Search Results

Search found 43110 results on 1725 pages for 'noob question'.

Page 377/1725 | < Previous Page | 373 374 375 376 377 378 379 380 381 382 383 384  | Next Page >

  • postfix relaying all mail through office365 problems

    - by amrith
    This is a rather long question with a long list of things tried and travails so please bear with me. The summary is this. I am able to relay email from ubuntu through office365 using postfix; the configuration works. It only works as one of the users; more specifically the user who authenticates against office365 is the only valid "from" More details follow. I have a machine in Amazon's cloud on which I run a bunch of jobs and would like to have statuses mailed over to me. I use office365 at work so I want to relay mail through office365. I'm most familiar with postfix so I used that as the MTA. Configuration is ubuntu 12.04LTS; I've installed postfix and mail-utils. For this example, let me say my company is "company.com" and the machine in question (through an elastic IP and a DNS entry) is called "plaything.company.com". hostname is set to "plaything.company.com", so is /etc/mailname On plaything, I have the following users registered alpha, bravo, and charlie. I have the following configuration files. alias_database = hash:/etc/aliases alias_maps = hash:/etc/aliases append_dot_mydomain = no biff = no config_directory = /etc/postfix inet_interfaces = all inet_protocols = ipv4 mailbox_size_limit = 0 mydestination = plaything.company.com, localhost.company.com, , localhost myhostname = plaything.company.com mynetworks = 127.0.0.0/8 [::ffff:127.0.0.0]/104 [::1]/128 myorigin = /etc/mailname readme_directory = no recipient_delimiter = + relayhost = [smtp.office365.com]:587 sender_canonical_maps = hash:/etc/postfix/sender_canonical smtp_sasl_auth_enable = yes smtp_sasl_password_maps = hash:/etc/postfix/sasl_passwd smtp_sasl_security_options = noanonymous smtp_sasl_tls_security_options = noanonymous smtp_tls_CAfile = /etc/ssl/certs/ca-certificates.crt smtp_tls_session_cache_database = btree:${data_directory}/smtp_scache smtp_use_tls = yes smtpd_banner = $myhostname ESMTP $mail_name (Ubuntu) smtpd_tls_cert_file = /etc/ssl/certs/ssl-cert-snakeoil.pem smtpd_tls_key_file = /etc/ssl/private/ssl-cert-snakeoil.key smtpd_tls_session_cache_database = btree:${data_directory}/smtpd_scache smtpd_use_tls = yes As the machine is called plaything.company.com I went through the exercise of registering all the appropriate DNS entries to make office365 recognize that I owned plaything.company.com and allowed me to create a user called [email protected] in office365. In office365, I setup [email protected] as having another email address of [email protected]. Then, I made the following sender_canonical [email protected] [email protected] I created a sasl_passwd file that reads: smtp.office365.com [email protected]:123456password123456 let's just say that the password for [email protected] is 1234...456 With all this setup, login as alpha and mail [email protected] Cc: Subject: test test and the whole thing works wonderfully. email gets sent off by postfix, TLS works like a champ, authenticates as daemon@... and [email protected] in Office365 gets an email message. The issue comes up when logged in as bravo to the machine. sender is [email protected] and office365 says: status=bounced (host smtp.office365.com[132.245.12.25] said: 550 5.7.1 Client does not have permissions to send as this sender (in reply to end of DATA command)) this is because I'm trying to send mail as bravo@... and authenticating with office365 as daemon@.... The reason it works with alpha@... is because in office365, I setup [email protected] as having another email address of [email protected]. In Postfix Relay to Office365, Miles Erickson answers the question thusly: Don't send mail to Office365 as a user from your Office365-hosted e-mail domain. Use a subdomain instead, e.g. [email protected] instead of [email protected]. It wouldn't hurt to set up an SPF record for services.mydomain.com or whatever you decide to use. Don't authenticate against mail.messaging.microsoft.com as an Office365 user. Just connect on port 25 and deliver the mail to your domain as any foreign SMTP agent would do. OK, I've done #1, I have those records on DNS but for the most part they are not relevant once Office365 recognizes that I own the domain. Here are those records: CNAME records: - msoid.plaything.company.com - autodiscover.plaything.company.com MX record: - plaything.company.com (plaything-company-com.mail.protection.outlook.com) TXT record: - plaything.company.com (v=spf1 include:spf.protection.outlook.com -all) I've tried #2 but no matter what I do, office365 just blows away the connection with "not authenticated". I can try even a simple telnet to port 25 and attempt to send and it doesn't work. 250 BY2PR01CA007.outlook.office365.com Hello [54.221.245.236] 530 5.7.1 Client was not authenticated Connection closed by foreign host. Is there someone out there who has this kind of a configuration working where multiple users on a linux machine are able to relay mail using postfix through office365? There has to be someone out there doing this who can tell me what is wrong with my setup ...

    Read the article

  • Strategy for using snapshots to back up Ubuntu Linux server?

    - by MountainX
    I need some backup advice for my home file server. Here are the mount points, volume groups, logical volumes and used/total space of all the volumes on my Ubuntu 8.10 home file server. / vgA/lvRoot [7.5G/50G] /tmp vgB/lvTmp [195M/30G] /var vgB/lvVar [780M/30G] swap vgB/lvSwap [16.00 GB] /media1 vgC/lvMedia1 [400G/975G] /media2 vgC/lvMedia2 [75G/295G] /boot partition (no volume group) [95M/200M] /video partition (no volume group) [450G/950G] /backups vgD/lvBackupTarget [800G/925G] /home vgE/lvHome [85G/200G] I have just added a 2.0 TB external USB drive that I would like to use to backup everything. (It will be a close fit to get it all on one 2.0 TB drive. I actually have a 2nd external USB drive if needed.) I'd like to backup "/", var, /media1, media2 and /home. I'll deal with /boot and /video separately since they are not logical volumes. For all the logical volumes I'm anticipating taking snapshots and then copying those snapshots to the 2.0 TB external USB drive. I have never done a task like that before. If I do that, I could use the tutorial I found here: http://www.howtoforge.com/linux_lvm_snapshots My questions are: What is the best overall strategy? Is it LVM snapshots, as I'm assuming? How should I prepare, subdivide and mount the 2.0 TB external USB drive? 2.a. Should I create one or more regular partitions or should I create a physical volume with one or more logical volumes? 2.b. Would it be advisable to extactly mirror the source pv/lv layout on the external drive, and if so, is this a good strategy? What's the best way to get the snapshots onto the external drive? dd? Even though this is a strategy question, feedback with actual commands is appreciated. I need step-by-step cookbook-style help because I don't do much server admin work. (Background: This is a home file server that I have rarely had to touch in about 2 years. It has done its job without much intervention. The really old PC that I used to back everything up recently failed, so I'm replacing that with the external USB drive(s) and I'd like to upgrade my backup strategy at the same time. Previously, I just copied stuff from /backups over to the other computer and that would not have made things very easy in a real restore situation. The /backups mount point contains backup copies of "most" of the important data on a file by file basis, but it does not contain copies of /boot, etc. BTW, the actual internal HDD that holds /backups is separate from the other storage devices.) EDIT: I'll propose a strategy... The idea came from a comment here: LVM mirroring VS RAID1 "LVM mirrors are for replication of a logical volume to a different physical volume. It's essentially meant to "move the data to a different disk". The mirror is then broken..." That would fit my requirements well. Here is an ideal situation: establish the LV mirror on the external drive break the link with the mirror create a (persistent) snapshot on the mirror after a week, resync the mirror with the original source and update the mirror break the link and create another snapshot on the mirror. Obviously, the mirror will be like a weekly full backup. And the snapshots on the mirror will represent earlier points in time. If this would work and if it would be time efficient, it would give a nice full & differential type backup on the external drive based on LVM. I have not heard of a strategy like this before. Will it work? Could it be scripted? Thoughts? EDIT 2: Creating Portable DiskSafes With LoopbackFS And LVM Snapshots This article seems intriguing: http://www.howtoforge.com/creating-portable-disksafes-with-loopbackfs-and-lvm-snapshots Unfortunately, I don't understand exactly how to map those ideas to the strategy I'm proposing above. I'm going to ask this last bit as a separate question. I will leave my original question in place because I still desire feedback on the overall best strategy. At this moment I'm assuming it is LVM mirroring in the style of "Creating Portable DiskSafes with LVM Snapshots" but that might be wrong.

    Read the article

  • How to disable Mac OS X from using swap when there still is "Inactive" memory?

