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  • elffile: ELF Specific File Identification Utility

    - by user9154181
    Solaris 11 has a new standard user level command, /usr/bin/elffile. elffile is a variant of the file utility that is focused exclusively on linker related files: ELF objects, archives, and runtime linker configuration files. All other files are simply identified as "non-ELF". The primary advantage of elffile over the existing file utility is in the area of archives — elffile examines the archive members and can produce a summary of the contents, or per-member details. The impetus to add elffile to Solaris came from the effort to extend the format of Solaris archives so that they could grow beyond their previous 32-bit file limits. That work introduced a new archive symbol table format. Now that there was more than one possible format, I thought it would be useful if the file utility could identify which format a given archive is using, leading me to extend the file utility: % cc -c ~/hello.c % ar r foo.a hello.o % file foo.a foo.a: current ar archive, 32-bit symbol table % ar r -S foo.a hello.o % file foo.a foo.a: current ar archive, 64-bit symbol table In turn, this caused me to think about all the things that I would like the file utility to be able to tell me about an archive. In particular, I'd like to be able to know what's inside without having to unpack it. The end result of that train of thought was elffile. Much of the discussion in this article is adapted from the PSARC case I filed for elffile in December 2010: PSARC 2010/432 elffile Why file Is No Good For Archives And Yet Should Not Be Fixed The standard /usr/bin/file utility is not very useful when applied to archives. When identifying an archive, a user typically wants to know 2 things: Is this an archive? Presupposing that the archive contains objects, which is by far the most common use for archives, what platform are the objects for? Are they for sparc or x86? 32 or 64-bit? Some confusing combination from varying platforms? The file utility provides a quick answer to question (1), as it identifies all archives as "current ar archive". It does nothing to answer the more interesting question (2). To answer that question, requires a multi-step process: Extract all archive members Use the file utility on the extracted files, examine the output for each file in turn, and compare the results to generate a suitable summary description. Remove the extracted files It should be easier and more efficient to answer such an obvious question. It would be reasonable to extend the file utility to examine archive contents in place and produce a description. However, there are several reasons why I decided not to do so: The correct design for this feature within the file utility would have file examine each archive member in turn, applying its full abilities to each member. This would be elegant, but also represents a rather dramatic redesign and re-implementation of file. Archives nearly always contain nothing but ELF objects for a single platform, so such generality in the file utility would be of little practical benefit. It is best to avoid adding new options to standard utilities for which other implementations of interest exist. In the case of the file utility, one concern is that we might add an option which later appears in the GNU version of file with a different and incompatible meaning. Indeed, there have been discussions about replacing the Solaris file with the GNU version in the past. This may or may not be desirable, and may or may not ever happen. Either way, I don't want to preclude it. Examining archive members is an O(n) operation, and can be relatively slow with large archives. The file utility is supposed to be a very fast operation. I decided that extending file in this way is overkill, and that an investment in the file utility for better archive support would not be worth the cost. A solution that is more narrowly focused on ELF and other linker related files is really all that we need. The necessary code for doing this already exists within libelf. All that is missing is a small user-level wrapper to make that functionality available at the command line. In that vein, I considered adding an option for this to the elfdump utility. I examined elfdump carefully, and even wrote a prototype implementation. The added code is small and simple, but the conceptual fit with the rest of elfdump is poor. The result complicates elfdump syntax and documentation, definite signs that this functionality does not belong there. And so, I added this functionality as a new user level command. The elffile Command The syntax for this new command is elffile [-s basic | detail | summary] filename... Please see the elffile(1) manpage for additional details. To demonstrate how output from elffile looks, I will use the following files: FileDescription configA runtime linker configuration file produced with crle dwarf.oAn ELF object /etc/passwdA text file mixed.aArchive containing a mixture of ELF and non-ELF members mixed_elf.aArchive containing ELF objects for different machines not_elf.aArchive containing no ELF objects same_elf.aArchive containing a collection of ELF objects for the same machine. This is the most common type of archive. The file utility identifies these files as follows: % file config dwarf.o /etc/passwd mixed.a mixed_elf.a not_elf.a same_elf.a config: Runtime Linking Configuration 64-bit MSB SPARCV9 dwarf.o: ELF 64-bit LSB relocatable AMD64 Version 1 /etc/passwd: ascii text mixed.a: current ar archive, 32-bit symbol table mixed_elf.a: current ar archive, 32-bit symbol table not_elf.a: current ar archive same_elf.a: current ar archive, 32-bit symbol table By default, elffile uses its "summary" output style. This output differs from the output from the file utility in 2 significant ways: Files that are not an ELF object, archive, or runtime linker configuration file are identified as "non-ELF", whereas the file utility attempts further identification for such files. When applied to an archive, the elffile output includes a description of the archive's contents, without requiring member extraction or other additional steps. Applying elffile to the above files: % elffile config dwarf.o /etc/passwd mixed.a mixed_elf.a not_elf.a same_elf.a config: Runtime Linking Configuration 64-bit MSB SPARCV9 dwarf.o: ELF 64-bit LSB relocatable AMD64 Version 1 /etc/passwd: non-ELF mixed.a: current ar archive, 32-bit symbol table, mixed ELF and non-ELF content mixed_elf.a: current ar archive, 32-bit symbol table, mixed ELF content not_elf.a: current ar archive, non-ELF content same_elf.a: current ar archive, 32-bit symbol table, ELF 64-bit LSB relocatable AMD64 Version 1 The output for same_elf.a is of particular interest: The vast majority of archives contain only ELF objects for a single platform, and in this case, the default output from elffile answers both of the questions about archives posed at the beginning of this discussion, in a single efficient step. This makes elffile considerably more useful than file, within the realm of linker-related files. elffile can produce output in two other styles, "basic", and "detail". The basic style produces output that is the same as that from 'file', for linker-related files. The detail style produces per-member identification of archive contents. This can be useful when the archive contents are not homogeneous ELF object, and more information is desired than the summary output provides: % elffile -s detail mixed.a mixed.a: current ar archive, 32-bit symbol table mixed.a(dwarf.o): ELF 32-bit LSB relocatable 80386 Version 1 mixed.a(main.c): non-ELF content mixed.a(main.o): ELF 64-bit LSB relocatable AMD64 Version 1 [SSE]

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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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  • DBA Best Practices - A Blog Series: Episode 1 - Backups

    - by Argenis
      This blog post is part of the DBA Best Practices series, on which various topics of concern for daily database operations are discussed. Your feedback and comments are very much welcome, so please drop by the comments section and be sure to leave your thoughts on the subject. Morning Coffee When I was a DBA, the first thing I did when I sat down at my desk at work was checking that all backups had completed successfully. It really was more of a ritual, since I had a dual system in place to check for backup completion: 1) the scheduled agent jobs to back up the databases were set to alert the NOC in failure, and 2) I had a script run from a central server every so often to check for any backup failures. Why the redundancy, you might ask. Well, for one I was once bitten by the fact that database mail doesn't work 100% of the time. Potential causes for failure include issues on the SMTP box that relays your server email, firewall problems, DNS issues, etc. And so to be sure that my backups completed fine, I needed to rely on a mechanism other than having the servers do the taking - I needed to interrogate the servers and ask each one if an issue had occurred. This is why I had a script run every so often. Some of you might have monitoring tools in place like Microsoft System Center Operations Manager (SCOM) or similar 3rd party products that would track all these things for you. But at that moment, we had no resort but to write our own Powershell scripts to do it. Now it goes without saying that if you don't have backups in place, you might as well find another career. Your most sacred job as a DBA is to protect the data from a disaster, and only properly safeguarded backups can offer you peace of mind here. "But, we have a cluster...we don't need backups" Sadly I've heard this line more than I would have liked to. You need to understand that a cluster is comprised of shared storage, and that is precisely your single point of failure. A cluster will protect you from an issue at the Operating System level, and also under an outage of any SQL-related service or dependent devices. But it will most definitely NOT protect you against corruption, nor will it protect you against somebody deleting data from a table - accidentally or otherwise. Backup, fine. How often do I take a backup? The answer to this is something you will hear frequently when working with databases: it depends. What does it depend on? For one, you need to understand how much data your business is willing to lose. This is what's called Recovery Point Objective, or RPO. If you don't know how much data your business is willing to lose, you need to have an honest and realistic conversation about data loss expectations with your customers, internal or external. From my experience, their first answer to the question "how much data loss can you withstand?" will be "zero". In that case, you will need to explain how zero data loss is very difficult and very costly to achieve, even in today's computing environments. Do you want to go ahead and take full backups of all your databases every hour, or even every day? Probably not, because of the impact that taking a full backup can have on a system. That's what differential and transaction log backups are for. Have I answered the question of how often to take a backup? No, and I did that on purpose. You need to think about how much time you have to recover from any event that requires you to restore your databases. This is what's called Recovery Time Objective. Again, if you go ask your customer how long of an outage they can withstand, at first you will get a completely unrealistic number - and that will be your starting point for discussing a solution that is cost effective. The point that I'm trying to get across is that you need to have a plan. This plan needs to be practiced, and tested. Like a football playbook, you need to rehearse the moves you'll perform when the time comes. How often is up to you, and the objective is that you feel better about yourself and the steps you need to follow when emergency strikes. A backup is nothing more than an untested restore Backups are files. Files are prone to corruption. Put those two together and realize how you feel about those backups sitting on that network drive. When was the last time you restored any of those? Restoring your backups on another box - that, by the way, doesn't have to match the specs of your production server - will give you two things: 1) peace of mind, because now you know that your backups are good and 2) a place to offload your consistency checks with DBCC CHECKDB or any of the other DBCC commands like CHECKTABLE or CHECKCATALOG. This is a great strategy for VLDBs that cannot withstand the additional load created by the consistency checks. If you choose to offload your consistency checks to another server though, be sure to run DBCC CHECKDB WITH PHYSICALONLY on the production server, and if you're using SQL Server 2008 R2 SP1 CU4 and above, be sure to enable traceflags 2562 and/or 2549, which will speed up the PHYSICALONLY checks further - you can read more about this enhancement here. Back to the "How Often" question for a second. If you have the disk, and the network latency, and the system resources to do so, why not backup the transaction log often? As in, every 5 minutes, or even less than that? There's not much downside to doing it, as you will have to clear the log with a backup sooner than later, lest you risk running out space on your tlog, or even your drive. The one drawback to this approach is that you will have more files to deal with at restore time, and processing each file will add a bit of extra time to the entire process. But it might be worth that time knowing that you minimized the amount of data lost. Again, test your plan to make sure that it matches your particular needs. Where to back up to? Network share? Locally? SAN volume? This is another topic where everybody has a favorite choice. So, I'll stick to mentioning what I like to do and what I consider to be the best practice in this regard. I like to backup to a SAN volume, i.e., a drive that actually lives in the SAN, and can be easily attached to another server in a pinch, saving you valuable time - you wouldn't need to restore files on the network (slow) or pull out drives out a dead server (been there, done that, it’s also slow!). The key is to have a copy of those backup files made quickly, and, if at all possible, to a remote target on a different datacenter - or even the cloud. There are plenty of solutions out there that can help you put such a solution together. That right there is the first step towards a practical Disaster Recovery plan. But there's much more to DR, and that's material for a different blog post in this series.

