VirtualDisks/0004fb000012000055e0fc4c5c8a35ee.img ** ** = to identify your virtual disk you have verify its name under the "vm.cfg" file of your new guest.Clone VI virtual-disk to OVS virtual-diskdd if=/dev/mapper/6000F4B00000000000210135bef64994-6000F4B00000000000610135c3a0b8cb of=/OVS/Repositories/0004fb00000300006cfeb81c12f12f00/VirtualDisks/0004fb000012000055e0fc4c5c8a35ee.img Clean unsupported parameters and changes on OVS.1. Restore original /etc/lvm/lvm.conf    # By default for OVS we restrict every block device:     filter = [ "r/.*/" ]    and uncomment the line starting with "filter" as above.2. Force-stop lvm2-monitor service  # service lvm2-monitor force-stop 3. Restore original /etc/lvm directories ( archive, backup and cache )  # cd /etc/lvm  # rm -fr archive backup cache; mkdir archive backup cache4. Reboot OVSRefresh OVS repository and start your guest.By OracleVM Manager refresh your repository:By OracleVM Manager start your "migrated" guest: Comments and corrections are welcome.  Simon COTER 

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  • Default permission for newly-created files/folders using ACLs not respected by commands like "unzip"

    - by Ngoc Pham
    I am having trouble with setting up a system for multiple users accessing the same set of files. I've read tuts and docs around and played with ACLs but haven't succeeded yet. MY SCENARIO: Have multiple users, for example, user1 and user2, which is belong to a group called sharedusers. They must have all WRITE permission to a same set of files and directories, say underlying in /userdata/sharing/. I have the folder's group set to sharedusers and SGID to have all newly created files/dirs inside set to same group. ubuntu@home:/userdata$ ll drwxr-sr-x 2 ubuntu sharedusers 4096 Nov 24 03:51 sharing/ I set ACLs for this directory so I can have permission of sub dirs/files inheritted from its parents. ubuntu@home:/userdata$ setfacl -m group:sharedusers:rwx sharing/ ubuntu@home:/userdata$ setfacl -d -m group:sharedusers:rwx sharing/ Here's what I've got: ubuntu@home:/userdata$ getfacl sharing/ # file: sharing/ # owner: ubuntu # group: sharedusers # flags: -s- user::rwx group::r-x group:sharedusers:rwx mask::rwx other::r-x default:user::rwx default:group::r-x default:group:sharedusers:rwx default:mask::rwx default:other::r-x Seems okay as when I create new folder with new files inside and the permission is correct. ubuntu@home:/userdata/sharing$ mkdir a && cd a ubuntu@home:/userdata/sharing/a$ touch a_test ubuntu@home:/userdata/sharing/a$ getfacl a_test # file: a_test # owner: ubuntu # group: sharedusers user::rw- group::r-x #effective:r-- group:sharedusers:rwx #effective:rw- mask::rw- other::r-- As you can see, the sharedusers group has effective permission rw-. HOWEVER, if I have a zip file, and use unzip -q command to unzip the file inside the folder sharing, the extracted folders don't have group write permisison. Therefore, the users from group sharedusers cannot modify files under those extracted folders. ubuntu@home:/userdata/sharing$ unzip -q Joomla_3.0.2-Stable-Full_Package.zip ubuntu@home:/userdata/sharing$ ll drwxrwsr-x+ 2 ubuntu sharedusers 4096 Nov 24 04:00 a/ drwxr-xr-x+ 10 ubuntu sharedusers 4096 Nov 7 01:52 administrator/ drwxr-xr-x+ 13 ubuntu sharedusers 4096 Nov 7 01:52 components/ You an spot the difference in permissions between folder a (created before) and folder administrator extracted by unzip. And the ACLs of a files inside administrator: ubuntu@home:/userdata/sharing$ getfacl administrator/index.php # file: administrator/index.php # owner: ubuntu # group: ubuntu user::rw- group::r-x #effective:r-- group:sharedusers:rwx #effective:r-- mask::r-- other::r-- It also has ubuntu group, not sharedusers group as expected. Could someone please explain the problem and give me advice? Thank you in advance!

