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  • 7 Good Reasons to Upgrade E-Business Suite to the cloud

    - by Lisa Schwartz
    v\:* {behavior:url(#default#VML);} o\:* {behavior:url(#default#VML);} w\:* {behavior:url(#default#VML);} .shape {behavior:url(#default#VML);} As promised here is blog Part 2: Why Upgrade to Oracle E-Business Suite 12 in the cloud? 7 Good Reasons to Upgrade to E-Business Suite 12 in the Cloud: 1)   Take advantage of new and improved features: from global sub-ledger accounting to mobile access for supply chain management to built-in extensions for information search and discovery. If you haven’t checked out the latest features yet, there are over 1000 EBS 12 enhancements. 2) Plan now to address any ongoing Oracle Support considerations and regulatory compliance requirements. EBS Release 11 support is ending soon. Based upon that information alone, you should have an EBS upgrade strategy and planning well underway. 3) Customizations got you worried? Expedite your next Oracle E-Business Suite upgrade – have Oracle identify all customizations, reduce un-needed customizations (EBS 12 has built-in many of your customizations) and during the upgrade keep all necessary customizations to run your business. 4) Migrating EBS to the cloud allows parallel migration and testing. Therefore no extra hardware purchases for the testing and upgrade. Business disruption is minimized. And, by moving to the cloud, this provides for smoother future upgrades that are based on your own timeline. 5) Oracle Experts will upgrade and run your EBS applications for you in the cloud. Free your IT resources to develop new services and work on projects that are critical to business innovation and competitiveness. Your IT resources will not be inundated with upgrade tasks!      6) Reallocate precious IT dollars to other projects, eliminate CapEx costs. 7) Oracle minimizes business risk by having enterprise class cloud services under stringent SLAs designed to run your business applications for you such as: a. Enterprise grade infrastructure b. World-class security and identity management c. Best practices in regulatory compliance: from classified federal gov’t standards, to healthcare HIPPA standards to meeting Financial Services requirements (PCI DSS) Normal 0 false false false EN-US X-NONE X-NONE /* Style Definitions */ table.MsoNormalTable {mso-style-name:"Table Normal"; mso-tstyle-rowband-size:0; mso-tstyle-colband-size:0; mso-style-noshow:yes; mso-style-priority:99; mso-style-parent:""; mso-padding-alt:0in 5.4pt 0in 5.4pt; mso-para-margin-top:0in; mso-para-margin-right:0in; mso-para-margin-bottom:10.0pt; mso-para-margin-left:0in; line-height:115%; mso-pagination:widow-orphan; font-size:11.0pt; font-family:"Calibri","sans-serif"; mso-ascii-font-family:Calibri; mso-ascii-theme-font:minor-latin; mso-hansi-font-family:Calibri; mso-hansi-theme-font:minor-latin; mso-bidi-font-family:"Times New Roman"; mso-bidi-theme-font:minor-bidi;} 7 Normal 0 false false false EN-US X-NONE X-NONE /* Style Definitions */ table.MsoNormalTable {mso-style-name:"Table Normal"; mso-tstyle-rowband-size:0; mso-tstyle-colband-size:0; mso-style-noshow:yes; mso-style-priority:99; mso-style-parent:""; mso-padding-alt:0in 5.4pt 0in 5.4pt; mso-para-margin-top:0in; mso-para-margin-right:0in; mso-para-margin-bottom:10.0pt; mso-para-margin-left:0in; line-height:115%; mso-pagination:widow-orphan; font-size:11.0pt; font-family:"Calibri","sans-serif"; mso-ascii-font-family:Calibri; mso-ascii-theme-font:minor-latin; mso-hansi-font-family:Calibri; mso-hansi-theme-font:minor-latin; mso-bidi-font-family:"Times New Roman"; mso-bidi-theme-font:minor-bidi;} Next Step: To help you upgrade and get to the cloud in the shortest period of  time, Oracle has a program called Oracle Upgrade Factory for Oracle E-Business Suite 12. It offers a unique approach, seamlessly bundling Managed Cloud Services and Oracle Consulting Services together for an entire Oracle E-Business Suite upgrade and migration to a managed private  cloud. Read the Oracle Upgrade Factory Solution Brief here. Normal 0 false false false false EN-US X-NONE X-NONE /* Style Definitions */ table.MsoNormalTable {mso-style-name:"Table Normal"; mso-tstyle-rowband-size:0; mso-tstyle-colband-size:0; mso-style-noshow:yes; mso-style-priority:99; mso-style-parent:""; mso-padding-alt:0in 5.4pt 0in 5.4pt; mso-para-margin-top:0in; mso-para-margin-right:0in; mso-para-margin-bottom:10.0pt; mso-para-margin-left:0in; line-height:115%; mso-pagination:widow-orphan; font-size:11.0pt; font-family:"Calibri","sans-serif"; mso-ascii-font-family:Calibri; mso-ascii-theme-font:minor-latin; mso-hansi-font-family:Calibri; mso-hansi-theme-font:minor-latin; mso-bidi-font-family:"Times New Roman"; mso-bidi-theme-font:minor-bidi;}

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  • Solaris 11 SRU / Update relationship explained, and blackout period on delivery of new bug fixes eliminated

    - by user12244672
    Relationship between SRUs and Update releases As you may know, Support Repository Updates (SRUs) for Oracle Solaris 11 are released monthly and are available to customers with an appropriate support contract.  SRUs primarily deliver bug fixes.  They may also deliver low risk feature enhancements. Solaris Update are typically released once or twice a year, containing support for new hardware, new software feature enhancements, and all bug fixes available at the time the Update content was finalized.  They also contain a significant number of new bug fixes, for issues found internally in Oracle and complex customer bug fixes which  require significant "soak" time to ensure their efficacy prior to release. Changes to SRU and Update Naming Conventions We're changing the naming convention of Update releases from a date based format such as Oracle Solaris 10 8/11 to a simpler "dot" version numbering, e.g. Oracle Solaris 11.1. Oracle Solaris 11 11/11 (i.e. the initial Oracle Solaris 11 release) may be referred to as 11.0. SRUs will simply be named as "dot.dot" releases, e.g. Oracle Solaris 11.1.1, for SRU1 after Oracle Solaris 11.1. Many Oracle products and infrastructure tools such as BugDB and MOS are tailored towards this "dot.dot" style of release naming, so these name changes align Oracle Solaris with these conventions. No Blackout Periods on Bug Fix Releases The Oracle Solaris 11 release process has been enhanced to eliminate blackout periods on the delivery of new bug fixes to customers. Previously, Oracle Solaris Updates were a superset of all preceding bug fix deliveries.  This made for a very simple update message - that which releases later is always a superset of that which was delivered previously. However, it had a downside.  Once the contents of an Update release were frozen prior to release, the release of new bug fixes for customer issues was also frozen to maintain the Update's superset relationship. Since the amount of change allowed into the final internal builds of an Update release is reduced to mitigate risk, this throttling back also impacted the release of new bug fixes to customers. This meant that there was effectively a 6 to 9 week hiatus on the release of new bug fixes prior to the release of each Update.  That wasn't good for customers awaiting critical bug fixes. We've eliminated this hiatus on the delivery of new bug fixes in Oracle Solaris 11 by allowing new bug fixes to continue to be released in SRUs even after the contents of the next Update release have been frozen. The release of SRUs will remain contiguous, with the first SRU released after the Update release effectively being a superset of both the the Update release and all preceding SRUs*.  That is, later SRUs are supersets of the content of previous SRUs. Therefore, the progression path from the final SRUs prior to the Update release is to the first SRU after the Update release, rather than to the Update release itself. The timeline / logical sequence of releases can be shown as follows: Updates: 11.0                                                11.1                               11.2     etc.                  \                                                         \                                    \ SRUs:       11.0.1, 11.0.2,...,11.0.12, 11.0.13, 11.1.1, 11.1.2,...,11.1.x, 11.2.1, etc. For example, for systems with Oracle Solaris 11 11/11 SRU12.4 or later installed, the recommended update path is to Oracle Solaris 11.1.1 (i.e. SRU1 after Solaris 11.1) or later rather than to the Solaris 11.1 release itself.  This will ensure no bug fixes are "lost" during the update. If for any reason you do wish to update from SRU12.4 or later to the 11.1 release itself - for example to update a test system - the instructions to do so are in the SRU12.4 README, https://updates.oracle.com/Orion/Services/download?type=readme&aru=15564533 For systems with Oracle Solaris 11 11/11 SRU11.4 or earlier installed, customers can update to either the 11.1 release or any 11.1 SRU as both will be supersets of their current version. Please do read the README of the SRU you are updating to, as it will contain important installation instructions which will save you time and effort. *Nerdy details: SRUs only contain the latest change delta relative to the Update on which they are based.  Their dependencies will, however, effectively pull in the Update content.  Customers maintaining a local Repo (e.g. behind their firewall), need to add both the 11.1 content and the relevant SRU content to their Repo, to enable the SRU's dependencies to be resolved.  Both will be available from the standard Support Repo and from MOS.  This is no different to existing SRUs for Oracle Solaris 11.0, whereby you may often get away with using just the SRU content to update, but the original 11.0 content may be needed in the Repo to resolve dependencies.

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  • Why people don't patch and upgrade?!?

    - by Mike Dietrich
    Discussing the topic "Why Upgrade" or "Why not Upgrade" is not always fun. Actually the arguments repeat from customer to customer. Typically we hear things such as: A PSU or Patch Set introduces new bugs A new PSU or Patch Set introduces new features which lead to risk and require application verification  Patching means risk Patching changes the execution plans Patching requires too much testing Patching is too much work for our DBAs Patching costs a lot of money and doesn't pay out And to be very honest sometimes it's hard for me to stay calm in such discussions. Let's discuss some of these points a bit more in detail. A PSU or Patch Set introduces new bugsWell, yes, that is true as no software containing more than some lines of code is bug free. This applies to Oracle's code as well as too any application or operating system code. But first of all, does that mean you never patch your OS because the patch may introduce new flaws? And second, what is the point of saying "it introduces new bugs"? Does that mean you will never get rid of the mean issues we know about and we fixed already? Scroll down from MOS Note:161818.1 to the patch release you are on, no matter if it's 10.2.0.4 or 11.2.0.3 and check for the Known Issues And Alerts.Will you take responsibility to know about all these issues and refuse to upgrade to 11.2.0.4? I won't. A new PSU or Patch Set introduces new featuresOk, we can discuss that. Offering new functionality within a database patch set is a dubious thing. It has advantages such as in 11.2.0.4 where we backported Database Redaction to. But this is something you will only use once you have an Advanced Security license. I interpret that statement I've heard quite often from customers in a different way: People don't want to get surprises such as new behaviour. This certainly gives everybody a hard time. And we've had many examples in the past (SESSION_CACHED_CURSROS in 10.2.0.4,  _DATAFILE_WRITE_ERRORS_CRASH_INSTANCE in 11.2.0.2 and others) where those things weren't documented, not even in the README. Thanks to many friends out there I learned about those as well. So new behaviour is the topic people consider as risky - not really new features. And just to point this out: A PSU never brings in new features or new behaviour by definition! Patching means riskDoes it really mean risk? Yes, there were issues in the past (and sometimes in the present as well) where a patch didn't get installed correctly. But personally I consider it way more risky to not patch. Keep that in mind: The day Oracle publishes an PSU (or CPU) containing security fixes all the great security experts out there go public with their findings as well. So from that day on even my grandma can find out about those issues and try to attack somebody. Now a lot of people say: "My database does not face the internet." And I will answer: "The enemy is sitting already behind your firewalls. And knows potentially about these things." My statement: Not patching introduces way more risk to your environment than patching. Seriously! Patching changes the execution plansDo they really? I agree - there's a very small risk for this happening with Patch Sets. But not with PSUs or CPUs as they contain no optimizer fixes changing behaviour (but they may contain fixes curing wrong-query-result-bugs). But what's the point of a changing execution plan? In Oracle Database 11g it is so simple to be prepared. SQL Plan Management is a free EE feature - so once that occurs you'll put the plan into the Plan Baseline. Basta! Yes, you wouldn't like to get such surprises? Than please use the SQL Performance Analyzer (SPA) from Real Application Testing and you'll detect that easily upfront in minutes. And not to forget this, a plan change can also be very positive!Yes, there's a little risk with a database patchset - and we have many possibilites to detect this before patching. Patching requires too much testingWell, does it really? I have seen in the past 12 years how people test. There are very different efforts and approaches on this. I have seen people spending a hell of money on licenses or on project team staffing. And I have seen people sailing blindly without any tests just going the John-Wayne-approach.Proper tools will allow you to test easily without too much efforts. See the paragraph above. We have used Real Application Testing in so many customer projects reducing the amount of work spend on testing by over 50%. But apart from that at some point you will have to stop testing. If you don't you'll get lost and you'll burn money. There's no 100% guaranty. You will have to deal with a little risk as reaching the final 5% of certainty will cost you the same as it did cost to reach 95%. And doing this will lead to abnormal long product cycles that you'll run behind forever. And this will cost even more money. Patching is too much work for our DBAsPatching is a lot of work. I agree. And it's no fun work. It's boring, annoying. You don't learn much from that. That's why you should try to automate this task. Use the Database's Lifecycle Management Pack. And don't cry about the fact that it costs money. Yes it does. But it will ease the process and you'll save a lot of costs as you don't waste your valuable time with patching. Or use Oracle Database 12c Oracle Multitenant and patch either by unplug/plug or patch an entire container database with all PDBs with one patch in one task. We have customer reference cases proofing it saved them 75% of time, effort and cost since they've used Lifecycle Management Pack. So why don't you use it? Patching costs a lot of money and doesn't pay outWell, see my statements in the paragraph above. And it pays out as flying with a database with 100 known critical flaws in it which are already fixed by Oracle (such as in the Oct 2013 PSU for Oracle Database 12c) will cost ways more in case of failure or even data loss. Bet with me? Let me finally ask you some questions. What cell phone are you using and which OS does it run? Do you have an iPhone 5 and did you upgrade already to iOS 7.0.3? I've just encountered on mine that the alarm (which I rely on when traveling) has gotten now a dependency on the physical switch "sound on/off". If it is switched to "off" physically the alarm rings "silently". What a wonderful example of a behaviour change coming in with a patch set. Will this push you to stay with iOS5 or iOS6? No, because those have security flaws which won't be fixed anymore. What browser are you surfing with? Do you use Mozilla 3.6? Well, congratulations to all the hackers. It will be easy for them to attack you and harm your system. I'd guess you have the auto updater on.  Same for Google Chrome, Safari, IE. Right? -Mike The T.htmtableborders, .htmtableborders td, .htmtableborders th {border : 1px dashed lightgrey ! important;} html, body { border: 0px; } body { background-color: #ffffff; } img, hr { cursor: default }

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  • Real tortoises keep it slow and steady. How about the backups?

    - by Maria Zakourdaev
      … Four tortoises were playing in the backyard when they decided they needed hibiscus flower snacks. They pooled their money and sent the smallest tortoise out to fetch the snacks. Two days passed and there was no sign of the tortoise. "You know, she is taking a lot of time", said one of the tortoises. A little voice from just out side the fence said, "If you are going to talk that way about me I won't go." Is it too much to request from the quite expensive 3rd party backup tool to be a way faster than the SQL server native backup? Or at least save a respectable amount of storage by producing a really smaller backup files?  By saying “really smaller”, I mean at least getting a file in half size. After Googling the internet in an attempt to understand what other “sql people” are using for database backups, I see that most people are using one of three tools which are the main players in SQL backup area:  LiteSpeed by Quest SQL Backup by Red Gate SQL Safe by Idera The feedbacks about those tools are truly emotional and happy. However, while reading the forums and blogs I have wondered, is it possible that many are accustomed to using the above tools since SQL 2000 and 2005.  This can easily be understood due to the fact that a 300GB database backup for instance, using regular a SQL 2005 backup statement would have run for about 3 hours and have produced ~150GB file (depending on the content, of course).  Then you take a 3rd party tool which performs the same backup in 30 minutes resulting in a 30GB file leaving you speechless, you run to management persuading them to buy it due to the fact that it is definitely worth the price. In addition to the increased speed and disk space savings you would also get backup file encryption and virtual restore -  features that are still missing from the SQL server. But in case you, as well as me, don’t need these additional features and only want a tool that performs a full backup MUCH faster AND produces a far smaller backup file (like the gain you observed back in SQL 2005 days) you will be quite disappointed. SQL Server backup compression feature has totally changed the market picture. Medium size database. Take a look at the table below, check out how my SQL server 2008 R2 compares to other tools when backing up a 300GB database. It appears that when talking about the backup speed, SQL 2008 R2 compresses and performs backup in similar overall times as all three other tools. 3rd party tools maximum compression level takes twice longer. Backup file gain is not that impressive, except the highest compression levels but the price that you pay is very high cpu load and much longer time. Only SQL Safe by Idera was quite fast with it’s maximum compression level but most of the run time have used 95% cpu on the server. Note that I have used two types of destination storage, SATA 11 disks and FC 53 disks and, obviously, on faster storage have got my backup ready in half time. Looking at the above results, should we spend money, bother with another layer of complexity and software middle-man for the medium sized databases? I’m definitely not going to do so.  Very large database As a next phase of this benchmark, I have moved to a 6 terabyte database which was actually my main backup target. Note, how multiple files usage enables the SQL Server backup operation to use parallel I/O and remarkably increases it’s speed, especially when the backup device is heavily striped. SQL Server supports a maximum of 64 backup devices for a single backup operation but the most speed is gained when using one file per CPU, in the case above 8 files for a 2 Quad CPU server. The impact of additional files is minimal.  However, SQLsafe doesn’t show any speed improvement between 4 files and 8 files. Of course, with such huge databases every half percent of the compression transforms into the noticeable numbers. Saving almost 470GB of space may turn the backup tool into quite valuable purchase. Still, the backup speed and high CPU are the variables that should be taken into the consideration. As for us, the backup speed is more critical than the storage and we cannot allow a production server to sustain 95% cpu for such a long time. Bottomline, 3rd party backup tool developers, we are waiting for some breakthrough release. There are a few unanswered questions, like the restore speed comparison between different tools and the impact of multiple backup files on restore operation. Stay tuned for the next benchmarks.    Benchmark server: SQL Server 2008 R2 sp1 2 Quad CPU Database location: NetApp FC 15K Aggregate 53 discs Backup statements: No matter how good that UI is, we need to run the backup tasks from inside of SQL Server Agent to make sure they are covered by our monitoring systems. I have used extended stored procedures (command line execution also is an option, I haven’t noticed any impact on the backup performance). SQL backup LiteSpeed SQL Backup SQL safe backup database <DBNAME> to disk= '\\<networkpath>\par1.bak' , disk= '\\<networkpath>\par2.bak', disk= '\\<networkpath>\par3.bak' with format, compression EXECUTE master.dbo.xp_backup_database @database = N'<DBName>', @backupname= N'<DBName> full backup', @desc = N'Test', @compressionlevel=8, @filename= N'\\<networkpath>\par1.bak', @filename= N'\\<networkpath>\par2.bak', @filename= N'\\<networkpath>\par3.bak', @init = 1 EXECUTE master.dbo.sqlbackup '-SQL "BACKUP DATABASE <DBNAME> TO DISK= ''\\<networkpath>\par1.sqb'', DISK= ''\\<networkpath>\par2.sqb'', DISK= ''\\<networkpath>\par3.sqb'' WITH DISKRETRYINTERVAL = 30, DISKRETRYCOUNT = 10, COMPRESSION = 4, INIT"' EXECUTE master.dbo.xp_ss_backup @database = 'UCMSDB', @filename = '\\<networkpath>\par1.bak', @backuptype = 'Full', @compressionlevel = 4, @backupfile = '\\<networkpath>\par2.bak', @backupfile = '\\<networkpath>\par3.bak' If you still insist on using 3rd party tools for the backups in your production environment with maximum compression level, you will definitely need to consider limiting cpu usage which will increase the backup operation time even more: RedGate : use THREADPRIORITY option ( values 0 – 6 ) LiteSpeed : use  @throttle ( percentage, like 70%) SQL safe :  the only thing I have found was @Threads option.   Yours, Maria