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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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  • I Know What I Did This Summer: Put Down Trex Decking

    - by thatjeffsmith
    If you’re wondering why I would bore everyone with my pictures and frequent status updates/tweets from the past week – it’s so I could document the process of refurbishing my deck, or what some would call a porch. When we go to take a vacation, buy a car, do anything – we also read personal blogs to get the real story. So, if you’re curious about what it takes to tackle this sort of project, read on. Skills/Equipment/Manpower We Possessed I took the old decking out by myself. I’m about 230 lbs, more than 6′ tall, and I’m pretty healthy. This took about 8 hours over two afternoons. Three of us put the deck back together. My wife has two engineering degrees. Her father also has two engineering degrees. Lots of brainpower available here. Also, her dad ran the public works department for a country for more than 20 years – so lots and lots of practical experience on hand. We had a compound mitre saw, a skilsaw, 2-3 crowbars, a framing hammer, 3 cordless drills, a corded drill, lots of sawhorses, a power sander, an angle grinder, a 10×10 Coleman canopy tent, a Ford F-150 pickup truck, outdoor speakers and lots of iTunes playlists, plenty of water and cold beer. Why We Did This Our deck was relatively young – it was built in 2005. However, the pressure treated boards must not have been adequately maintained before we bought the house. I had powerwashed the deck every other year and had it stained a few times. The boards just rotted. We’re going to be in the house for a long time, and we wanted something that would look nice and require little maintenance. More bad deck boards The deck boards were in bad shape Things We Learned The two most important things: The hidden fasteners have to be put in JUST right. Wedge them into the grooved board, then bend down the bit that is screwed down. We didn’t do this on the first board and couldn’t get the second board to fit nearly close enough. Watching the official TREX YouTube video helped immensely, and we should have watched that first. When pre-drilling holes for the boards that need screwed down – DO NOT pre-drill through the underlying framing wood. ONLY pre-drill through the TREX itself. The screw won’t seat in the board properly. Instead of sitting down flush with the board, it will stop at the top of the board and just spin. I had to call the the place that sold me the screws to find this out. So about a third of our screws look like crap. If it doesn’t look or feel right – stop everything and pick up your computer or your phone. It’s not right, and it will be much easier to stop and find out why. We didn’t do this, and now I’m going to see every screw that’s not flush with the boards and get upset. Oh well. The Process How much time did it take? Well I spent about 8 hours taking the deck apart. And then the 3 of use spent 8 hours the first day, 10 hours the second day, 8 hours the third, and another 6 hours on the fourth day. That’s like 104 man-hours. We supposedly saved four or five thousand dollars in labor, but don’t do the math here or you might get a bit upset. The main thing is that we got what we wanted, and there won’t be any surprises later. Now for some pictures… This 6”+ pry bar made the destruction of the old deck much easier Most of the joists, once exposed, were OK. This joist wasn’t sitting on ANYTHING before. We think a lazy gas person cut the board to sneak a gas line in. Awesome… These monster lag bolts had to be accounted for when putting in the additional framing The border pattern Sheri wanted to put in required a lot more framing. These were the first boards to go down – we screwed them in as there was no way to attach clips I sat, kicked in the boards, and then drilled these clips in – but my wife was able to go MUCH faster by using her hands to lock the boards in and drill on her knees. I liked locking the board in with my feet when they needed to be ‘encouraged’ to go straight. The first board took FOREVER to go in, but then when we got rolling, we were able to put in a 20′ board in less than 10 minutes. This was end of construction day #2 – we got much further than we thought we would. Ah, the dreaded last 10% – what to do here? Remember those ‘floating’ stringers? Yeah, we fixed that up a bit, too. My wife used a website (and her brain) to calculate exactly how to cut the stringers to give us the rise/run we needed with the proper clearance and all that jazz. The stairs with stringers and toe kicks – this was worth the effort It started raining on us as I screwed down the steps – this we managed to get our shade tent up on the deck to protect us from the rain too The stairs, finished Finished, mostly Good corner shot The top of the stairs Stairs, looking down Celebratory beer In Summary There are a few things we’re not happy with. I think we can fix them up – but later. I have a few things left to finish, rewire the lighting, get the gas grille put back in, and rehang some screen doors. I was expecting this to be a lot worse than it was. If I didn’t have the help, I would have never done it myself. But I’m glad that I did have that help and did do that project. It’s not often you get to spend that kind of qualify time with family and building cool stuff.

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  • DBA Best Practices - A Blog Series: Episode 1 - Backups

    - by Argenis
      This blog post is part of the DBA Best Practices series, on which various topics of concern for daily database operations are discussed. Your feedback and comments are very much welcome, so please drop by the comments section and be sure to leave your thoughts on the subject. Morning Coffee When I was a DBA, the first thing I did when I sat down at my desk at work was checking that all backups have completed successfully. It really was more of a ritual, since I had a dual system in place to check for backup completion: 1) the scheduled agent jobs to back up the databases were set to alert the NOC in failure, and 2) I had a script run from a central server every so often to check for any backup failures. Why the redundancy, you might ask. Well, for one I was once bitten by the fact that database mail doesn't work 100% of the time. Potential causes for failure include issues on the SMTP box that relays your server email, firewall problems, DNS issues, etc. And so to be sure that my backups completed fine, I needed to rely on a mechanism other than having the servers do the taking - I needed to interrogate the servers and ask each one if an issue had occurred. This is why I had a script run every so often. Some of you might have monitoring tools in place like Microsoft System Center Operations Manager (SCOM) or similar 3rd party products that would track all these things for you. But at that moment, we had no resort but to write our own Powershell scripts to do it. Now it goes without saying that if you don't have backups in place, you might as well find another career. Your most sacred job as a DBA is to protect the data from a disaster, and only properly safeguarded backups can offer you peace of mind here. "But, we have a cluster...we don't need backups" Sadly I've heard this line more than I would have liked to. You need to understand that a cluster is comprised of shared storage, and that is precisely your single point of failure. A cluster will protect you from an issue at the Operating System level, and also under an outage of any SQL-related service or dependent devices. But it will most definitely NOT protect you against corruption, nor will it protect you against somebody deleting data from a table - accidentally or otherwise. Backup, fine. How often do I take a backup? The answer to this is something you will hear frequently when working with databases: it depends. What does it depend on? For one, you need to understand how much data your business is willing to lose. This is what's called Recovery Point Objective, or RPO. If you don't know how much data your business is willing to lose, you need to have an honest and realistic conversation about data loss expectations with your customers, internal or external. From my experience, their first answer to the question "how much data loss can you withstand?" will be "zero". In that case, you will need to explain how zero data loss is very difficult and very costly to achieve, even in today's computing environments. Do you want to go ahead and take full backups of all your databases every hour, or even every day? Probably not, because of the impact that taking a full backup can have on a system. That's what differential and transaction log backups are for. Have I answered the question of how often to take a backup? No, and I did that on purpose. You need to think about how much time you have to recover from any event that requires you to restore your databases. This is what's called Recovery Time Objective. Again, if you go ask your customer how long of an outage they can withstand, at first you will get a completely unrealistic number - and that will be your starting point for discussing a solution that is cost effective. The point that I'm trying to get across is that you need to have a plan. This plan needs to be practiced, and tested. Like a football playbook, you need to rehearse the moves you'll perform when the time comes. How often is up to you, and the objective is that you feel better about yourself and the steps you need to follow when emergency strikes. A backup is nothing more than an untested restore Backups are files. Files are prone to corruption. Put those two together and realize how you feel about those backups sitting on that network drive. When was the last time you restored any of those? Restoring your backups on another box - that, by the way, doesn't have to match the specs of your production server - will give you two things: 1) peace of mind, because now you know that your backups are good and 2) a place to offload your consistency checks with DBCC CHECKDB or any of the other DBCC commands like CHECKTABLE or CHECKCATALOG. This is a great strategy for VLDBs that cannot withstand the additional load created by the consistency checks. If you choose to offload your consistency checks to another server though, be sure to run DBCC CHECKDB WITH PHYSICALONLY on the production server, and if you're using SQL Server 2008 R2 SP1 CU4 and above, be sure to enable traceflags 2562 and/or 2549, which will speed up the PHYSICALONLY checks further - you can read more about this enhancement here. Back to the "How Often" question for a second. If you have the disk, and the network latency, and the system resources to do so, why not backup the transaction log often? As in, every 5 minutes, or even less than that? There's not much downside to doing it, as you will have to clear the log with a backup sooner than later, lest you risk running out space on your tlog, or even your drive. The one drawback to this approach is that you will have more files to deal with at restore time, and processing each file will add a bit of extra time to the entire process. But it might be worth that time knowing that you minimized the amount of data lost. Again, test your plan to make sure that it matches your particular needs. Where to back up to? Network share? Locally? SAN volume? This is another topic where everybody has a favorite choice. So, I'll stick to mentioning what I like to do and what I consider to be the best practice in this regard. I like to backup to a SAN volume, i.e., a drive that actually lives in the SAN, and can be easily attached to another server in a pinch, saving you valuable time - you wouldn't need to restore files on the network (slow) or pull out drives out a dead server (been there, done that, it’s also slow!). The key is to have a copy of those backup files made quickly, and, if at all possible, to a remote target on a different datacenter - or even the cloud. There are plenty of solutions out there that can help you put such a solution together. That right there is the first step towards a practical Disaster Recovery plan. But there's much more to DR, and that's material for a different blog post in this series.