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  • #OOW 2012 : IaaS, Private Cloud, Multitenant Database, and X3H2M2

    - by Eric Bezille
    The title of this post is a summary of the 4 announcements made by Larry Ellison today, during the opening session of Oracle Open World 2012... To know what's behind X3H2M2, you will have to wait a little, as I will go in order, beginning with the IaaS - Infrastructure as a Service - announcement. Oracle IaaS goes Public... and Private... Starting in 2004 with Fusion development, Oracle Cloud was launch last year to provide not only SaaS Application, based on standard development, but also the underlying PaaS, required to build the specifics, and required interconnections between applications, in and outside of the Cloud. Still, to cover the end-to-end Cloud  Services spectrum, we had to provide an Infrastructure as a Service, leveraging our Servers, Storage, OS, and Virtualization Technologies, all "Engineered Together". This Cloud Infrastructure, was already available for our customers to build rapidly their own Private Cloud either on SPARC/Solaris or x86/Linux... The second announcement made today bring that proposition a big step further : for cautious customers (like Banks, or sensible industries) who would like to benefits from the Cloud value of "as a Service", but don't want their Data out in the Cloud... We propose to them to operate the same systems, Exadata, Exalogic & SuperCluster, that are providing our Public Cloud Infrastructure, behind their firewall, in a Private Cloud model. Oracle 12c Multitenant Database This is also a major announcement made today, on what's coming with Oracle Database 12c : the ability to consolidate multiple databases with no extra additional  cost especially in terms of memory needed on the server node, which is often THE consolidation limiting factor. The principle could be compare to Solaris Zones, where, you will have a Database Container, who is "owning" the memory and Database background processes, and "Pluggable" Database in this Database Container. This particular feature is a strong compelling event to evaluate rapidly Oracle Database 12c once it will be available, as this is major step forward into true Database consolidation with Multitenancy on a shared (optimized) infrastructure. X3H2M2, enabling the new Exadata X3 in-Memory Database Here we are :  X3H2M2 stands for X3 (the new version of Exadata announced also today) Heuristic Hierarchical Mass Memory, providing the capability to keep most if not all the Data in the memory cache hierarchy. Of course, this is the major software enhancement of the new X3 Exadata machine, but as this is a software, our current customers would be able to benefit from it on their existing systems by upgrading to the new release. But that' not the only thing that we did with X3, at the same time we have upgraded everything : the CPUs, adding more cores per server node (16 vs. 12, with the arrival of Intel E5 / Sandy Bridge), the memory with 512GB memory as well per node,  and the new Flash Fire card, bringing now up to 22 TB of Flash cache. All of this 4TB of RAM + 22TB of Flash being use cleverly not only for read but also for write by the X3H2M2 algorithm... making a very big difference compare to traditional storage flash extension. But what does those extra performances brings to you on an already very efficient system: double your performances compare to the fastest storage array on the market today (including flash) and divide you storage price x10 at the same time... Something to consider closely this days... Especially that we also announced the availability of a new Exadata X3-2 8th rack : a good starting point. As you have seen a major opening for this year again with true innovation. But that was not the only thing that we saw today, as before Larry's talk, Fujitsu did introduce more in deep the up coming new SPARC processor, that they are co-developing with us. And as such Andrew Mendelsohn - Senior Vice President Database Server Technologies came on stage to explain that the next step after I/O optimization for Database with Exadata, was to accelerate the Database at execution level by bringing functions in the SPARC processor silicium. All in all, to process more and more Data... The big theme of the day... and of the Oracle User Groups Conferences that were also happening today and where I had the opportunity to attend some interesting sessions on practical use cases of Big Data one in Finances and Fraud profiling and the other one on practical deployment of Oracle Exalytics for Data Analytics. In conclusion, one picture to try to size Oracle Open World ... and you can understand why, with such a rich content... and this only the first day !