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  • Best Practices - which domain types should be used to run applications

    - by jsavit
    This post is one of a series of "best practices" notes for Oracle VM Server for SPARC (formerly named Logical Domains) One question that frequently comes up is "which types of domain should I use to run applications?" There used to be a simple answer in most cases: "only run applications in guest domains", but enhancements to T-series servers, Oracle VM Server for SPARC and the advent of SPARC SuperCluster have made this question more interesting and worth qualifying differently. This article reviews the relevant concepts and provides suggestions on where to deploy applications in a logical domains environment. Review: division of labor and types of domain Oracle VM Server for SPARC offloads many functions from the hypervisor to domains (also called virtual machines). This is a modern alternative to using a "thick" hypervisor that provides all virtualization functions, as in traditional VM designs, This permits a simpler hypervisor design, which enhances reliability, and security. It also reduces single points of failure by assigning responsibilities to multiple system components, which further improves reliability and security. In this architecture, management and I/O functionality are provided within domains. Oracle VM Server for SPARC does this by defining the following types of domain, each with their own roles: Control domain - management control point for the server, used to configure domains and manage resources. It is the first domain to boot on a power-up, is an I/O domain, and is usually a service domain as well. I/O domain - has been assigned physical I/O devices: a PCIe root complex, a PCI device, or a SR-IOV (single-root I/O Virtualization) function. It has native performance and functionality for the devices it owns, unmediated by any virtualization layer. Service domain - provides virtual network and disk devices to guest domains. Guest domain - a domain whose devices are all virtual rather than physical: virtual network and disk devices provided by one or more service domains. In common practice, this is where applications are run. Typical deployment A service domain is generally also an I/O domain: otherwise it wouldn't have access to physical device "backends" to offer to its clients. Similarly, an I/O domain is also typically a service domain in order to leverage the available PCI busses. Control domains must be I/O domains, because they boot up first on the server and require physical I/O. It's typical for the control domain to also be a service domain too so it doesn't "waste" the I/O resources it uses. A simple configuration consists of a control domain, which is also the one I/O and service domain, and some number of guest domains using virtual I/O. In production, customers typically use multiple domains with I/O and service roles to eliminate single points of failure: guest domains have virtual disk and virtual devices provisioned from more than one service domain, so failure of a service domain or I/O path or device doesn't result in an application outage. This is also used for "rolling upgrades" in which service domains are upgraded one at a time while their guests continue to operate without disruption. (It should be noted that resiliency to I/O device failures can also be provided by the single control domain, using multi-path I/O) In this type of deployment, control, I/O, and service domains are used for virtualization infrastructure, while applications run in guest domains. Changing application deployment patterns The above model has been widely and successfully used, but more configuration options are available now. Servers got bigger than the original T2000 class machines with 2 I/O busses, so there is more I/O capacity that can be used for applications. Increased T-series server capacity made it attractive to run more vertical applications, such as databases, with higher resource requirements than the "light" applications originally seen. This made it attractive to run applications in I/O domains so they could get bare-metal native I/O performance. This is leveraged by the SPARC SuperCluster engineered system, announced a year ago at Oracle OpenWorld. In SPARC SuperCluster, I/O domains are used for high performance applications, with native I/O performance for disk and network and optimized access to the Infiniband fabric. Another technical enhancement is the introduction of Direct I/O (DIO) and Single Root I/O Virtualization (SR-IOV), which make it possible to give domains direct connections and native I/O performance for selected I/O devices. A domain with either a DIO or SR-IOV device is an I/O domain. In summary: not all I/O domains own PCI complexes, and there are increasingly more I/O domains that are not service domains. They use their I/O connectivity for performance for their own applications. However, there are some limitations and considerations: at this time, a domain using physical I/O cannot be live-migrated to another server. There is also a need to plan for security and introducing unneeded dependencies: if an I/O domain is also a service domain providing virtual I/O go guests, it has the ability to affect the correct operation of its client guest domains. This is even more relevant for the control domain. where the ldm has to be protected from unauthorized (or even mistaken) use that would affect other domains. As a general rule, running applications in the service domain or the control domain should be avoided. To recap: Guest domains with virtual I/O still provide the greatest operational flexibility, including features like live migration. I/O domains can be used for applications with high performance requirements. This is used to great effect in SPARC SuperCluster and in general T4 deployments. Direct I/O (DIO) and Single Root I/O Virtualization (SR-IOV) make this more attractive by giving direct I/O access to more domains. Service domains should in general not be used for applications, because compromised security in the domain, or an outage, can affect other domains that depend on it. This concern can be mitigated by providing guests' their virtual I/O from more than one service domain, so an interruption of service in the service domain does not cause an application outage. The control domain should in general not be used to run applications, for the same reason. SPARC SuperCluster use the control domain for applications, but it is an exception: it's not a general purpose environment; it's an engineered system with specifically configured applications and optimization for optimal performance. These are recommended "best practices" based on conversations with a number of Oracle architects. Keep in mind that "one size does not fit all", so you should evaluate these practices in the context of your own requirements. Summary Higher capacity T-series servers have made it more attractive to use them for applications with high resource requirements. New deployment models permit native I/O performance for demanding applications by running them in I/O domains with direct access to their devices. This is leveraged in SPARC SuperCluster, and can be leveraged in T-series servers to provision high-performance applications running in domains. Carefully planned, this can be used to provide higher performance for critical applications.

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  • To SYNC or not to SYNC – Part 3

    - by AshishRay
    I can't believe it has been almost a year since my last blog post. I know, that's an absolute no-no in the blogosphere. And I know that "I have been busy" is not a good excuse. So - without trying to come up with an excuse - let me state this - my apologies for taking such a long time to write the next Part. Without further ado, here goes. This is Part 3 of a multi-part blog article where we are discussing various aspects of setting up Data Guard synchronous redo transport (SYNC). In Part 1 of this article, I debunked the myth that Data Guard SYNC is similar to a two-phase commit operation. In Part 2, I discussed the various ways that network latency may or may not impact a Data Guard SYNC configuration. In this article, I will talk in details regarding why Data Guard SYNC is a good thing. I will also talk about distance implications for setting up such a configuration. So, Why Good? Why is Data Guard SYNC a good thing? Because, at the end of the day, this gives you the assurance of zero data loss - it doesn’t matter what outage may befall your primary system. Befall! Boy, that sounds theatrical. But seriously - think about this - it minimizes your data risks. That’s a big deal. Whether you have an outage due to bad disks, faulty hardware components, hardware / software bugs, physical data corruptions, power failures, lightning that takes out significant part of your data center, fire that melts your assets, water leakage from the cooling system, human errors such as accidental deletion of online redo log files - it doesn’t matter - you can have that “Om - peace” look on your face and then you can failover to the standby system, without losing a single bit of data in your Oracle database. You will be a hero, as shown in this not so imaginary conversation: IT Manager: Well, what’s the status? You: John is doing the trace analysis on the storage array. IT Manager: So? How long is that gonna take? You: Well, he is stuck, waiting for a response from <insert your not-so-favorite storage vendor here>. IT Manager: So, no root cause yet? You: I told you, he is stuck. We have escalated with their Support, but you know how long these things take. IT Manager: Darn it - the site is down! You: Not really … IT Manager: What do you mean? You: John is stuck, but Sreeni has already done a failover to the Data Guard standby. IT Manager: Whoa, whoa - wait! Failover means we lost some data, why did you do this without letting the Business group know? You: We didn’t lose any data. Remember, we had set up Data Guard with SYNC? So now, any problems on the production – we just failover. No data loss, and we are up and running in minutes. The Business guys don’t need to know. IT Manager: Wow! Are we great or what!! You: I guess … Ok, so you get it - SYNC is good. But as my dear friend Larry Carpenter says, “TANSTAAFL”, or "There ain't no such thing as a free lunch". Yes, of course - investing in Data Guard SYNC means that you have to invest in a low-latency network, you have to monitor your applications and database especially in peak load conditions, and you cannot under-provision your standby systems. But all these are good and necessary things, if you are supporting mission-critical apps that are supposed to be running 24x7. The peace of mind that this investment will give you is priceless, especially if you are serious about HA. How Far Can We Go? Someone may say at this point - well, I can’t use Data Guard SYNC over my coast-to-coast deployment. Most likely - true. So how far can you go? Well, we have customers who have deployed Data Guard SYNC over 300+ miles! Does this mean that you can also deploy over similar distances? Duh - no! I am going to say something here that most IT managers don’t like to hear - “It depends!” It depends on your application design, application response time / throughput requirements, network topology, etc. However, because of the optimal way we do SYNC, customers have been able to stretch Data Guard SYNC deployments over longer distances compared to traditional, storage-centric ways of doing this. The MAA Database 10.2 best practices paper Data Guard Redo Transport & Network Configuration, and Oracle Database 11.2 High Availability Best Practices Manual talk about some of these SYNC-related metrics. For example, a test deployment of Data Guard SYNC over 330 miles with 10ms latency showed an impact less than 5% for a busy OLTP application. Even if you can’t deploy Data Guard SYNC over your WAN distance, or if you already have an ASYNC standby located 1000-s of miles away, here’s another nifty way to boost your HA. Have a local standby, configured SYNC. How local is “local”? Again - it depends. One customer runs a local SYNC standby across the campus. Another customer runs it across 15 miles in another data center. Both of these customers are running Data Guard SYNC as their HA standard. If a localized outage affects their primary system, no problem! They have all the data available on the standby, to which they can failover. Very fast. In seconds. Wait - did I say “seconds”? Yes, Virginia, there is a Santa Claus. But you have to wait till the next blog article to find out more. I assure you tho’ that this time you won’t have to wait for another year for this.

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  • SQL Server Developer Tools &ndash; Codename Juneau vs. Red-Gate SQL Source Control

    - by Ajarn Mark Caldwell
    So how do the new SQL Server Developer Tools (previously code-named Juneau) stack up against SQL Source Control?  Read on to find out. At the PASS Community Summit a couple of weeks ago, it was announced that the previously code-named Juneau software would be released under the name of SQL Server Developer Tools with the release of SQL Server 2012.  This replacement for Database Projects in Visual Studio (also known in a former life as Data Dude) has some great new features.  I won’t attempt to describe them all here, but I will applaud Microsoft for making major improvements.  One of my favorite changes is the way database elements are broken down.  Previously every little thing was in its own file.  For example, indexes were each in their own file.  I always hated that.  Now, SSDT uses a pattern similar to Red-Gate’s and puts the indexes and keys into the same file as the overall table definition. Of course there are really cool features to keep your database model in sync with the actual source scripts, and the rename refactoring feature is now touted as being more than just a search and replace, but rather a “semantic-aware” search and replace.  Funny, it reminds me of SQL Prompt’s Smart Rename feature.  But I’m not writing this just to criticize Microsoft and argue that they are late to the party with this feature set.  Instead, I do see it as a viable alternative for folks who want all of their source code to be version controlled, but there are a couple of key trade-offs that you need to know about when you choose which tool set to use. First, the basics Both tool sets integrate with a wide variety of source control systems including the most popular: Subversion, GIT, Vault, and Team Foundation Server.  Both tools have integrated functionality to produce objects to upgrade your target database when you are ready (DACPACs in SSDT, integration with SQL Compare for SQL Source Control).  If you regularly live in Visual Studio or the Business Intelligence Development Studio (BIDS) then SSDT will likely be comfortable for you.  Like BIDS, SSDT is a Visual Studio Project Type that comes with SQL Server, and if you don’t already have Visual Studio installed, it will install the shell for you.  If you already have Visual Studio 2010 installed, then it will just add this as an available project type.  On the other hand, if you regularly live in SQL Server Management Studio (SSMS) then you will really enjoy the SQL Source Control integration from within SSMS.  Both tool sets store their database model in script files.  In SSDT, these are on your file system like other source files; in SQL Source Control, these are stored in the folder structure in your source control system, and you can always GET them to your file system if you want to browse them directly. For me, the key differentiating factors are 1) a single, unified check-in, and 2) migration scripts.  How you value those two features will likely make your decision for you. Unified Check-In If you do a continuous-integration (CI) style of development that triggers an automated build with unit testing on every check-in of source code, and you use Visual Studio for the rest of your development, then you will want to really consider SSDT.  Because it is just another project in Visual Studio, it can be added to your existing Solution, and you can then do a complete, or unified single check-in of all changes whether they are application or database changes.  This is simply not possible with SQL Source Control because it is in a different development tool (SSMS instead of Visual Studio) and there is no way to do one unified check-in between the two.  You CAN do really fast back-to-back check-ins, but there is the possibility that the automated build that is triggered from the first check-in will cause your unit tests to fail and the CI tool to report that you broke the build.  Of course, the automated build that is triggered from the second check-in which contains the “other half” of your changes should pass and so the amount of time that the build was broken may be very, very short, but if that is very, very important to you, then SQL Source Control just won’t work; you’ll have to use SSDT. Refactoring and Migrations If you work on a mature system, or on a not-so-mature but also not-so-well-designed system, where you want to refactor the database schema as you go along, but you can’t have data suddenly disappearing from your target system, then you’ll probably want to go with SQL Source Control.  As I wrote previously, there are a number of changes which you can make to your database that the comparison tools (both from Microsoft and Red Gate) simply cannot handle without the possibility (or probability) of data loss.  Currently, SSDT only offers you the ability to inject PRE and POST custom deployment scripts.  There is no way to insert your own script in the middle to override the default behavior of the tool.  In version 3.0 of SQL Source Control (Early Access version now available) you have that ability to create your own custom migration script to take the place of the commands that the tool would have done, and ensure the preservation of your data.  Or, even if the default tool behavior would have worked, but you simply know a better way then you can take control and do things your way instead of theirs. You Decide In the environment I work in, our automated builds are not triggered off of check-ins, but off of the clock (currently once per night) and so there is no point at which the automated build and unit tests will be triggered without having both sides of the development effort already checked-in.  Therefore having a unified check-in, while handy, is not critical for us.  As for migration scripts, these are critically important to us.  We do a lot of new development on systems that have already been in production for years, and it is not uncommon for us to need to do a refactoring of the database.  Because of the maturity of the existing system, that often involves data migrations or other additional SQL tasks that the comparison tools just can’t detect on their own.  Therefore, the ability to create a custom migration script to override the tool’s default behavior is very important to us.  And so, you can see why we will continue to use Red Gate SQL Source Control for the foreseeable future.

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  • Alcatel-Lucent: Enterprise 2.0: The Top 5 Things I would Do Over

    - by Kellsey Ruppel
    Happy Monday! Does anyone else feel as if the weekend went entirely too quickly? At least for those of us in the United States, we have the 4th of July Holiday next week to look forward to This week on the blog, we are going to focus on "WebCenter by Example" and highlight best practices from customers and partners. I recently came across this article and I think this is a great example of how we can learn from one another when it comes to social collaboration adoption. Do you agree with Jem? What things or best practices have you learned in your organizations?  By Jem Janik, Enterprise community manager, Alcatel-Lucent  Not so long ago, Engage, the Alcatel-Lucent employee social network and collaboration platform, celebrated its third birthday. With more than 25,000 members actively interacting each month, Engage has been a big enough success that it’s been the subject of external articles, and often those of us who helped launch it will go out and speak about what aspects contributed to that success. Hindsight is still 20/20 and what it takes to successfully launch an enterprise 2.0 community is fairly well-known now.  Today I want to tell you what I suspect you really want to know about.  As the enterprise community manager for Engage, after three years in, what are the top 5 things I wish we (and I mostly mean me) could do over? #5 Define your analytics solution from the start There is so much to do when you launch a community and initially growing it without complete chaos is quite a task.  It doesn’t take too long to get to a point where you want to focus your continued efforts in growing company collaboration.  Do people truly talk across regional boundaries or have we shifted siloed conversations to a new platform.  Is there one organization that doesn’t interact with another? If you are lucky you’ll have someone in your community team well versed in the world of databases and SQL queries, but it takes time to figure out what backend analytics data actually means. Professional support can be expensive and it may be hard to justify later as it typically has the community manager as the only main customer.  Figure out what you think you’ll want to know and how to get it early on. The sooner the better even if it doesn’t seem that critical at the time. #4 Lobbies guide you to the right places One piece of feedback that comes up more and more as we keep growing Engage is it’s hard to find stuff, or new people are not sure where to start. Something we’re doing now is defining some general topic areas of interest to be like “lobbies” into the platform and some common hashtags to go with them. I liken this to walking into a large medical or professional building for the first time.  There are hundreds of offices, and you look to a sign in the lobby to get guided to the right place for you.  We’re building that sign for members now, but again we missed the boat as the majority of the company has had their initial Engage experience. #3 Clean up, clean up, clean up Knowledge work and folksonomies are messy! The day we opened the doors to Engage I would have said we should keep everything ever created in Engage with an argument that it was a window into our collective knowledge so nothing should go.  Well, 6000+ groups and 200,000+ pieces of content later, I’ve changed my mind.  As previously mentioned, with too much “stuff” the system can be overwhelming to new members and it makes it harder to get what you’re looking for.   Do we need that help document about a tool we no longer have? NO!  Do we need that group that had 1 document and 2 discussions in the last two years? NO! Should we only have one group about a given topic instead of 4?  YES! Last fall, Engage defined a cleanup process for groups not used for a long time.  We also formed a volunteer cleaning army who are extra eyes on the hunt for “stuff” that should be updated, merged, or deleted.  It’s better late than never, but in line with what’s becoming a theme I wish these efforts had started earlier. #2 Communications & local community management One of the most important aspects of my job is to make sure people who should be talking to each other are actually doing it.  Connecting people to the other people they should know, the groups they should join, a piece of content that shouldn’t be missed.   I have worked both inside and outside of communications teams, and they are the best informed people in your company.  They know when something big is coming, how it impacts employees, how it fits with strategy, who else knows more, etc.  Having communications professionals who are power users can help scale up community management because they are already so well connected.  They also need to have the platform skills to pay attention without suffering email overload, how to grab someone’s attention, etc.  I wish I’d had figured this out much earlier.  If I had I would have groomed more communications colleagues into advocates and power members right at the start. #1 Grooming advocates vs. natural advocates I’ve just alluded to this above already. The very best advocates are those who naturally embrace your platform and automatically start to see new ways to work within it.  Those advocates seem to come out of the woodwork naturally since some of them are early adopters.  Not surprisingly, our best advocates today are those same people who were willing to come kick the tires when the community was completely empty.  Unfortunately, we didn’t get a global spread of those natural advocates.  I did ask around when we first launched for other people who might be good candidates, but didn’t push too hard as there were so many other things to get ready.  That was a mistake.  If I could get a redo I would have formally asked for people to be assigned where there were gaps and groomed them into an advocate.  Today as we find new advocates to fill the gaps, people are hesitant as the initial set has three years of practice are ahead of the curve power members; it definitely would have been easier earlier on. As fairly early adopters to corporate scale enterprise collaboration, there hasn’t been a roadmap to follow as we’ve grown Engage, which is part of the fun! It’s clear a lot of issues are more easily tackled the earlier you identify and begin to correct them, and I’ve identified the main five I wish I could redo.  In the spirit of collaboration, I hope someone else learns from my mistakes! View the original article by Jem here. 