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  • Why JSF Matters (to You)

    - by reza_rahman
          "Those who have knowledge, don’t predict. Those who predict, don’t have knowledge."                                                                                                    – Lao Tzu You may have noticed Thoughtworks recently crowned the likes AngularJS, etc imminent successors to server-side web frameworks. They apparently also deemed it necessary to single out JSF for righteous scorn. I have to say as I was reading the analysis I couldn't help but remember they also promptly jumped on the Ruby, Rails, Clojure, etc bandwagon a good few years ago seemingly similarly crowing these dynamic languages imminent successors to Java. I remember thinking then as I do now whether the folks at Thoughtworks are really that much smarter than me or if they are simply more prone to the Hipster buzz of the day. I'll let you make the final call on that one. I also noticed mention of "J2EE" in the context of JSF and had to wonder how up-to-date or knowledgeable the person writing the analysis actually was given that the term was basically retired almost a decade ago. There's one thing that I am absolutely sure about though - as a long time pretty happy user of JSF, I had no choice but to speak up on what I believe JSF offers. If you feel the same way, I would encourage you to support the team behind JSF whose hard work you may have benefited from over the years. True to his outspoken character PrimeFaces lead Cagatay Civici certainly did not mince words making the case for the JSF ecosystem - his excellent write-up is well worth a read. He specifically pointed out the practical problems in going whole hog with bare metal JavaScript, CSS, HTML for many development teams. I'll admit I had to smile when I read his closing sentence as well as the rather cheerful comments to the post from actual current JSF/PrimeFaces users that are apparently supposed to be on a gloomy death march. In a similar vein, OmniFaces developer Arjan Tijms did a great job pointing out the fact that despite the extremely competitive server-side Java Web UI space, JSF seems to manage to always consistently come out in either the number one or number two spot over many years and many data sources - do give his well-written message in the JAX-RS user forum a careful read. I don't think it's really reasonable to expect this to be the case for so many years if JSF was not at least a capable if not outstanding technology. If fact if you've ever wondered, Oracle itself is one of the largest JSF users on the planet. As Oracle's Shay Shmeltzer explains in a recent JSF Central interview, many of Oracle's strategic products such as ADF, ADF Mobile and Fusion Applications itself is built on JSF. There are well over 3,000 active developers working on these codebases. I don't think anyone can think of a more compelling reason to make sure that a technology is as effective as possible for practical development under real world conditions. Standing on the shoulders of the above giants, I feel like I can be pretty brief in making my own case for JSF: JSF is a powerful abstraction that brings the original Smalltalk MVC pattern to web development. This means cutting down boilerplate code to the bare minimum such that you really can think of just writing your view markup and then simply wire up some properties and event handlers on a POJO. The best way to see what this really means is to compare JSF code for a pretty small case to other approaches. You should then multiply the additional work for the typical enterprise project to try to understand what the productivity trade-offs are. This is reason alone for me to personally never take any other approach seriously as my primary web UI solution unless it can match the sheer productivity of JSF. Thanks to JSF's focus on components from the ground-up JSF has an extremely strong ecosystem that includes projects like PrimeFaces, RichFaces, OmniFaces, ICEFaces and of course ADF Faces/Mobile. These component libraries taken together constitute perhaps the largest widget set ever developed and optimized for a single web UI technology. To begin to grasp what this really means, just briefly browse the excellent PrimeFaces showcase and think about the fact that you can readily use the widgets on that showcase by just using some simple markup and knowing near to nothing about AJAX, JavaScript or CSS. JSF has the fair and legitimate advantage of being an open vendor neutral standard. This means that no single company, individual or insular clique controls JSF - openness, transparency, accountability, plurality, collaboration and inclusiveness is virtually guaranteed by the standards process itself. You have the option to choose between compatible implementations, escape any form of lock-in or even create your own compatible implementation! As you might gather from the quote at the top of the post, I am not a fan of crystal ball gazing and certainly don't want to engage in it myself. Who knows? However far-fetched it may seem maybe AngularJS is the only future we all have after all. If that is the case, so be it. Unlike what you might have been told, Java EE is about choice at heart and it can certainly work extremely well as a back-end for AngularJS. Likewise, you are also most certainly not limited to just JSF for working with Java EE - you have a rich set of choices like Struts 2, Vaadin, Errai, VRaptor 4, Wicket or perhaps even the new action-oriented web framework being considered for Java EE 8 based on the work in Jersey MVC... Please note that any views expressed here are my own only and certainly does not reflect the position of Oracle as a company.

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  • Java EE 7 Survey Results!

    - by reza_rahman
    On November 8th, the Java EE EG posted a survey to gather broad community feedback on a number of critical open issues. For reference, you can find the original survey here. We kept the survey open for about three weeks until November 30th. To our delight, over 1100 developers took time out of their busy lives to let their voices be heard! The results of the survey were sent to the EG on December 12th. The subsequent EG discussion is available here. The exact summary sent to the EG is available here. We would like to take this opportunity to thank each and every one the individuals who took the survey. It is very appreciated, encouraging and worth it's weight in gold. In particular, I tried to capture just some of the high-quality, intelligent, thoughtful and professional comments in the summary to the EG. I highly encourage you to continue to stay involved, perhaps through the Adopt-a-JSR program. We would also like to sincerely thank java.net, JavaLobby, TSS and InfoQ for helping spread the word about the survey. Below is a brief summary of the results... APIs to Add to Java EE 7 Full/Web Profile The first question asked which of the four new candidate APIs (WebSocket, JSON-P, JBatch and JCache) should be added to the Java EE 7 Full and Web profile respectively. As the following graph shows, there was significant support for adding all the new APIs to the full profile: Support is relatively the weakest for Batch 1.0, but still good. A lot of folks saw WebSocket 1.0 as a critical technology with comments such as this one: "A modern web application needs Web Sockets as first class citizens" While it is clearly seen as being important, a number of commenters expressed dissatisfaction with the lack of a higher-level JSON data binding API as illustrated by this comment: "How come we don't have a Data Binding API for JSON" JCache was also seen as being very important as expressed with comments like: "JCache should really be that foundational technology on which other specs have no fear to depend on" The results for the Web Profile is not surprising. While there is strong support for adding WebSocket 1.0 and JSON-P 1.0 to the Web Profile, support for adding JCache 1.0 and Batch 1.0 is relatively weak. There was actually significant opposition to adding Batch 1. 0 (with 51.8% casting a 'No' vote). Enabling CDI by Default The second question asked was whether CDI should be enabled in Java EE environments by default. A significant majority of 73.3% developers supported enabling CDI, only 13.8% opposed. Comments such as these two reflect a strong general support for CDI as well as a desire for better Java EE alignment with CDI: "CDI makes Java EE quite valuable!" "Would prefer to unify EJB, CDI and JSF lifecycles" There is, however, a palpable concern around the performance impact of enabling CDI by default as exemplified by this comment: "Java EE projects in most cases use CDI, hence it is sensible to enable CDI by default when creating a Java EE application. However, there are several issues if CDI is enabled by default: scanning can be slow - not all libs use CDI (hence, scanning is not needed)" Another significant concern appears to be around backwards compatibility and conflict with other JSR 330 implementations like Spring: "I am leaning towards yes, however can easily imagine situations where errors would be caused by automatically activating CDI, especially in cases of backward compatibility where another DI engine (such as Spring and the like) happens to use the same mechanics to inject dependencies and in that case there would be an overlap in injections and probably an uncertain outcome" Some commenters such as this one attempt to suggest solutions to these potential issues: "If you have Spring in use and use javax.inject.Inject then you might get some unexpected behavior that could be equally confusing. I guess there will be a way to switch CDI off. I'm tempted to say yes but am cautious for this reason" Consistent Usage of @Inject The third question was around using CDI/JSR 330 @Inject consistently vs. allowing JSRs to create their own injection annotations. A slight majority of 53.3% developers supported using @Inject consistently across JSRs. 28.8% said using custom injection annotations is OK, while 18.0% were not sure. The vast majority of commenters were strongly supportive of CDI and general Java EE alignment with CDI as illistrated by these comments: "Dependency Injection should be standard from now on in EE. It should use CDI as that is the DI mechanism in EE and is quite powerful. Having a new JSR specific DI mechanism to deal with just means more reflection, more proxies. JSRs should also be constructed to allow some of their objects Injectable. @Inject @TransactionalCache or @Inject @JMXBean etc...they should define the annotations and stereotypes to make their code less procedural. Dog food it. If there is a shortcoming in CDI for a JSR fix it and we will all be grateful" "We're trying to make this a comprehensive platform, right? Injection should be a fundamental part of the platform; everything else should build on the same common infrastructure. Each-having-their-own is just a recipe for chaos and having to learn the same thing 10 different ways" Expanding the Use of @Stereotype The fourth question was about expanding CDI @Stereotype to cover annotations across Java EE beyond just CDI. A significant majority of 62.3% developers supported expanding the use of @Stereotype, only 13.3% opposed. A majority of commenters supported the idea as well as the theme of general CDI/Java EE alignment as expressed in these examples: "Just like defining new types for (compositions of) existing classes, stereotypes can help make software development easier" "This is especially important if many EJB services are decoupled from the EJB component model and can be applied via individual annotations to Java EE components. @Stateless is a nicely compact annotation. Code will not improve if that will have to be applied in the future as @Transactional, @Pooled, @Secured, @Singlethreaded, @...." Some, however, expressed concerns around increased complexity such as this commenter: "Could be very convenient, but I'm afraid if it wouldn't make some important class annotations less visible" Expanding Interceptor Use The final set of questions was about expanding interceptors further across Java EE... A very solid 96.3% of developers wanted to expand interceptor use to all Java EE components. 35.7% even wanted to expand interceptors to other Java EE managed classes. Most developers (54.9%) were not sure if there is any place that injection is supported that should not support interceptors. 32.8% thought any place that supports injection should also support interceptors. Only 12.2% were certain that there are places where injection should be supported but not interceptors. The comments reflected the diversity of opinions, generally supportive of interceptors: "I think interceptors are as fundamental as injection and should be available anywhere in the platform" "The whole usage of interceptors still needs to take hold in Java programming, but it is a powerful technology that needs some time in the Sun. Basically it should become part of Java SE, maybe the next step after lambas?" A distinct chain of thought separated interceptors from filters and listeners: "I think that the Servlet API already provides a rich set of possibilities to hook yourself into different Servlet container events. I don't find a need to 'pollute' the Servlet model with the Interceptors API"