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  • Data Source Security Part 5

    - by Steve Felts
    If you read through the first four parts of this series on data source security, you should be an expert on this focus area.  There is one more small topic to cover related to WebLogic Resource permissions.  After that comes the test, I mean example, to see with a real set of configuration parameters what the results are with some concrete values. WebLogic Resource Permissions All of the discussion so far has been about database credentials that are (eventually) used on the database side.  WLS has resource credentials to control what WLS users are allowed to access JDBC resources.  These can be defined on the Policies tab on the Security tab associated with the data source.  There are four permissions: “reserve” (get a new connection), “admin”, “shrink”, and reset (plus the all-inclusive “ALL”); we will focus on “reserve” here because we are talking about getting connections.  By default, JDBC resource permissions are completely open – anyone can do anything.  As soon as you add one policy for a permission, then all other users are restricted.  For example, if I add a policy so that “weblogic” can reserve a connection, then all other users will fail to reserve connections unless they are also explicitly added.  The validation is done for WLS user credentials only, not database user credentials.  Configuration of resources in general is described at “Create policies for resource instances” http://docs.oracle.com/cd/E24329_01/apirefs.1211/e24401/taskhelp/security/CreatePoliciesForResourceInstances.html.  This feature can be very useful to restrict what code and users can get to your database. There are the three use cases: API Use database credentials User for permission checking getConnection() True or false Current WLS user getConnection(user,password) False User/password from API getConnection(user,password) True Current WLS user If a simple getConnection() is used or database credentials are enabled, the current user that is authenticated to the WLS system is checked. If database credentials are not enabled, then the user and password on the API are used. Example The following is an actual example of the interactions between identity-based-connection-pooling-enabled, oracle-proxy-session, and use-database-credentials. On the database side, the following objects are configured.- Database users scott; jdbcqa; jdbcqa3- Permission for proxy: alter user jdbcqa3 grant connect through jdbcqa;- Permission for proxy: alter user jdbcqa grant connect through jdbcqa; The following WebLogic Data Source objects are configured.- Users weblogic, wluser- Credential mapping “weblogic” to “scott”- Credential mapping "wluser" to "jdbcqa3"- Data source descriptor configured with user “jdbcqa”- All tests are run with Set Client ID set to true (more about that below).- All tests are run with oracle-proxy-session set to false (more about that below). The test program:- Runs in servlet- Authenticates to WLS as user “weblogic” Use DB Credentials Identity based getConnection(scott,***) getConnection(weblogic,***) getConnection(jdbcqa3,***) getConnection()  true  true Identity scottClient weblogicProxy null weblogic fails - not a db user User jdbcqa3Client weblogicProxy null Default user jdbcqaClient weblogicProxy null  false  true scott fails - not a WLS user User scottClient scottProxy null jdbcqa3 fails - not a WLS user User scottClient scottProxy null  true  false Proxy for scott fails weblogic fails - not a db user User jdbcqa3Client weblogicProxy jdbcqa Default user jdbcqaClient weblogicProxy null  false  false scott fails - not a WLS user Default user jdbcqaClient scottProxy null jdbcqa3 fails - not a WLS user Default user jdbcqaClient scottProxy null If Set Client ID is set to false, all cases would have Client set to null. If this was not an Oracle thin driver, the one case with the non-null Proxy in the above table would throw an exception because proxy session is only supported, implicitly or explicitly, with the Oracle thin driver. When oracle-proxy-session is set to true, the only cases that will pass (with a proxy of "jdbcqa") are the following.1. Setting use-database-credentials to true and doing getConnection(jdbcqa3,…) or getConnection().2. Setting use-database-credentials to false and doing getConnection(wluser, …) or getConnection(). Summary There are many options to choose from for data source security.  Considerations include the number and volatility of WLS and Database users, the granularity of data access, the depth of the security identity (property on the connection or a real user), performance, coordination of various components in the software stack, and driver capabilities.  Now that you have the big picture (remember that table in part 1), you can make a more informed choice.

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  • Mysql innoDB corruption after server crash