    - by Motin
    A common phenomena in my day to day usage (and several other's according to various posts throughout the internet) of OS X, the system seems to become slow whenever there is no more "Free" memory available. Supposedly, this is due to swapping, since heavy disk activity is apparent and that vm_stat reports many pageouts. (Correct me from wrong) However, the amount of "Inactive" ram is typically around 12.5%-25% of all available memory (^1.) when swapping starts/occurs/ends. According to http://support.apple.com/kb/ht1342 : Inactive memory This information in memory is not actively being used, but was recently used. For example, if you've been using Mail and then quit it, the RAM that Mail was using is marked as Inactive memory. This Inactive memory is available for use by another application, just like Free memory. However, if you open Mail before its Inactive memory is used by a different application, Mail will open quicker because its Inactive memory is converted to Active memory, instead of loading Mail from the slower hard disk. And according to http://developer.apple.com/library/mac/#documentation/Performance/Conceptual/ManagingMemory/Articles/AboutMemory.html : The inactive list contains pages that are currently resident in physical memory but have not been accessed recently. These pages contain valid data but may be released from memory at any time. So, basically: When a program has quit, it's memory becomes marked as Inactive and should be claimable at any time. Still, OS X will prefer to start swapping out memory to the Swap file instead of just claiming this memory, whenever the "Free" memory gets to low. Why? What is the advantage of this behavior over, say, instantly releasing Inactive memory and not even touch the swap file? Some sources (^2.) indicate that OS X would page out the "Inactive" memory to swap before releasing it, but that doesn't make sense now does it if the memory may be released from memory at any time? Swapping is expensive, releasing is cheap, right? Can this behavior be changed using some preference or known hack? (Preferably one that doesn't include disabling swap/dynamic_pager altogether and restarting...) I do appreciate the purge command, as well as the concept of Repairing disk permissions to force some Free memory, but those are ways to painfully force more Free memory than to actually fixing the swap/release decision logic... Btw a similar question was asked here: http://forums.macnn.com/90/mac-os-x/434650/why-does-os-x-swap-when/ and here: http://hintsforums.macworld.com/showthread.php?t=87688 but even though the OPs re-asked the core question, none of the replies addresses an answer to it... ^1. UPDATE 17-mar-2012 Since I first posted this question, I have gone from 4gb to 8gb of installed ram, and the problem remains. The amount of "Inactive" ram was 0.5gb-1.0gb before and is now typically around 1.0-2.0GB when swapping starts/occurs/ends, ie it seems that around 12.5%-25% of the ram is preserved as Inactive by osx kernel logic. ^2. For instance http://apple.stackexchange.com/questions/4288/what-does-it-mean-if-i-have-lots-of-inactive-memory-at-the-end-of-a-work-day : Once all your memory is used (free memory is 0), the OS will write out inactive memory to the swapfile to make more room in active memory. UPDATE 17-mar-2012 Here is a round-up of the methods that have been suggested to help so far: The purge command "Used to approximate initial boot conditions with a cold disk buffer cache for performance analysis. It does not affect anonymous memory that has been allocated through malloc, vm_allocate, etc". This is useful to prevent osx to swap-out the disk cache (which is ridiculous that osx actually does so in the first place), but with the downside that the disk cache is released, meaning that if the disk cache was not about to be swapped out, one would simply end up with a cold disk buffer cache, probably affecting performance negatively. The FreeMemory app and/or Repairing disk permissions to force some Free memory Doesn't help releasing any memory, only moving some gigabytes of memory contents from ram to the hd. In the end, this causes lots of swap-ins when I attempt to use the applications that were open while freeing memory, as a lot of its vm is now on swap. Speeding up swap-allocation using dynamicpagerwrapper Seems a good thing to do in order to speed up swap-usage, but does not address the problem of osx swapping in the first place while there is still inactive memory. Disabling swap by disabling dynamicpager and restarting This will force osx not to use swap to the price of the system hanging when all memory is used. Not a viable alternative... Disabling swap using a hacked dynamicpager Similar to disabling dynamicpager above, some excerpts from the comments to the blog post indicate that this is not a viable solution: "The Inactive Memory is high as usual". "when your system is running out of memory, the whole os hangs...", "if you consume the whole amount of memory of the mac, the machine will likely hang" To sum up, I am still unaware of a way of disabling Mac OS X from using swap when there still is "Inactive" memory. If it isn't possible, maybe at least there is an explanation somewhere of why osx prefers to swap out memory that may be released from memory at any time?

    Read the article

  • Managing BES Software Configurations

    - by DaveJohnston
    Hi, I am having problems with OTA deployment of a bespoke application that we have written. I have read loads of threads elsewhere and I have got mixed help, but for my particular case none of it has really helped. So I thought I would explain my exact situation and try and get some help here. I am running BES version 4.1.5 (Bundle 79) for Microsoft Exchange. The application we have written is split into 5 modules, which we control, and another 4 modules which are 3rd party libraries that we require. So for our modules the version numbers are regularly changing but for the others they are pretty much always going to remain the same. We have an alx file set up that identifies all of the files required and in fact I am able to create a software configuration and deploy the application with no problems. What I am trying to do however is maintain multiple versions of our application on the BES and be able to select which version I want to deploy to each user. I have tried this a number of ways (as I said I have read lots of other threads with solutions to this problem) but each seems to come with its own problem. First of all I tried just creating different configurations for each version of the application, but because they each had the same application ID the BES informed me that I couldn't do this. I read somewhere that the solution was to create a second shared folder (e.g. \Program Files\Common Files\RIM) and add the apploader stuff and the new version of the app to this folder. I could then create a second software configuration that would have the same application ID. The result of this seemed promising to start with. When I changed the config that was assigned to a user the new version was pushed out fine. But afterwards the BES reported that the device state was invalid, which meant I couldn't push anything else until I reactivated the device. I guess this is because the first config was never set to disallowed so the old version wasn't removed and the device essentially reported that it had multiple versions of the same application installed. The next suggestion I got was to change the application ID for each version, e.g. to include the version number. This meant that each version of the application could be included in a single configuration and I could set one to disallowed and the other to required. Initially this worked and the first version was deployed. But when I switched (i.e. the old version became disallowed and the new version required) the BES reported upgrade required and removed the old version. The device restarts and the old version is gone but the new version is not pushed out. I checked the BES and it still said Upgrade Required. I checked the log files and found: [40000] (11/12 09:50:27.397):{0xEB8} {[email protected], PIN=1234, UserId=2}SCS::PollDBQueueNewRequests - Queuing POLL_FOR_MISSING_APPS request [40000] (11/12 09:50:28.241):{0xE9C} RequestHandler::PollForMissingApps: Starting Poll For Missing Apps. [40304] (11/12 09:50:28.241):{0xE90} WorkerThreadPool:: ThreadProc(): Thread released with empty queue [40000] (11/12 09:50:28.241):{0xE9C} SCS::RemoveAppDeliveryRequests - No App Delivery Requests purged for User id 2 [30000] (11/12 09:50:28.960):{0xE9C} Discard duplicate module group "name" on device [30000] (11/12 09:50:28.960):{0xE9C} Discard duplicate module group "name" on device [40000] (11/12 09:50:29.163):{0xE9C} RequestHandler::PollForMissingApps: Completed Poll For Missing Apps, elapsed time 0.922 seconds. (You will notice I have removed actual names and email addresses etc for privacy reasons. But one question: where does the name of the module group come from? In my case it is close to the application ID but doesn't include the version number that I added at the end in order to get it to work. Is that information embedded in a COD file or something??) So it is reporting a duplicate module group on the device? What does this mean? I checked the device properties (as reported on the BES) and it confirms that the modules with the old version numbers are still present on the device. So the application has been removed but not the modules?? I checked the device and the modules are gone, so it is just the BES reporting that they are still there?? I checked the database and it has the modules in questions in the SyncDeviceMgmt table. If I delete these from the DB the BES changes to report Install Required, and low and behold the new version of the app is pushed out. So at the end of all that, my question is: does anyone have any other suggestions of how to handle upgrading our bespoke application OTA from the BES? Or can anyone point out something I am doing wrong in what I described above that might solve the problems I am having? I guess the question is why does the database maintain that the modules are on the device after they are removed? Thanks for any help you can provide.

    Read the article

  • How does this main domain have a CNAME record?