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  • Using Unity – Part 5

    - by nmarun
    In the previous article of the series, I talked about constructor and property (setter) injection. I wanted to write about how to work with arrays and generics in Unity in this blog, after seeing how lengthy this one got, I’ve decided to write about generics in the next one. This one will only concentrate on arrays. My Product4 class has the following definition: 1: public interface IProduct 2: { 3: string WriteProductDetails(); 4: } 5:  6: public class Product4 : IProduct 7: { 8: public string Name { get; set; } 9: public ILogger[] Loggers { get; set; } 10:  11: public Product4(string productName, ILogger[] loggers) 12: { 13: Name = productName; 14: Loggers = loggers; 15: } 16:  17: public string WriteProductDetails() 18: { 19: StringBuilder productDetails = new StringBuilder(); 20: productDetails.AppendFormat("{0}<br/>", Name); 21: for (int i = 0; i < Loggers.Count(); i++) 22: { 23: productDetails.AppendFormat("{0}<br/>", Loggers[i].WriteLog()); 24: } 25: 26: return productDetails.ToString(); 27: } 28: } The key parts are line 4 where we declare an array of ILogger and line 5 where-in the constructor passes an instance of an array of ILogger objects. I’ve created another class – FakeLogger: 1: public class FakeLogger : ILogger 2: { 3: public string WriteLog() 4: { 5: return string.Format("Type: {0}", GetType()); 6: } 7: } It’s implementation is the same as what we had for the FileLogger class. Coming to the web.config file, first add the following aliases. The alias for FakeLogger should make sense right away. ILoggerArray defines an array of ILogger objects. I’ll tell why we need an alias for System.String data type. 1: <typeAlias alias="string" type="System.String, mscorlib" /> 2: <typeAlias alias="ILoggerArray" type="ProductModel.ILogger[], ProductModel" /> 3: <typeAlias alias="FakeLogger" type="ProductModel.FakeLogger, ProductModel"/> Next is to create mappings for the FileLogger and FakeLogger classes: 1: <type type="ILogger" mapTo="FileLogger" name="logger1"> 2: <lifetime type="singleton" /> 3: </type> 4: <type type="ILogger" mapTo="FakeLogger" name="logger2"> 5: <lifetime type="singleton" /> 6: </type> Finally, for the real deal: 1: <type type="IProduct" mapTo="Product4" name="ArrayProduct"> 2: <typeConfig extensionType="Microsoft.Practices.Unity.Configuration.TypeInjectionElement,Microsoft.Practices.Unity.Configuration, Version=1.2.0.0, Culture=neutral, PublicKeyToken=31bf3856ad364e35"> 3: <constructor> 4: <param name="productName" parameterType="string" > 5: <value value="Product name from config file" type="string"/> 6: </param> 7: <param name="loggers" parameterType="ILoggerArray"> 8: <array> 9: <dependency name="logger2" /> 10: <dependency name="logger1" /> 11: </array> 12: </param> 13: </constructor> 14: </typeConfig> 15: </type> Here’s where I’m saying, that if a type of IProduct is requested to be resolved, map it to type Product4. Furthermore, the Product4 has two constructor parameters – a string and an array of type ILogger. You might have observed the first parameter of the constructor is named ‘productName’ and that matches the value in the name attribute of the param element. The parameterType of ‘string’ maps to ‘System.String, mscorlib’ and is defined in the type alias above. The set up is similar for the second constructor parameter. The name matches the name of the parameter (loggers) and is of type ILoggerArray, which maps to an array of ILogger objects. We’ve also decided to add two elements to this array when unity resolves it – an instance of FileLogger and one of FakeLogger. The click event of the button does the following: 1: //unityContainer.RegisterType<IProduct, Product4>(); 2: //IProduct product4 = unityContainer.Resolve<IProduct>(); 3: IProduct product4 = unityContainer.Resolve<IProduct>("ArrayConstructor"); 4: productDetailsLabel.Text = product4.WriteProductDetails(); It’s worth mentioning here about the change in the format of resolving the IProduct to create an instance of Product4. You cannot use the regular way (the commented lines) to get an instance of Product4. The reason is due to the behavior of Unity which Alex Ermakov has brilliantly explained here. The corresponding output of the action is: You have a couple of options when it comes to adding dependency elements in the array node. You can: - leave it empty (no dependency elements declared): This will only create an empty array of loggers. This way you can check for non-null condition, in your mock classes. - add multiple dependency elements with the same name 1: <param name="loggers" parameterType="ILoggerArray"> 2: <array> 3: <dependency name="logger2" /> 4: <dependency name="logger2" /> 5: </array> 6: </param> With this you’ll see two instances of FakeLogger in the output. This article shows how Unity allows you to instantiate objects with arrays. Find the code here.

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  • rm on a directory with millions of files

    - by BMDan
    Background: physical server, about two years old, 7200-RPM SATA drives connected to a 3Ware RAID card, ext3 FS mounted noatime and data=ordered, not under crazy load, kernel 2.6.18-92.1.22.el5, uptime 545 days. Directory doesn't contain any subdirectories, just millions of small (~100 byte) files, with some larger (a few KB) ones. We have a server that has gone a bit cuckoo over the course of the last few months, but we only noticed it the other day when it started being unable to write to a directory due to it containing too many files. Specifically, it started throwing this error in /var/log/messages: ext3_dx_add_entry: Directory index full! The disk in question has plenty of inodes remaining: Filesystem Inodes IUsed IFree IUse% Mounted on /dev/sda3 60719104 3465660 57253444 6% / So I'm guessing that means we hit the limit of how many entries can be in the directory file itself. No idea how many files that would be, but it can't be more, as you can see, than three million or so. Not that that's good, mind you! But that's part one of my question: exactly what is that upper limit? Is it tunable? Before I get yelled at--I want to tune it down; this enormous directory caused all sorts of issues. Anyway, we tracked down the issue in the code that was generating all of those files, and we've corrected it. Now I'm stuck with deleting the directory. A few options here: rm -rf (dir)I tried this first. I gave up and killed it after it had run for a day and a half without any discernible impact. unlink(2) on the directory: Definitely worth consideration, but the question is whether it'd be faster to delete the files inside the directory via fsck than to delete via unlink(2). That is, one way or another, I've got to mark those inodes as unused. This assumes, of course, that I can tell fsck not to drop entries to the files in /lost+found; otherwise, I've just moved my problem. In addition to all the other concerns, after reading about this a bit more, it turns out I'd probably have to call some internal FS functions, as none of the unlink(2) variants I can find would allow me to just blithely delete a directory with entries in it. Pooh. while [ true ]; do ls -Uf | head -n 10000 | xargs rm -f 2/dev/null; done ) This is actually the shortened version; the real one I'm running, which just adds some progress-reporting and a clean stop when we run out of files to delete, is: export i=0; time ( while [ true ]; do ls -Uf | head -n 3 | grep -qF '.png' || break; ls -Uf | head -n 10000 | xargs rm -f 2/dev/null; export i=$(($i+10000)); echo "$i..."; done ) This seems to be working rather well. As I write this, it's deleted 260,000 files in the past thirty minutes or so. Now, for the questions: As mentioned above, is the per-directory entry limit tunable? Why did it take "real 7m9.561s / user 0m0.001s / sys 0m0.001s" to delete a single file which was the first one in the list returned by "ls -U", and it took perhaps ten minutes to delete the first 10,000 entries with the command in #3, but now it's hauling along quite happily? For that matter, it deleted 260,000 in about thirty minutes, but it's now taken another fifteen minutes to delete 60,000 more. Why the huge swings in speed? Is there a better way to do this sort of thing? Not store millions of files in a directory; I know that's silly, and it wouldn't have happened on my watch. Googling the problem and looking through SF and SO offers a lot of variations on "find" that obviously have the wrong idea; it's not going to be faster than my approach for several self-evident reasons. But does the delete-via-fsck idea have any legs? Or something else entirely? I'm eager to hear out-of-the-box (or inside-the-not-well-known-box) thinking. Thanks for reading the small novel; feel free to ask questions and I'll be sure to respond. I'll also update the question with the final number of files and how long the delete script ran once I have that. Final script output!: 2970000... 2980000... 2990000... 3000000... 3010000... real 253m59.331s user 0m6.061s sys 5m4.019s So, three million files deleted in a bit over four hours.

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

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

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  • optimizing iPhone OpenGL ES fill rate

    - by NateS
    I have an Open GL ES game on the iPhone. My framerate is pretty sucky, ~20fps. Using the Xcode OpenGL ES performance tool on an iPhone 3G, it shows: Renderer Utilization: 95% to 99% Tiler Utilization: ~27% I am drawing a lot of pretty large images with a lot of blending. If I reduce the number of images drawn, framerates go from ~20 to ~40, though the performance tool results stay about the same (renderer still maxed). I think I'm being limited by the fill rate of the iPhone 3G, but I'm not sure. My questions are: How can I determine with more granularity where the bottleneck is? That is my biggest problem, I just don't know what is taking all the time. If it is fillrate, is there anything I do to improve it besides just drawing less? I am using texture atlases. I have tried to minimize image binds, though it isn't always possible (drawing order, not everything fits on one 1024x1024 texture, etc). Every frame I do 10 image binds. This seem pretty reasonable, but I could be mistaken. I'm using vertex arrays and glDrawArrays. I don't really have a lot of geometry. I can try to be more precise if needed. Each image is 2 triangles and I try to batch things were possible, though often (maybe half the time) images are drawn with individual glDrawArrays calls. Besides the images, I have ~60 triangles worth of geometry being rendered in ~6 glDrawArrays calls. I often glTranslate before calling glDrawArrays. Would it improve the framerate to switch to VBOs? I don't think it is a huge amount of geometry, but maybe it is faster for other reasons? Are there certain things to watch out for that could reduce performance? Eg, should I avoid glTranslate, glColor4g, etc? I'm using glScissor in a 3 places per frame. Each use consists of 2 glScissor calls, one to set it up, and one to reset it to what it was. I don't know if there is much of a performance impact here. If I used PVRTC would it be able to render faster? Currently all my images are GL_RGBA. I don't have memory issues. Here is a rough idea of what I'm drawing, in this order: 1) Switch to perspective matrix. 2) Draw a full screen background image 3) Draw a full screen image with translucency (this one has a scrolling texture). 4) Draw a few sprites. 5) Switch to ortho matrix. 6) Draw a few sprites. 7) Switch to perspective matrix. 8) Draw sprites and some other textured geometry. 9) Switch to ortho matrix. 10) Draw a few sprites (eg, game HUD). Steps 1-6 draw a bunch of background stuff. 8 draws most of the game content. 10 draws the HUD. As you can see, there are many layers, some of them full screen and some of the sprites are pretty large (1/4 of the screen). The layers use translucency, so I have to draw them in back-to-front order. This is further complicated by needing to draw various layers in ortho and others in perspective. I will gladly provide additional information if reqested. Thanks in advance for any performance tips or general advice on my problem!

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  • Does Apache ever give incorrect "out of threads" errors?