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  • C++0x rvalue references - lvalues-rvalue binding

    - by Doug
    This is a follow-on question to http://stackoverflow.com/questions/2748866/c0x-rvalue-references-and-temporaries In the previous question, I asked how this code should work: void f(const std::string &); //less efficient void f(std::string &&); //more efficient void g(const char * arg) { f(arg); } It seems that the move overload should probably be called because of the implicit temporary, and this happens in GCC but not MSVC (or the EDG front-end used in MSVC's Intellisense). What about this code? void f(std::string &&); //NB: No const string & overload supplied void g1(const char * arg) { f(arg); } void g2(const std::string & arg) { f(arg); } It seems that, based on the answers to my previous question that function g1 is legal (and is accepted by GCC 4.3-4.5, but not by MSVC). However, GCC and MSVC both reject g2 because of clause 13.3.3.1.4/3, which prohibits lvalues from binding to rvalue ref arguments. I understand the rationale behind this - it is explained in N2831 "Fixing a safety problem with rvalue references". I also think that GCC is probably implementing this clause as intended by the authors of that paper, because the original patch to GCC was written by one of the authors (Doug Gregor). However, I don't this is quite intuitive. To me, (a) a const string & is conceptually closer to a string && than a const char *, and (b) the compiler could create a temporary string in g2, as if it were written like this: void g2(const std::string & arg) { f(std::string(arg)); } Indeed, sometimes the copy constructor is considered to be an implicit conversion operator. Syntactically, this is suggested by the form of a copy constructor, and the standard even mentions this specifically in clause 13.3.3.1.2/4, where the copy constructor for derived-base conversions is given a higher conversion rank than other implicit conversions: A conversion of an expression of class type to the same class type is given Exact Match rank, and a conversion of an expression of class type to a base class of that type is given Conversion rank, in spite of the fact that a copy/move constructor (i.e., a user-defined conversion function) is called for those cases. (I assume this is used when passing a derived class to a function like void h(Base), which takes a base class by value.) Motivation My motivation for asking this is something like the question asked in http://stackoverflow.com/questions/2696156/how-to-reduce-redundant-code-when-adding-new-c0x-rvalue-reference-operator-over ("How to reduce redundant code when adding new c++0x rvalue reference operator overloads"). If you have a function that accepts a number of potentially-moveable arguments, and would move them if it can (e.g. a factory function/constructor: Object create_object(string, vector<string>, string) or the like), and want to move or copy each argument as appropriate, you quickly start writing a lot of code. If the argument types are movable, then one could just write one version that accepts the arguments by value, as above. But if the arguments are (legacy) non-movable-but-swappable classes a la C++03, and you can't change them, then writing rvalue reference overloads is more efficient. So if lvalues did bind to rvalues via an implicit copy, then you could write just one overload like create_object(legacy_string &&, legacy_vector<legacy_string> &&, legacy_string &&) and it would more or less work like providing all the combinations of rvalue/lvalue reference overloads - actual arguments that were lvalues would get copied and then bound to the arguments, actual arguments that were rvalues would get directly bound. Questions My questions are then: Is this a valid interpretation of the standard? It seems that it's not the conventional or intended one, at any rate. Does it make intuitive sense? Is there a problem with this idea that I"m not seeing? It seems like you could get copies being quietly created when that's not exactly expected, but that's the status quo in places in C++03 anyway. Also, it would make some overloads viable when they're currently not, but I don't see it being a problem in practice. Is this a significant enough improvement that it would be worth making e.g. an experimental patch for GCC?

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  • ndd on Solaris 10

    - by user12620111
    This is mostly a repost of LaoTsao's Weblog with some tweaks. Last time that I tried to cut & paste directly off of his page, some of the XML was messed up. I run this from my MacBook. It should also work from your windows laptop if you use cygwin. ================If not already present, create a ssh key on you laptop================ # ssh-keygen -t rsa ================ Enable passwordless ssh from my laptop. Need to type in the root password for the remote machines. Then, I no longer need to type in the password when I ssh or scp from my laptop to servers. ================ #!/usr/bin/env bash for server in `cat servers.txt` do   echo root@$server   cat ~/.ssh/id_rsa.pub | ssh root@$server "cat >> .ssh/authorized_keys" done ================ servers.txt ================ testhost1testhost2 ================ etc_system_addins ================ set rpcmod:clnt_max_conns=8 set zfs:zfs_arc_max=0x1000000000 set nfs:nfs3_bsize=131072 set nfs:nfs4_bsize=131072 ================ ndd-nettune.txt ================ #!/sbin/sh # # ident   "@(#)ndd-nettune.xml    1.0     01/08/06 SMI" . /lib/svc/share/smf_include.sh . /lib/svc/share/net_include.sh # Make sure that the libraries essential to this stage of booting  can be found. LD_LIBRARY_PATH=/lib; export LD_LIBRARY_PATH echo "Performing Directory Server Tuning..." >> /tmp/smf.out # # Standard SuperCluster Tunables # /usr/sbin/ndd -set /dev/tcp tcp_max_buf 2097152 /usr/sbin/ndd -set /dev/tcp tcp_xmit_hiwat 1048576 /usr/sbin/ndd -set /dev/tcp tcp_recv_hiwat 1048576 # Reset the library path now that we are past the critical stage unset LD_LIBRARY_PATH ================ ndd-nettune.xml ================ <?xml version="1.0"?> <!DOCTYPE service_bundle SYSTEM "/usr/share/lib/xml/dtd/service_bundle.dtd.1"> <!-- ident "@(#)ndd-nettune.xml 1.0 04/09/21 SMI" --> <service_bundle type='manifest' name='SUNWcsr:ndd'>   <service name='network/ndd-nettune' type='service' version='1'>     <create_default_instance enabled='true' />     <single_instance />     <dependency name='fs-minimal' type='service' grouping='require_all' restart_on='none'>       <service_fmri value='svc:/system/filesystem/minimal' />     </dependency>     <dependency name='loopback-network' grouping='require_any' restart_on='none' type='service'>       <service_fmri value='svc:/network/loopback' />     </dependency>     <dependency name='physical-network' grouping='optional_all' restart_on='none' type='service'>       <service_fmri value='svc:/network/physical' />     </dependency>     <exec_method type='method' name='start' exec='/lib/svc/method/ndd-nettune' timeout_seconds='3' > </exec_method>     <exec_method type='method' name='stop'  exec=':true'                       timeout_seconds='3' > </exec_method>     <property_group name='startd' type='framework'>       <propval name='duration' type='astring' value='transient' />     </property_group>     <stability value='Unstable' />     <template>       <common_name>     <loctext xml:lang='C'> ndd network tuning </loctext>       </common_name>       <documentation>     <manpage title='ndd' section='1M' manpath='/usr/share/man' />       </documentation>     </template>   </service> </service_bundle> ================ system_tuning.sh ================ #!/usr/bin/env bash for server in `cat servers.txt` do   cat etc_system_addins | ssh root@$server "cat >> /etc/system"   scp ndd-nettune.xml root@${server}:/var/svc/manifest/site/ndd-nettune.xml   scp ndd-nettune.txt root@${server}:/lib/svc/method/ndd-nettune   ssh root@$server chmod +x /lib/svc/method/ndd-nettune   ssh root@$server svccfg validate /var/svc/manifest/site/ndd-nettune.xml   ssh root@$server svccfg import /var/svc/manifest/site/ndd-nettune.xml done

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  • Is multithreading the right way to go for my case?

    - by Julien Lebosquain
    Hello, I'm currently designing a multi-client / server application. I'm using plain good old sockets because WCF or similar technology is not what I need. Let me explain: it isn't the classical case of a client simply calling a service; all clients can 'interact' with each other by sending a packet to the server, which will then do some action, and possible re-dispatch an answer message to one or more clients. Although doable with WCF, the application will get pretty complex with hundreds of different messages. For each connected client, I'm of course using asynchronous methods to send and receive bytes. I've got the messages fully working, everything's fine. Except that for each line of code I'm writing, my head just burns because of multithreading issues. Since there could be around 200 clients connected at the same time, I chose to go the fully multithreaded way: each received message on a socket is immediately processed on the thread pool thread it was received, not on a single consumer thread. Since each client can interact with other clients, and indirectly with shared objects on the server, I must protect almost every object that is mutable. I first went with a ReaderWriterLockSlim for each resource that must be protected, but quickly noticed that there are more writes overall than reads in the server application, and switched to the well-known Monitor to simplify the code. So far, so good. Each resource is protected, I have helper classes that I must use to get a lock and its protected resource, so I can't use an object without getting a lock. Moreover, each client has its own lock that is entered as soon as a packet is received from its socket. It's done to prevent other clients from making changes to the state of this client while it has some messages being processed, which is something that will happen frequently. Now, I don't just need to protect resources from concurrent accesses. I must keep every client in sync with the server for some collections I have. One tricky part that I'm currently struggling with is the following: I have a collection of clients. Each client has its own unique ID. When a client connects, it must receive the IDs of every connected client, and each one of them must be notified of the newcomer's ID. When a client disconnects, every other client must know it so that its ID is no longer valid for them. Every client must always have, at a given time, the same clients collection as the server so that I can assume that everybody knows everybody. This way if I'm sending a message to client #1 telling "Client #2 has done something", I know that it will always be correctly interpreted: Client 1 will never wonder "but who is Client 2 anyway?". My first attempt for handling the connection of a new client (let's call it X) was this pseudo-code (remember that newClient is already locked here): lock (clients) { foreach (var client in clients) { lock (client) { client.Send("newClient with id X has connected"); } } clients.Add(newClient); newClient.Send("the list of other clients"); } Now imagine that in the same time, another client has sent a packet that translates into a message that must be broadcasted to every connected client, the pseudo-code will be something like this (remember that the current client - let's call it Y - is already locked here): lock (clients) { foreach (var client in clients) { lock (client) { client.Send("something"); } } } An obvious deadlock occurs here: on one thread X is locked, the clients lock has been entered, started looping through the clients, and at one moment must get Y's lock... which is already acquired on the second thread, itself waiting for the clients collection lock to be released! This is not the only case like this in the server application. There are other collections which must be kept in sync with the clients, some properties on a client can be changed by another one, etc. I tried other types of locks, lock-free mechanisms and a bunch of other things. Either there were obvious deadlocks when I'm using too much locks for safety, or obvious race conditions otherwise. When I finally find a good middle point between the two, it usually comes with very subtle race conditions / dead locks and other multi-threading issues... my head hurts very quickly since for any single line of code I'm writing I have to review almost the whole application to ensure everything will behave correctly with any number of threads. So here's my final question: how would you resolve this specific case, the general case, and more importantly: aren't I going the wrong way here? I have little problems with the .NET framework, C#, simple concurrency or algorithms in general. Still, I'm lost here. I know I could use only one thread processing the incoming requests and everything will be fine. However, that won't scale well at all with more clients... But I'm thinking more and more to go this simple way. What do you think? Thanks in advance to you, StackOverflow people which have taken the time to read this huge question. I really had to explain the whole context if I want to get some help.

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  • Using JS script for "raining images". Can't seem to hide pre-loaded image

    - by user1813605
    I am trying to hide an image in a script pre-loading on the page. Below script makes images "rain" down the screen onClick. It functions well, but it displays the pre-loaded image itself on the page before the button is clicked. I'm trying to hide the image until the button is pressed. If anyone has any insight on how to hide the image until the function dispenseMittens() runs, I'd be eternally grateful :) Thanks! <script language="javascript"> var pictureSrc = 'mitten.gif'; //the location of the mittens var pictureWidth = 40; //the width of the mittens var pictureHeight = 46; //the height of the mittens var numFlakes = 10; //the number of mittens var downSpeed = 0.01; var lrFlakes = 10; var EmergencyMittens = false; //safety checks. Browsers will hang if this is wrong. If other values are wrong there will just be errors if( typeof( numFlakes ) != 'number' || Math.round( numFlakes ) != numFlakes || numFlakes < 1 ) { numFlakes = 10; } //draw the snowflakes for( var x = 0; x < numFlakes; x++ ) { if( document.layers ) { //releave NS4 bug document.write('<layer id="snFlkDiv'+x+'"><img src="'+pictureSrc+'" height="'+pictureHeight+'" width="'+pictureWidth+'" alt="*" border="0"></layer>'); } else { document.write('<div style="position:absolute;" id="snFlkDiv'+x+'"><img src="'+pictureSrc+'" height="'+pictureHeight+'" width="'+pictureWidth+'" alt="*" border="0"></div>'); } } //calculate initial positions (in portions of browser window size) var xcoords = new Array(), ycoords = new Array(), snFlkTemp; for( var x = 0; x < numFlakes; x++ ) { xcoords[x] = ( x + 1 ) / ( numFlakes + 1 ); do { snFlkTemp = Math.round( ( numFlakes - 1 ) * Math.random() ); } while( typeof( ycoords[snFlkTemp] ) == 'number' ); ycoords[snFlkTemp] = x / numFlakes; } //now animate function mittensFall() { if( !getRefToDivNest('snFlkDiv0') ) { return; } var scrWidth = 0, scrHeight = 0, scrollHeight = 0, scrollWidth = 0; //find screen settings for all variations. doing this every time allows for resizing and scrolling if( typeof( window.innerWidth ) == 'number' ) { scrWidth = window.innerWidth; scrHeight = window.innerHeight; } else { if( document.documentElement && ( document.documentElement.clientWidth || document.documentElement.clientHeight ) ) { scrWidth = document.documentElement.clientWidth; scrHeight = document.documentElement.clientHeight; } else { if( document.body && ( document.body.clientWidth || document.body.clientHeight ) ) { scrWidth = document.body.clientWidth; scrHeight = document.body.clientHeight; } } } if( typeof( window.pageYOffset ) == 'number' ) { scrollHeight = pageYOffset; scrollWidth = pageXOffset; } else { if( document.body && ( document.body.scrollLeft || document.body.scrollTop ) ) { scrollHeight = document.body.scrollTop; scrollWidth = document.body.scrollLeft; } else { if( document.documentElement && ( document.documentElement.scrollLeft || document.documentElement.scrollTop ) ) { scrollHeight = document.documentElement.scrollTop; scrollWidth = document.documentElement.scrollLeft; } } } //move the snowflakes to their new position for( var x = 0; x < numFlakes; x++ ) { if( ycoords[x] * scrHeight > scrHeight - pictureHeight ) { ycoords[x] = 0; } var divRef = getRefToDivNest('snFlkDiv'+x); if( !divRef ) { return; } if( divRef.style ) { divRef = divRef.style; } var oPix = document.childNodes ? 'px' : 0; divRef.top = ( Math.round( ycoords[x] * scrHeight ) + scrollHeight ) + oPix; divRef.left = ( Math.round( ( ( xcoords[x] * scrWidth ) - ( pictureWidth / 2 ) ) + ( ( scrWidth / ( ( numFlakes + 1 ) * 4 ) ) * ( Math.sin( lrFlakes * ycoords[x] ) - Math.sin( 3 * lrFlakes * ycoords[x] ) ) ) ) + scrollWidth ) + oPix; ycoords[x] += downSpeed; } } //DHTML handlers function getRefToDivNest(divName) { if( document.layers ) { return document.layers[divName]; } //NS4 if( document[divName] ) { return document[divName]; } //NS4 also if( document.getElementById ) { return document.getElementById(divName); } //DOM (IE5+, NS6+, Mozilla0.9+, Opera) if( document.all ) { return document.all[divName]; } //Proprietary DOM - IE4 return false; } function dispenseMittens() { if (EmergencyMittens) { window.clearInterval(EmergencyMittens); } else { EmergencyMittens = window.setInterval('mittensFall();',100); } } </script>

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  • apache2 doesn't start with location