    - by Ward Loockx
    Yesterday my server died because an outage in the data center. Today it's back up, but having some problems with mysql. First of all my mysql server was not able to start. For this reason I deleted the files ib_logfile0 and ib_logfile1 in /var/lib/mysql folder (I still have the old failing files). After this my server was able to startup again. But now I see a lot of issues in the mysql log file. Sep 1 09:43:55 * mysqld: 120901 9:43:55 InnoDB: Error: page 70944 log sequence number 8 1483471899 Sep 1 09:43:55 * mysqld: InnoDB: is in the future! Current system log sequence number 5 612394935. Sep 1 09:43:55 * mysqld: InnoDB: Your database may be corrupt or you may have copied the InnoDB Sep 1 09:43:55 * mysqld: InnoDB: tablespace but not the InnoDB log files. See Sep 1 09:43:55 * mysqld: InnoDB: http://dev.mysql.com/doc/refman/5.1/en/forcing-recovery.html When I check the docs on mysql.com, I found that I need to recover my database with backups. I have a backup but not sure what's the good way on importing it. Or is there a way to recover without having to re-import the database again? So if I'm correct I need to put innodb_force_recovery to 4 in mysql and delete all current data and re-import? Is there a way to do this without having downtime? I also have one slave running. This slave has the current status now: Last_Error: Relay log read failure: Could not parse relay log event entry. The possible reasons are: the master's binary log is corrupted (you can check this by running 'mysqlbinlog' on the binary log), the slave's relay log is corrupted (you can check this by running 'mysqlbinlog' on the relay log), a network problem, or a bug in the master's or slave's MySQL code. If you want to check the master's binary log or slave's relay log, you will be able to know their names by issuing 'SHOW SLAVE STATUS' on this slave. How can I totally reset the slave after the new import on the master has happend? Hopefully we can find a solution without not to much downtime. Thanks!

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  • How do I organize a GUI application for passing around events and for setting up reads from a shared resource

    - by Savanni D'Gerinel
    My tools involved here are GTK and Haskell. My questions are probably pretty trivial for anyone who has done significant GUI work, but I've been off in the equivalent of CGI applications for my whole career. I'm building an application that displays tabular data, displays the same data in a graph form, and has an edit field for both entering new data and for editing existing data. After asking about sharing resources, I decided that all of the data involved will be stored in an MVar so that every component can just read the current state from the MVar. All of that works, but now it is time for me to rearrange the application so that it can be interactive. With that in mind, I have three widgets: a TextView (for editing), a TreeView (for displaying the data), and a DrawingArea (for displaying the data as a graph). I THINK I need to do two things, and the core of my question is, are these the right things, or is there a better way. Thing the first: All event handlers, those functions that will be called any time a redisplay is needed, need to be written at a high level and then passed into the function that actually constructs the widget to begin with. For instance: drawStatData :: DrawingArea -> MVar Core.ST -> (Core.ST -> SetRepWorkout.WorkoutStore) -> IO () createStatView :: (DrawingArea -> IO ()) -> IO VBox createUI :: MVar Core.ST -> (Core.ST -> SetRepWorkout.WorkoutStore) -> IO HBox createUI storeMVar field = do graphs <- createStatView (\area -> drawStatData area storeMVar field) hbox <- hBoxNew False 10 boxPackStart hbox graphs PackNatural 0 return hbox In this case, createStatView builds up a VBox that contains a DrawingArea to graph the data and potentially other widgets. It attaches drawStatData to the realize and exposeEvent events for the DrawingArea. I would do something similar for the TreeView, but I am not completely sure what since I have not yet done it and what I am thinking of would involve replacing the TreeModel every time the TreeView needs to be updated. My alternative to the above would be... drawStatData :: DrawingArea -> MVar Core.ST -> (Core.ST -> SetRepWorkout.WorkoutStore) -> IO () createStatView :: IO (VBox, DrawingArea) ... but in this case, I would arrange createUI like so: createUI :: MVar Core.ST -> (Core.ST -> SetRepWorkout.WorkoutStore) -> IO HBox createUI storeMVar field = do (graphbox, graph) <- createStatView (\area -> drawStatData area storeMVar field) hbox <- hBoxNew False 10 boxPackStart hbox graphs PackNatural 0 on graph realize (drawStatData graph storeMVar field) on graph exposeEvent (do liftIO $ drawStatData graph storeMVar field return ()) return hbox I'm not sure which is better, but that does lead me to... Thing the second: it will be necessary for me to rig up an event system so that various events can send signals all the way to my widgets. I'm going to need a mediator of some kind to pass events around and to translate application-semantic events to the actual events that my widgets respond to. Is it better for me to pass my addressable widgets up the call stack to the level where the mediator lives, or to pass the mediator down the call stack and have the widgets register directly with it? So, in summary, my two questions: 1) pass widgets up the call stack to a global mediator, or pass the global mediator down and have the widgets register themselves to it? 2) pass my redraw functions to the builders and have the builders attach the redraw functions to the constructed widgets, or pass the constructed widgets back and have a higher level attach the redraw functions (and potentially link some widgets together)? Okay, and... 3) Books or wikis about GUI application architecture, preferably coherent architectures where people aren't arguing about minute details? The application in its current form (displays data but does not write data or allow for much interaction) is available at https://bitbucket.org/savannidgerinel/fitness . You can run the application by going to the root directory and typing runhaskell -isrc src/Main.hs data/ or... cabal build dist/build/fitness/fitness data/ You may need to install libraries, but cabal should tell you which ones.