    - by Eli Courtwright
    Lately our Apache web server has been giving us this error multiple times per day: [Tue Apr 06 01:07:10 2010] [error] Server ran out of threads to serve requests. Consider raising the ThreadsPerChild setting We raised our ThreadsPerChild setting from 50 to 100, but we still get the error. Our access logs indicate that these errors never even happen at periods of high load. For example, here's an excerpt from our access log (ip addresses and some urls are edited for privacy). As you can see, the above error happened at 1:07 and only a small handful of requests occurred in the several minutes leading up to the error: 99.88.77.66 - - [06/Apr/2010:00:59:33 -0400] "GET /WebRepository/jquery/jquery-ui-1.7.1.custom/css/smoothness/images/ui-icons_222222_256x240.png HTTP/1.1" 304 - 99.88.77.66 - - [06/Apr/2010:00:59:34 -0400] "GET /WebRepository/jquery/jquery-ui-1.7.1.custom/css/smoothness/images/ui-bg_glass_75_dadada_1x400.png HTTP/1.1" 200 111 99.88.77.66 - - [06/Apr/2010:00:59:34 -0400] "GET /WebRepository/jquery/jquery-ui-1.7.1.custom/css/smoothness/images/ui-bg_glass_75_dadada_1x400.png HTTP/1.1" 200 111 99.88.77.66 - mpeu [06/Apr/2010:00:59:40 -0400] "GET /some/dynamic/content HTTP/1.1" 200 145049 55.44.33.22 - mpeu [06/Apr/2010:01:06:56 -0400] "GET /other/dynamic/content HTTP/1.1" 200 12311 55.44.33.22 - - [06/Apr/2010:01:06:56 -0400] "GET /WebRepository/jquery/jquery-ui-1.7.1.custom/css/smoothness/jquery-ui-1.7.1.custom.css HTTP/1.1" 304 - 55.44.33.22 - - [06/Apr/2010:01:06:56 -0400] "GET /WebRepository/jquery/jquery-ui-1.7.1.custom/js/jquery-1.3.2.min.js HTTP/1.1" 304 - 55.44.33.22 - - [06/Apr/2010:01:06:56 -0400] "GET /WebRepository/jquery/jquery-ui-1.7.1.custom/js/jquery-ui-1.7.1.custom.min.js HTTP/1.1" 304 - 55.44.33.22 - - [06/Apr/2010:01:06:56 -0400] "GET /WebRepository/jquery.tablesorter.min.js HTTP/1.1" 304 - 55.44.33.22 - - [06/Apr/2010:01:06:56 -0400] "GET /WebRepository/date.js HTTP/1.1" 304 - 55.44.33.22 - - [06/Apr/2010:01:06:56 -0400] "GET /WebRepository/pdfs/image1.gif HTTP/1.1" 304 - 55.44.33.22 - - [06/Apr/2010:01:06:56 -0400] "GET /WebRepository/pdfs/image2.png HTTP/1.1" 304 - 55.44.33.22 - - [06/Apr/2010:01:06:56 -0400] "GET /WebRepository/pdfs/image3.png HTTP/1.1" 304 - 55.44.33.22 - - [06/Apr/2010:01:06:56 -0400] "GET /WebRepository/pdfs/image4.png HTTP/1.1" 304 - 55.44.33.22 - - [06/Apr/2010:01:06:56 -0400] "GET /WebRepository/pdfs/image5.png HTTP/1.1" 304 - 55.44.33.22 - - [06/Apr/2010:01:06:56 -0400] "GET /WebRepository/pdfs/image6.png HTTP/1.1" 304 - 55.44.33.22 - - [06/Apr/2010:01:06:56 -0400] "GET /WebRepository/pdfs/image7.png HTTP/1.1" 304 - 55.44.33.22 - - [06/Apr/2010:01:06:57 -0400] "GET /WebRepository/pdfs/image8.png HTTP/1.1" 304 - 55.44.33.22 - - [06/Apr/2010:01:06:57 -0400] "GET /WebRepository/pdfs/image9.png HTTP/1.1" 304 - 55.44.33.22 - - [06/Apr/2010:01:06:57 -0400] "GET /WebRepository/pdfs/imageA.png HTTP/1.1" 304 - 55.44.33.22 - - [06/Apr/2010:01:06:57 -0400] "GET /WebRepository/jquery/jquery-ui-1.7.1.custom/css/smoothness/images/ui-bg_flat_75_ffffff_40x100.png HTTP/1.1" 304 - 55.44.33.22 - - [06/Apr/2010:01:06:59 -0400] "GET /WebRepository/jquery/jquery-ui-1.7.1.custom/css/smoothness/images/ui-bg_highlight-soft_75_cccccc_1x100.png HTTP/1.1" 304 - 55.44.33.22 - - [06/Apr/2010:01:06:59 -0400] "GET /WebRepository/jquery/jquery-ui-1.7.1.custom/css/smoothness/images/ui-bg_glass_75_e6e6e6_1x400.png HTTP/1.1" 200 110 55.44.33.22 - - [06/Apr/2010:01:06:59 -0400] "GET /WebRepository/jquery/jquery-ui-1.7.1.custom/css/smoothness/images/ui-bg_glass_75_e6e6e6_1x400.png HTTP/1.1" 200 110 11.22.33.44 - mpeu [06/Apr/2010:01:18:03 -0400] "GET /other/dynamic/content HTTP/1.1" 200 12311 11.22.33.44 - - [06/Apr/2010:01:18:03 -0400] "GET /WebRepository/jquery/jquery-ui-1.7.1.custom/js/jquery-1.3.2.min.js HTTP/1.1" 304 - 11.22.33.44 - - [06/Apr/2010:01:18:04 -0400] "GET /WebRepository/jquery/jquery-ui-1.7.1.custom/css/smoothness/jquery-ui-1.7.1.custom.css HTTP/1.1" 200 27374 11.22.33.44 - - [06/Apr/2010:01:18:04 -0400] "GET /WebRepository/jquery/jquery-ui-1.7.1.custom/js/jquery-ui-1.7.1.custom.min.js HTTP/1.1" 304 - 11.22.33.44 - - [06/Apr/2010:01:18:04 -0400] "GET /WebRepository/jquery.tablesorter.min.js HTTP/1.1" 200 12795 11.22.33.44 - - [06/Apr/2010:01:18:04 -0400] "GET /WebRepository/date.js HTTP/1.1" 200 25809 For what it's worth, we're running the version of Apache that ships with Oracle 10g (some 2.0 version), and we're using mod_plsql to generate our dynamic content. Since the Apache server runs as a separate process and the database doesn't record any problems when this error occurs, I'm doubtful that Oracle is the problem. Unfortunately, the errors are freaking out our sysadmins, who are inclined to blame any and all problems which occur with the server on this error. Is this a known bug in Apache that I simply haven't been able to find any reference to through Google?

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  • Summer Programming Plans

    - by Gabe
    I've wanted to start "hacking" for many months now. But I put it off in favor of school and other things. Now, though, I'm free for the summer and want to learn as much as I can. I have a rough idea of what I want to try my hand at, but need some guidance as to what specifically - and how - I should learn. This is my plan so far: 1) Get good at programming in general. I plan to read up on how to think/work like a programmer. I'm waiting for the Pragmatic Programmer to arrive, which will be the first book I read. Q: What other books/ebooks should I look at? What more can I do here? 2) Learn/Improve at HTML/CSS. My first project will be to make a personal website/blog for myself using HTML and CSS. ----Then I hope to write/design articles like Dustin Curtis. After I finish this (and learn a programming language) I'll try to create user-based a user-focused website. Q: It's my understanding that just trying to design/manage websites is a good way to learn/improve at HTML/CSS. Is that all correct? 3) Try music development. This might be a sort of stretch for stackoverflow, but I'm interested in mixing/making techno songs. (Think Justice, or Daft Punk, or MSTRKRFT.) Q: I have a Mac. Any ideas on how I could start/learn music making? Any programs I should download, for instance? 4) My main goal: Learning a web development language/framework. I'm a year into learning/using C++. But what I really want to do is develop websites and web apps. I've searched online, and there seems to be great debate over which language/framework to learn first (and which is best). I think I've narrowed it down to three: Ruby (Rails), Python (Django), and PHP (?). Q #1: Which should I learn and use first? (Reasons?) Q #2: One reason I was leaning towards PHP is that I'm taking a PHP development course next semester. Learning it now would make that course easy. If PHP was not the answer to Q #1, is it worth learning both? Or, would it be better to just focus on PHP for this summer and next semester, and then transition thereafter to a better language? 5) iPhone/iPad Programming (Maybe). I've a number of simple, useful app ideas that I'd like to eventually get too. I just bought a Mac, as well as a few app development books. Q #1: Am I spreading myself thin trying to learn all of the above, and objective-C? Q #2: How much harder/easier is objective-C compared to the above languages? Also, how easy is it to learn obj-C after learning a web development language (and some C++)? Q #3: Yes or no? Should I go for it, or just keeep with #1-4 for now? Also: If you have any tips on how I should learn (or how you learned to hack), I'm all ears. I'd be especially interested in how you planned out learning: did you just hack whenever you felt like it, or did you "study" the language a few hours a day, or something else? Thanks so much, guys.

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  • FMOD.net streaming, callback and exinfo parameters

    - by Tesserex
    I posted a question on gamedev about how to play nsf files (NES console music) in FMOD. It didn't get any results, but since then I made some progress. I decided that the easiest method was just to compile an existing player into a dll and then call it from C# to populate my buffer. The problem now is getting it to sound right, and making sure all my paremeters are correct. Here are the facts so far: The nsf dll is dealing with shorts, so the data is PCM16. The sample nsf I'm using has a playback rate of 60 Hz. Just for playing around now, I'm using a frequency of 48000. Based on 2 and 3, the dll calculates a necessary buffer size of 48000 / 60hz = 800. This means it will render 800 shorts worth of buffer for every simulated NES frame. I've so far got my C# code to play the nsf, at the correct pitch and tempo, but it's very grainy / fuzzy, which I'm attributing to the fact that the FMOD read callback is giving a data length of 1600, whereas I should be expecting 800. I've tried playing around with all the numbers and it either crashes, or the music changes pitch, tempo, or both. Here's some of my C# code: uint channels = 1, frequency = 48000; FMOD.MODE mode = (FMOD.MODE.DEFAULT | FMOD.MODE.OPENUSER | FMOD.MODE.LOOP_NORMAL); FMOD.Sound sound = new FMOD.Sound(); FMOD.CREATESOUNDEXINFO ex = new FMOD.CREATESOUNDEXINFO(); ex.cbsize = Marshal.SizeOf(ex); ex.fileoffset = 0; ex.format = FMOD.SOUND_FORMAT.PCM16; // does this even matter? It doesn't change my results as long as it's long enough for one update ex.length = frequency; ex.numchannels = (int)channels; ex.defaultfrequency = (int)frequency; ex.pcmreadcallback = pcmreadcallback; ex.dlsname = null; // eventually I will calculate this with frequency / nsf hz, but I'm just testing for now ex.decodebuffersize = 800; // from the dll load_nsf_file("file.nsf", 8, (int)frequency); // 8 is the track number to play var result = system.createSound( (string)null, (mode | FMOD.MODE.CREATESTREAM), ref ex, ref sound); channel = new FMOD.Channel(); result = system.playSound(FMOD.CHANNELINDEX.FREE, sound, false, ref channel); private FMOD.RESULT PCMREADCALLBACK(IntPtr soundraw, IntPtr data, uint datalen) { // from the dll process_buffer(data, (int)800); // if I use datalen, it usually crashes (I can't get datalen to = 800 safely) return FMOD.RESULT.OK; } So here are some of my questions: What is the relationship between exinfo.decodebuffersize, frequency, and the datalen parameter of the read callback? With this code sample, it's coming in as 3200. I don't know where that factor of 4 between it and the decodebuffersize comes from. Is datalen in the callback referring to number of bytes, or shorts? The process_buffer function takes a short array and its length. I would expect fmod is talking about shorts as well because I told it PCM16. Maybe my playback quality is bad for some totally different reason. If so I have no idea where to begin solving that. Any ideas there?