    - by Geod24
    I have a small domain, which I use only for personal purposes. I'm the main user, and have at most 3-4 users at the same time. I use apache2 with passenger to serve redmine. So I start with an empty apache2: root@xxxxx:/home/# service apache2 start [ ok ] Starting web server: apache2. root@xxxxx:/home/# a2dissite Your choices are: Which site(s) do you want to disable (wildcards ok)? Then enable my site, and restart (not reload) apache2: root@xxxxx:/home/# a2ensite 200-redmine Enabling site 200-redmine. To activate the new configuration, you need to run: service apache2 reload root@xxxxx:/home/# service apache2 restart [FAIL] Restarting web server: apache2 failed! [warn] The apache2 instance did not start within 20 seconds. Please read the log files to discover problems ... (warning). root@xxxxx:/home/# service apache2 restart [FAIL] Restarting web server: apache2 failed! [warn] There are processes named 'apache2' running which do not match your pid file which are left untouched in the name of safety, Please review the situation by hand. ... (warning). root@xxxxx:/home/# pidof apache2 20948 Here's my 200-redmine.conf: PerlLoadModule Apache::Redmine <VirtualHost *:80> ServerName redmine.xxxxx.xxx DocumentRoot /var/www/redmine/public/ ErrorLog ${APACHE_LOG_DIR}/redmine.error.log CustomLog ${APACHE_LOG_DIR}/redmine.access.log common MaxRequestLen 20971520 <Directory "/var/www/redmine/public/"> Options Indexes ExecCGI FollowSymLinks Order allow,deny Allow from all AllowOverride all </Directory> SetEnv GIT_PROJECT_ROOT /opt/git/ SetEnv GIT_HTTP_EXPORT_ALL ScriptAlias /git/ /usr/lib/git-core/git-http-backend/ <Location /git> PerlAuthenHandler Apache::Authn::Redmine::authen_handler PerlAccessHandler Apache::Authn::Redmine::access_handler AuthType Basic Require valid-user AuthName "Redmine Git Repository" RedmineDSN "DBI:mysql:database=redmine;host=localhost:3306" RedmineDbUser "redmine" RedmineDbPass "password" RedmineCacheCredsMax 50 </Location> </VirtualHost> Now if I comment out the ScriptAlias / stuff, it works ! In addition, starting the server with 200-redmine disabled, then enabling it works. But apache2 will die randomly. Plus the location doesn't work. The logs show nothing: root@xxxxx:/home/# ll /var/log/apache2/ total 8 drwxr-xr-x 2 root root 4096 Oct 30 07:52 coredump -rw-r--r-- 1 root root 0 Nov 4 02:39 default.access.log -rw-r--r-- 1 root root 2356 Nov 4 02:39 default.error.log -rw-r--r-- 1 root root 0 Nov 4 02:39 other_vhosts_access.log -rw-r--r-- 1 root root 0 Nov 4 02:39 redmine.access.log -rw-r--r-- 1 root root 0 Nov 4 02:39 redmine.error.log root@xxxxx:/home/# ll /var/log/apache2/coredump/ total 0 root@xxxxx:/home/# cat /var/log/apache2/default.error.log [ 2013-11-04 02:39:36.0130 21471/7fcf090f4740 agents/Watchdog/Main.cpp:452 ]: Options: { 'analytics_log_user' => 'nobody', 'default_group' => 'nogroup', 'default_python' => 'python', 'default_ruby' => '/usr/bin/ruby', 'default_user' => 'nobody', 'log_level' => '0', 'max_instances_per_app' => '0', 'max_pool_size' => '6', 'passenger_root' => '/usr/lib/ruby/vendor_ruby/phusion_passenger/locations.ini', 'pool_idle_time' => '300', 'temp_dir' => '/tmp', 'union_station_gateway_address' => 'gateway.unionstationapp.com', 'union_station_gateway_port' => '443', 'user_switching' => 'true', 'web_server_pid' => '21470', 'web_server_type' => 'apache', 'web_server_worker_gid' => '33', 'web_server_worker_uid' => '33' } [ 2013-11-04 02:39:36.0255 21474/7f9a99fda740 agents/HelperAgent/Main.cpp:597 ]: PassengerHelperAgent online, listening at unix:/tmp/passenger.1.0.21470/generation-0/request [ 2013-11-04 02:39:36.0507 21479/7f8316b0f740 agents/LoggingAgent/Main.cpp:330 ]: PassengerLoggingAgent online, listening at unix:/tmp/passenger.1.0.21470/generation-0/logging [ 2013-11-04 02:39:36.0511 21471/7fcf090f4740 agents/Watchdog/Main.cpp:635 ]: All Phusion Passenger agents started! [ 2013-11-04 02:39:36.3158 21495/7fba6f686740 agents/Watchdog/Main.cpp:452 ]: Options: { 'analytics_log_user' => 'nobody', 'default_group' => 'nogroup', 'default_python' => 'python', 'default_ruby' => '/usr/bin/ruby', 'default_user' => 'nobody', 'log_level' => '0', 'max_instances_per_app' => '0', 'max_pool_size' => '6', 'passenger_root' => '/usr/lib/ruby/vendor_ruby/phusion_passenger/locations.ini', 'pool_idle_time' => '300', 'temp_dir' => '/tmp', 'union_station_gateway_address' => 'gateway.unionstationapp.com', 'union_station_gateway_port' => '443', 'user_switching' => 'true', 'web_server_pid' => '21491', 'web_server_type' => 'apache', 'web_server_worker_gid' => '33', 'web_server_worker_uid' => '33' } [ 2013-11-04 02:39:36.3304 21498/7f0106d9b740 agents/HelperAgent/Main.cpp:597 ]: PassengerHelperAgent online, listening at unix:/tmp/passenger.1.0.21491/generation-0/request [ 2013-11-04 02:39:36.3522 21503/7f92ad392740 agents/LoggingAgent/Main.cpp:330 ]: PassengerLoggingAgent online, listening at unix:/tmp/passenger.1.0.21491/generation-0/logging [ 2013-11-04 02:39:36.3525 21495/7fba6f686740 agents/Watchdog/Main.cpp:635 ]: All Phusion Passenger agents started! And at last: root@xxxxx:/home/# apache2ctl -t -D DUMP_VHOSTS VirtualHost configuration: *:80 is a NameVirtualHost default server redmine.xxxx.xxx (/etc/apache2/sites-enabled/200-redmine.conf:5) port 80 namevhost redmine.xxxx.xxx (/etc/apache2/sites-enabled/200-redmine.conf:5) port 80 namevhost redmine.xxxxx.xxx (/etc/apache2/sites-enabled/200-redmine.conf:5) root@xxxxx:/home/# uname -a Linux xxxx.xxx 3.2.0-4-amd64 #1 SMP Debian 3.2.51-1 x86_64 GNU/Linux root@xxxxx:/home/# dpkg --list | grep apache2 ii apache2 2.4.6-3 amd64 Apache HTTP Server ii apache2-bin 2.4.6-3 amd64 Apache HTTP Server (binary files and modules) ii apache2-data 2.4.6-3 all Apache HTTP Server (common files) ii apache2-utils 2.4.6-3 amd64 Apache HTTP Server (utility programs for web servers) ii libapache2-mod-fcgid 1:2.3.9-1 amd64 FastCGI interface module for Apache 2 ii libapache2-mod-passenger 4.0.10-1 amd64 Rails and Rack support for Apache2 ii libapache2-mod-perl2 2.0.8+httpd24-r1449661-6+b1 amd64 Integration of perl with the Apache2 web server ii libapache2-mod-perl2-dev 2.0.8+httpd24-r1449661-6 all Integration of perl with the Apache2 web server - development files ii libapache2-mod-perl2-doc 2.0.8+httpd24-r1449661-6 all Integration of perl with the Apache2 web server - documentation ii libapache2-mod-proxy-html 1:2.4.6-3 amd64 Transitional package for apache2-bin ii libapache2-mod-svn 1.7.13-2 amd64 Apache Subversion server modules for Apache httpd ii libapache2-reload-perl 0.12-2 all module for reloading Perl modules when changed on disk ii libapache2-svn 1.7.13-2 all Apache Subversion server modules for Apache httpd (dummy package) root@xxxxx:/home/# a2dismod Your choices are: access_compat alias auth_basic authn_core authn_file authz_core authz_host authz_svn authz_user autoindex dav dav_svn deflate dir env fcgid filter mime mpm_event negotiation passenger perl proxy proxy_http rewrite setenvif status Which module(s) do you want to disable (wildcards ok)?

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  • Integrating JavaScript Unit Tests with Visual Studio

    - by Stephen Walther
    Modern ASP.NET web applications take full advantage of client-side JavaScript to provide better interactivity and responsiveness. If you are building an ASP.NET application in the right way, you quickly end up with lots and lots of JavaScript code. When writing server code, you should be writing unit tests. One big advantage of unit tests is that they provide you with a safety net that enable you to safely modify your existing code – for example, fix bugs, add new features, and make performance enhancements -- without breaking your existing code. Every time you modify your code, you can execute your unit tests to verify that you have not broken anything. For the same reason that you should write unit tests for your server code, you should write unit tests for your client code. JavaScript is just as susceptible to bugs as C#. There is no shortage of unit testing frameworks for JavaScript. Each of the major JavaScript libraries has its own unit testing framework. For example, jQuery has QUnit, Prototype has UnitTestJS, YUI has YUI Test, and Dojo has Dojo Objective Harness (DOH). The challenge is integrating a JavaScript unit testing framework with Visual Studio. Visual Studio and Visual Studio ALM provide fantastic support for server-side unit tests. You can easily view the results of running your unit tests in the Visual Studio Test Results window. You can set up a check-in policy which requires that all unit tests pass before your source code can be committed to the source code repository. In addition, you can set up Team Build to execute your unit tests automatically. Unfortunately, Visual Studio does not provide “out-of-the-box” support for JavaScript unit tests. MS Test, the unit testing framework included in Visual Studio, does not support JavaScript unit tests. As soon as you leave the server world, you are left on your own. The goal of this blog entry is to describe one approach to integrating JavaScript unit tests with MS Test so that you can execute your JavaScript unit tests side-by-side with your C# unit tests. The goal is to enable you to execute JavaScript unit tests in exactly the same way as server-side unit tests. You can download the source code described by this project by scrolling to the end of this blog entry. Rejected Approach: Browser Launchers One popular approach to executing JavaScript unit tests is to use a browser as a test-driver. When you use a browser as a test-driver, you open up a browser window to execute and view the results of executing your JavaScript unit tests. For example, QUnit – the unit testing framework for jQuery – takes this approach. The following HTML page illustrates how you can use QUnit to create a unit test for a function named addNumbers(). <!DOCTYPE HTML PUBLIC "-//W3C//DTD HTML 4.01 Transitional//EN" "http://www.w3.org/TR/html4/loose.dtd"> <html> <head> <title>Using QUnit</title> <link rel="stylesheet" href="http://github.com/jquery/qunit/raw/master/qunit/qunit.css" type="text/css" /> </head> <body> <h1 id="qunit-header">QUnit example</h1> <h2 id="qunit-banner"></h2> <div id="qunit-testrunner-toolbar"></div> <h2 id="qunit-userAgent"></h2> <ol id="qunit-tests"></ol> <div id="qunit-fixture">test markup, will be hidden</div> <script type="text/javascript" src="http://code.jquery.com/jquery-latest.js"></script> <script type="text/javascript" src="http://github.com/jquery/qunit/raw/master/qunit/qunit.js"></script> <script type="text/javascript"> // The function to test function addNumbers(a, b) { return a+b; } // The unit test test("Test of addNumbers", function () { equals(4, addNumbers(1,3), "1+3 should be 4"); }); </script> </body> </html> This test verifies that calling addNumbers(1,3) returns the expected value 4. When you open this page in a browser, you can see that this test does, in fact, pass. The idea is that you can quickly refresh this QUnit HTML JavaScript test driver page in your browser whenever you modify your JavaScript code. In other words, you can keep a browser window open and keep refreshing it over and over while you are developing your application. That way, you can know very quickly whenever you have broken your JavaScript code. While easy to setup, there are several big disadvantages to this approach to executing JavaScript unit tests: You must view your JavaScript unit test results in a different location than your server unit test results. The JavaScript unit test results appear in the browser and the server unit test results appear in the Visual Studio Test Results window. Because all of your unit test results don’t appear in a single location, you are more likely to introduce bugs into your code without noticing it. Because your unit tests are not integrated with Visual Studio – in particular, MS Test -- you cannot easily include your JavaScript unit tests when setting up check-in policies or when performing automated builds with Team Build. A more sophisticated approach to using a browser as a test-driver is to automate the web browser. Instead of launching the browser and loading the test code yourself, you use a framework to automate this process. There are several different testing frameworks that support this approach: · Selenium – Selenium is a very powerful framework for automating browser tests. You can create your tests by recording a Firefox session or by writing the test driver code in server code such as C#. You can learn more about Selenium at http://seleniumhq.org/. LTAF – The ASP.NET team uses the Lightweight Test Automation Framework to test JavaScript code in the ASP.NET framework. You can learn more about LTAF by visiting the project home at CodePlex: http://aspnet.codeplex.com/releases/view/35501 jsTestDriver – This framework uses Java to automate the browser. jsTestDriver creates a server which can be used to automate multiple browsers simultaneously. This project is located at http://code.google.com/p/js-test-driver/ TestSwam – This framework, created by John Resig, uses PHP to automate the browser. Like jsTestDriver, the framework creates a test server. You can open multiple browsers that are automated by the test server. Learn more about TestSwarm by visiting the following address: https://github.com/jeresig/testswarm/wiki Yeti – This is the framework introduced by Yahoo for automating browser tests. Yeti uses server-side JavaScript and depends on Node.js. Learn more about Yeti at http://www.yuiblog.com/blog/2010/08/25/introducing-yeti-the-yui-easy-testing-interface/ All of these frameworks are great for integration tests – however, they are not the best frameworks to use for unit tests. In one way or another, all of these frameworks depend on executing tests within the context of a “living and breathing” browser. If you create an ASP.NET Unit Test then Visual Studio will launch a web server before executing the unit test. Why is launching a web server so bad? It is not the worst thing in the world. However, it does introduce dependencies that prevent your code from being tested in isolation. One of the defining features of a unit test -- versus an integration test – is that a unit test tests code in isolation. Another problem with launching a web server when performing unit tests is that launching a web server can be slow. If you cannot execute your unit tests quickly, you are less likely to execute your unit tests each and every time you make a code change. You are much more likely to fall into the pit of failure. Launching a browser when performing a JavaScript unit test has all of the same disadvantages as launching a web server when performing an ASP.NET unit test. Instead of testing a unit of JavaScript code in isolation, you are testing JavaScript code within the context of a particular browser. Using the frameworks listed above for integration tests makes perfect sense. However, I want to consider a different approach for creating unit tests for JavaScript code. Using Server-Side JavaScript for JavaScript Unit Tests A completely different approach to executing JavaScript unit tests is to perform the tests outside of any browser. If you really want to test JavaScript then you should test JavaScript and leave the browser out of the testing process. There are several ways that you can execute JavaScript on the server outside the context of any browser: Rhino – Rhino is an implementation of JavaScript written in Java. The Rhino project is maintained by the Mozilla project. Learn more about Rhino at http://www.mozilla.org/rhino/ V8 – V8 is the open-source Google JavaScript engine written in C++. This is the JavaScript engine used by the Chrome web browser. You can download V8 and embed it in your project by visiting http://code.google.com/p/v8/ JScript – JScript is the JavaScript Script Engine used by Internet Explorer (up to but not including Internet Explorer 9), Windows Script Host, and Active Server Pages. Internet Explorer is still the most popular web browser. Therefore, I decided to focus on using the JScript Script Engine to execute JavaScript unit tests. Using the Microsoft Script Control There are two basic ways that you can pass JavaScript to the JScript Script Engine and execute the code: use the Microsoft Windows Script Interfaces or use the Microsoft Script Control. The difficult and proper way to execute JavaScript using the JScript Script Engine is to use the Microsoft Windows Script Interfaces. You can learn more about the Script Interfaces by visiting http://msdn.microsoft.com/en-us/library/t9d4xf28(VS.85).aspx The main disadvantage of using the Script Interfaces is that they are difficult to use from .NET. There is a great series of articles on using the Script Interfaces from C# located at http://www.drdobbs.com/184406028. I picked the easier alternative and used the Microsoft Script Control. The Microsoft Script Control is an ActiveX control that provides a higher level abstraction over the Window Script Interfaces. You can download the Microsoft Script Control from here: http://www.microsoft.com/downloads/en/details.aspx?FamilyID=d7e31492-2595-49e6-8c02-1426fec693ac After you download the Microsoft Script Control, you need to add a reference to it to your project. Select the Visual Studio menu option Project, Add Reference to open the Add Reference dialog. Select the COM tab and add the Microsoft Script Control 1.0. Using the Script Control is easy. You call the Script Control AddCode() method to add JavaScript code to the Script Engine. Next, you call the Script Control Run() method to run a particular JavaScript function. The reference documentation for the Microsoft Script Control is located at the MSDN website: http://msdn.microsoft.com/en-us/library/aa227633%28v=vs.60%29.aspx Creating the JavaScript Code to Test To keep things simple, let’s imagine that you want to test the following JavaScript function named addNumbers() which simply adds two numbers together: MvcApplication1\Scripts\Math.js function addNumbers(a, b) { return 5; } Notice that the addNumbers() method always returns the value 5. Right-now, it will not pass a good unit test. Create this file and save it in your project with the name Math.js in your MVC project’s Scripts folder (Save the file in your actual MVC application and not your MVC test application). Creating the JavaScript Test Helper Class To make it easier to use the Microsoft Script Control in unit tests, we can create a helper class. This class contains two methods: LoadFile() – Loads a JavaScript file. Use this method to load the JavaScript file being tested or the JavaScript file containing the unit tests. ExecuteTest() – Executes the JavaScript code. Use this method to execute a JavaScript unit test. Here’s the code for the JavaScriptTestHelper class: JavaScriptTestHelper.cs   using System; using System.IO; using Microsoft.VisualStudio.TestTools.UnitTesting; using MSScriptControl; namespace MvcApplication1.Tests { public class JavaScriptTestHelper : IDisposable { private ScriptControl _sc; private TestContext _context; /// <summary> /// You need to use this helper with Unit Tests and not /// Basic Unit Tests because you need a Test Context /// </summary> /// <param name="testContext">Unit Test Test Context</param> public JavaScriptTestHelper(TestContext testContext) { if (testContext == null) { throw new ArgumentNullException("TestContext"); } _context = testContext; _sc = new ScriptControl(); _sc.Language = "JScript"; _sc.AllowUI = false; } /// <summary> /// Load the contents of a JavaScript file into the /// Script Engine. /// </summary> /// <param name="path">Path to JavaScript file</param> public void LoadFile(string path) { var fileContents = File.ReadAllText(path); _sc.AddCode(fileContents); } /// <summary> /// Pass the path of the test that you want to execute. /// </summary> /// <param name="testMethodName">JavaScript function name</param> public void ExecuteTest(string testMethodName) { dynamic result = null; try { result = _sc.Run(testMethodName, new object[] { }); } catch { var error = ((IScriptControl)_sc).Error; if (error != null) { var description = error.Description; var line = error.Line; var column = error.Column; var text = error.Text; var source = error.Source; if (_context != null) { var details = String.Format("{0} \r\nLine: {1} Column: {2}", source, line, column); _context.WriteLine(details); } } throw new AssertFailedException(error.Description); } } public void Dispose() { _sc = null; } } }     Notice that the JavaScriptTestHelper class requires a Test Context to be instantiated. For this reason, you can use the JavaScriptTestHelper only with a Visual Studio Unit Test and not a Basic Unit Test (These are two different types of Visual Studio project items). Add the JavaScriptTestHelper file to your MVC test application (for example, MvcApplication1.Tests). Creating the JavaScript Unit Test Next, we need to create the JavaScript unit test function that we will use to test the addNumbers() function. Create a folder in your MVC test project named JavaScriptTests and add the following JavaScript file to this folder: MvcApplication1.Tests\JavaScriptTests\MathTest.js /// <reference path="JavaScriptUnitTestFramework.js"/> function testAddNumbers() { // Act var result = addNumbers(1, 3); // Assert assert.areEqual(4, result, "addNumbers did not return right value!"); }   The testAddNumbers() function takes advantage of another JavaScript library named JavaScriptUnitTestFramework.js. This library contains all of the code necessary to make assertions. Add the following JavaScriptnitTestFramework.js to the same folder as the MathTest.js file: MvcApplication1.Tests\JavaScriptTests\JavaScriptUnitTestFramework.js var assert = { areEqual: function (expected, actual, message) { if (expected !== actual) { throw new Error("Expected value " + expected + " is not equal to " + actual + ". " + message); } } }; There is only one type of assertion supported by this file: the areEqual() assertion. Most likely, you would want to add additional types of assertions to this file to make it easier to write your JavaScript unit tests. Deploying the JavaScript Test Files This step is non-intuitive. When you use Visual Studio to run unit tests, Visual Studio creates a new folder and executes a copy of the files in your project. After you run your unit tests, your Visual Studio Solution will contain a new folder named TestResults that includes a subfolder for each test run. You need to configure Visual Studio to deploy your JavaScript files to the test run folder or Visual Studio won’t be able to find your JavaScript files when you execute your unit tests. You will get an error that looks something like this when you attempt to execute your unit tests: You can configure Visual Studio to deploy your JavaScript files by adding a Test Settings file to your Visual Studio Solution. It is important to understand that you need to add this file to your Visual Studio Solution and not a particular Visual Studio project. Right-click your Solution in the Solution Explorer window and select the menu option Add, New Item. Select the Test Settings item and click the Add button. After you create a Test Settings file for your solution, you can indicate that you want a particular folder to be deployed whenever you perform a test run. Select the menu option Test, Edit Test Settings to edit your test configuration file. Select the Deployment tab and select your MVC test project’s JavaScriptTest folder to deploy. Click the Apply button and the Close button to save the changes and close the dialog. Creating the Visual Studio Unit Test The very last step is to create the Visual Studio unit test (the MS Test unit test). Add a new unit test to your MVC test project by selecting the menu option Add New Item and selecting the Unit Test project item (Do not select the Basic Unit Test project item): The difference between a Basic Unit Test and a Unit Test is that a Unit Test includes a Test Context. We need this Test Context to use the JavaScriptTestHelper class that we created earlier. Enter the following test method for the new unit test: [TestMethod] public void TestAddNumbers() { var jsHelper = new JavaScriptTestHelper(this.TestContext); // Load JavaScript files jsHelper.LoadFile("JavaScriptUnitTestFramework.js"); jsHelper.LoadFile(@"..\..\..\MvcApplication1\Scripts\Math.js"); jsHelper.LoadFile("MathTest.js"); // Execute JavaScript Test jsHelper.ExecuteTest("testAddNumbers"); } This code uses the JavaScriptTestHelper to load three files: JavaScripUnitTestFramework.js – Contains the assert functions. Math.js – Contains the addNumbers() function from your MVC application which is being tested. MathTest.js – Contains the JavaScript unit test function. Next, the test method calls the JavaScriptTestHelper ExecuteTest() method to execute the testAddNumbers() JavaScript function. Running the Visual Studio JavaScript Unit Test After you complete all of the steps described above, you can execute the JavaScript unit test just like any other unit test. You can use the keyboard combination CTRL-R, CTRL-A to run all of the tests in the current Visual Studio Solution. Alternatively, you can use the buttons in the Visual Studio toolbar to run the tests: (Unfortunately, the Run All Impacted Tests button won’t work correctly because Visual Studio won’t detect that your JavaScript code has changed. Therefore, you should use either the Run Tests in Current Context or Run All Tests in Solution options instead.) The results of running the JavaScript tests appear side-by-side with the results of running the server tests in the Test Results window. For example, if you Run All Tests in Solution then you will get the following results: Notice that the TestAddNumbers() JavaScript test has failed. That is good because our addNumbers() function is hard-coded to always return the value 5. If you double-click the failing JavaScript test, you can view additional details such as the JavaScript error message and the line number of the JavaScript code that failed: Summary The goal of this blog entry was to explain an approach to creating JavaScript unit tests that can be easily integrated with Visual Studio and Visual Studio ALM. I described how you can use the Microsoft Script Control to execute JavaScript on the server. By taking advantage of the Microsoft Script Control, we were able to execute our JavaScript unit tests side-by-side with all of our other unit tests and view the results in the standard Visual Studio Test Results window. You can download the code discussed in this blog entry from here: http://StephenWalther.com/downloads/Blog/JavaScriptUnitTesting/JavaScriptUnitTests.zip Before running this code, you need to first install the Microsoft Script Control which you can download from here: http://www.microsoft.com/downloads/en/details.aspx?FamilyID=d7e31492-2595-49e6-8c02-1426fec693ac