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  • SQL SERVER – Backing Up and Recovering the Tail End of a Transaction Log – Notes from the Field #042

    - by Pinal Dave
    [Notes from Pinal]: The biggest challenge which people face is not taking backup, but the biggest challenge is to restore a backup successfully. I have seen so many different examples where users have failed to restore their database because they made some mistake while they take backup and were not aware of the same. Tail Log backup was such an issue in earlier version of SQL Server but in the latest version of SQL Server, Microsoft team has fixed the confusion with additional information on the backup and restore screen itself. Now they have additional information, there are a few more people confused as they have no clue about this. Previously they did not find this as a issue and now they are finding tail log as a new learning. Linchpin People are database coaches and wellness experts for a data driven world. In this 42nd episode of the Notes from the Fields series database expert Tim Radney (partner at Linchpin People) explains in a very simple words, Backing Up and Recovering the Tail End of a Transaction Log. Many times when restoring a database over an existing database SQL Server will warn you about needing to make a tail end of the log backup. This might be your reminder that you have to choose to overwrite the database or could be your reminder that you are about to write over and lose any transactions since the last transaction log backup. You might be asking yourself “What is the tail end of the transaction log”. The tail end of the transaction log is simply any committed transactions that have occurred since the last transaction log backup. This is a very crucial part of a recovery strategy if you are lucky enough to be able to capture this part of the log. Most organizations have chosen to accept some amount of data loss. You might be shaking your head at this statement however if your organization is taking transaction logs backup every 15 minutes, then your potential risk of data loss is up to 15 minutes. Depending on the extent of the issue causing you to have to perform a restore, you may or may not have access to the transaction log (LDF) to be able to back up those vital transactions. For example, if the storage array or disk that holds your transaction log file becomes corrupt or damaged then you wouldn’t be able to recover the tail end of the log. If you do have access to the physical log file then you can still back up the tail end of the log. In 2013 I presented a session at the PASS Summit called “The Ultimate Tail Log Backup and Restore” and have been invited back this year to present it again. During this session I demonstrate how you can back up the tail end of the log even after the data file becomes corrupt. In my demonstration I set my database offline and then delete the data file (MDF). The database can’t become more corrupt than that. I attempt to bring the database back online to change the state to RECOVERY PENDING and then backup the tail end of the log. I can do this by specifying WITH NO_TRUNCATE. Using NO_TRUNCATE is equivalent to specifying both COPY_ONLY and CONTINUE_AFTER_ERROR. It as its name says, does not try to truncate the log. This is a great demo however how could I achieve backing up the tail end of the log if the failure destroys my entire instance of SQL and all I had was the LDF file? During my demonstration I also demonstrate that I can attach the log file to a database on another instance and then back up the tail end of the log. If I am performing proper backups then my most recent full, differential and log files should be on a server other than the one that crashed. I am able to achieve this task by creating new database with the same name as the failed database. I then set the database offline, delete my data file and overwrite the log with my good log file. I attempt to bring the database back online and then backup the log with NO_TRUNCATE just like in the first example. I encourage each of you to view my blog post and watch the video demonstration on how to perform these tasks. I really hope that none of you ever have to perform this in production, however it is a really good idea to know how to do this just in case. It really isn’t a matter of “IF” you will have to perform a restore of a production system but more of a “WHEN”. Being able to recover the tail end of the log in these sever cases could be the difference of having to notify all your business customers of data loss or not. If you want me to take a look at your server and its settings, or if your server is facing any issue we can Fix Your SQL Server. Note: Tim has also written an excellent book on SQL Backup and Recovery, a must have for everyone. Reference: Pinal Dave (http://blog.sqlauthority.com)Filed under: Notes from the Field, PostADay, SQL, SQL Authority, SQL Performance, SQL Query, SQL Server, SQL Tips and Tricks, T SQL

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  • Why doesn't video run smoothly on my laptop anymore?