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  • Understanding the memory consumption on iPhone

    - by zoul
    Hello! I am working on a 2D iPhone game using OpenGL ES and I keep hitting the 24 MB memory limit – my application keeps crashing with the error code 101. I tried real hard to find where the memory goes, but the numbers in Instruments are still much bigger than what I would expect. I ran the application with the Memory Monitor, Object Alloc, Leaks and OpenGL ES instruments. When the application gets loaded, free physical memory drops from 37 MB to 23 MB, the Object Alloc settles around 7 MB, Leaks show two or three leaks a few bytes in size, the Gart Object Size is about 5 MB and Memory Monitor says the application takes up about 14 MB of real memory. I am perplexed as where did the memory go – when I dig into the Object Allocations, most of the memory is in the textures, exactly as I would expect. But both my own texture allocation counter and the Gart Object Size agree that the textures should take up somewhere around 5 MB. I am not aware of allocating anything else that would be worth mentioning, and the Object Alloc agrees. Where does the memory go? (I would be glad to supply more details if this is not enough.) Update: I really tried to find where I could allocate so much memory, but with no results. What drives me wild is the difference between the Object Allocations (~7 MB) and real memory usage as shown by Memory Monitor (~14 MB). Even if there were huge leaks or huge chunks of memory I forget about, the should still show up in the Object Allocations, shouldn’t they? I’ve already tried the usual suspects, ie. the UIImage with its caching, but that did not help. Is there a way to track memory usage “debugger-style”, line by line, watching each statement’s impact on memory usage? What I have found so far: I really am using that much memory. It is not easy to measure the real memory consumption, but after a lot of counting I think the memory consumption is really that high. My fault. I found no easy way to measure the memory used. The Memory Monitor numbers are accurate (these are the numbers that really matter), but the Memory Monitor can’t tell you where exactly the memory goes. The Object Alloc tool is almost useless for tracking the real memory usage. When I create a texture, the allocated memory counter goes up for a while (reading the texture into the memory), then drops (passing the texture data to OpenGL, freeing). This is OK, but does not always happen – sometimes the memory usage stays high even after the texture has been passed on to OpenGL and freed from “my” memory. This means that the total amount of memory allocated as shown by the Object Alloc tool is smaller than the real total memory consumption, but bigger than the real consumption minus textures (real – textures < object alloc < real). Go figure. I misread the Programming Guide. The memory limit of 24 MB applies to textures and surfaces, not the whole application. The actual red line lies a bit further, but I could not find any hard numbers. The consensus is that 25–30 MB is the ceiling. When the system gets short on memory, it starts sending the memory warning. I have almost nothing to free, but other applications do release some memory back to the system, especially Safari (which seems to be caching the websites). When the free memory as shown in the Memory Monitor goes zero, the system starts killing. I had to bite the bullet and rewrite some parts of the code to be more efficient on memory, but I am probably still pushing it. I

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  • Odd tcp deadlock under windows

    - by John Robertson
    We are moving large amounts of data on a LAN and it has to happen very rapidly and reliably. Currently we use windows TCP as implemented in C++. Using large (synchronous) sends moves the data much faster than a bunch of smaller (synchronous) sends but will frequently deadlock for large gaps of time (.15 seconds) causing the overall transfer rate to plummet. This deadlock happens in very particular circumstances which makes me believe it should be preventable altogether. More importantly if we don't really know the cause we don't really know it won't happen some time with smaller sends anyway. Can anyone explain this deadlock? Deadlock description (OK, zombie-locked, it isn't dead, but for .15 or so seconds it stops, then starts again) The receiving side sends an ACK. The sending side sends a packet containing the end of a message (push flag is set) The call to socket.recv takes about .15 seconds(!) to return About the time the call returns an ACK is sent by the receiving side The the next packet from the sender is finally sent (why is it waiting? the tcp window is plenty big) The odd thing about (3) is that typically that call doesn't take much time at all and receives exactly the same amount of data. On a 2Ghz machine that's 300 million instructions worth of time. I am assuming the call doesn't (heaven forbid) wait for the received data to be acked before it returns, so the ack must be waiting for the call to return, or both must be delayed by something else. The problem NEVER happens when there is a second packet of data (part of the same message) arriving between 1 and 2. That part very clearly makes it sound like it has to do with the fact that windows TCP will not send back a no-data ACK until either a second packet arrives or a 200ms timer expires. However the delay is less than 200 ms (its more like 150 ms). The third unseemly character (and to my mind the real culprit) is (5). Send is definitely being called well before that .15 seconds is up, but the data NEVER hits the wire before that ack returns. That is the most bizarre part of this deadlock to me. Its not a tcp blockage because the TCP window is plenty big since we set SO_RCVBUF to something like 500*1460 (which is still under a meg). The data is coming in very fast (basically there is a loop spinning out data via send) so the buffer should fill almost immediately. According to msdn the buffer being full and at least one pending send should cause the data to be sent (though in another place it mentions that there various "heuristics" used in deciding when a send hits the wire). Anway, why the sender doesn't actually send more data during that .15 second pause is the most bizarre part to me. The information above was captured on the receiving side via wireshark (except of course the socket.recv return times which were logged in a text file). We tried changing the send buffer to zero and turning off Nagle on the sender (yes, I know Nagle is about not sending small packets - but we tried turning Nagle off in case that was part of the unstated "heuristics" affecting whether the message would be posted to the wire. Technically microsoft's Nagle is that a small packet isn't sent if the buffer is full and there is an outstanding ACK, so it seemed like a possibility).

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  • Why C# doesn't implement indexed properties ?

    - by Thomas Levesque
    I know, I know... Eric Lippert's answer to this kind of question is usually something like "because it wasn't worth the cost of designing, implementing, testing and documenting it". But still, I'd like a better explanation... I was reading this blog post about new C# 4 features, and in the section about COM Interop, the following part caught my attention : By the way, this code uses one more new feature: indexed properties (take a closer look at those square brackets after Range.) But this feature is available only for COM interop; you cannot create your own indexed properties in C# 4.0. OK, but why ? I already knew and regretted that it wasn't possible to create indexed properties in C#, but this sentence made me think again about it. I can see several good reasons to implement it : the CLR supports it (for instance, PropertyInfo.GetValue has an index parameter), so it's a pity we can't take advantage of it in C# it is supported for COM interop, as shown in the article (using dynamic dispatch) it is implemented in VB.NET it is already possible to create indexers, i.e. to apply an index to the object itself, so it would probably be no big deal to extend the idea to properties, keeping the same syntax and just replacing this with a property name It would allow to write that kind of things : public class Foo { private string[] _values = new string[3]; public string Values[int index] { get { return _values[index]; } set { _values[index] = value; } } } Currently the only workaround that I know is to create an inner class (ValuesCollection for instance) that implements an indexer, and change the Values property so that it returns an instance of that inner class. This is very easy to do, but annoying... So perhaps the compiler could do it for us ! An option would be to generate an inner class that implements the indexer, and expose it through a public generic interface : // interface defined in the namespace System public interface IIndexer<TIndex, TValue> { TValue this[TIndex index] { get; set; } } public class Foo { private string[] _values = new string[3]; private class <>c__DisplayClass1 : IIndexer<int, string> { private Foo _foo; public <>c__DisplayClass1(Foo foo) { _foo = foo; } public string this[int index] { get { return _foo._values[index]; } set { _foo._values[index] = value; } } } private IIndexer<int, string> <>f__valuesIndexer; public IIndexer<int, string> Values { get { if (<>f__valuesIndexer == null) <>f__valuesIndexer = new <>c__DisplayClass1(this); return <>f__valuesIndexer; } } } But of course, in that case the property would actually return a IIndexer<int, string>, and wouldn't really be an indexed property... It would be better to generate a real CLR indexed property. What do you think ? Would you like to see this feature in C# ? If not, why ?

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  • image processing algorithm in MATLAB

    - by user261002
    I am trying to reconstruct an algorithm belong to this paper: Decomposition of biospeckle images in temporary spectral bands Here is an explanation of the algorithm: We recorded a sequence of N successive speckle images with a sampling frequency fs. In this way it was possible to observe how a pixel evolves through the N images. That evolution can be treated as a time series and can be processed in the following way: Each signal corresponding to the evolution of every pixel was used as input to a bank of filters. The intensity values were previously divided by their temporal mean value to minimize local differences in reflectivity or illumination of the object. The maximum frequency that can be adequately analyzed is determined by the sampling theorem and s half of sampling frequency fs. The latter is set by the CCD camera, the size of the image, and the frame grabber. The bank of filters is outlined in Fig. 1. In our case, ten 5° order Butterworth11 filters were used, but this number can be varied according to the required discrimination. The bank was implemented in a computer using MATLAB software. We chose the Butter-worth filter because, in addition to its simplicity, it is maximally flat. Other filters, an infinite impulse response, or a finite impulse response could be used. By means of this bank of filters, ten corresponding signals of each filter of each temporary pixel evolution were obtained as output. Average energy Eb in each signal was then calculated: where pb(n) is the intensity of the filtered pixel in the nth image for filter b divided by its mean value and N is the total number of images. In this way, en values of energy for each pixel were obtained, each of hem belonging to one of the frequency bands in Fig. 1. With these values it is possible to build ten images of the active object, each one of which shows how much energy of time-varying speckle there is in a certain frequency band. False color assignment to the gray levels in the results would help in discrimination. and here is my MATLAB code base on that : clear all for i=0:39 str = num2str(i); str1 = strcat(str,'.mat'); load(str1); D{i+1}=A; end new_max = max(max(A)); new_min = min(min(A)); for i=20:180 for j=20:140 ts = []; for k=1:40 ts = [ts D{k}(i,j)]; %%% kth image pixel i,j --- ts is time series end ts = double(ts); temp = mean(ts); ts = ts-temp; ts = ts/temp; N = 5; % filter order W = [0.00001 0.05;0.05 0.1;0.1 0.15;0.15 0.20;0.20 0.25;0.25 0.30;0.30 0.35;0.35 0.40;0.40 0.45;0.45 0.50]; N1 = 5; for ind = 1:10 Wn = W(ind,:); [B,A] = butter(N1,Wn); ts_f(ind,:) = filter(B,A,ts); end for ind=1:10 imag_test1{ind}(i,j) =sum((ts_f(ind,:)./mean(ts_f(ind,:))).^2); end end end for i=1:10 temp_imag = imag_test1{i}(:,:); x=isnan(temp_imag); temp_imag(x)=0; temp_imag=medfilt2(temp_imag); t_max = max(max(temp_imag)); t_min = min(min(temp_imag)); temp_imag = (temp_imag-t_min).*(double(new_max-new_min)/double(t_max-t_min))+double(new_min); imag_test2{i}(:,:) = temp_imag; end for i=1:10 A=imag_test2{i}(:,:); B=A/max(max(A)); B=histeq(B); figure,imshow(B) colorbar end but I am not getting the same result as paper. has anybody has aby idea why? or where I have gone wrong? Refrence Link to the paper