    - by TRiG
    I was under the impression that only subdomains could have CNAME records: main domains need to define all their own records. However, apt-get.com seems to have only a CNAME record. How can this work? $ dig apt-get.com ; <<>> DiG 9.8.1-P1 <<>> apt-get.com ;; global options: +cmd ;; Got answer: ;; ->>HEADER<<- opcode: QUERY, status: NOERROR, id: 45743 ;; flags: qr rd ra; QUERY: 1, ANSWER: 9, AUTHORITY: 0, ADDITIONAL: 0 ;; QUESTION SECTION: ;apt-get.com. IN A ;; ANSWER SECTION: apt-get.com. 86336 IN CNAME thie5ku9.dsgeneration.com. thie5ku9.dsgeneration.com. 60 IN A 208.73.211.242 thie5ku9.dsgeneration.com. 60 IN A 208.73.211.246 thie5ku9.dsgeneration.com. 60 IN A 208.73.211.166 thie5ku9.dsgeneration.com. 60 IN A 208.73.211.232 thie5ku9.dsgeneration.com. 60 IN A 208.73.211.161 thie5ku9.dsgeneration.com. 60 IN A 208.73.210.233 thie5ku9.dsgeneration.com. 60 IN A 208.73.211.186 thie5ku9.dsgeneration.com. 60 IN A 208.73.211.188 ;; Query time: 59 msec ;; SERVER: 127.0.0.1#53(127.0.0.1) ;; WHEN: Tue Jun 10 15:05:48 2014 ;; MSG SIZE rcvd: 193 $ dig apt-get.com ns ; <<>> DiG 9.8.1-P1 <<>> apt-get.com ns ;; global options: +cmd ;; Got answer: ;; ->>HEADER<<- opcode: QUERY, status: SERVFAIL, id: 43831 ;; flags: qr rd ra; QUERY: 1, ANSWER: 0, AUTHORITY: 0, ADDITIONAL: 0 ;; QUESTION SECTION: ;apt-get.com. IN NS ;; Query time: 26 msec ;; SERVER: 127.0.0.1#53(127.0.0.1) ;; WHEN: Tue Jun 10 15:12:37 2014 ;; MSG SIZE rcvd: 29 $ dig apt-get.com ns @b.gtld-servers.net ; <<>> DiG 9.8.1-P1 <<>> apt-get.com ns @b.gtld-servers.net ;; global options: +cmd ;; Got answer: ;; ->>HEADER<<- opcode: QUERY, status: NOERROR, id: 38228 ;; flags: qr rd; QUERY: 1, ANSWER: 0, AUTHORITY: 2, ADDITIONAL: 2 ;; WARNING: recursion requested but not available ;; QUESTION SECTION: ;apt-get.com. IN NS ;; AUTHORITY SECTION: apt-get.com. 172800 IN NS ns1.domainrecover.com. apt-get.com. 172800 IN NS ns2.domainrecover.com. ;; ADDITIONAL SECTION: ns1.domainrecover.com. 172800 IN A 66.45.232.66 ns2.domainrecover.com. 172800 IN A 65.23.159.179 ;; Query time: 70 msec ;; SERVER: 192.33.14.30#53(192.33.14.30) ;; WHEN: Tue Jun 10 15:07:05 2014 ;; MSG SIZE rcvd: 111 The domain does resolve. I get the following headers: GET / HTTP/1.1 User-Agent: Testing_Sniffer/4.15 Host: apt-get.com Accept: */* HTTP/1.0 200 (OK) Cache-Control: private, no-cache, must-revalidate Connection: Keep-Alive Pragma: no-cache Server: Oversee Turing v1.0.0 Content-Length: 1347 Content-Type: text/html Expires: Mon, 26 Jul 1997 05:00:00 GMT Keep-Alive: timeout=3, max=96 P3P: policyref="http://www.dsparking.com/w3c/p3p.xml", CP="NOI DSP COR ADMa OUR NOR STA" Set-Cookie: parkinglot=1; domain=.apt-get.com; path=/; expires=Wed, 11-Jun-2014 14:10:37 GMT <!DOCTYPE HTML PUBLIC "-//W3C//DTD HTML 4.01 Frameset//EN" "http://www.w3.org/TR/html4/frameset.dtd"> <!-- turing_cluster_prod --> <html> <head> <meta http-equiv="Content-Type" content="text/html; charset=utf-8" /> <title>apt-get.com</title> <meta name="keywords" content="apt-get.com" /> <meta name="description" content="apt-get.com" /> <meta name="robots" content="index, follow" /> <meta name="revisit-after" content="10" /> <meta name="viewport" content="width=device-width, initial-scale=1.0" /> <script type="text/javascript"> document.cookie = "jsc=1"; </script> </head> <frameset rows="100%,*" frameborder="no" border="0" framespacing="0"> <frame src="http://apt-get.com?epl=5PfLSSqWrYDAt-gbwMDK_rA3b1UJCYVTJHfxTzr9FTDQV84b6vAgVhU3FTeCRQNiuRNv79Ni0V3mkEVNRhpqo2gpMjp5iOIR1w2_EISPENaqzoXohVXl2QI3ryXlRCB4FaIIaxynnWXWY6QBgBgNiIZ6agD1NBoNGg0ajXpUCXUAIJDer78AAOB_AwAAQIDbCwAAe_NWlVlTJllBMTZoWkKPAAAA8A" name="apt-get.com"> </frameset> <noframes> <body><a href="http://apt-get.com?epl=5PfLSSqWrYDAt-gbwMDK_rA3b1UJCYVTJHfxTzr9FTDQV84b6vAgVhU3FTeCRQNiuRNv79Ni0V3mkEVNRhpqo2gpMjp5iOIR1w2_EISPENaqzoXohVXl2QI3ryXlRCB4FaIIaxynnWXWY6QBgBgNiIZ6agD1NBoNGg0ajXpUCXUAIJDer78AAOB_AwAAQIDbCwAAe_NWlVlTJllBMTZoWkKPAAAA8A">Click here to go to apt-get.com</a>.</body> </noframes> </html>

    Read the article

  • SQL Server and Hyper-V Dynamic Memory - Part 1

    - by SQLOS Team
    SQL and Dynamic Memory Blog Post Series   Hyper-V Dynamic Memory is a new feature in Windows Server 2008 R2 SP1 that allows the memory assigned to guest virtual machines to vary according to demand. Using this feature with SQL Server is supported, but how well does it work in an environment where available memory can vary dynamically, especially since SQL Server likes memory, and is not very eager to let go of it? The next three posts will look at this question in detail. In Part 1 Serdar Sutay, a program manager in the Windows Hyper-V team, introduces Dynamic Memory with an overview of the basic architecture, configuration and monitoring concepts. In subsequent parts we will look at SQL Server memory handling, and develop some guidelines on using SQL Server with Dynamic Memory.   Part 1: Dynamic Memory Introduction   In virtualized environments memory is often the bottleneck for reaching higher VM densities. In Windows Server 2008 R2 SP1 Hyper-V introduced a new feature “Dynamic Memory” to improve VM densities on Hyper-V hosts. Dynamic Memory increases the memory utilization in virtualized environments by enabling VM memory to be changed dynamically when the VM is running.   This brings up the question of how to utilize this feature with SQL Server VMs as SQL Server performance is very sensitive to the memory being used. In the next three posts we’ll discuss the internals of Dynamic Memory, SQL Server Memory Management and how to use Dynamic Memory with SQL Server VMs.   Memory Utilization Efficiency in Virtualized Environments   The primary reason memory is usually the bottleneck for higher VM densities is that users tend to be generous when assigning memory to their VMs. Here are some memory sizing practices we’ve heard from customers:   ·         I assign 4 GB of memory to my VMs. I don’t know if all of it is being used by the applications but no one complains. ·         I take the minimum system requirements and add 50% more. ·         I go with the recommendations provided by my software vendor.   In reality correctly sizing a virtual machine requires significant effort to monitor the memory usage of the applications. Since this is not done in most environments, VMs are usually over-provisioned in terms of memory. In other words, a SQL Server VM that is assigned 4 GB of memory may not need to use 4 GB.   How does Dynamic Memory help?   Dynamic Memory improves the memory utilization by removing the requirement to determine the memory need for an application. Hyper-V determines the memory needed by applications in the VM by evaluating the memory usage information in the guest with Dynamic Memory. VMs can start with a small amount of memory and they can be assigned more memory dynamically based on the workload of applications running inside.   Overview of Dynamic Memory Concepts   ·         Startup Memory: Startup Memory is the starting amount of memory when Dynamic Memory is enabled for a VM. Dynamic Memory will make sure that this amount of memory is always assigned to the VMs by default.   ·         Maximum Memory: Maximum Memory specifies the maximum amount of memory that a VM can grow to with Dynamic Memory. ·         Memory Demand: Memory Demand is the amount determined by Dynamic Memory as the memory needed by the applications in the VM. In Windows Server 2008 R2 SP1, this is equal to the total amount of committed memory of the VM. ·         Memory Buffer: Memory Buffer is the amount of memory assigned to the VMs in addition to their memory demand to satisfy immediate memory requirements and file cache needs.   Once Dynamic Memory is enabled for a VM, it will start with the “Startup Memory”. After the boot process Dynamic Memory will determine the “Memory Demand” of the VM. Based on this memory demand it will determine the amount of “Memory Buffer” that needs to be assigned to the VM. Dynamic Memory will assign the total of “Memory Demand” and “Memory Buffer” to the VM as long as this value is less than “Maximum Memory” and as long as physical memory is available on the host.   What happens when there is not enough physical memory available on the host?   Once there is not enough physical memory on the host to satisfy VM needs, Dynamic Memory will assign less than needed amount of memory to the VMs based on their importance. A concept known as “Memory Weight” is used to determine how much VMs should be penalized based on their needed amount of memory. “Memory Weight” is a configuration setting on the VM. It can be configured to be higher for the VMs with high performance requirements. Under high memory pressure on the host, the “Memory Weight” of the VMs are evaluated in a relative manner and the VMs with lower relative “Memory Weight” will be penalized more than the ones with higher “Memory Weight”.   Dynamic Memory Configuration   Based on these concepts “Startup Memory”, “Maximum Memory”, “Memory Buffer” and “Memory Weight” can be configured as shown below in Windows Server 2008 R2 SP1 Hyper-V Manager. Memory Demand is automatically calculated by Dynamic Memory once VMs start running.     Dynamic Memory Monitoring    In Windows Server 2008 R2 SP1, Hyper-V Manager displays the memory status of VMs in the following three columns:         ·         Assigned Memory represents the current physical memory assigned to the VM. In regular conditions this will be equal to the sum of “Memory Demand” and “Memory Buffer” assigned to the VM. When there is not enough memory on the host, this value can go below the Memory Demand determined for the VM. ·         Memory Demand displays the current “Memory Demand” determined for the VM. ·         Memory Status displays the current memory status of the VM. This column can represent three values for a VM: o   OK: In this condition the VM is assigned the total of Memory Demand and Memory Buffer it needs. o   Low: In this condition the VM is assigned all the Memory Demand and a certain percentage of the Memory Buffer it needs. o   Warning: In this condition the VM is assigned a lower memory than its Memory Demand. When VMs are running in this condition, it’s likely that they will exhibit performance problems due to internal paging happening in the VM.    So far so good! But how does it work with SQL Server?   SQL Server is aggressive in terms of memory usage for good reasons. This raises the question: How do SQL Server and Dynamic Memory work together? To understand the full story, we’ll first need to understand how SQL Server Memory Management works. This will be covered in our second post in “SQL and Dynamic Memory” series. Meanwhile if you want to dive deeper into Dynamic Memory you can check the below posts from the Windows Virtualization Team Blog:   http://blogs.technet.com/virtualization/archive/2010/03/18/dynamic-memory-coming-to-hyper-v.aspx   http://blogs.technet.com/virtualization/archive/2010/03/25/dynamic-memory-coming-to-hyper-v-part-2.aspx   http://blogs.technet.com/virtualization/archive/2010/04/07/dynamic-memory-coming-to-hyper-v-part-3.aspx   http://blogs.technet.com/b/virtualization/archive/2010/04/21/dynamic-memory-coming-to-hyper-v-part-4.aspx   http://blogs.technet.com/b/virtualization/archive/2010/05/20/dynamic-memory-coming-to-hyper-v-part-5.aspx   http://blogs.technet.com/b/virtualization/archive/2010/07/12/dynamic-memory-coming-to-hyper-v-part-6.aspx   - Serdar Sutay   Originally posted at http://blogs.msdn.com/b/sqlosteam/