    - by andygrunt
    This might be an impossible question to answer remotely but I figure there may be some common causes that people can suggest so I think it's worth asking... Video no longer plays smoothly on my laptop. It used to but not for a while now. For example, playing a video on YouTube is pretty typical: I press play (making sure it's not on HD or even HQ) and the video buffers a little then starts to play. At first it plays fine then the video starts to stutter, turning into a slideshow while the sound continues to play smoothly. If I try playing the same video on my Playstation 3 (which is linked to the same network) it plays smoothly so it can't be the connection. Another example is streaming DivX videos. Again, I wait while it buffers and it starts but very soon, instead of a slideshow, this time the video just plays slowly while the sound continues as normal (instantly getting out of sync). Even if I let the video fully load before pressing play (i.e. it's no longer streaming), it still behaves the same way. I can even let it load 100% then save the file to hard disk and use VLC player to view it, and the same thing happens. I'm using an old laptop running Windows XP. For the past several years it's been connected to the router via Wi-Fi but in the past few days I've changed that to a network cable (like my PS3) but that hasn't helped. Yes, I regularly install various bits and pieces of software but nothing that I can identify as being the cause. So, are there known causes of this sort of behaviour and if so, what can I do to fix it? Thanks. Update to answer a few questions... Laptop Spec' (note: video has played back fine for the majority of time I've had the laptop) Toshiba Satellite 1900-603 (possibly called something else outside the UK) Intel Pentium 4 2.2 Ghz Processor Originally had 512Mb memory but recently doubled that to 1 Gig of memory Graphics: ATI Mobility Radeon 16Mb DDR VRAM Windows XP SP3 (Home edition) Over the years I've done several things to speed it up (disabling indexing etc) and am generally happy with the performance. I also regularly have a clear out of old software (if for no other reason than the laptop only has a 40Gb hard disk) and use CCleaner and Glary Utilities to strip out much of the crap from my system. Also recently (after doubling the memory), I've tried a few new things which might be likely candidates for slowing the video down such as Rocketdock, Jingle keyboard (which gives an old style 'clacky' typewriter sound when I type - love it), SugarSync, Taskbar Shuffle. However, the video doesn't play smoothly even when I try quit all these apps.

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  • Local admin password recovery: Windows Vista

    - by Jim Dennis
    I am faced with an unsettling situation. A friend of my father's has rather suddenly become a widower. Naturally they've taken care of the bank accounts and all the normal mundane things that people have been doing for a century or so. However, she was the computer user of the household. He was aware that they had some online banking stuff and bill paying stuff ... and that she spent lots of time on FaceBook and stuff like that. However, he doesn't know what her local passwords were (actually only vaguely aware that her couple of desktop and couple of laptop system even had passwords). He's never heard of "admin" passwords so that's no good either. In the past I've used KNOPPIX and the old LinuxCare "bootable business card" to recover NT passwords. I've never done this with MS Windows Vista. So, I'm looking for the best advice on how to do this. Naturally I do have physical access to the systems (the two laptops are charging across the room from me; and her old desktop systems are, naturally, still back at his place). Getting it right is much more important than fast or easy (I don't want to mess up those filesystems and possibly lose some photos or other stuff that he or his kids or grandkids will want). (BTW: if anyone things this is some social engineering hack to play upon the sympathies of the community to get the information I'm asking for ... think about it for a minute. I know about IRC and the "warez" boards. I know I can find this stuff out there if I dig enough. I'm just asking here because it'll hopefully be faster and, secondarily to raise awareness. As more of us put more of our lives online ... as we get older and as places like FaceBook continue to widen the appeal of computing to a broader segment of older people ... we are, as computer nerds, going to see a lot more of this. Survivors will needs us to be careful, sensitive and ethically responsible as they try to recover those bits of legacy during their bereavement. I can now tell you, first hand, it sucks!)

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  • Why bother writing an Windows 8 app?