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  • How to sanely configure security policy in Tomcat 6

    - by Chas Emerick
    I'm using Tomcat 6.0.24, as packaged for Ubuntu Karmic. The default security policy of Ubuntu's Tomcat package is pretty stringent, but appears straightforward. In /var/lib/tomcat6/conf/policy.d, there are a variety of files that establish default policy. Worth noting at the start: I've not changed the stock tomcat install at all -- no new jars into its common lib directory(ies), no server.xml changes, etc. Putting the .war file in the webapps directory is the only deployment action. the web application I'm deploying fails with thousands of access denials under this default policy (as reported to the log thanks to the -Djava.security.debug="access,stack,failure" system property). turning off the security manager entirely results in no errors whatsoever, and proper app functionality What I'd like to do is add an application-specific security policy file to the policy.d directory, which seems to be the recommended practice. I added this to policy.d/100myapp.policy (as a starting point -- I would like to eventually trim back the granted permissions to only what the app actually needs): grant codeBase "file:${catalina.base}/webapps/ROOT.war" { permission java.security.AllPermission; }; grant codeBase "file:${catalina.base}/webapps/ROOT/-" { permission java.security.AllPermission; }; grant codeBase "file:${catalina.base}/webapps/ROOT/WEB-INF/-" { permission java.security.AllPermission; }; grant codeBase "file:${catalina.base}/webapps/ROOT/WEB-INF/lib/-" { permission java.security.AllPermission; }; grant codeBase "file:${catalina.base}/webapps/ROOT/WEB-INF/classes/-" { permission java.security.AllPermission; }; Note the thrashing around attempting to find the right codeBase declaration. I think that's likely my fundamental problem. Anyway, the above (really only the first two grants appear to have any effect) almost works: the thousands of access denials are gone, and I'm left with just one. Relevant stack trace: java.security.AccessControlException: access denied (java.io.FilePermission /var/lib/tomcat6/webapps/ROOT/WEB-INF/classes/com/foo/some-file-here.txt read) java.security.AccessControlContext.checkPermission(AccessControlContext.java:323) java.security.AccessController.checkPermission(AccessController.java:546) java.lang.SecurityManager.checkPermission(SecurityManager.java:532) java.lang.SecurityManager.checkRead(SecurityManager.java:871) java.io.File.exists(File.java:731) org.apache.naming.resources.FileDirContext.file(FileDirContext.java:785) org.apache.naming.resources.FileDirContext.lookup(FileDirContext.java:206) org.apache.naming.resources.ProxyDirContext.lookup(ProxyDirContext.java:299) org.apache.catalina.loader.WebappClassLoader.findResourceInternal(WebappClassLoader.java:1937) org.apache.catalina.loader.WebappClassLoader.findResource(WebappClassLoader.java:973) org.apache.catalina.loader.WebappClassLoader.getResource(WebappClassLoader.java:1108) java.lang.ClassLoader.getResource(ClassLoader.java:973) I'm pretty convinced that the actual file that's triggering the denial is irrelevant -- it's just some properties file that we check for optional configuration parameters. What's interesting is that: it doesn't exist in this context the fact that the file doesn't exist ends up throwing a security exception, rather than java.io.File.exists() simply returning false (although I suppose that's just a matter of the semantics of the read permission). Another workaround (besides just disabling the security manager in tomcat) is to add an open-ended permission to my policy file: grant { permission java.security.AllPermission; }; I presume this is functionally equivalent to turning off the security manager. I suppose I must be getting the codeBase declaration in my grants subtly wrong, but I'm not seeing it at the moment.

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  • Overriding Object.Equals() instance method in C#; now Code Analysis / FxCop warning CA2218: "should

    - by Chris W. Rea
    I've got a complex class in my C# project on which I want to be able to do equality tests. It is not a trivial class; it contains a variety of scalar properties as well as references to other objects and collections (e.g. IDictionary). For what it's worth, my class is sealed. To enable a performance optimization elsewhere in my system (an optimization that avoids a costly network round-trip), I need to be able to compare instances of these objects to each other for equality – other than the built-in reference equality – and so I'm overriding the Object.Equals() instance method. However, now that I've done that, Visual Studio 2008's Code Analysis a.k.a. FxCop, which I keep enabled by default, is raising the following warning: warning : CA2218 : Microsoft.Usage : Since 'MySuperDuperClass' redefines Equals, it should also redefine GetHashCode. I think I understand the rationale for this warning: If I am going to be using such objects as the key in a collection, the hash code is important. i.e. see this question. However, I am not going to be using these objects as the key in a collection. Ever. Feeling justified to suppress the warning, I looked up code CA2218 in the MSDN documentation to get the full name of the warning so I could apply a SuppressMessage attribute to my class as follows: [SuppressMessage("Microsoft.Naming", "CA2218:OverrideGetHashCodeOnOverridingEquals", Justification="This class is not to be used as key in a hashtable.")] However, while reading further, I noticed the following: How to Fix Violations To fix a violation of this rule, provide an implementation of GetHashCode. For a pair of objects of the same type, you must ensure that the implementation returns the same value if your implementation of Equals returns true for the pair. When to Suppress Warnings ----- Do not suppress a warning from this rule. [arrow & emphasis mine] So, I'd like to know: Why shouldn't I suppress this warning as I was planning to? Doesn't my case warrant suppression? I don't want to code up an implementation of GetHashCode() for this object that will never get called, since my object will never be the key in a collection. If I wanted to be pedantic, instead of suppressing, would it be more reasonable for me to override GetHashCode() with an implementation that throws a NotImplementedException? Update: I just looked this subject up again in Bill Wagner's good book Effective C#, and he states in "Item 10: Understand the Pitfalls of GetHashCode()": If you're defining a type that won't ever be used as the key in a container, this won't matter. Types that represent window controls, web page controls, or database connections are unlikely to be used as keys in a collection. In those cases, do nothing. All reference types will have a hash code that is correct, even if it is very inefficient. [...] In most types that you create, the best approach is to avoid the existence of GetHashCode() entirely. ... that's where I originally got this idea that I need not be concerned about GetHashCode() always.

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  • Android launches system settings instead of my app

    - by jsundin
    Hi, For some reason whenever I (try to) start my app the phone decides to launch system settings instead of my "main activity". And yes, I am referring to the "Android system settings", and not something from my app. This only happens on my phone, and I suppose it probably could be related to the fact that my app had just opened system settings when I decided to re-launch with a new version from Eclipse. It is possible to start the app from within Eclipse, but when I navigate back from the app it returns to the system settings rather than the home screen, as if the settings activity was started first and then my activity. If I then start the app from the phone all I get is system settings yet again. The app is listening to the VIEW-action for a specific URL substring, and when I start the app using a matching URL I get the same result as when I start it from Eclipse, app starts, but when I return I return to settings. I have tried googling for this problem, and all I could find was something about Android saving state when an app gets killed, but without any information on how to reset this state. I have tried uninstalling the app, killing system settings, rebooting the phone, reinstalling, clearing application data.. no luck.. For what it's worth, here's the definition of my main activity from the manifest, <activity android:name=".HomeActivity" android:label="@string/app_name" android:screenOrientation="portrait" android:clearTaskOnLaunch="true" android:launchMode="singleTop"> <intent-filter> <action android:name="android.intent.action.MAIN" /> <category android:name="android.intent.category.LAUNCHER" /> </intent-filter> <intent-filter> <action android:name="android.intent.action.VIEW"></action> <category android:name="android.intent.category.DEFAULT"></category> <category android:name="android.intent.category.BROWSABLE"></category> <data android:pathPrefix="/isak-web-mobile/smart/" android:scheme="http" android:host="*"></data> </intent-filter> </activity> And here is the logcat-line from when I try to start my app, nothing about any settings anywhere. I/ActivityManager( 1301): Starting activity: Intent { act=android.intent.action.MAIN cat=[android.intent.category.LAUNCHER] flg=0x10200000 cmp=se.opencare.isak/.HomeActivity } When I launch from Eclipse I also get this line (as one would expect), I/ActivityManager( 1301): Start proc se.opencare.isak for activity se.opencare.isak/.HomeActivity: pid=23068 uid=10163 gids={3003, 1007, 1015} If it matters the phone is a HTC Desire Z running 2.2.1. Currently, this is my HomeActivity, public class HomeActivity extends Activity { public static final String TAG = "HomeActivity"; @Override protected void onActivityResult(int requestCode, int resultCode, Intent data) { Log.d(TAG, "onActivityResult(" + requestCode + ", " + resultCode + ", " + data + ")"); super.onActivityResult(requestCode, resultCode, data); } @Override protected void onCreate(Bundle savedInstanceState) { Log.d(TAG, "onCreate(" + savedInstanceState + ")"); super.onCreate(savedInstanceState); } @Override protected void onDestroy() { Log.d(TAG, "onDestroy()"); super.onDestroy(); } @Override protected void onPause() { Log.d(TAG, "onPause()"); super.onPause(); } @Override protected void onPostCreate(Bundle savedInstanceState) { Log.d(TAG, "onPostCreate(" + savedInstanceState + ")"); super.onPostCreate(savedInstanceState); } @Override protected void onPostResume() { Log.d(TAG, "onPostResume()"); super.onPostResume(); } @Override protected void onRestart() { Log.d(TAG, "onRestart()"); super.onRestart(); } @Override protected void onRestoreInstanceState(Bundle savedInstanceState) { Log.d(TAG, "onRestoreInstanceState(" + savedInstanceState + ")"); super.onRestoreInstanceState(savedInstanceState); } @Override protected void onResume() { Log.d(TAG, "onResume()"); super.onResume(); } @Override protected void onStart() { Log.d(TAG, "onStart()"); super.onStart(); } @Override protected void onStop() { Log.d(TAG, "onStop()"); super.onStop(); } @Override protected void onUserLeaveHint() { Log.d(TAG, "onUserLeaveHint()"); super.onUserLeaveHint(); } } Nothing (of the above) is written to the log.