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  • Security Issues with Single Page Apps

    - by Stephen.Walther
    Last week, I was asked to do a code review of a Single Page App built using the ASP.NET Web API, Durandal, and Knockout (good stuff!). In particular, I was asked to investigate whether there any special security issues associated with building a Single Page App which are not present in the case of a traditional server-side ASP.NET application. In this blog entry, I discuss two areas in which you need to exercise extra caution when building a Single Page App. I discuss how Single Page Apps are extra vulnerable to both Cross-Site Scripting (XSS) attacks and Cross-Site Request Forgery (CSRF) attacks. This goal of this blog post is NOT to persuade you to avoid writing Single Page Apps. I’m a big fan of Single Page Apps. Instead, the goal is to ensure that you are fully aware of some of the security issues related to Single Page Apps and ensure that you know how to guard against them. Cross-Site Scripting (XSS) Attacks According to WhiteHat Security, over 65% of public websites are open to XSS attacks. That’s bad. By taking advantage of XSS holes in a website, a hacker can steal your credit cards, passwords, or bank account information. Any website that redisplays untrusted information is open to XSS attacks. Let me give you a simple example. Imagine that you want to display the name of the current user on a page. To do this, you create the following server-side ASP.NET page located at http://MajorBank.com/SomePage.aspx: <%@Page Language="C#" %> <html> <head> <title>Some Page</title> </head> <body> Welcome <%= Request["username"] %> </body> </html> Nothing fancy here. Notice that the page displays the current username by using Request[“username”]. Using Request[“username”] displays the username regardless of whether the username is present in a cookie, a form field, or a query string variable. Unfortunately, by using Request[“username”] to redisplay untrusted information, you have now opened your website to XSS attacks. Here’s how. Imagine that an evil hacker creates the following link on another website (hackers.com): <a href="/SomePage.aspx?username=<script src=Evil.js></script>">Visit MajorBank</a> Notice that the link includes a query string variable named username and the value of the username variable is an HTML <SCRIPT> tag which points to a JavaScript file named Evil.js. When anyone clicks on the link, the <SCRIPT> tag will be injected into SomePage.aspx and the Evil.js script will be loaded and executed. What can a hacker do in the Evil.js script? Anything the hacker wants. For example, the hacker could display a popup dialog on the MajorBank.com site which asks the user to enter their password. The script could then post the password back to hackers.com and now the evil hacker has your secret password. ASP.NET Web Forms and ASP.NET MVC have two automatic safeguards against this type of attack: Request Validation and Automatic HTML Encoding. Protecting Coming In (Request Validation) In a server-side ASP.NET app, you are protected against the XSS attack described above by a feature named Request Validation. If you attempt to submit “potentially dangerous” content — such as a JavaScript <SCRIPT> tag — in a form field or query string variable then you get an exception. Unfortunately, Request Validation only applies to server-side apps. Request Validation does not help in the case of a Single Page App. In particular, the ASP.NET Web API does not pay attention to Request Validation. You can post any content you want – including <SCRIPT> tags – to an ASP.NET Web API action. For example, the following HTML page contains a form. When you submit the form, the form data is submitted to an ASP.NET Web API controller on the server using an Ajax request: <!DOCTYPE html> <html xmlns="http://www.w3.org/1999/xhtml"> <head> <title></title> </head> <body> <form data-bind="submit:submit"> <div> <label> User Name: <input data-bind="value:user.userName" /> </label> </div> <div> <label> Email: <input data-bind="value:user.email" /> </label> </div> <div> <input type="submit" value="Submit" /> </div> </form> <script src="Scripts/jquery-1.7.1.js"></script> <script src="Scripts/knockout-2.1.0.js"></script> <script> var viewModel = { user: { userName: ko.observable(), email: ko.observable() }, submit: function () { $.post("/api/users", ko.toJS(this.user)); } }; ko.applyBindings(viewModel); </script> </body> </html> The form above is using Knockout to bind the form fields to a view model. When you submit the form, the view model is submitted to an ASP.NET Web API action on the server. Here’s the server-side ASP.NET Web API controller and model class: public class UsersController : ApiController { public HttpResponseMessage Post(UserViewModel user) { var userName = user.UserName; return Request.CreateResponse(HttpStatusCode.OK); } } public class UserViewModel { public string UserName { get; set; } public string Email { get; set; } } If you submit the HTML form, you don’t get an error. The “potentially dangerous” content is passed to the server without any exception being thrown. In the screenshot below, you can see that I was able to post a username form field with the value “<script>alert(‘boo’)</script”. So what this means is that you do not get automatic Request Validation in the case of a Single Page App. You need to be extra careful in a Single Page App about ensuring that you do not display untrusted content because you don’t have the Request Validation safety net which you have in a traditional server-side ASP.NET app. Protecting Going Out (Automatic HTML Encoding) Server-side ASP.NET also protects you from XSS attacks when you render content. By default, all content rendered by the razor view engine is HTML encoded. For example, the following razor view displays the text “<b>Hello!</b>” instead of the text “Hello!” in bold: @{ var message = "<b>Hello!</b>"; } @message   If you don’t want to render content as HTML encoded in razor then you need to take the extra step of using the @Html.Raw() helper. In a Web Form page, if you use <%: %> instead of <%= %> then you get automatic HTML Encoding: <%@ Page Language="C#" %> <% var message = "<b>Hello!</b>"; %> <%: message %> This automatic HTML Encoding will prevent many types of XSS attacks. It prevents <script> tags from being rendered and only allows &lt;script&gt; tags to be rendered which are useless for executing JavaScript. (This automatic HTML encoding does not protect you from all forms of XSS attacks. For example, you can assign the value “javascript:alert(‘evil’)” to the Hyperlink control’s NavigateUrl property and execute the JavaScript). The situation with Knockout is more complicated. If you use the Knockout TEXT binding then you get HTML encoded content. On the other hand, if you use the HTML binding then you do not: <!-- This JavaScript DOES NOT execute --> <div data-bind="text:someProp"></div> <!-- This Javacript DOES execute --> <div data-bind="html:someProp"></div> <script src="Scripts/jquery-1.7.1.js"></script> <script src="Scripts/knockout-2.1.0.js"></script> <script> var viewModel = { someProp : "<script>alert('Evil!')<" + "/script>" }; ko.applyBindings(viewModel); </script>   So, in the page above, the DIV element which uses the TEXT binding is safe from XSS attacks. According to the Knockout documentation: “Since this binding sets your text value using a text node, it’s safe to set any string value without risking HTML or script injection.” Just like server-side HTML encoding, Knockout does not protect you from all types of XSS attacks. For example, there is nothing in Knockout which prevents you from binding JavaScript to a hyperlink like this: <a data-bind="attr:{href:homePageUrl}">Go</a> <script src="Scripts/jquery-1.7.1.min.js"></script> <script src="Scripts/knockout-2.1.0.js"></script> <script> var viewModel = { homePageUrl: "javascript:alert('evil!')" }; ko.applyBindings(viewModel); </script> In the page above, the value “javascript:alert(‘evil’)” is bound to the HREF attribute using Knockout. When you click the link, the JavaScript executes. Cross-Site Request Forgery (CSRF) Attacks Cross-Site Request Forgery (CSRF) attacks rely on the fact that a session cookie does not expire until you close your browser. In particular, if you visit and login to MajorBank.com and then you navigate to Hackers.com then you will still be authenticated against MajorBank.com even after you navigate to Hackers.com. Because MajorBank.com cannot tell whether a request is coming from MajorBank.com or Hackers.com, Hackers.com can submit requests to MajorBank.com pretending to be you. For example, Hackers.com can post an HTML form from Hackers.com to MajorBank.com and change your email address at MajorBank.com. Hackers.com can post a form to MajorBank.com using your authentication cookie. After your email address has been changed, by using a password reset page at MajorBank.com, a hacker can access your bank account. To prevent CSRF attacks, you need some mechanism for detecting whether a request is coming from a page loaded from your website or whether the request is coming from some other website. The recommended way of preventing Cross-Site Request Forgery attacks is to use the “Synchronizer Token Pattern” as described here: https://www.owasp.org/index.php/Cross-Site_Request_Forgery_%28CSRF%29_Prevention_Cheat_Sheet When using the Synchronizer Token Pattern, you include a hidden input field which contains a random token whenever you display an HTML form. When the user opens the form, you add a cookie to the user’s browser with the same random token. When the user posts the form, you verify that the hidden form token and the cookie token match. Preventing Cross-Site Request Forgery Attacks with ASP.NET MVC ASP.NET gives you a helper and an action filter which you can use to thwart Cross-Site Request Forgery attacks. For example, the following razor form for creating a product shows how you use the @Html.AntiForgeryToken() helper: @model MvcApplication2.Models.Product <h2>Create Product</h2> @using (Html.BeginForm()) { @Html.AntiForgeryToken(); <div> @Html.LabelFor( p => p.Name, "Product Name:") @Html.TextBoxFor( p => p.Name) </div> <div> @Html.LabelFor( p => p.Price, "Product Price:") @Html.TextBoxFor( p => p.Price) </div> <input type="submit" /> } The @Html.AntiForgeryToken() helper generates a random token and assigns a serialized version of the same random token to both a cookie and a hidden form field. (Actually, if you dive into the source code, the AntiForgeryToken() does something a little more complex because it takes advantage of a user’s identity when generating the token). Here’s what the hidden form field looks like: <input name=”__RequestVerificationToken” type=”hidden” value=”NqqZGAmlDHh6fPTNR_mti3nYGUDgpIkCiJHnEEL59S7FNToyyeSo7v4AfzF2i67Cv0qTB1TgmZcqiVtgdkW2NnXgEcBc-iBts0x6WAIShtM1″ /> And here’s what the cookie looks like using the Google Chrome developer toolbar: You use the [ValidateAntiForgeryToken] action filter on the controller action which is the recipient of the form post to validate that the token in the hidden form field matches the token in the cookie. If the tokens don’t match then validation fails and you can’t post the form: public ActionResult Create() { return View(); } [ValidateAntiForgeryToken] [HttpPost] public ActionResult Create(Product productToCreate) { if (ModelState.IsValid) { // save product to db return RedirectToAction("Index"); } return View(); } How does this all work? Let’s imagine that a hacker has copied the Create Product page from MajorBank.com to Hackers.com – the hacker grabs the HTML source and places it at Hackers.com. Now, imagine that the hacker trick you into submitting the Create Product form from Hackers.com to MajorBank.com. You’ll get the following exception: The Cross-Site Request Forgery attack is blocked because the anti-forgery token included in the Create Product form at Hackers.com won’t match the anti-forgery token stored in the cookie in your browser. The tokens were generated at different times for different users so the attack fails. Preventing Cross-Site Request Forgery Attacks with a Single Page App In a Single Page App, you can’t prevent Cross-Site Request Forgery attacks using the same method as a server-side ASP.NET MVC app. In a Single Page App, HTML forms are not generated on the server. Instead, in a Single Page App, forms are loaded dynamically in the browser. Phil Haack has a blog post on this topic where he discusses passing the anti-forgery token in an Ajax header instead of a hidden form field. He also describes how you can create a custom anti-forgery token attribute to compare the token in the Ajax header and the token in the cookie. See: http://haacked.com/archive/2011/10/10/preventing-csrf-with-ajax.aspx Also, take a look at Johan’s update to Phil Haack’s original post: http://johan.driessen.se/posts/Updated-Anti-XSRF-Validation-for-ASP.NET-MVC-4-RC (Other server frameworks such as Rails and Django do something similar. For example, Rails uses an X-CSRF-Token to prevent CSRF attacks which you generate on the server – see http://excid3.com/blog/rails-tip-2-include-csrf-token-with-every-ajax-request/#.UTFtgDDkvL8 ). For example, if you are creating a Durandal app, then you can use the following razor view for your one and only server-side page: @{ Layout = null; } <!DOCTYPE html> <html> <head> <title>Index</title> </head> <body> @Html.AntiForgeryToken() <div id="applicationHost"> Loading app.... </div> @Scripts.Render("~/scripts/vendor") <script type="text/javascript" src="~/App/durandal/amd/require.js" data-main="/App/main"></script> </body> </html> Notice that this page includes a call to @Html.AntiForgeryToken() to generate the anti-forgery token. Then, whenever you make an Ajax request in the Durandal app, you can retrieve the anti-forgery token from the razor view and pass the token as a header: var csrfToken = $("input[name='__RequestVerificationToken']").val(); $.ajax({ headers: { __RequestVerificationToken: csrfToken }, type: "POST", dataType: "json", contentType: 'application/json; charset=utf-8', url: "/api/products", data: JSON.stringify({ name: "Milk", price: 2.33 }), statusCode: { 200: function () { alert("Success!"); } } }); Use the following code to create an action filter which you can use to match the header and cookie tokens: using System.Linq; using System.Net.Http; using System.Web.Helpers; using System.Web.Http.Controllers; namespace MvcApplication2.Infrastructure { public class ValidateAjaxAntiForgeryToken : System.Web.Http.AuthorizeAttribute { protected override bool IsAuthorized(HttpActionContext actionContext) { var headerToken = actionContext .Request .Headers .GetValues("__RequestVerificationToken") .FirstOrDefault(); ; var cookieToken = actionContext .Request .Headers .GetCookies() .Select(c => c[AntiForgeryConfig.CookieName]) .FirstOrDefault(); // check for missing cookie or header if (cookieToken == null || headerToken == null) { return false; } // ensure that the cookie matches the header try { AntiForgery.Validate(cookieToken.Value, headerToken); } catch { return false; } return base.IsAuthorized(actionContext); } } } Notice that the action filter derives from the base AuthorizeAttribute. The ValidateAjaxAntiForgeryToken only works when the user is authenticated and it will not work for anonymous requests. Add the action filter to your ASP.NET Web API controller actions like this: [ValidateAjaxAntiForgeryToken] public HttpResponseMessage PostProduct(Product productToCreate) { // add product to db return Request.CreateResponse(HttpStatusCode.OK); } After you complete these steps, it won’t be possible for a hacker to pretend to be you at Hackers.com and submit a form to MajorBank.com. The header token used in the Ajax request won’t travel to Hackers.com. This approach works, but I am not entirely happy with it. The one thing that I don’t like about this approach is that it creates a hard dependency on using razor. Your single page in your Single Page App must be generated from a server-side razor view. A better solution would be to generate the anti-forgery token in JavaScript. Unfortunately, until all browsers support a way to generate cryptographically strong random numbers – for example, by supporting the window.crypto.getRandomValues() method — there is no good way to generate anti-forgery tokens in JavaScript. So, at least right now, the best solution for generating the tokens is the server-side solution with the (regrettable) dependency on razor. Conclusion The goal of this blog entry was to explore some ways in which you need to handle security differently in the case of a Single Page App than in the case of a traditional server app. In particular, I focused on how to prevent Cross-Site Scripting and Cross-Site Request Forgery attacks in the case of a Single Page App. I want to emphasize that I am not suggesting that Single Page Apps are inherently less secure than server-side apps. Whatever type of web application you build – regardless of whether it is a Single Page App, an ASP.NET MVC app, an ASP.NET Web Forms app, or a Rails app – you must constantly guard against security vulnerabilities.