    - by Dennis Vroegop
    So you want to know more about development for Window 8. Great! There are lots of reasons you should be excited about this. Since I don’t know why YOU are interested in this, I’ll make a list of reasons people can choose from. (as a side note: whenever I talk about Win8 development I am referring to the Metro Style / WinRt side of things. Apps for the ‘classic’ desktop side of Win8 on Intel are business as usual…) So… Why would you care about making an app for Windows 8? 1. It’s cool. Let’s not beat around the bush: if you like development for a hobby then you’ll love to work on this new platform. You can create apps in a relative short time (short time as in compared to writing a new CRM system) and that makes it great for a hobby product. 2. You’ll stand out. Hey, we all need an ego boost every now and then. We all need to feel special. So if you can manage to be one of the first to have you app in the Store then you’ll likely to be noticed. Just close your eyes for a moment and image you standing in a bar. It’s crowded, and then you casually say “Oh yeah, I just had my app certified and it’s in the Win8 store now”. People will stop talking, will offer you drinks and beautiful women / gorgeous man / furry creatures from Alpha Centauri (whatever your preferences are) will propose. Or maybe not. Anyway…. 3. Make some cash! IDC predicts there will be about 350,000,000 Windows 8 licenses sold in the next year. Think about that number. 350,000,000. And they all have access to the Store. Where you’re app will be. With one little click they can select it, download and somehow magically $1.00 or $2.00 from their bank account is transferred to yours. Now, I am not saying that all of those people will download and buy your app but what if only 1% of them did? Remember: there aren’t that many apps available yet….. 4. Learn. Creating new small apps is a great way to learn new stuff. Yes, you could read about it (on this blog for instance) but the only way to learn something is to do it. So be prepared for the future and learn something new by doing it.Write an app! Now! 5. The biggie (for me at least): it’s fun. Even if you remove the points above it’s still fun to write for these devices and this platform. Now some of you will say : “But why not write a great app for IOS or Android?” I think this is a valid question. Of course the novelty of the platform wears out and points 2 and 3 from above list will not be as relevant as it is today. But still 1 4 and 5 remain. And don’t forget: if you already work on the Microsoft platform it’s not that hard to learn this new Win8 stuff. If you have done some XAML development (be it WPF or Silverlight) you are almost there in becoming a good Win8 developer. So you’ll be more productive much sooner than when you have to learn Objective C or Java. Even if you’re a HTML / Javascript developer (I say developer here, not designer) you’ll be up to speed on Win8 development pretty soon. Yes, you, that funky Web Developer who lives and breathes HTML5, CSS3 and JavaScript / Node.Js / JQuery: you too can be a Win8 developer. A first class Win8 developer! So.. Download the stuff you need from http://dev.windows.com install Windows 8 and Visual Studio 12 and by the time you’re ready I’ll be working on the next article: how to do all this? Happy coding!

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  • Windows Phone 8 Announcement