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  • Why isn't my algorithm for find the biggest and smallest inputs working?

    - by Matt Ellen
    I have started a new job, and with it comes a new language: Ironpython. Thankfully a good language :D Before starting I got to grips with Python on the whole, but that was only a week's worth of learning. Now I'm writing actual code. I've been charged with writing an algorithm that finds the best input parameter to collect data with. The basic algorithm is (as I've been instructed): Set the input parameter to a good guess Start collecting data When data is available stop collecting find the highest point If the point before this (i.e. for the previous parameter value) was higher and the point before that was lower then we've found the max otherwise the input parameter is increased by the initial guess. goto 2 If the max is found then the min needs to be found. To do this the algorithm carries on increasing the input, but by 1/10 of the max, until the current point is greater than the previous point and the point before that is also greater. Once the min is found then the algorithm stops. Currently I have a simplified data generator outputting the sin of the input, so that I know that the min value should be PI and the max value should be PI/2 The main Python code looks like this (don't worry, this is just for my edification, I don't write real code like this): import sys sys.path.append(r"F:\Programming Source\C#\PythonHelp\PythonHelp\bin\Debug") import clr clr.AddReferenceToFile("PythonHelpClasses.dll") import PythonHelpClasses from PHCStruct import Helper from System import Math helper = Helper() def run(): b = PythonHelpClasses.Executor() a = PythonHelpClasses.HasAnEvent() b.Input = 0.0 helper.__init__() def AnEventHandler(e): b.Stop() h = helper h.lastLastVal, h.lastVal, h.currentVal = h.lastVal, h.currentVal, e.Number if h.lastLastVal < h.lastVal and h.currentVal < h.lastVal and h.NotPast90: h.NotPast90 = False h.bestInput = h.lastInput inputInc = 0.0 if h.NotPast90: inputInc = Math.PI/10.0 else: inputInc = h.bestInput/10.0 if h.lastLastVal > h.lastVal and h.currentVal > h.lastVal and h.NotPast180: h.NotPast180 = False if h.NotPast180: h.lastInput, b.Input = b.Input, b.Input + inputInc b.Start(a) else: print "Best input:", h.bestInput print "Last input:", h.lastInput b.Stop() a.AnEvent += AnEventHandler b.Start(a) PHCStruct.py: class Helper(): def __init__(self): self.currentVal = 0 self.lastVal = 0 self.lastLastVal = 0 self.NotPast90 = True self.NotPast180 = True self.bestInput = 0 self.lastInput = 0 PythonHelpClasses has two small classes I wrote in C# before I realised how to do it in Ironpython. Executor runs a delegate asynchronously while it's running member is true. The important code: public void Start(HasAnEvent hae) { running = true; RunDelegate r = new RunDelegate(hae.UpdateNumber); AsyncCallback ac = new AsyncCallback(UpdateDone); IAsyncResult ar = r.BeginInvoke(Input, ac, null); } public void Stop() { running = false; } public void UpdateDone(IAsyncResult ar) { RunDelegate r = (RunDelegate)((AsyncResult)ar).AsyncDelegate; r.EndInvoke(ar); if (running) { AsyncCallback ac = new AsyncCallback(UpdateDone); IAsyncResult ar2 = r.BeginInvoke(Input, ac, null); } } HasAnEvent has a function that generates the sin of its input and fires an event with that result as its argument. i.e.: public void UpdateNumber(double val) { AnEventArgs e = new AnEventArgs(Math.Sin(val)); System.Threading.Thread.Sleep(1000); if (null != AnEvent) { AnEvent(e); } } The sleep is in there just to slow things down a bit. The problem I am getting is that the algorithm is not coming up with the best input being PI/2 and the final input being PI, but I can't see why. Also the best and final inputs are different each time I run the programme. Can anyone see why? Also when the algorithm terminates the best and final inputs are printed to the screen multiple times, not just once. Can someone explain why?

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  • Multidimensional array problem in VHDL?

    - by Nektarios
    I'm trying to use a multidimensional array in VHDL and I'm having a lot of trouble getting it to work properly. My issue is that I've got an array of 17, of 16 vectors, of a given size. What I want to do is create 17 registers that are array of 16 * std_logic_vector of 32 bits (which = my b, 512). So, I'm trying to pass in something to input and output on the register instantiation that tells the compiler/synthesizer that I want to pass in something that is 512 bits worth... Similar to in C if I had: int var[COLS][ROWS][ELEMENTS]; memcpy(&var[3].. // I'm talking about 3rd COL here, passing in memory that is ROWS*ELEMENTS long (My actual declaration is here:) type partial_pipeline_registers_type is array (0 to 16, 0 to 15) of std_logic_vector(iw - 1 downto 0); signal h_blk_pipelined_input : partial_pipeline_registers_type; I tried simply using h_blk_pipelined_input(0) .. up to (16) but this doesn't work. I get the following error, which makes me see that I need to double index in to the array: ERROR:HDLParsers:821 - (at the register) Wrong index type for h_blk_pipelined_input. So then I tried what's below, and I get this error: ERROR:HDLParsers:164 - (at the register code). parse error, unexpected TO, expecting COMMA or CLOSEPAR instantiate_h_pipelined_reg : regn generic map ( N=> b, init => bzeros ) port map ( clk => clk , rst => '0', en => '1', input => h_blk_pipelined_input((i - 1), 0 to 15), output=> h_blk_pipelined_input((i), 0 to 15)); -- Changing 0 to 15 to (0 to 15) has no effect... I'm using XST, and from their documentation (http://www.xilinx.com/itp/xilinx6/books/data/docs/xst/xst0067_9.html), the above should have worked: ...declaration: subtype MATRIX15 is array(4 downto 0, 2 downto 0) of STD_LOGIC_VECTOR (7 downto 0); A multi-dimensional array signal or variable can be completely used: Just a slice of one row can be specified: MATRIX15 (4,4 downto 1) <= TAB_B (3 downto 0); One alternative is that I can create more registers that are 16 times smaller, and instead of trying to do all '0 to 15' at once, I would just do that 15 additional times. However, I think this may lead to inefficiency in synthesis and I don't feel like this is the right solution. EDIT: Tried what Ben said, instantiate_h_m_qa_pipeline_registers: for i in 1 to 16 generate instantiate_h_pipelined_reg : regn generic map ( N=> b, init => bzeros ) port map ( clk => clk , rst => '0', en => '1', input => h_blk_pipelined_input(i - 1), output=> h_blk_pipelined_input(i)); end generate instantiate_h_m_qa_pipeline_registers; The signals are now defined as: type std_logic_block is array (0 to 15) of std_logic_vector(iw - 1 downto 0) ; type partial_pipeline_registers_type is array (0 to 16) of std_logic_block; signal h_blk_pipelined_input : partial_pipeline_registers_type; And the error I get from XST is: ERROR:HDLParsers:800 - ((where the register part is)) Type of input is incompatible with type of h_blk_pipelined_input. I'm able to do everything I was able to do before, using ()() syntax instead of ( , ) so I haven't lost anything going this way, but it still doesn't resolve my problem.

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  • Help with infrequent segmentation fault in accessing boost::unordered_multimap or struct

    - by Sarah
    I'm having trouble debugging a segmentation fault. I'd appreciate tips on how to go about narrowing in on the problem. The error appears when an iterator tries to access an element of a struct Infection, defined as: struct Infection { public: explicit Infection( double it, double rt ) : infT( it ), recT( rt ) {} double infT; // infection start time double recT; // scheduled recovery time }; These structs are kept in a special structure, InfectionMap: typedef boost::unordered_multimap< int, Infection > InfectionMap; Every member of class Host has an InfectionMap carriage. Recovery times and associated host identifiers are kept in a priority queue. When a scheduled recovery event arises in the simulation for a particular strain s in a particular host, the program searches through carriage of that host to find the Infection whose recT matches the recovery time (double recoverTime). (For reasons that aren't worth going into, it's not as expedient for me to use recT as the key to InfectionMap; the strain s is more useful, and coinfections with the same strain are possible.) assert( carriage.size() > 0 ); pair<InfectionMap::iterator,InfectionMap::iterator> ret = carriage.equal_range( s ); InfectionMap::iterator it; for ( it = ret.first; it != ret.second; it++ ) { if ( ((*it).second).recT == recoverTime ) { // produces seg fault carriage.erase( it ); } } I get a "Program received signal EXC_BAD_ACCESS, Could not access memory. Reason: KERN_INVALID_ADDRESS at address..." on the line specified above. The recoverTime is fine, and the assert(...) in the code is not tripped. As I said, this seg fault appears 'randomly' after thousands of successful recovery events. How would you go about figuring out what's going on? I'd love ideas about what could be wrong and how I can further investigate the problem. Update I added a new assert and a check just inside the for loop: assert( carriage.size() > 0 ); assert( carriage.count( s ) > 0 ); pair<InfectionMap::iterator,InfectionMap::iterator> ret = carriage.equal_range( s ); InfectionMap::iterator it; cout << "carriage.count(" << s << ")=" << carriage.count(s) << endl; for ( it = ret.first; it != ret.second; it++ ) { cout << "(*it).first=" << (*it).first << endl; // error here if ( ((*it).second).recT == recoverTime ) { carriage.erase( it ); } } The EXC_BAD_ACCESS error now appears at the (*it).first call, again after many thousands of successful recoveries. Can anyone give me tips on how to figure out how this problem arises? I'm trying to use gdb. Frame 0 from the backtrace reads "#0 0x0000000100001d50 in Host::recover (this=0x100530d80, s=0, recoverTime=635.91148029170529) at Host.cpp:317" I'm not sure what useful information I can extract here. Update 2 I added a break; after the carriage.erase(it). This works, but I have no idea why (e.g., why it would remove the seg fault at (*it).first.

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