    Read the article

  • SQL Server and Hyper-V Dynamic Memory - Part 1

    - by SQLOS Team
    SQL and Dynamic Memory Blog Post Series   Hyper-V Dynamic Memory is a new feature in Windows Server 2008 R2 SP1 that allows the memory assigned to guest virtual machines to vary according to demand. Using this feature with SQL Server is supported, but how well does it work in an environment where available memory can vary dynamically, especially since SQL Server likes memory, and is not very eager to let go of it? The next three posts will look at this question in detail. In Part 1 Serdar Sutay, a program manager in the Windows Hyper-V team, introduces Dynamic Memory with an overview of the basic architecture, configuration and monitoring concepts. In subsequent parts we will look at SQL Server memory handling, and develop some guidelines on using SQL Server with Dynamic Memory.   Part 1: Dynamic Memory Introduction   In virtualized environments memory is often the bottleneck for reaching higher VM densities. In Windows Server 2008 R2 SP1 Hyper-V introduced a new feature “Dynamic Memory” to improve VM densities on Hyper-V hosts. Dynamic Memory increases the memory utilization in virtualized environments by enabling VM memory to be changed dynamically when the VM is running.   This brings up the question of how to utilize this feature with SQL Server VMs as SQL Server performance is very sensitive to the memory being used. In the next three posts we’ll discuss the internals of Dynamic Memory, SQL Server Memory Management and how to use Dynamic Memory with SQL Server VMs.   Memory Utilization Efficiency in Virtualized Environments   The primary reason memory is usually the bottleneck for higher VM densities is that users tend to be generous when assigning memory to their VMs. Here are some memory sizing practices we’ve heard from customers:   ·         I assign 4 GB of memory to my VMs. I don’t know if all of it is being used by the applications but no one complains. ·         I take the minimum system requirements and add 50% more. ·         I go with the recommendations provided by my software vendor.   In reality correctly sizing a virtual machine requires significant effort to monitor the memory usage of the applications. Since this is not done in most environments, VMs are usually over-provisioned in terms of memory. In other words, a SQL Server VM that is assigned 4 GB of memory may not need to use 4 GB.   How does Dynamic Memory help?   Dynamic Memory improves the memory utilization by removing the requirement to determine the memory need for an application. Hyper-V determines the memory needed by applications in the VM by evaluating the memory usage information in the guest with Dynamic Memory. VMs can start with a small amount of memory and they can be assigned more memory dynamically based on the workload of applications running inside.   Overview of Dynamic Memory Concepts   ·         Startup Memory: Startup Memory is the starting amount of memory when Dynamic Memory is enabled for a VM. Dynamic Memory will make sure that this amount of memory is always assigned to the VMs by default.   ·         Maximum Memory: Maximum Memory specifies the maximum amount of memory that a VM can grow to with Dynamic Memory. ·         Memory Demand: Memory Demand is the amount determined by Dynamic Memory as the memory needed by the applications in the VM. In Windows Server 2008 R2 SP1, this is equal to the total amount of committed memory of the VM. ·         Memory Buffer: Memory Buffer is the amount of memory assigned to the VMs in addition to their memory demand to satisfy immediate memory requirements and file cache needs.   Once Dynamic Memory is enabled for a VM, it will start with the “Startup Memory”. After the boot process Dynamic Memory will determine the “Memory Demand” of the VM. Based on this memory demand it will determine the amount of “Memory Buffer” that needs to be assigned to the VM. Dynamic Memory will assign the total of “Memory Demand” and “Memory Buffer” to the VM as long as this value is less than “Maximum Memory” and as long as physical memory is available on the host.   What happens when there is not enough physical memory available on the host?   Once there is not enough physical memory on the host to satisfy VM needs, Dynamic Memory will assign less than needed amount of memory to the VMs based on their importance. A concept known as “Memory Weight” is used to determine how much VMs should be penalized based on their needed amount of memory. “Memory Weight” is a configuration setting on the VM. It can be configured to be higher for the VMs with high performance requirements. Under high memory pressure on the host, the “Memory Weight” of the VMs are evaluated in a relative manner and the VMs with lower relative “Memory Weight” will be penalized more than the ones with higher “Memory Weight”.   Dynamic Memory Configuration   Based on these concepts “Startup Memory”, “Maximum Memory”, “Memory Buffer” and “Memory Weight” can be configured as shown below in Windows Server 2008 R2 SP1 Hyper-V Manager. Memory Demand is automatically calculated by Dynamic Memory once VMs start running.     Dynamic Memory Monitoring    In Windows Server 2008 R2 SP1, Hyper-V Manager displays the memory status of VMs in the following three columns:         ·         Assigned Memory represents the current physical memory assigned to the VM. In regular conditions this will be equal to the sum of “Memory Demand” and “Memory Buffer” assigned to the VM. When there is not enough memory on the host, this value can go below the Memory Demand determined for the VM. ·         Memory Demand displays the current “Memory Demand” determined for the VM. ·         Memory Status displays the current memory status of the VM. This column can represent three values for a VM: o   OK: In this condition the VM is assigned the total of Memory Demand and Memory Buffer it needs. o   Low: In this condition the VM is assigned all the Memory Demand and a certain percentage of the Memory Buffer it needs. o   Warning: In this condition the VM is assigned a lower memory than its Memory Demand. When VMs are running in this condition, it’s likely that they will exhibit performance problems due to internal paging happening in the VM.    So far so good! But how does it work with SQL Server?   SQL Server is aggressive in terms of memory usage for good reasons. This raises the question: How do SQL Server and Dynamic Memory work together? To understand the full story, we’ll first need to understand how SQL Server Memory Management works. This will be covered in our second post in “SQL and Dynamic Memory” series. Meanwhile if you want to dive deeper into Dynamic Memory you can check the below posts from the Windows Virtualization Team Blog:   http://blogs.technet.com/virtualization/archive/2010/03/18/dynamic-memory-coming-to-hyper-v.aspx   http://blogs.technet.com/virtualization/archive/2010/03/25/dynamic-memory-coming-to-hyper-v-part-2.aspx   http://blogs.technet.com/virtualization/archive/2010/04/07/dynamic-memory-coming-to-hyper-v-part-3.aspx   http://blogs.technet.com/b/virtualization/archive/2010/04/21/dynamic-memory-coming-to-hyper-v-part-4.aspx   http://blogs.technet.com/b/virtualization/archive/2010/05/20/dynamic-memory-coming-to-hyper-v-part-5.aspx   http://blogs.technet.com/b/virtualization/archive/2010/07/12/dynamic-memory-coming-to-hyper-v-part-6.aspx   - Serdar Sutay   Originally posted at http://blogs.msdn.com/b/sqlosteam/