    - by Tim Murphy
    As if the Surface announcement on Monday wasn’t exciting enough, today Microsoft announce that Windows Phone 8 will be coming this fall.  That itself is great news, but the features coming were like confetti flying in all different directions.  Given this speed I couldn’t capture every feature they covered.  A summary of what I did capture is listed below starting with their eight main features. Common Core The first thing that they covered is that Windows Phone 8 will share a core OS with Windows 8.  It will also run natively on multiple cores.  They mentioned that they have run it on up to 64 cores to this point.  The phones as you might expect will at least start as dual core.  If you remember there were metrics saying that Windows Phone 7 performed operations faster on a single core than other platforms did with dual cores.  The metrics they showed here indicate that Windows Phone 8 runs faster on comparable dual core hardware than other platforms. New Screen Resolutions Screen resolution has never been an issue for me, but it has been a criticism of Windows Phone 7 in the media.  Windows Phone 8 will supports three screen resolutions: WVGA 800 x 480, WXGA 1280 x 768, and 720 1280x720.  Hopefully this makes pixel counters a little happier. MicroSD Support This was one of my pet peeves when I got my Samsung Focus. With Windows Phone 8 the operating system will support adding MicroSD cards after initial setup.  Of course this is dependent on the hardware company on implementing it, but I think we have seen that even feature phone manufacturers have not had a problem supporting this in the past. NFC NFC has been an anticipated feature for some time.  What Microsoft showed today included the fact that they didn’t just want it to be for the phone.  There is cross platform NFC functionality between Windows Phone 8 and Windows 8.  The demos , while possibly a bit fanciful, showed would could be achieved even in a retail environment.  We are getting closer and closer to a Minority Report world with these technologies. Wallet Windows Phone 8 isn’t the first platform to have a wallet concept.  What they have done to differentiate themselves is to make it sot that it is not dependent on a SIM type chip like other platforms.  They have also expanded the concept beyond just banks to other types of credits such as airline miles. Nokia Mapping People have been envious of the Lumia phones having the Nokia mapping software.  Now all Windows Phone 8 devices will use NavTeq data and will have the capability to run in an offline fashion.  This is a major step forward from the Bing “touch for the next turn” maps. IT Administration The lack of features for enterprise administration and deployment was a complaint even before the Windows Phone 7 was released.  With the Windows Phone 8 release such features as Bitlocker and Secure boot will be baked into the OS. We will also have the ability to privately sign and distribute applications. Changing Start Screen Joe Belfiore made a big deal about this aspect of the new release.  Users will have more color themes available to them and the live tiles will be highly customizable. You will have the ability to resize and organize the tiles in a more dynamic way.  This allows for less important tiles or ones with less information to be made smaller.  And There Is More So what other tidbits came out of the presentation?  Later this summer the API for WP8 will be available.  There will be developer events coming to a city near you.  Another announcement of interest to developers is the ability to write applications at a native code level.  This is a boon for game developers and those who need highly efficient applications. As a topper on the cake there was mention of in app payment. On the consumer side we also found out that all updates will be available over the air.  Along with this came the fact that Microsoft will support all devices with updates for at least 18 month and you will be able to subscribe for early updates.  Update coming for Windows Phone 7.5 customers to WP7.8.  The main enhancement will be the new live tile features.  The big bonus is that the update will bypass the carriers.  I would assume though that you will be brought up to date with all previous patches that your carrier may not have released. There is so much more, but that is enough for one post.  Needless to say, EXCITING! del.icio.us Tags: Windows Phone 8,WP8,Windows Phone 7,WP7,Announcements,Microsoft

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  • Excessive Outbound DNS Traffic

    - by user1318414
    I have a VPS system which I have had for 3 years on one host without issue. Recently, the host started sending an extreme amount of outbound DNS traffic to 31.193.132.138. Due to the way that Linode responded to this, I have recently left Linode and moved to 6sync. The server was completely rebuilt on 6sync with the exception of postfix mail configurations. Currently, the daemons run are as follows: sshd nginx postfix dovecot php5-fpm (localhost only) spampd (localhost only) clamsmtpd (localhost only) Given that the server was 100% rebuilt, I can't find any serious exploits against the above stated daemons, passwords have changed, ssh keys don't even exist on the rebuild yet, etc... it seems extremely unlikely that this is a compromise which is being used to DoS the address. The provided IP is noted online as a known SPAM source. My initial assumption was that it was attempting to use my postfix server as a relay, and the bogus addresses it was providing were domains with that IP registered as their nameservers. I would imagine given my postfix configuration that DNS queries for things such as SPF information would come in with equal or greater amount than the number of attempted spam e-mails sent. Both Linode and 6Sync have said that the outbound traffic is extremely disproportionate. The following is all the information I received from Linode regarding the outbound traffic: 21:28:28.647263 IP 97.107.134.33.32775 > 31.193.132.138.53: 28720 op8+% [b2&3=0x4134] [17267a] [30550q] [28773n] [14673au][|domain] 21:28:28.647264 IP 97.107.134.33 > 31.193.132.138: udp 21:28:28.647264 IP 97.107.134.33.32775 > 31.193.132.138.53: 28720 op8+% [b2&3=0x4134] [17267a] [30550q] [28773n] [14673au][|domain] 21:28:28.647265 IP 97.107.134.33 > 31.193.132.138: udp 21:28:28.647265 IP 97.107.134.33.32775 > 31.193.132.138.53: 28720 op8+% [b2&3=0x4134] [17267a] [30550q] [28773n] [14673au][|domain] 21:28:28.647266 IP 97.107.134.33 > 31.193.132.138: udp 6sync cannot confirm whether or not the recent spike in outbound traffic was to the same IP or over DNS, but I have presumed as such. For now my server is blocking the entire 31.0.0.0/8 subnet to help deter this while I figure it out. Anyone have any idea what is going on?

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