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  • RHEL - blocked FC remote port time out: saving binding

    - by Dev G
    My Server went into a faulty state since the database could not write on the partition. I found out that the partition went into Read Only mode. Finally to fix it, I had to do a hard reboot. Linux 2.6.18-164.el5PAE #1 SMP Tue Aug 18 15:59:11 EDT 2009 i686 i686 i386 GNU/Linux /var/log/messages Oct 31 00:56:45 ota3g1 Had[17275]: VCS ERROR V-16-1-10214 Concurrency Violation:CurrentCount increased above 1 for failover group sg_network Oct 31 00:57:05 ota3g1 Had[17275]: VCS CRITICAL V-16-1-50086 CPU usage on ota3g1.mtsallstream.com is 100% Oct 31 01:01:47 ota3g1 Had[17275]: VCS ERROR V-16-1-10214 Concurrency Violation:CurrentCount increased above 1 for failover group sg_network Oct 31 01:06:50 ota3g1 Had[17275]: VCS ERROR V-16-1-10214 Concurrency Violation:CurrentCount increased above 1 for failover group sg_network Oct 31 01:11:52 ota3g1 Had[17275]: VCS ERROR V-16-1-10214 Concurrency Violation:CurrentCount increased above 1 for failover group sg_network Oct 31 01:12:10 ota3g1 kernel: lpfc 0000:29:00.1: 1:1305 Link Down Event x2 received Data: x2 x20 x80000 x0 x0 Oct 31 01:12:10 ota3g1 kernel: lpfc 0000:29:00.1: 1:1303 Link Up Event x3 received Data: x3 x1 x10 x1 x0 x0 0 Oct 31 01:12:12 ota3g1 kernel: lpfc 0000:29:00.1: 1:1305 Link Down Event x4 received Data: x4 x20 x80000 x0 x0 Oct 31 01:12:40 ota3g1 kernel: rport-8:0-0: blocked FC remote port time out: saving binding Oct 31 01:12:40 ota3g1 kernel: lpfc 0000:29:00.1: 1:(0):0203 Devloss timeout on WWPN 20:25:00:a0:b8:74:f5:65 NPort x0000e4 Data: x0 x7 x0 Oct 31 01:12:40 ota3g1 kernel: sd 8:0:0:4: SCSI error: return code = 0x00010000 Oct 31 01:12:40 ota3g1 kernel: end_request: I/O error, dev sdi, sector 38617577 Oct 31 01:12:40 ota3g1 kernel: sd 8:0:0:4: SCSI error: return code = 0x00010000 Oct 31 01:12:40 ota3g1 kernel: end_request: I/O error, dev sdi, sector 283532153 Oct 31 01:12:40 ota3g1 kernel: sd 8:0:0:4: SCSI error: return code = 0x00010000 Oct 31 01:12:40 ota3g1 kernel: end_request: I/O error, dev sdi, sector 90825 Oct 31 01:12:40 ota3g1 kernel: Aborting journal on device dm-16. Oct 31 01:12:40 ota3g1 kernel: sd 8:0:0:4: SCSI error: return code = 0x00010000 Oct 31 01:12:40 ota3g1 kernel: end_request: I/O error, dev sdi, sector 868841 Oct 31 01:12:40 ota3g1 kernel: sd 8:0:0:4: SCSI error: return code = 0x00010000 Oct 31 01:12:40 ota3g1 kernel: Aborting journal on device dm-10. Oct 31 01:12:40 ota3g1 kernel: end_request: I/O error, dev sdi, sector 37759889 Oct 31 01:12:40 ota3g1 kernel: sd 8:0:0:4: SCSI error: return code = 0x00010000 Oct 31 01:12:40 ota3g1 kernel: end_request: I/O error, dev sdi, sector 283349449 Oct 31 01:12:40 ota3g1 kernel: printk: 6 messages suppressed. Oct 31 01:12:40 ota3g1 kernel: Aborting journal on device dm-12. Oct 31 01:12:40 ota3g1 kernel: EXT3-fs error (device dm-12) in ext3_reserve_inode_write: Journal has aborted Oct 31 01:12:40 ota3g1 kernel: Buffer I/O error on device dm-16, logical block 1545 Oct 31 01:12:40 ota3g1 kernel: lost page write due to I/O error on dm-16 Oct 31 01:12:40 ota3g1 kernel: sd 8:0:0:4: SCSI error: return code = 0x00010000 Oct 31 01:12:40 ota3g1 kernel: end_request: I/O error, dev sdi, sector 12745 Oct 31 01:12:40 ota3g1 kernel: Buffer I/O error on device dm-10, logical block 1545 Oct 31 01:12:40 ota3g1 kernel: EXT3-fs error (device dm-16) in ext3_reserve_inode_write: Journal has aborted Oct 31 01:12:40 ota3g1 kernel: lost page write due to I/O error on dm-10 Oct 31 01:12:40 ota3g1 kernel: sd 8:0:0:4: SCSI error: return code = 0x00010000 Oct 31 01:12:40 ota3g1 kernel: end_request: I/O error, dev sdi, sector 37749121 Oct 31 01:12:40 ota3g1 kernel: Buffer I/O error on device dm-12, logical block 0 Oct 31 01:12:40 ota3g1 kernel: lost page write due to I/O error on dm-12 Oct 31 01:12:40 ota3g1 kernel: sd 8:0:0:4: SCSI error: return code = 0x00010000 Oct 31 01:12:40 ota3g1 kernel: EXT3-fs error (device dm-12) in ext3_dirty_inode: Journal has aborted Oct 31 01:12:40 ota3g1 kernel: end_request: I/O error, dev sdi, sector 37757897 Oct 31 01:12:40 ota3g1 kernel: Buffer I/O error on device dm-12, logical block 1097 Oct 31 01:12:40 ota3g1 kernel: lost page write due to I/O error on dm-12 Oct 31 01:12:40 ota3g1 kernel: sd 8:0:0:4: SCSI error: return code = 0x00010000 Oct 31 01:12:40 ota3g1 kernel: end_request: I/O error, dev sdi, sector 283337089 Oct 31 01:12:40 ota3g1 kernel: Buffer I/O error on device dm-16, logical block 0 Oct 31 01:12:40 ota3g1 kernel: lost page write due to I/O error on dm-16 Oct 31 01:12:40 ota3g1 kernel: sd 8:0:0:4: SCSI error: return code = 0x00010000 Oct 31 01:12:40 ota3g1 kernel: EXT3-fs error (device dm-16) in ext3_dirty_inode: Journal has aborted Oct 31 01:12:40 ota3g1 kernel: end_request: I/O error, dev sdi, sector 37749121 Oct 31 01:12:40 ota3g1 kernel: Buffer I/O error on device dm-12, logical block 0 Oct 31 01:12:41 ota3g1 kernel: lost page write due to I/O error on dm-12 Oct 31 01:12:41 ota3g1 kernel: sd 8:0:0:4: SCSI error: return code = 0x00010000 Oct 31 01:12:41 ota3g1 kernel: end_request: I/O error, dev sdi, sector 283337089 Oct 31 01:12:41 ota3g1 kernel: Buffer I/O error on device dm-16, logical block 0 Oct 31 01:12:41 ota3g1 kernel: lost page write due to I/O error on dm-16 Oct 31 01:12:41 ota3g1 kernel: sd 8:0:0:4: SCSI error: return code = 0x00010000 df -h Filesystem Size Used Avail Use% Mounted on /dev/mapper/cciss-root 4.9G 730M 3.9G 16% / /dev/mapper/cciss-home 9.7G 1.2G 8.1G 13% /home /dev/mapper/cciss-var 9.7G 494M 8.8G 6% /var /dev/mapper/cciss-usr 15G 2.6G 12G 19% /usr /dev/mapper/cciss-tmp 3.9G 153M 3.6G 5% /tmp /dev/sda1 996M 43M 902M 5% /boot tmpfs 5.9G 0 5.9G 0% /dev/shm /dev/mapper/cciss-product 25G 16G 7.4G 68% /product /dev/mapper/cciss-opt 20G 4.5G 14G 25% /opt /dev/mapper/dg_db1-vol_db1_system 18G 2.2G 15G 14% /database/OTADB/sys /dev/mapper/dg_db1-vol_db1_undo 18G 5.8G 12G 35% /database/OTADB/undo /dev/mapper/dg_db1-vol_db1_redo 8.9G 4.3G 4.2G 51% /database/OTADB/redo /dev/mapper/dg_db1-vol_db1_sgbd 8.9G 654M 7.8G 8% /database/OTADB/admin /dev/mapper/dg_db1-vol_db1_arch 98G 24G 69G 26% /database/OTADB/arch /dev/mapper/dg_db1-vol_db1_indexes 240G 14G 214G 6% /database/OTADB/index /dev/mapper/dg_db1-vol_db1_data 275G 47G 215G 18% /database/OTADB/data /dev/mapper/dg_dbrman-vol_db_rman 8.9G 351M 8.1G 5% /database/RMAN /dev/mapper/dg_app1-vol_app1 151G 113G 31G 79% /files/ota /etc/fstab /dev/cciss/root / ext3 defaults 1 1 /dev/cciss/home /home ext3 defaults 1 2 /dev/cciss/var /var ext3 defaults 1 2 /dev/cciss/usr /usr ext3 defaults 1 2 /dev/cciss/tmp /tmp ext3 defaults 1 2 LABEL=/boot /boot ext3 defaults 1 2 tmpfs /dev/shm tmpfs defaults 0 0 devpts /dev/pts devpts gid=5,mode=620 0 0 sysfs /sys sysfs defaults 0 0 proc /proc proc defaults 0 0 /dev/cciss/swap swap swap defaults 0 0 /dev/cciss/product /product ext3 defaults 1 2 /dev/cciss/opt /opt ext3 defaults 1 2 /dev/dg_db1/vol_db1_system /database/OTADB/sys ext3 defaults 1 2 /dev/dg_db1/vol_db1_undo /database/OTADB/undo ext3 defaults 1 2 /dev/dg_db1/vol_db1_redo /database/OTADB/redo ext3 defaults 1 2 /dev/dg_db1/vol_db1_sgbd /database/OTADB/admin ext3 defaults 1 2 /dev/dg_db1/vol_db1_arch /database/OTADB/arch ext3 defaults 1 2 /dev/dg_db1/vol_db1_indexes /database/OTADB/index ext3 defaults 1 2 /dev/dg_db1/vol_db1_data /database/OTADB/data ext3 defaults 1 2 /dev/dg_dbrman/vol_db_rman /database/RMAN ext3 defaults 1 2 /dev/dg_app1/vol_app1 /files/ota ext3 defaults 1 2 Thanks for all the help.

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  • "power limit notification" clobbering on 12G Dell servers with RHEL6

    - by Andrew B
    Server: Poweredge r620 OS: RHEL 6.4 Kernel: 2.6.32-358.18.1.el6.x86_64 I'm experiencing application alarms in my production environment. Critical CPU hungry processes are being starved of resources and causing a processing backlog. The problem is happening on all the 12th Generation Dell servers (r620s) in a recently deployed cluster. As near as I can tell, instances of this happening are matching up to peak CPU utilization, accompanied by massive amounts of "power limit notification" spam in dmesg. An excerpt of one of these events: Nov 7 10:15:15 someserver [.crit] CPU12: Core power limit notification (total events = 14) Nov 7 10:15:15 someserver [.crit] CPU0: Core power limit notification (total events = 14) Nov 7 10:15:15 someserver [.crit] CPU6: Core power limit notification (total events = 14) Nov 7 10:15:15 someserver [.crit] CPU14: Core power limit notification (total events = 14) Nov 7 10:15:15 someserver [.crit] CPU18: Core power limit notification (total events = 14) Nov 7 10:15:15 someserver [.crit] CPU2: Core power limit notification (total events = 14) Nov 7 10:15:15 someserver [.crit] CPU4: Core power limit notification (total events = 14) Nov 7 10:15:15 someserver [.crit] CPU16: Core power limit notification (total events = 14) Nov 7 10:15:15 someserver [.crit] CPU0: Package power limit notification (total events = 11) Nov 7 10:15:15 someserver [.crit] CPU6: Package power limit notification (total events = 13) Nov 7 10:15:15 someserver [.crit] CPU14: Package power limit notification (total events = 14) Nov 7 10:15:15 someserver [.crit] CPU18: Package power limit notification (total events = 14) Nov 7 10:15:15 someserver [.crit] CPU20: Core power limit notification (total events = 14) Nov 7 10:15:15 someserver [.crit] CPU8: Core power limit notification (total events = 14) Nov 7 10:15:15 someserver [.crit] CPU2: Package power limit notification (total events = 12) Nov 7 10:15:15 someserver [.crit] CPU10: Core power limit notification (total events = 14) Nov 7 10:15:15 someserver [.crit] CPU22: Core power limit notification (total events = 14) Nov 7 10:15:15 someserver [.crit] CPU4: Package power limit notification (total events = 14) Nov 7 10:15:15 someserver [.crit] CPU16: Package power limit notification (total events = 13) Nov 7 10:15:15 someserver [.crit] CPU20: Package power limit notification (total events = 14) Nov 7 10:15:15 someserver [.crit] CPU8: Package power limit notification (total events = 14) Nov 7 10:15:15 someserver [.crit] CPU10: Package power limit notification (total events = 14) Nov 7 10:15:15 someserver [.crit] CPU22: Package power limit notification (total events = 14) Nov 7 10:15:15 someserver [.crit] CPU15: Core power limit notification (total events = 369) Nov 7 10:15:15 someserver [.crit] CPU3: Core power limit notification (total events = 369) Nov 7 10:15:15 someserver [.crit] CPU1: Core power limit notification (total events = 369) Nov 7 10:15:15 someserver [.crit] CPU5: Core power limit notification (total events = 369) Nov 7 10:15:15 someserver [.crit] CPU17: Core power limit notification (total events = 369) Nov 7 10:15:15 someserver [.crit] CPU13: Core power limit notification (total events = 369) Nov 7 10:15:15 someserver [.crit] CPU15: Package power limit notification (total events = 375) Nov 7 10:15:15 someserver [.crit] CPU3: Package power limit notification (total events = 374) Nov 7 10:15:15 someserver [.crit] CPU1: Package power limit notification (total events = 376) Nov 7 10:15:15 someserver [.crit] CPU5: Package power limit notification (total events = 376) Nov 7 10:15:15 someserver [.crit] CPU7: Core power limit notification (total events = 369) Nov 7 10:15:15 someserver [.crit] CPU19: Core power limit notification (total events = 369) Nov 7 10:15:15 someserver [.crit] CPU17: Package power limit notification (total events = 377) Nov 7 10:15:15 someserver [.crit] CPU9: Core power limit notification (total events = 369) Nov 7 10:15:15 someserver [.crit] CPU21: Core power limit notification (total events = 369) Nov 7 10:15:15 someserver [.crit] CPU23: Core power limit notification (total events = 369) Nov 7 10:15:15 someserver [.crit] CPU11: Core power limit notification (total events = 369) Nov 7 10:15:15 someserver [.crit] CPU13: Package power limit notification (total events = 376) Nov 7 10:15:15 someserver [.crit] CPU7: Package power limit notification (total events = 375) Nov 7 10:15:15 someserver [.crit] CPU19: Package power limit notification (total events = 375) Nov 7 10:15:15 someserver [.crit] CPU9: Package power limit notification (total events = 374) Nov 7 10:15:15 someserver [.crit] CPU21: Package power limit notification (total events = 375) Nov 7 10:15:15 someserver [.crit] CPU23: Package power limit notification (total events = 374) A little Google Fu reveals that this is typically associated with the CPU running hot, or voltage regulation kicking in. I don't think that's what is happening though. Temperature sensors for all servers in the cluster are running fine, Power Cap Policy is disabled in the iDRAC, and my System Profile is set to "Performance" on all of these servers: # omreport chassis biossetup | grep -A10 'System Profile' System Profile Settings ------------------------------------------ System Profile : Performance CPU Power Management : Maximum Performance Memory Frequency : Maximum Performance Turbo Boost : Enabled C1E : Disabled C States : Disabled Monitor/Mwait : Enabled Memory Patrol Scrub : Standard Memory Refresh Rate : 1x Memory Operating Voltage : Auto Collaborative CPU Performance Control : Disabled A Dell mailing list post describes the symptoms almost perfectly. Dell suggested that the author try using the Performance profile, but that didn't help. He ended up applying some settings in Dell's guide for configuring a server for low latency environments and one of those settings (or a combination thereof) seems to have fixed the problem. Kernel.org bug #36182 notes that power-limit interrupt debugging was enabled by default, which is causing performance degradation in scenarios where CPU voltage regulation is kicking in. A RHN KB article (RHN login required) mentions a problem impacting PE r620 and r720 servers not running the Performance profile, and recommends an update to a kernel released two weeks ago. ...Except we are running the Performance profile... Everything I can find online is running me in circles here. What's the heck is going on?

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  • Hyper-V File Server Clustering - at my wit’s end

    - by René Kåbis
    I am at my wit’s end with File Server clustering under Hyper-V. I am hoping that someone might be able to help me figure out this Gordian Knot of a technology that seems to have dead ends (like forcing cluster VMs to use iSCSI drives where normally-attached VHDX drives could suffice) where logic and reason would normally provide a logical solution. My hardware: I will be running three servers (in the end), but right now everything is taking place on one server. One of the secondary servers will exist purely as a witness/quorum, and another slightly more powerful one will be acting as an emergency backup (with additional storage, just not redundant) to hold the secondary AD VM and the other halves of a set of clustered VMs: the SQL VM and the file system VM. Please note, these each are the depreciated nodes of a cluster, the main nodes will be on the most powerful first machine. My heavy lifter is a machine that also contains all of the truly redundant storage on the network. If this gives anyone the heebie-geebies, too bad. It has a 6TB (usable) RAID-10 array, and will (in the end) hold the primary nodes of both aforementioned clusters, but is right now holding all VMs. This is, right now: DC01, DC02, SQL01, SQL02, FS01 & FS02. Eventually, I will be adding additional VMs to handle Exchange, Sharepoint and Lync, but only to this main server (the secondary server won't be able to handle more than three or four VMs, so why burden it? The AD, SQL & FS VMs are the most critical for the business). If anyone is now saying, “wait, what about a SAN or a NAS for the file servers?”, well too bad. What exists on the main machine is what I have to deal with. I followed these instructions, but I seem to be unable to get things to work. In order to make the file server truly redundant, I cannot trust any one machine to hold the only data store on the network. Therefore, I have created a set of iSCSI drives on the VM-host of the main machine, and attached one to each file server VM. The end result is that I want my FS01 to sit on the heavy lifter, along with its iSCSI “drive”, and FS02 will sit on the secondary machine with its own iSCSI “drive” there as well. That is, neither iSCSI drive will end up sitting on the same machine as the other. As such, the clustered FS will utterly duplicate the contents of the iSCSI drives between each other, so that if one physical machine (or the FS VM) goes toes-up, the other has got a full copy of the data on its own iSCSI drive. My problem occurs when I try to apply the file server role within the failover cluster manager. Actually, it is even before that -- it occurs when adding the disks. Since I have added each disk preferentially to a specific VM (by limiting the initiator by DNS hostname, and by adding two-way CHAP authentication), this forces each VM to be in control of its own iSCSI disk. However, when I try to add the disks to the Disks section of Storage within Failover Cluster Manager, the entire process fails for a random disk of the pair. That is, one will get online, but the other will remain offline because it does not have the correct “owner node”. I mean, really -- WTF? Of course it doesn’t have the right owner node, both drives are showing the same node name!! I cannot seem to have one drive show up with one node name as owner, and the other drive show up with the other node name as owner. And because both drives are not “online”, I cannot create a pool to apply to a cluster role. Talk about getting stuck between a rock and a hard place! I’ve got more to add, but my work is closing for the day and I have to wrap things up. I will try to add more tomorrow morning when I get in. My main objective is to have a file server VM on each machine, the storage on each machine, but a transparent failover in case one physical machine fails. Essentially, a failover FS that doesn’t care which machine fails -- the storage contents are replicated equally on each machine. Am I even heading in the right direction?