    Read the article

  • NRF Online Merchandising Workshop: Where Online Retailers Are Focusing for Holiday and Beyond

    - by Rose Spicer-Oracle
    0 0 1 1204 6863 Oracle Corporation 57 16 8051 14.0 Normal 0 false false false EN-US JA X-NONE /* Style Definitions */ table.MsoNormalTable {mso-style-name:"Table Normal"; mso-tstyle-rowband-size:0; mso-tstyle-colband-size:0; mso-style-noshow:yes; mso-style-priority:99; mso-style-parent:""; mso-padding-alt:0in 5.4pt 0in 5.4pt; mso-para-margin:0in; mso-para-margin-bottom:.0001pt; mso-pagination:widow-orphan; font-size:12.0pt; font-family:Cambria; mso-ascii-font-family:Cambria; mso-ascii-theme-font:minor-latin; mso-hansi-font-family:Cambria; mso-hansi-theme-font:minor-latin;} Last month we attended the NRF Online Merchandising Workshop in LA, and it was a great opportunity to catch up with our customers, meet new retailers, and hear some great presentations from VF Corporation, Zazzle, Julep Beauty, Backcountry, eBags and more. The one-on-one conversations with Merchants and the keynote presentations carry the same themes across companies of all sizes and across verticals. With only 125 days left (and counting) until Black Friday, these conversations provided some great insight in to what’s top of mind for retailers during the most stressful time of their year, and a sneak peek in to what they will deliver this holiday season.  Some of the most popular topics were: When to start promoting for holiday: seems like a funny conversation to have in July, but a number of retailers said they already had their holiday shopping gift guides live on their site, and it was attracting a significant portion of their onsite traffic. When it comes to timing, most retailers were questioning when to begin their holiday promotions -- carefully balancing when to release pricing and specials, and knowing that customers are holding out for last-minute deals and price drops. Many retailers noted the frustrations around transparent pricing by Amazon and a few other mega-retailers last year, publishing their “lowest prices of the season” as early as October – ensuring shoppers that those prices were the best they could get all season long. Many retailers felt their hands were forced to drop prices. Others kept their set pricing with negative customer reaction, causing some to miss their holiday goals. The pressure is on, and most retailers identified November 1 as their target start date for the holiday promotions blitz. Some are even waiting for the big guys to release their “lowest prices of the season” guides and will then follow suit.      Attribution is tough – and a huge focus: understanding the path to conversion is a tough nut to crack, especially in the new omnichannel world where consumers use multiple touchpoints to make a single purchase, and internal management wants to know hard data. This has lead many retailers to invest in attribution; carefully tracking their online marketing efforts to determine what gets “credit” for the sale, instead of giving credit to the “last click.” Retailers noted that it is very difficult to determine the numbers when online and offline worlds collide – like when a shopper uses digital channels for research and then makes a purchase in a store. As one of the presenters from The North Face mentioned in her keynote, a key to enabling better customer service and satisfaction when it comes to converged online and offline sales is training the in-store staff, and creating a culture where it eventually “doesn’t matter what group gets the credit” if they all add to the sale. No doubt, the area of attribution will be a big area of retail investment in the coming years.      How to plan for the converged world: planning to ensure inventory gets where it needs to be was another concern. In conversations with retailers, we advised them to analyze customer patterns: where shoppers purchase items, where the items were sourced from and even where items are returned. This analysis is very valuable in determining inventory plans. From there, retailers can more accurately plan and allocate inventory to support both the online and offline customer behavior. As we head into the holiday season, the need for accurate enterprise-wide inventory visibility, and providing that information to associates, is even more critical to the brand-wide customer experience.       Improving the search / navigation / usability of the site(s): Aside from some of the big ideas and standard holiday pricing pressure, most conversations we had centered around continuing to improve the basics of the site. Reinvesting in search and navigation came up time and time again (FitForCommerce blogged about what a big topic it was at the event as well). Obviously getting shoppers on their path quickly and allowing them to find what they need fast is critical, but it was definitely interesting to hear just how much effort is still going in to honing the search and navigation experience. Adding new elements to search and navigation like typeahed, inventive navigation refinements, and new navigation categories like gift guides, specialized boutiques and flash sales were top of mind, in addition to searchandising and making search-driven product recommendations. (Oracle can help!)       Reducing cart abandonment: always a hot topic that is top of mind for every online retailer. Getting shoppers to the cart is often less then half the battle; getting them to click “buy” and complete the transaction is much more difficult. While retailers carefully study the checkout process and where shoppers tend to bounce, they know that how they design their checkout page is critical. We’re all online shoppers in our personal lives and we know how frustrating it can be when total prices are not transparent (i.e. shipping, processing, taxes is not included until the very last possible screen before clicking that buy button). Online retailers are struggling with where in the checkout process to surface the total price to be charged to reduce cart abandonment, while not showing the total figure too early in the process that it keeps shoppers from getting to checkout altogether. Recent research shows that providing total pricing prior to the checkout process dramatically reduces cart abandonment – as it serves as a filter to those shopping within a specific price band. Much of the cart abandonment discussion leads us to…       The free shipping / free returns question: it’s no secret that because of Amazon and programs like Prime, consumers expect free shipping, much to the chagrin of the smaller retailer. The reality is that if you’re not a mega-retailer, shipping is an expensive part of doing business that doesn’t allow most retailers to keep their prices low and offer free shipping. This has many retailers venturing out on the “free returns” path, especially in apparel. A number of retailers we spoke with are testing a flat rate shipping fee with free returns to see if they can crack the price threshold where shoppers are willing to pay for shipping with an added service. But, free shipping remains king.      Social ads and retargeting: they are working, but do they turn off consumers? That’s the big question. Every retailer we spoke with during a roundtable on the topic said that social ads and retargeting (where that pair of boots you’re been eyeing on a site magically follows you around the Internet) work and are meeting campaign goals. The larger question many retailers are asking is if this type of tactic is turning off a large number of shoppers, even if these campaigns are meeting their early goals. Retailers also mentioned that Facebook ads are working very well for them, especially when it comes to new customer acquisition, serving as a complimentary a channel to SEO when it comes to engaging new customers. While there are always new things to experiment with in retail, standard challenges are top of mind as retailers scramble to get ready for holiday. It will undoubtedly be another record-breaking online shopping season, but as retailers get more and more advanced with each Black Friday, expect some exciting things. This excitement needs to be backed by sound solutions and optimized operations. Then again, consumers are expecting more than ever, so I don’t doubt that retailers are already thinking about the possibilities of holiday 2015… and beyond. Customers who read this article, also found value in the following stories: Personalization for Retail: http://blogs.oracle.com/retail/entry/personalization_for_retailShop Direct User Experience Focus Drives Sales:https://blogs.oracle.com/retail/entry/shop_direct_user_experience_focusMaking Waves: Australian Online Retailer SurfStitch: https://blogs.oracle.com/oracleretail/entry/surf_stitchWhat’s new in Oracle Commerce v11.1 for RetailWhat the Content+Commerce Equation is Missing

    Read the article

  • exporting bind and keyframe bone poses from blender to use in OpenGL

    - by SaldaVonSchwartz
    I'm having a hard time trying to understand how exactly Blender's concept of bone transforms maps to the usual math of skinning (which I'm implementing in an OpenGL-based engine of sorts). Or I'm missing out something in the math.. It's gonna be long, but here's as much background as I can think of. First, a few notes and assumptions: I'm using column-major order and multiply from right to left. So for instance, vertex v transformed by matrix A and then further transformed by matrix B would be: v' = BAv. This also means whenever I export a matrix from blender through python, I export it (in text format) in 4 lines, each representing a column. This is so I can then I can read them back into my engine like this: if (fscanf(fileHandle, "%f %f %f %f", &skeleton.joints[currentJointIndex].inverseBindTransform.m[0], &skeleton.joints[currentJointIndex].inverseBindTransform.m[1], &skeleton.joints[currentJointIndex].inverseBindTransform.m[2], &skeleton.joints[currentJointIndex].inverseBindTransform.m[3])) { if (fscanf(fileHandle, "%f %f %f %f", &skeleton.joints[currentJointIndex].inverseBindTransform.m[4], &skeleton.joints[currentJointIndex].inverseBindTransform.m[5], &skeleton.joints[currentJointIndex].inverseBindTransform.m[6], &skeleton.joints[currentJointIndex].inverseBindTransform.m[7])) { if (fscanf(fileHandle, "%f %f %f %f", &skeleton.joints[currentJointIndex].inverseBindTransform.m[8], &skeleton.joints[currentJointIndex].inverseBindTransform.m[9], &skeleton.joints[currentJointIndex].inverseBindTransform.m[10], &skeleton.joints[currentJointIndex].inverseBindTransform.m[11])) { if (fscanf(fileHandle, "%f %f %f %f", &skeleton.joints[currentJointIndex].inverseBindTransform.m[12], &skeleton.joints[currentJointIndex].inverseBindTransform.m[13], &skeleton.joints[currentJointIndex].inverseBindTransform.m[14], &skeleton.joints[currentJointIndex].inverseBindTransform.m[15])) { I'm simplifying the code I show because otherwise it would make things unnecessarily harder (in the context of my question) to explain / follow. Please refrain from making remarks related to optimizations. This is not final code. Having said that, if I understand correctly, the basic idea of skinning/animation is: I have a a mesh made up of vertices I have the mesh model-world transform W I have my joints, which are really just transforms from each joint's space to its parent's space. I'll call these transforms Bj meaning matrix which takes from joint j's bind pose to joint j-1's bind pose. For each of these, I actually import their inverse to the engine, Bj^-1. I have keyframes each containing a set of current poses Cj for each joint J. These are initially imported to my engine in TQS format but after (S)LERPING them I compose them into Cj matrices which are equivalent to the Bjs (not the Bj^-1 ones) only that for the current spacial configurations of each joint at that frame. Given the above, the "skeletal animation algorithm is" On each frame: check how much time has elpased and compute the resulting current time in the animation, from 0 meaning frame 0 to 1, meaning the end of the animation. (Oh and I'm looping forever so the time is mod(total duration)) for each joint: 1 -calculate its world inverse bind pose, that is Bj_w^-1 = Bj^-1 Bj-1^-1 ... B0^-1 2 -use the current animation time to LERP the componets of the TQS and come up with an interpolated current pose matrix Cj which should transform from the joints current configuration space to world space. Similar to what I did to get the world version of the inverse bind poses, I come up with the joint's world current pose, Cj_w = C0 C1 ... Cj 3 -now that I have world versions of Bj and Cj, I store this joint's world- skinning matrix K_wj = Cj_w Bj_w^-1. The above is roughly implemented like so: - (void)update:(NSTimeInterval)elapsedTime { static double time = 0; time = fmod((time + elapsedTime),1.); uint16_t LERPKeyframeNumber = 60 * time; uint16_t lkeyframeNumber = 0; uint16_t lkeyframeIndex = 0; uint16_t rkeyframeNumber = 0; uint16_t rkeyframeIndex = 0; for (int i = 0; i < aClip.keyframesCount; i++) { uint16_t keyframeNumber = aClip.keyframes[i].number; if (keyframeNumber <= LERPKeyframeNumber) { lkeyframeIndex = i; lkeyframeNumber = keyframeNumber; } else { rkeyframeIndex = i; rkeyframeNumber = keyframeNumber; break; } } double lTime = lkeyframeNumber / 60.; double rTime = rkeyframeNumber / 60.; double blendFactor = (time - lTime) / (rTime - lTime); GLKMatrix4 bindPosePalette[aSkeleton.jointsCount]; GLKMatrix4 currentPosePalette[aSkeleton.jointsCount]; for (int i = 0; i < aSkeleton.jointsCount; i++) { F3DETQSType& lPose = aClip.keyframes[lkeyframeIndex].skeletonPose.jointPoses[i]; F3DETQSType& rPose = aClip.keyframes[rkeyframeIndex].skeletonPose.jointPoses[i]; GLKVector3 LERPTranslation = GLKVector3Lerp(lPose.t, rPose.t, blendFactor); GLKQuaternion SLERPRotation = GLKQuaternionSlerp(lPose.q, rPose.q, blendFactor); GLKVector3 LERPScaling = GLKVector3Lerp(lPose.s, rPose.s, blendFactor); GLKMatrix4 currentTransform = GLKMatrix4MakeWithQuaternion(SLERPRotation); currentTransform = GLKMatrix4Multiply(currentTransform, GLKMatrix4MakeTranslation(LERPTranslation.x, LERPTranslation.y, LERPTranslation.z)); currentTransform = GLKMatrix4Multiply(currentTransform, GLKMatrix4MakeScale(LERPScaling.x, LERPScaling.y, LERPScaling.z)); if (aSkeleton.joints[i].parentIndex == -1) { bindPosePalette[i] = aSkeleton.joints[i].inverseBindTransform; currentPosePalette[i] = currentTransform; } else { bindPosePalette[i] = GLKMatrix4Multiply(aSkeleton.joints[i].inverseBindTransform, bindPosePalette[aSkeleton.joints[i].parentIndex]); currentPosePalette[i] = GLKMatrix4Multiply(currentPosePalette[aSkeleton.joints[i].parentIndex], currentTransform); } aSkeleton.skinningPalette[i] = GLKMatrix4Multiply(currentPosePalette[i], bindPosePalette[i]); } } At this point, I should have my skinning palette. So on each frame in my vertex shader, I do: uniform mat4 modelMatrix; uniform mat4 projectionMatrix; uniform mat3 normalMatrix; uniform mat4 skinningPalette[6]; attribute vec4 position; attribute vec3 normal; attribute vec2 tCoordinates; attribute vec4 jointsWeights; attribute vec4 jointsIndices; varying highp vec2 tCoordinatesVarying; varying highp float lIntensity; void main() { vec3 eyeNormal = normalize(normalMatrix * normal); vec3 lightPosition = vec3(0., 0., 2.); lIntensity = max(0.0, dot(eyeNormal, normalize(lightPosition))); tCoordinatesVarying = tCoordinates; vec4 skinnedVertexPosition = vec4(0.); for (int i = 0; i < 4; i++) { skinnedVertexPosition += jointsWeights[i] * skinningPalette[int(jointsIndices[i])] * position; } gl_Position = projectionMatrix * modelMatrix * skinnedVertexPosition; } The result: The mesh parts that are supposed to animate do animate and follow the expected motion, however, the rotations are messed up in terms of orientations. That is, the mesh is not translated somewhere else or scaled in any way, but the orientations of rotations seem to be off. So a few observations: In the above shader notice I actually did not multiply the vertices by the mesh modelMatrix (the one which would take them to model or world or global space, whichever you prefer, since there is no parent to the mesh itself other than "the world") until after skinning. This is contrary to what I implied in the theory: if my skinning matrix takes vertices from model to joint and back to model space, I'd think the vertices should already be premultiplied by the mesh transform. But if I do so, I just get a black screen. As far as exporting the joints from Blender, my python script exports for each armature bone in bind pose, it's matrix in this way: def DFSJointTraversal(file, skeleton, jointList): for joint in jointList: poseJoint = skeleton.pose.bones[joint.name] jointTransform = poseJoint.matrix.inverted() file.write('Joint ' + joint.name + ' Transform {\n') for col in jointTransform.col: file.write('{:9f} {:9f} {:9f} {:9f}\n'.format(col[0], col[1], col[2], col[3])) DFSJointTraversal(file, skeleton, joint.children) file.write('}\n') And for current / keyframe poses (assuming I'm in the right keyframe): def exportAnimations(filepath): # Only one skeleton per scene objList = [object for object in bpy.context.scene.objects if object.type == 'ARMATURE'] if len(objList) == 0: return elif len(objList) > 1: return #raise exception? dialog box? skeleton = objList[0] jointNames = [bone.name for bone in skeleton.data.bones] for action in bpy.data.actions: # One animation clip per action in Blender, named as the action animationClipFilePath = filepath[0 : filepath.rindex('/') + 1] + action.name + ".aClip" file = open(animationClipFilePath, 'w') file.write('target skeleton: ' + skeleton.name + '\n') file.write('joints count: {:d}'.format(len(jointNames)) + '\n') skeleton.animation_data.action = action keyframeNum = max([len(fcurve.keyframe_points) for fcurve in action.fcurves]) keyframes = [] for fcurve in action.fcurves: for keyframe in fcurve.keyframe_points: keyframes.append(keyframe.co[0]) keyframes = set(keyframes) keyframes = [kf for kf in keyframes] keyframes.sort() file.write('keyframes count: {:d}'.format(len(keyframes)) + '\n') for kfIndex in keyframes: bpy.context.scene.frame_set(kfIndex) file.write('keyframe: {:d}\n'.format(int(kfIndex))) for i in range(0, len(skeleton.data.bones)): file.write('joint: {:d}\n'.format(i)) joint = skeleton.pose.bones[i] jointCurrentPoseTransform = joint.matrix translationV = jointCurrentPoseTransform.to_translation() rotationQ = jointCurrentPoseTransform.to_3x3().to_quaternion() scaleV = jointCurrentPoseTransform.to_scale() file.write('T {:9f} {:9f} {:9f}\n'.format(translationV[0], translationV[1], translationV[2])) file.write('Q {:9f} {:9f} {:9f} {:9f}\n'.format(rotationQ[1], rotationQ[2], rotationQ[3], rotationQ[0])) file.write('S {:9f} {:9f} {:9f}\n'.format(scaleV[0], scaleV[1], scaleV[2])) file.write('\n') file.close() Which I believe follow the theory explained at the beginning of my question. But then I checked out Blender's directX .x exporter for reference.. and what threw me off was that in the .x script they are exporting bind poses like so (transcribed using the same variable names I used so you can compare): if joint.parent: jointTransform = poseJoint.parent.matrix.inverted() else: jointTransform = Matrix() jointTransform *= poseJoint.matrix and exporting current keyframe poses like this: if joint.parent: jointCurrentPoseTransform = joint.parent.matrix.inverted() else: jointCurrentPoseTransform = Matrix() jointCurrentPoseTransform *= joint.matrix why are they using the parent's transform instead of the joint in question's? isn't the join transform assumed to exist in the context of a parent transform since after all it transforms from this joint's space to its parent's? Why are they concatenating in the same order for both bind poses and keyframe poses? If these two are then supposed to be concatenated with each other to cancel out the change of basis? Anyway, any ideas are appreciated.