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  • IIS7 FTP Setup - An error occured during the authentication process. 530 End Login failed

    - by robmzd
    I'm having a problem very similar to IIS 7.5 FTP IIS Manager Users Login Fail (530) on Windows Server 2008 R2 Standard. I have created an FTP site and IIS Manager user but am having trouble logging in. I could really do with getting this working with the IIS Manager user rather than by creating a new system user since I'm fairly restricted with those accounts. Here is the output when connecting locally through command prompt: C:\Windows\system32>ftp localhost Connected to MYSERVER. 220 Microsoft FTP Service User (MYSERVER:(none)): MyFtpLogin 331 Password required for MyFtpLogin. Password: *** 530-User cannot log in. Win32 error: Logon failure: unknown user name or bad password. Error details: An error occured during the authentication process. 530 End Login failed. I have followed the guide to configure ftp with iis manager authentication in iis 7 and Adding FTP Publishing to a Web Site in IIS 7 Things I have done and checked: The FTP Service is installed (along with FTP Extensibility). Local Service and Network Service have been given access to the site folder Permission has been given to the config files Granted read/write permissions to the FTP Root folder The Management Service is installed and running Enable remote connections is ticked with 'Windows credentials or IIS manager credentials' selected The IIS Manager User has been added to the server (root connection in the IIS connections branch) The new FTP site has been added IIS Manager Authentication has been added to the FTP authentication providers The IIS Manager user has been added to the IIS Manager Permissions list for the site Added Read/Write permissions for the user in the FTP Authorization Rules Here's a section of the applicationHost config file associated with the FTP site <site name="MySite" id="8"> <application path="/" applicationPool="MyAppPool"> <virtualDirectory path="/" physicalPath="D:\Websites\MySite" /> </application> <bindings> <binding protocol="http" bindingInformation="*:80:www.mydomain.co.uk" /> <binding protocol="ftp" bindingInformation="*:21:www.mydomain.co.uk" /> </bindings> <ftpServer> <security> <ssl controlChannelPolicy="SslAllow" dataChannelPolicy="SslAllow" /> <authentication> <basicAuthentication enabled="true" /> <customAuthentication> <providers> <add name="IisManagerAuth" enabled="true" /> </providers> </customAuthentication> </authentication> </security> </ftpServer> </site> ... <location path="MySite"> <system.ftpServer> <security> <authorization> <add accessType="Allow" users="MyFtpLogin" permissions="Read, Write" /> </authorization> </security> </system.ftpServer> </location> If I connect to the Site (not FTP) from my local IIS Manager using the same IIS Manager account details then it connects fine, I can browse files and change settings as I would locally (though I don't seem to have an option to upload files). Trying to connect via FTP though either through the browser or FileZilla etc... gives me: Status: Resolving address of www.mydomain.co.uk Status: Connecting to 123.456.12.123:21... Status: Connection established, waiting for welcome message... Response: 220 Microsoft FTP Service Command: USER MyFtpLogin Response: 331 Password required for MyFtpLogin. Command: PASS ********* Response: 530 User cannot log in. Error: Critical error Error: Could not connect to server I have tried collecting etw traces for ftp sessions, in the logs I get a FailBasicLogon followed by a FailCustomLogon, but no other info: FailBasicLogon SessionId={cad26a97-225d-45ba-ab1f-f6acd9046e55} | ErrorCode=0x8007052E StartCustomLogon SessionId={cad26a97-225d-45ba-ab1f-f6acd9046e55} | LogonProvider=IisManagerAuth StartCallProvider SessionId={cad26a97-225d-45ba-ab1f-f6acd9046e55} | provider=IisManagerAuth EndCallProvider SessionId={cad26a97-225d-45ba-ab1f-f6acd9046e55} EndCustomLogon SessionId={cad26a97-225d-45ba-ab1f-f6acd9046e55} FailCustomLogon SessionId={cad26a97-225d-45ba-ab1f-f6acd9046e55} | ErrorCode=0x8007052E FailFtpCommand SessionId={cad26a97-225d-45ba-ab1f-f6acd9046e55} | ReturnValue=0x8007052E | SubStatus=ERROR_DURING_AUTHENTICATION In the normal FTP logs I just get: 2012-10-23 16:13:11 123.456.12.123 - 123.456.12.123 21 ControlChannelOpened - - 0 0 e2d4e935-fb31-4f2c-af79-78d75d47c18e - 2012-10-23 16:13:11 123.456.12.123 - 123.456.12.123 21 USER MyFtpLogin 331 0 0 e2d4e935-fb31-4f2c-af79-78d75d47c18e - 2012-10-23 16:13:11 123.456.12.123 - 123.456.12.123 21 PASS *** 530 1326 41 e2d4e935-fb31-4f2c-af79-78d75d47c18e - 2012-10-23 16:13:11 123.456.12.123 - 123.456.12.123 21 ControlChannelClosed - - 0 0 e2d4e935-fb31-4f2c-af79-78d75d47c18e - If anyone has any ideas than I would be very grateful to hear them. Many thanks.

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  • Application Event Log keeps getting corrupted

    - by yakatz
    I recently asked about repairing a corrupt event log, because it seemed to be a one-off event. The event log has since exhibited the same behavior 3 times. We have been trying to find patterns, but so far we have found nothing. The server runs several ASP.NET applications and three scheduled tasks written in .NET. The last modified date of the event log once happened to be the same time as one of the scheduled tasks, but the others have not been. Any suggestions of where to look next or a way we can get any information out of a corrupt evtx file? The server is running critical e-commerce applications, so we want to keep the number of restarts required to a minimum. Edit: I ran DUMPEL and got very strange results. 1/9/2012 4:14:05 PM 1 100 1000 Application Error N/A SERVERNAME Faulting application name: w3wp.exe, version: 7.5.7601.17514, time stamp: 0x4ce7a5f8 Faulting module name: ntdll.dll, version: 6.1.7601.17514, time stamp: 0x4ce7ba58 Exception code: 0xc0000374 Fault offset: 0x000ce653 Faulting process id: 0x1070 Faulting application start time: 0x01cccf1386d30991 Faulting application path: C:\Windows\SysWOW64\inetsrv\w3wp.exe Faulting module path: C:\Windows\SysWOW64\ntdll.dll Report Id: dbf4f691-3b06-11e1-9025-005056a602e6 1/9/2012 4:14:07 PM 4 0 1001 Windows Error Reporting N/A SERVERNAME Fault bucket , type 0 Event Name: APPCRASH Response: Not available Cab Id: 0 Problem signature: P1: w3wp.exe P2: 7.5.7601.17514 P3: 4ce7a5f8 P4: StackHash_79d9 P5: 6.1.7601.17514 P6: 4ce7ba58 P7: c0000374 P8: 000ce653 P9: P10: Attached files: C:\Windows\Temp\WER975.tmp.appcompat.txt C:\Windows\Temp\WERA03.tmp.WERInternalMetadata.xml C:\Windows\Temp\WERA13.tmp.hdmp C:\Windows\Temp\WERD21.tmp.mdmp These files may be available here: C:\ProgramData\Microsoft\Windows\WER\ReportQueue\AppCrash_w3wp.exe_cd7d09dfc84119d82a2ac6a789038bd5661acfb_cab_128f0e67 Analysis symbol: Rechecking for solution: 0 Report Id: dbf4f691-3b06-11e1-9025-005056a602e6 Report Status: 4 1/9/2012 4:14:07 PM 4 0 1001 Windows Error Reporting N/A SERVERNAME Fault bucket , type 0 Event Name: APPCRASH Response: Not available Cab Id: 0 Problem signature: P1: w3wp.exe P2: 7.5.7601.17514 P3: 4ce7a5f8 P4: StackHash_79d9 P5: 6.1.7601.17514 P6: 4ce7ba58 P7: c0000374 P8: 000ce653 P9: P10: Attached files: C:\Windows\Temp\WER975.tmp.appcompat.txt C:\Windows\Temp\WERA03.tmp.WERInternalMetadata.xml C:\Windows\Temp\WERA13.tmp.hdmp C:\Windows\Temp\WERD21.tmp.mdmp These files may be available here: C:\ProgramData\Microsoft\Windows\WER\ReportQueue\AppCrash_w3wp.exe_cd7d09dfc84119d82a2ac6a789038bd5661acfb_cab_128f0e67 Analysis symbol: Rechecking for solution: 0 Report Id: dbf4f691-3b06-11e1-9025-005056a602e6 Report Status: 0 1/9/2012 4:14:12 PM 1 100 1000 Application Error N/A SERVERNAME Faulting application name: w3wp.exe, version: 7.5.7601.17514, time stamp: 0x4ce7a5f8 Faulting module name: ntdll.dll, version: 6.1.7601.17514, time stamp: 0x4ce7ba58 Exception code: 0xc0000374 Fault offset: 0x000ce653 Faulting process id: 0x16ac Faulting application start time: 0x01cccf139f475c0c Faulting application path: C:\Windows\SysWOW64\inetsrv\w3wp.exe Faulting module path: C:\Windows\SysWOW64\ntdll.dll Report Id: e03bae70-3b06-11e1-9025-005056a602e6 1/9/2012 4:14:16 PM 4 0 1001 Windows Error Reporting N/A SERVERNAME Fault bucket , type 0 Event Name: APPCRASH Response: Not available Cab Id: 0 Problem signature: P1: w3wp.exe P2: 7.5.7601.17514 P3: 4ce7a5f8 P4: StackHash_9c6c P5: 6.1.7601.17514 P6: 4ce7ba58 P7: c0000374 P8: 000ce653 P9: P10: Attached files: C:\Windows\Temp\WER2579.tmp.appcompat.txt C:\Windows\Temp\WER25F7.tmp.WERInternalMetadata.xml C:\Windows\Temp\WER25F8.tmp.hdmp C:\Windows\Temp\WER28F6.tmp.mdmp These files may be available here: C:\ProgramData\Microsoft\Windows\WER\ReportQueue\AppCrash_w3wp.exe_c49a67649524ad11b64bbf809211bc5ba742a3d6_cab_0b63321b Analysis symbol: Rechecking for solution: 0 Report Id: e03bae70-3b06-11e1-9025-005056a602e6 Report Status: 4 1/9/2012 4:14:16 PM 4 0 1001 Windows Error Reporting N/A SERVERNAME Fault bucket , type 0 Event Name: APPCRASH Response: Not available Cab Id: 0 Problem signature: P1: w3wp.exe P2: 7.5.7601.17514 P3: 4ce7a5f8 P4: StackHash_9c6c P5: 6.1.7601.17514 P6: 4ce7ba58 P7: c0000374 P8: 000ce653 P9: P10: Attached files: C:\Windows\Temp\WER2579.tmp.appcompat.txt C:\Windows\Temp\WER25F7.tmp.WERInternalMetadata.xml C:\Windows\Temp\WER25F8.tmp.hdmp C:\Windows\Temp\WER28F6.tmp.mdmp These files may be available here: C:\ProgramData\Microsoft\Windows\WER\ReportQueue\AppCrash_w3wp.exe_c49a67649524ad11b64bbf809211bc5ba742a3d6_cab_0b63321b Analysis symbol: Rechecking for solution: 0 Report Id: e03bae70-3b06-11e1-9025-005056a602e6 Report Status: 0 1/9/2012 4:14:21 PM 1 100 1000 Application Error N/A SERVERNAME Faulting application name: w3wp.exe, version: 7.5.7601.17514, time stamp: 0x4ce7a5f8 Faulting module name: ntdll.dll, version: 6.1.7601.17514, time stamp: 0x4ce7ba58 Exception code: 0xc0000374 Fault offset: 0x000ce653 Faulting process id: 0x17f8 Faulting application start time: 0x01cccf13a4ba5126 Faulting application path: C:\Windows\SysWOW64\inetsrv\w3wp.exe Faulting module path: C:\Windows\SysWOW64\ntdll.dll Report Id: e57a0a85-3b06-11e1-9025-005056a602e6 1/9/2012 4:14:21 PM 4 0 1001 Windows Error Reporting N/A SERVERNAME Fault bucket , type 0 Event Name: APPCRASH Response: Not available Cab Id: 0 Problem signature: P1: w3wp.exe P2: 7.5.7601.17514 P3: 4ce7a5f8 P4: StackHash_9c6c P5: 6.1.7601.17514 P6: 4ce7ba58 P7: c0000374 P8: 000ce653 P9: P10: Attached files: These files may be available here: C:\ProgramData\Microsoft\Windows\WER\ReportQueue\AppCrash_w3wp.exe_c49a67649524ad11b64bbf809211bc5ba742a3d6_1cfb4872 Analysis symbol: Rechecking for solution: 0 Report Id: e57a0a85-3b06-11e1-9025-005056a602e6 Report Status: 4 1/9/2012 4:14:21 PM 4 0 1001 Windows Error Reporting N/A SERVERNAME Fault bucket , type 0 Event Name: APPCRASH Response: Not available Cab Id: 0 Problem signature: P1: w3wp.exe P2: 7.5.7601.17514 P3: 4ce7a5f8 P4: StackHash_9c6c P5: 6.1.7601.17514 P6: 4ce7ba58 P7: c0000374 P8: 000ce653 P9: P10: Attached files: These files may be available here: C:\ProgramData\Microsoft\Windows\WER\ReportQueue\AppCrash_w3wp.exe_c49a67649524ad11b64bbf809211bc5ba742a3d6_1cfb4872 Analysis symbol: Rechecking for solution: 0 Report Id: e57a0a85-3b06-11e1-9025-005056a602e6 Report Status: 0 None of the files referenced actually exist (not even in WER ReportArchive). These should not be the only events mentioned. The log file has been cleared twice since January 9, so those events should not even be listed at all.

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  • snmptt not translating traps, even with translate_log_trap_oid=1

    - by mbrownnyc
    I am having some trouble configuring snmptt to properly translate snmp traps. The following is a problem: /etc/snmp/snmptt.conf reflects: EVENT fgFmTrapIfChange .1.3.6.1.4.1.12356.101.6.0.1004 "Status Events" Critical FORMAT $* EXEC /usr/local/nagios/libexec/eventhandlers/submit_check_result $r "snmp_traps" 2 "$O: $+*" "$*" SDESC Trap is sent to the managing FortiManager if an interface IP is changed Variables: 1: fnSysSerial 2: ifName 3: fgManIfIp 4: fgManIfMask EDESC when a trap is received, /var/log/messages reflects: Sep 6 12:07:32 SNMPMANAGERHOST snmptrapd[15385]: 2012-09-06 12:07:32 <UNKNOWN> [UDP: [192.168.100.2]:162->[192.168.100.31]]: #012.1.3.6.1.2.1.1.3.0 = Timeticks: (707253943) 81 days, 20:35:39.43 #011.1.3.6.1.6.3.1.1.4.1.0 = OID: .1.3.6.1.4.1.12356.101.6.0.1004 #011.1.3.6.1.4.1.12356.100.1.1.1.0 = STRING: FGTNNNNNNNNN #011.1.3.6.1.2.1.31.1.1.1.1.10 = STRING: internal4 #011.1.3.6.1.4.1.12356.101.6.2.1.0 = IpAddress: 192.168.65.100 #011.1.3.6.1.4.1.12356.101.6.2.2.0 = IpAddress: 255.255.255.0 Sep 6 12:07:37 SNMPMANAGERHOST icinga: EXTERNAL COMMAND: PROCESS_SERVICE_CHECK_RESULT; 192.168.100.2; snmp_traps; 2; enterprises.12356.101.6.0.1004: enterprises.12356.100.1.1.1.0:FGTNNNNNNNNN ifName.10:internal4 enterprises.12356.101.6.2.1.0:192.168.65.100 enterprises.12356.101.6.2.2.0:255.255.255.0 Since the icinga entry reflects the EXEC, it's obvious there is no translations occurring by snmptt. I have verified that translate_log_trap_oid and net_snmp_perl_enable is enabled in snmptt.ini When using --debug=1 to start snmptt, I see the following in the --debugfile: ********** Net-SNMP version 5.05 Perl module enabled ********** The main NET-SNMP version is reported as NET-SNMP version: 5.5. What else can be done to verify that snmptt is configured properly to translate traps? I have run snmptt-net-snmp-test to verify whatever net-snmp-perl version I have installed properly supports translations. The output indicates it does. /root/snmptt_1.3/snmptt-net-snmp-test --best_guess=2 SNMPTT Net-SNMP Test v1.0 (c) 2003 Alex Burger http://snmptt.sourceforge.net MIBS:RFC1213-MIB best_guess: 2 Testing translateObj ******************** Testing: .1.3.6.1.2.1.1.1, long_names=disabled, include_module=disabled Test passed. Result: sysDescr Testing: .1.3.6.1.2.1.1.1, long_names=disabled, include_module=enabled Test passed. Result: RFC1213-MIB::sysDescr Testing: .1.3.6.1.2.1.1.1, long_names=enabled, include_module=disabled Test passed. Result: .iso.org.dod.internet.mgmt.mib-2.system.sysDescr Testing: .1.3.6.1.2.1.1.1, long_names=enabled, include_module=enabled Test passed. Result: RFC1213-MIB::.iso.org.dod.internet.mgmt.mib-2.system.sysDescr Testing: sysDescr, long_names=disabled, include_module=disabled Test passed. Result: .1.3.6.1.2.1.1.1 Testing: RFC1213-MIB::sysDescr, long_names=disabled, include_module=disabled Test passed. Result: .1.3.6.1.2.1.1.1 Testing: system.sysDescr, long_names=disabled, include_module=disabled Test passed. Result: .1.3.6.1.2.1.1.1 Testing: RFC1213-MIB::system.sysDescr, long_names=disabled, include_module=disabled Test passed. Result: .1.3.6.1.2.1.1.1 Testing: .iso.org.dod.internet.mgmt.mib-2.system.sysDescr, long_names=disabled, include_module=disabled Test passed. Result: .1.3.6.1.2.1.1.1 Testing getType *************** Testing: .1.3.6.1.2.1.4.1 Test passed. Result: INTEGER Testing: ipForwarding Test passed. Result: INTEGER Testing Description ******************* Test passed. Result: ------------------------------------------------- The indication of whether this entity is acting as an IP gateway in respect to the forwarding of datagrams received by, but not addressed to, this entity. IP gateways forward datagrams. IP hosts do not (except those source-routed via the host). Note that for some managed nodes, this object may take on only a subset of the values possible. Accordingly, it is appropriate for an agent to return a `badValue' response if a management station attempts to change this object to an inappropriate value. ------------------------------------------------- I have manually gone through the MIB with the definition that's not resolving, and verified that it is properly linking back to the proper resolved definition. It is: FORTINET-FORTIGATE-MIB.txt contains: fgFmTrapIfChange NOTIFICATION-TYPE OBJECTS { fnSysSerial, ifName, fgManIfIp, fgManIfMask } STATUS current DESCRIPTION "Trap is sent to the managing FortiManager if an interface IP is changed" ::= { fgFmTrapPrefix 1004 } fgFmTrapPrefix OBJECT IDENTIFIER ::= { fgMgmt 0 } fgMgmt OBJECT IDENTIFIER ::= { fnFortiGateMib 6 } fnFortiGateMib ::= { fortinet 101 } IMPORTS FnBoolState, FnIndex, fnAdminEntry, fnSysSerial, fortinet FROM FORTINET-CORE-MIB fortinet MODULE-IDENTITY ::= { enterprises 12356 } LOOKS GOOD!!!!! 1.3.6.1.4.1.12356.101.6.0.1004 I've exhausted all the documentation and even posted fruitlessly in the snmptt-users mailing list. I can not prove it is the MIB. Why would snmptt fail to translate traps? Thanks, Matt

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  • Mount TMPFS instead of ro /dev