    Read the article

  • Hello Operator, My Switch Is Bored

    - by Paul White
    This is a post for T-SQL Tuesday #43 hosted by my good friend Rob Farley. The topic this month is Plan Operators. I haven’t taken part in T-SQL Tuesday before, but I do like to write about execution plans, so this seemed like a good time to start. This post is in two parts. The first part is primarily an excuse to use a pretty bad play on words in the title of this blog post (if you’re too young to know what a telephone operator or a switchboard is, I hate you). The second part of the post looks at an invisible query plan operator (so to speak). 1. My Switch Is Bored Allow me to present the rare and interesting execution plan operator, Switch: Books Online has this to say about Switch: Following that description, I had a go at producing a Fast Forward Cursor plan that used the TOP operator, but had no luck. That may be due to my lack of skill with cursors, I’m not too sure. The only application of Switch in SQL Server 2012 that I am familiar with requires a local partitioned view: CREATE TABLE dbo.T1 (c1 int NOT NULL CHECK (c1 BETWEEN 00 AND 24)); CREATE TABLE dbo.T2 (c1 int NOT NULL CHECK (c1 BETWEEN 25 AND 49)); CREATE TABLE dbo.T3 (c1 int NOT NULL CHECK (c1 BETWEEN 50 AND 74)); CREATE TABLE dbo.T4 (c1 int NOT NULL CHECK (c1 BETWEEN 75 AND 99)); GO CREATE VIEW V1 AS SELECT c1 FROM dbo.T1 UNION ALL SELECT c1 FROM dbo.T2 UNION ALL SELECT c1 FROM dbo.T3 UNION ALL SELECT c1 FROM dbo.T4; Not only that, but it needs an updatable local partitioned view. We’ll need some primary keys to meet that requirement: ALTER TABLE dbo.T1 ADD CONSTRAINT PK_T1 PRIMARY KEY (c1);   ALTER TABLE dbo.T2 ADD CONSTRAINT PK_T2 PRIMARY KEY (c1);   ALTER TABLE dbo.T3 ADD CONSTRAINT PK_T3 PRIMARY KEY (c1);   ALTER TABLE dbo.T4 ADD CONSTRAINT PK_T4 PRIMARY KEY (c1); We also need an INSERT statement that references the view. Even more specifically, to see a Switch operator, we need to perform a single-row insert (multi-row inserts use a different plan shape): INSERT dbo.V1 (c1) VALUES (1); And now…the execution plan: The Constant Scan manufactures a single row with no columns. The Compute Scalar works out which partition of the view the new value should go in. The Assert checks that the computed partition number is not null (if it is, an error is returned). The Nested Loops Join executes exactly once, with the partition id as an outer reference (correlated parameter). The Switch operator checks the value of the parameter and executes the corresponding input only. If the partition id is 0, the uppermost Clustered Index Insert is executed, adding a row to table T1. If the partition id is 1, the next lower Clustered Index Insert is executed, adding a row to table T2…and so on. In case you were wondering, here’s a query and execution plan for a multi-row insert to the view: INSERT dbo.V1 (c1) VALUES (1), (2); Yuck! An Eager Table Spool and four Filters! I prefer the Switch plan. My guess is that almost all the old strategies that used a Switch operator have been replaced over time, using things like a regular Concatenation Union All combined with Start-Up Filters on its inputs. Other new (relative to the Switch operator) features like table partitioning have specific execution plan support that doesn’t need the Switch operator either. This feels like a bit of a shame, but perhaps it is just nostalgia on my part, it’s hard to know. Please do let me know if you encounter a query that can still use the Switch operator in 2012 – it must be very bored if this is the only possible modern usage! 2. Invisible Plan Operators The second part of this post uses an example based on a question Dave Ballantyne asked using the SQL Sentry Plan Explorer plan upload facility. If you haven’t tried that yet, make sure you’re on the latest version of the (free) Plan Explorer software, and then click the Post to SQLPerformance.com button. That will create a site question with the query plan attached (which can be anonymized if the plan contains sensitive information). Aaron Bertrand and I keep a close eye on questions there, so if you have ever wanted to ask a query plan question of either of us, that’s a good way to do it. The problem The issue I want to talk about revolves around a query issued against a calendar table. The script below creates a simplified version and adds 100 years of per-day information to it: USE tempdb; GO CREATE TABLE dbo.Calendar ( dt date NOT NULL, isWeekday bit NOT NULL, theYear smallint NOT NULL,   CONSTRAINT PK__dbo_Calendar_dt PRIMARY KEY CLUSTERED (dt) ); GO -- Monday is the first day of the week for me SET DATEFIRST 1;   -- Add 100 years of data INSERT dbo.Calendar WITH (TABLOCKX) (dt, isWeekday, theYear) SELECT CA.dt, isWeekday = CASE WHEN DATEPART(WEEKDAY, CA.dt) IN (6, 7) THEN 0 ELSE 1 END, theYear = YEAR(CA.dt) FROM Sandpit.dbo.Numbers AS N CROSS APPLY ( VALUES (DATEADD(DAY, N.n - 1, CONVERT(date, '01 Jan 2000', 113))) ) AS CA (dt) WHERE N.n BETWEEN 1 AND 36525; The following query counts the number of weekend days in 2013: SELECT Days = COUNT_BIG(*) FROM dbo.Calendar AS C WHERE theYear = 2013 AND isWeekday = 0; It returns the correct result (104) using the following execution plan: The query optimizer has managed to estimate the number of rows returned from the table exactly, based purely on the default statistics created separately on the two columns referenced in the query’s WHERE clause. (Well, almost exactly, the unrounded estimate is 104.289 rows.) There is already an invisible operator in this query plan – a Filter operator used to apply the WHERE clause predicates. We can see it by re-running the query with the enormously useful (but undocumented) trace flag 9130 enabled: Now we can see the full picture. The whole table is scanned, returning all 36,525 rows, before the Filter narrows that down to just the 104 we want. Without the trace flag, the Filter is incorporated in the Clustered Index Scan as a residual predicate. It is a little bit more efficient than using a separate operator, but residual predicates are still something you will want to avoid where possible. The estimates are still spot on though: Anyway, looking to improve the performance of this query, Dave added the following filtered index to the Calendar table: CREATE NONCLUSTERED INDEX Weekends ON dbo.Calendar(theYear) WHERE isWeekday = 0; The original query now produces a much more efficient plan: Unfortunately, the estimated number of rows produced by the seek is now wrong (365 instead of 104): What’s going on? The estimate was spot on before we added the index! Explanation You might want to grab a coffee for this bit. Using another trace flag or two (8606 and 8612) we can see that the cardinality estimates were exactly right initially: The highlighted information shows the initial cardinality estimates for the base table (36,525 rows), the result of applying the two relational selects in our WHERE clause (104 rows), and after performing the COUNT_BIG(*) group by aggregate (1 row). All of these are correct, but that was before cost-based optimization got involved :) Cost-based optimization When cost-based optimization starts up, the logical tree above is copied into a structure (the ‘memo’) that has one group per logical operation (roughly speaking). The logical read of the base table (LogOp_Get) ends up in group 7; the two predicates (LogOp_Select) end up in group 8 (with the details of the selections in subgroups 0-6). These two groups still have the correct cardinalities as trace flag 8608 output (initial memo contents) shows: During cost-based optimization, a rule called SelToIdxStrategy runs on group 8. It’s job is to match logical selections to indexable expressions (SARGs). It successfully matches the selections (theYear = 2013, is Weekday = 0) to the filtered index, and writes a new alternative into the memo structure. The new alternative is entered into group 8 as option 1 (option 0 was the original LogOp_Select): The new alternative is to do nothing (PhyOp_NOP = no operation), but to instead follow the new logical instructions listed below the NOP. The LogOp_GetIdx (full read of an index) goes into group 21, and the LogOp_SelectIdx (selection on an index) is placed in group 22, operating on the result of group 21. The definition of the comparison ‘the Year = 2013’ (ScaOp_Comp downwards) was already present in the memo starting at group 2, so no new memo groups are created for that. New Cardinality Estimates The new memo groups require two new cardinality estimates to be derived. First, LogOp_Idx (full read of the index) gets a predicted cardinality of 10,436. This number comes from the filtered index statistics: DBCC SHOW_STATISTICS (Calendar, Weekends) WITH STAT_HEADER; The second new cardinality derivation is for the LogOp_SelectIdx applying the predicate (theYear = 2013). To get a number for this, the cardinality estimator uses statistics for the column ‘theYear’, producing an estimate of 365 rows (there are 365 days in 2013!): DBCC SHOW_STATISTICS (Calendar, theYear) WITH HISTOGRAM; This is where the mistake happens. Cardinality estimation should have used the filtered index statistics here, to get an estimate of 104 rows: DBCC SHOW_STATISTICS (Calendar, Weekends) WITH HISTOGRAM; Unfortunately, the logic has lost sight of the link between the read of the filtered index (LogOp_GetIdx) in group 22, and the selection on that index (LogOp_SelectIdx) that it is deriving a cardinality estimate for, in group 21. The correct cardinality estimate (104 rows) is still present in the memo, attached to group 8, but that group now has a PhyOp_NOP implementation. Skipping over the rest of cost-based optimization (in a belated attempt at brevity) we can see the optimizer’s final output using trace flag 8607: This output shows the (incorrect, but understandable) 365 row estimate for the index range operation, and the correct 104 estimate still attached to its PhyOp_NOP. This tree still has to go through a few post-optimizer rewrites and ‘copy out’ from the memo structure into a tree suitable for the execution engine. One step in this process removes PhyOp_NOP, discarding its 104-row cardinality estimate as it does so. To finish this section on a more positive note, consider what happens if we add an OVER clause to the query aggregate. This isn’t intended to be a ‘fix’ of any sort, I just want to show you that the 104 estimate can survive and be used if later cardinality estimation needs it: SELECT Days = COUNT_BIG(*) OVER () FROM dbo.Calendar AS C WHERE theYear = 2013 AND isWeekday = 0; The estimated execution plan is: Note the 365 estimate at the Index Seek, but the 104 lives again at the Segment! We can imagine the lost predicate ‘isWeekday = 0’ as sitting between the seek and the segment in an invisible Filter operator that drops the estimate from 365 to 104. Even though the NOP group is removed after optimization (so we don’t see it in the execution plan) bear in mind that all cost-based choices were made with the 104-row memo group present, so although things look a bit odd, it shouldn’t affect the optimizer’s plan selection. I should also mention that we can work around the estimation issue by including the index’s filtering columns in the index key: CREATE NONCLUSTERED INDEX Weekends ON dbo.Calendar(theYear, isWeekday) WHERE isWeekday = 0 WITH (DROP_EXISTING = ON); There are some downsides to doing this, including that changes to the isWeekday column may now require Halloween Protection, but that is unlikely to be a big problem for a static calendar table ;)  With the updated index in place, the original query produces an execution plan with the correct cardinality estimation showing at the Index Seek: That’s all for today, remember to let me know about any Switch plans you come across on a modern instance of SQL Server! Finally, here are some other posts of mine that cover other plan operators: Segment and Sequence Project Common Subexpression Spools Why Plan Operators Run Backwards Row Goals and the Top Operator Hash Match Flow Distinct Top N Sort Index Spools and Page Splits Singleton and Range Seeks Bitmaps Hash Join Performance Compute Scalar © 2013 Paul White – All Rights Reserved Twitter: @SQL_Kiwi