    - by schiggn
    I am working on a ARM-Based embedded system with a custom Debian Linux based on kernel 2.6.31. In the final system, the Root file system is stored as squashfs on flash. Now, the folder /dev is created by udev, but since there is no hot plugging functionality needed and booting time is critical, I wanted to delete udev and "hard code" the /dev folder (read here, page 5). because i still need to change parameters of the devices (with ioctl /sysfs) this does not work for me in this case. so i thought of mounting a tmpfs on /dev and change the parameters there. is this possible? and how to do best? my approach would be: delete /dev from RFS create tar containing basic devices mount tmpfs /dev untar tar-file into /dev change parameters Could this work? Do you see any problems? I found out, that you can mount on top of already mounted mount point, is it somehow possible just to take data with while mounting the new file system? if so that would be very convenient! Thanks Update: I just tried that out, but I'm stuck at a certain point. I packed all my devices into devices.tar, packed it into /usr of my squashfs and added the following lines to mountkernfs.sh, which is executed right after INIT. #mount /dev on tmpfs echo -n "Mounting /dev on tmpfs..." mount -o size=5M,mode=0755 -t tmpfs tmpfs /dev mknod -m 600 /dev/console c 5 1 mknod -m 600 /dev/null c 1 3 echo "done." echo -n "Populating /dev..." tar -xf /usr/devices.tar -C /dev echo "done." This works fine on the version over NFS, if I place printf's in the code, I can see it executing, if I comment out the extracting part, its complaining about missing devices. Booting OK mmc0: new high speed SDHC card at address 0007 mmcblk0: mmc0:0007 SD04G 3.67 GiB mmcblk0: p1 IP-Config: Unable to set interface netmask (-22). Looking up port of RPC 100003/2 on 192.168.1.234 Looking up port of RPC 100005/1 on 192.168.1.234 VFS: Mounted root (nfs filesystem) on device 0:14. Freeing init memory: 136K INIT: version 2.86 booting Mounting /dev on tmpfs...done. Populating /dev...done. Initializing /var...done. Setting the system clock. System Clock set to: Thu Sep 13 11:26:23 UTC 2012. INIT: Entering runlevel: 2 UBI: attaching mtd8 to ubi0 Commenting out the extraction of the tar mmc0: new high speed SDHC card at address 0007 mmcblk0: mmc0:0007 SD04G 3.67 GiB mmcblk0: p1 IP-Config: Unable to set interface netmask (-22). Looking up port of RPC 100003/2 on 192.168.1.234 Looking up port of RPC 100005/1 on 192.168.1.234 VFS: Mounted root (nfs filesystem) on device 0:14. Freeing init memory: 136K INIT: version 2.86 booting Mounting /dev on tmpfs...done. Populating /dev...done. Initializing /var...done. Setting the system clock. Cannot access the Hardware Clock via any known method. Use the --debug option to see the details of our search for an access method. Unable to set System Clock to: Thu Sep 13 12:24:00 UTC 2012 ... (warning). INIT: Entering runlevel: 2 libubi: error!: cannot open "/dev/ubi_ctrl" So far so good. But if I pack the whole story into a squashfs and boot from there, it is acting strange. It's telling me while booting that it is unable to open an initial console and its throwing errors on mounting the UBIFS devices, but finally provides a login anyway. Over that my echo's are not executed. If I then log in, /dev is mounted as TMPFS as desired and all the devices reside inside. When I redo the "mount" command to mount the UBIFS partitions it is executed whitout problem and useable. From squashfs VFS: Mounted root (squashfs filesystem) readonly on device 31:15. Freeing init memory: 136K Warning: unable to open an initial console. mmc0: new high speed SDHC card at address 0007 mmcblk0: mmc0:0007 SD04G 3.67 GiB mmcblk0: p1 UBIFS error (pid 484): ubifs_get_sb: cannot open "ubi1_0", error -19 Additionally, a part of the rest of the bootscripts is still exexuted, but not all of them. Does anyone has a clue why? Other question, is 5MB enough/too much for /dev?

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  • Set up linux box for hosting a-z

    - by microchasm
    I am in the process of reinstalling the OS on a machine that will be used to host a couple of apps for our business. The apps will be local only; access from external clients will be via vpn only. The prior setup used a hosting control panel (Plesk) for most of the admin, and I was looking at using another similar piece of software for the reinstall - but I figured I should finally learn how it all works. I can do most of the things the software would do for me, but am unclear on the symbiosis of it all. This is all an attempt to further distance myself from the land of Configuration Programmer/Programmer, if at all possible. I can't find a full walkthrough anywhere for what I'm looking for, so I thought I'd put up this question, and if people can help me on the way I will edit this with the answers, and document my progress/pitfalls. Hopefully someday this will help someone down the line. The details: CentOS 5.5 x86_64 httpd: Apache/2.2.3 mysql: 5.0.77 (to be upgraded) php: 5.1 (to be upgraded) The requirements: SECURITY!! Secure file transfer Secure client access (SSL Certs and CA) Secure data storage Virtualhosts/multiple subdomains Local email would be nice, but not critical The Steps: Download latest CentOS DVD-iso (torrent worked great for me). Install CentOS: While going through the install, I checked the Server Components option thinking I was going to be using another Plesk-like admin. In hindsight, considering I've decided to try to go my own way, this probably wasn't the best idea. Basic config: Setup users, networking/ip address etc. Yum update/upgrade. Upgrade PHP/MySQL: To upgrade PHP and MySQL to the latest versions, I had to look to another repo outside CentOS. IUS looks great and I'm happy I found it! Add IUS repository to our package manager cd /tmp wget http://dl.iuscommunity.org/pub/ius/stable/Redhat/5/x86_64/epel-release-1-1.ius.el5.noarch.rpm rpm -Uvh epel-release-1-1.ius.el5.noarch.rpm wget http://dl.iuscommunity.org/pub/ius/stable/Redhat/5/x86_64/ius-release-1-4.ius.el5.noarch.rpm rpm -Uvh ius-release-1-4.ius.el5.noarch.rpm yum list | grep -w \.ius\. # list all the packages in the IUS repository; use this to find PHP/MySQL version and libraries you want to install Remove old version of PHP and install newer version from IUS rpm -qa | grep php # to list all of the installed php packages we want to remove yum shell # open an interactive yum shell remove php-common php-mysql php-cli #remove installed PHP components install php53 php53-mysql php53-cli php53-common #add packages you want transaction solve #important!! checks for dependencies transaction run #important!! does the actual installation of packages. [control+d] #exit yum shell php -v PHP 5.3.2 (cli) (built: Apr 6 2010 18:13:45) Upgrade MySQL from IUS repository /etc/init.d/mysqld stop rpm -qa | grep mysql # to see installed mysql packages yum shell remove mysql mysql-server #remove installed MySQL components install mysql51 mysql51-server mysql51-devel transaction solve #important!! checks for dependencies transaction run #important!! does the actual installation of packages. [control+d] #exit yum shell service mysqld start mysql -v Server version: 5.1.42-ius Distributed by The IUS Community Project Upgrade instructions courtesy of IUS wiki: http://wiki.iuscommunity.org/Doc/ClientUsageGuide Install rssh (restricted shell) to provide scp and sftp access, without allowing ssh login cd /tmp wget http://dag.wieers.com/rpm/packages/rssh/rssh-2.3.2-1.2.el5.rf.x86_64.rpm rpm -ivh rssh-2.3.2-1.2.el5.rf.x86_64.rpm useradd -m -d /home/dev -s /usr/bin/rssh dev passwd dev Edit /etc/rssh.conf to grant access to SFTP to rssh users. vi /etc/rssh.conf Uncomment or add: allowscp allowsftp This allows me to connect to the machine via SFTP protocol in Transmit (my FTP program of choice; I'm sure it's similar with other FTP apps). rssh instructions appropriated (with appreciation!) from http://www.cyberciti.biz/tips/linux-unix-restrict-shell-access-with-rssh.html Set up virtual interfaces ifconfig eth1:1 192.168.1.3 up #start up the virtual interface cd /etc/sysconfig/network-scripts/ cp ifcfg-eth1 ifcfg-eth1:1 #copy default script and match name to our virtual interface vi ifcfg-eth1:1 #modify eth1:1 script #ifcfg-eth1:1 | modify so it looks like this: DEVICE=eth1:1 IPADDR=192.168.1.3 NETMASK=255.255.255.0 NETWORK=192.168.1.0 ONBOOT=yes NAME=eth1:1 Add more Virtual interfaces as needed by repeating. Because of the ONBOOT=yes line in the ifcfg-eth1:1 file, this interface will be brought up when the system boots, or the network starts/restarts. service network restart Shutting down interface eth0: [ OK ] Shutting down interface eth1: [ OK ] Shutting down loopback interface: [ OK ] Bringing up loopback interface: [ OK ] Bringing up interface eth0: [ OK ] Bringing up interface eth1: [ OK ] ping 192.168.1.3 64 bytes from 192.168.1.3: icmp_seq=1 ttl=64 time=0.105 ms And this is where I'm at. I will keep editing this as I make progress. Any tips on how to Configure virtual interfaces/ip based virtual hosts for SSL, setting up a CA, or anything else would be appreciated.

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  • IPV6 causing issue with DNS

    - by Mike Wells
    I have set up an 'internal' DNS at my work, basically we have ourdomain.com that is for internet, email etc and I have created on one of our linux network servers (debian) a DNS using bind9 with the domain ourdomain.inc. So based on my files below and the symptoms I'm describing; what effect could IPV6 be having on my setup? What can I do to fix this? I assume it is not actually the IPV6 causing the issue, but rather something in my setup. These are the critical (I think) files I have modified: named.conf.local zone "ourdomain.inc" { type master; file "/etc/bind/zones/ourdomain.inc.db"; }; zone "201.168.192.in-addr.arpa" { type master; file "/etc/bind/zones/rev.201.168.192.in-addr.arpa"; }; named.conf.options options { directory "/var/cache/bind"; // If there is a firewall between you and nameservers you want // to talk to, you may need to fix the firewall to allow multiple // ports to talk. See http://www.kb.cert.org/vuls/id/800113 // If your ISP provided one or more IP addresses for stable // nameservers, you probably want to use them as forwarders. // Uncomment the following block, and insert the addresses replacing // the all-0's placeholder. forwarders { 1.2.3.4; //IP of our external DNS provider }; auth-nxdomain no; # conform to RFC1035 listen-on-v6 { any; }; }; ourdomain.inc.db $TTL 86400 ourdomain.inc. IN SOA ns1.ipower.com. admin.ourdomain.inc. ( 2006081401 28800 3600 604800 38400 ) serv1 IN A 192.168.201.223 serv2 IN A 192.168.201.220 serv3 IN A 192.168.201.219 ns1.ipower.com. IN A 1.2.3.4 ns2.ipower.com. IN A 1.2.3.5 @ IN NS ns1.ipower.com. @ IN NS ns2.ipower.com. svn IN CNAME serv1 docs IN CNAME serv2 jira IN CNAME serv3 confluence IN CNAME serv3 fisheye IN CNAME serv3 rev.201.168.192.in-addr.arpa $TTL 86400 201.168.192.in-addr.arpa. IN SOA ns1.ipower.com. admin.ourdomain.inc. ( 2006081401; 28800; 604800; 604800; 86400 ) 223 IN PTR serv1 @ IN NS ns1.ipower.com. @ IN NS ns2.ipower.com. named.conf include "/etc/bind/named.conf.options"; include "/etc/bind/named.conf.local"; include "/etc/bind/named.conf.default-zones"; I then made our internal DNS my preferred DNS with the two external DNSs the next in-line. More the most part this seems to work, I can ping svn.ourdomain.inc and it resolves to the correct IP, I can also ping google.com and it also resolves no problem. So all seem good. However, periodically (couple of times a day at least), I loose the ability to ping the svn.domain.inc (and all others defined under the internal DNS). What seem to fix the issue temporarily is to disable IPV6 on the network adapter of the client machine and then re-enable it. Then it works for a bit but will always fail again. System Info Internal DNS Distributor ID: Debian Description: Debian GNU/Linux 6.0.6 (squeeze) Release: 6.0.6 Codename: squeeze Linux 2.6.32-5-686 i686 BIND 9.7.3 PC OS Name: Microsoft Windows 7 Professional OS Version: 6.1.7601 Service Pack 1 Build 7601 System Type: x64-based PC Network Card(s): 2 NIC(s) Installed. [01]: Realtek PCIe GBE Family Controller Connection Name: WORK LAN DHCP Enabled: No IP address(es) [01]: the.ipv4.address [02]: the:ipv6:address The question... So based on my files above and the symptoms I described; what effect could IPV6 be having on my setup? What can I do to fix this? I assume it is not actually the IPV6 causing the issue, but rather something in my setup.

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  • Apache DS fails to list users

    - by CuriousMind
    Apache ds fails to list the users INFO | jvm 1 | 2012/03/28 15:54:04 | java.lang.Error: ERR_546 CRITICAL: page header magic for block 59 not OK 0 INFO | jvm 1 | 2012/03/28 15:54:04 | at jdbm.recman.PageHeader.(PageHeader.java:95) INFO | jvm 1 | 2012/03/28 15:54:04 | at jdbm.recman.PageHeader.getView(PageHeader.java:124) INFO | jvm 1 | 2012/03/28 15:54:04 | at jdbm.recman.PageManager.getNext(PageManager.java:234) INFO | jvm 1 | 2012/03/28 15:54:04 | at jdbm.recman.PageCursor.next(PageCursor.java:104) INFO | jvm 1 | 2012/03/28 15:54:04 | at jdbm.recman.PhysicalRowIdManager.fetch(PhysicalRowIdManager.java:158) INFO | jvm 1 | 2012/03/28 15:54:04 | at jdbm.recman.BaseRecordManager.fetch(BaseRecordManager.java:324) INFO | jvm 1 | 2012/03/28 15:54:04 | at jdbm.recman.CacheRecordManager.fetch(CacheRecordManager.java:262) INFO | jvm 1 | 2012/03/28 15:54:04 | at jdbm.btree.BPage.loadBPage(BPage.java:899) INFO | jvm 1 | 2012/03/28 15:54:04 | at jdbm.btree.BPage.childBPage(BPage.java:890) INFO | jvm 1 | 2012/03/28 15:54:04 | at jdbm.btree.BPage.find(BPage.java:284) INFO | jvm 1 | 2012/03/28 15:54:04 | at jdbm.btree.BPage.find(BPage.java:285) INFO | jvm 1 | 2012/03/28 15:54:04 | at jdbm.btree.BTree.find(BTree.java:408) INFO | jvm 1 | 2012/03/28 15:54:04 | at org.apache.directory.server.core.partition.impl.btree.jdbm.JdbmTable.get(JdbmTable.java:395) INFO | jvm 1 | 2012/03/28 15:54:04 | at org.apache.directory.server.core.partition.impl.btree.jdbm.JdbmMasterTable.get(JdbmMasterTable.java:155) INFO | jvm 1 | 2012/03/28 15:54:04 | at org.apache.directory.server.core.partition.impl.btree.jdbm.JdbmStore.lookup(JdbmStore.java:1332) INFO | jvm 1 | 2012/03/28 15:54:04 | at org.apache.directory.server.core.partition.impl.btree.jdbm.JdbmStore.lookup(JdbmStore.java:70) INFO | jvm 1 | 2012/03/28 15:54:04 | at org.apache.directory.server.xdbm.search.impl.EqualityEvaluator.evaluate(EqualityEvaluator.java:126) INFO | jvm 1 | 2012/03/28 15:54:04 | at org.apache.directory.server.xdbm.search.impl.AndCursor.matches(AndCursor.java:234) INFO | jvm 1 | 2012/03/28 15:54:04 | at org.apache.directory.server.xdbm.search.impl.AndCursor.next(AndCursor.java:143) INFO | jvm 1 | 2012/03/28 15:54:04 | at org.apache.directory.server.xdbm.search.impl.AndCursor.next(AndCursor.java:139) INFO | jvm 1 | 2012/03/28 15:54:04 | at org.apache.directory.server.core.partition.impl.btree.ServerEntryCursorAdaptor.next(ServerEntryCursorAdaptor.java:178) INFO | jvm 1 | 2012/03/28 15:54:04 | at org.apache.directory.server.core.filtering.BaseEntryFilteringCursor.next(BaseEntryFilteringCursor.java:499) INFO | jvm 1 | 2012/03/28 15:54:04 | at org.apache.directory.server.ldap.handlers.SearchHandler.readResults(SearchHandler.java:314) INFO | jvm 1 | 2012/03/28 15:54:04 | at org.apache.directory.server.ldap.handlers.SearchHandler.doSimpleSearch(SearchHandler.java:749) INFO | jvm 1 | 2012/03/28 15:54:04 | at org.apache.directory.server.ldap.handlers.SearchHandler.handleIgnoringReferrals(SearchHandler.java:978) INFO | jvm 1 | 2012/03/28 15:54:04 | at org.apache.directory.server.ldap.handlers.SearchHandler.handleIgnoringReferrals(SearchHandler.java:78) INFO | jvm 1 | 2012/03/28 15:54:04 | at org.apache.directory.server.ldap.handlers.ReferralAwareRequestHandler.handle(ReferralAwareRequestHandler.java:83) INFO | jvm 1 | 2012/03/28 15:54:04 | at org.apache.directory.server.ldap.handlers.ReferralAwareRequestHandler.handle(ReferralAwareRequestHandler.java:57) INFO | jvm 1 | 2012/03/28 15:54:04 | at org.apache.directory.server.ldap.handlers.LdapRequestHandler.handleMessage(LdapRequestHandler.java:208) INFO | jvm 1 | 2012/03/28 15:54:04 | at org.apache.directory.server.ldap.handlers.LdapRequestHandler.handleMessage(LdapRequestHandler.java:58) INFO | jvm 1 | 2012/03/28 15:54:04 | at org.apache.mina.handler.demux.DemuxingIoHandler.messageReceived(DemuxingIoHandler.java:232) INFO | jvm 1 | 2012/03/28 15:54:04 | at org.apache.directory.server.ldap.LdapProtocolHandler.messageReceived(LdapProtocolHandler.java:193) INFO | jvm 1 | 2012/03/28 15:54:04 | at org.apache.mina.core.filterchain.DefaultIoFilterChain$TailFilter.messageReceived(DefaultIoFilterChain.java:713) INFO | jvm 1 | 2012/03/28 15:54:04 | at org.apache.mina.core.filterchain.DefaultIoFilterChain.callNextMessageReceived(DefaultIoFilterChain.java:434) INFO | jvm 1 | 2012/03/28 15:54:04 | at org.apache.mina.core.filterchain.DefaultIoFilterChain.access$1200(DefaultIoFilterChain.java:46) INFO | jvm 1 | 2012/03/28 15:54:04 | at org.apache.mina.core.filterchain.DefaultIoFilterChain$EntryImpl$1.messageReceived(DefaultIoFilterChain.java:793) INFO | jvm 1 | 2012/03/28 15:54:04 | at org.apache.mina.core.filterchain.IoFilterEvent.fire(IoFilterEvent.java:71) INFO | jvm 1 | 2012/03/28 15:54:04 | at org.apache.mina.core.session.IoEvent.run(IoEvent.java:63) INFO | jvm 1 | 2012/03/28 15:54:04 | at org.apache.mina.filter.executor.UnorderedThreadPoolExecutor$Worker.runTask(UnorderedThreadPoolExecutor.java:480) INFO | jvm 1 | 2012/03/28 15:54:04 | at org.apache.mina.filter.executor.UnorderedThreadPoolExecutor$Worker.run(UnorderedThreadPoolExecutor.java:434) INFO | jvm 1 | 2012/03/28 15:54:04 | at java.lang.Thread.run(Thread.java:619) INFO | jvm 1 | 2012/03/28 15:54:04 | [15:54:04] WARN [org.apache.directory.server.ldap.LdapProtocolHandler] - Null LdapSession given to cleanUpSession. INFO | jvm 1 | 2012/03/28 15:55:20 | [15:55:20] WARN [org.apache.directory.server.ldap.LdapProtocolHandler] - Unexpected exception forcing session to close: sending disconnect notice to client.

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