    Read the article

  • Who writes the words? A rant with graphs.

    - by Roger Hart
    If you read my rant, you'll know that I'm getting a bit of a bee in my bonnet about user interface text. But rather than just yelling about the way the world should be (short version: no UI text would suck), it seemed prudent to actually gather some data. Rachel Potts has made an excellent first foray, by conducting a series of interviews across organizations about how they write user interface text. You can read Rachel's write up here. She presents the facts as she found them, and doesn't editorialise. The result is insightful, but impartial isn't really my style. So here's a rant with graphs. My method, and how it sucked I sent out a short survey. Survey design is one of my hobby-horses, and since some smartarse in the comments will mention it if I don't, I'll step up and confess: I did not design this one well. It was potentially ambiguous, implicitly excluded people, and since I only really advertised it on Twitter and a couple of mailing lists the sample will be chock full of biases. Regardless, these were the questions: What do you do? Select the option that best describes your role What kind of software does your organization make? (optional) In your organization, who writes the text on your software user interfaces? (for example: button names, static text, tooltips, and so on) Tick all that apply. In your organization who is responsible for user interface text? Who "owns" it? The most glaring issue (apart from question 3 being a bit broken) was that I didn't make it clear that I was asking about applications. Desktop, mobile, or web, I wouldn't have minded. In fact, it might have been interesting to categorize and compare. But a few respondents commented on the seeming lack of relevance, since they didn't really make software. There were some other issues too. It wasn't the best survey. So, you know, pinch of salt time with what follows. Despite this, there were 100 or so respondents. This post covers the overview, and you can look at the raw data in this spreadsheet What did people do? Boring graph number one: I wasn't expecting that. Given I pimped the survey on twitter and a couple of Tech Comms discussion lists, I was more banking on and even Content Strategy/Tech Comms split. What the "Others" specified: Three people chipped in with Technical Writer. Author, apparently, doesn't cut it. There's a "nobody reads the instructions" joke in there somewhere, I'm sure. There were a couple of hybrid roles, including Tech Comms and Testing, which sounds gruelling and thankless. There was also, an Intranet Manager, a Creative Director, a Consultant, a CTO, an Information Architect, and a Translator. That's a pretty healthy slice through the industry. Who wrote UI text? Boring graph number two: Annoyingly, I made this a "tick all that apply" question, so I can't make crude and inflammatory generalizations about percentages. This is more about who gets involved in user interface wording. So don't panic about the number of developers writing UI text. First off, it just means they're involved. Second, they might be good at it. What? It could happen. Ours are involved - they write a placeholder and flag it to me for changes. Sometimes I don't make any. It's also not surprising that there's so much UX in the mix. Some of that will be people taking care, and crafting an understandable interface. Some of it will be whatever text goes on the wireframe making it into production. I'm going to assume that's what happened at eBay, when their iPhone app purportedly shipped with the placeholder text "Some crappy content goes here". Ahem. Listing all 17 "other" responses would make this post lengthy indeed, but you can read them in the raw data spreadsheet. The award for the approach that sounds the most like a good idea yet carries the highest risk of ending badly goes to whoever offered up "External agencies using focus groups". If you're reading this, and that actually works, leave a comment. I'm fascinated. Who owned UI text Stop. Bar chart time: Wow. Let's cut to the chase, and by "chase", I mean those inflammatory generalizations I was talking about: In around 60% of cases the person responsible for user interface text probably lacks the relevant expertise. Even in the categories I count as being likely to have relevant skills (Marketing Copywriters, Content Strategists, Technical Authors, and User Experience Designers) there's a case for each role being unsuited, as you'll see in Rachel's blog post So it's not as simple as my headline. Does that mean that you personally, Mr Developer reading this, write bad button names? Of course not. I know nothing about you. It rather implies that as a category, the majority of people looking after UI text have neither communication nor user experience as their primary skill set, and as such will probably only be good at this by happy accident. I don't have a way of measuring those frequency of those accidents. What the Others specified: I don't know who owns it. I assume the project manager is responsible. "copywriters" when they wish to annoy me. the client's web maintenance person, often PR or MarComm That last one chills me to the bone. Still, at least nobody said "the work experience kid". You can see the rest in the spreadsheet. My overwhelming impression here is of user interface text as an unloved afterthought. There were fewer "nobody" responses than I expected, and a much broader split. But the relative predominance of developers owning and writing UI text suggests to me that organizations don't see it as something worth dedicating attention to. If true, that's bothersome. Because the words on the screen, particularly the names of things, are fundamental to the ability to understand an use software. It's also fascinating that Technical Authors and Content Strategists are neck and neck. For such a nascent discipline, Content Strategy appears to have made a mark on software development. Or my sample is skewed. But it feels like a bit of validation for my rant: Content Strategy is eating Tech Comms' lunch. That's not a bad thing. Well, not if the UI text is getting done well. And that's the caveat to this whole post. I couldn't care less who writes UI text, provided they consider the user and don't suck at it. I care that it may be falling by default to people poorly disposed to doing it right. And I care about that because so much user interface text sucks. The most interesting question Was one I forgot to ask. It's this: Does your organization have technical authors/writers? Like a lot of survey data, that doesn't tell you much on its own. But once we get a bit dimensional, it become more interesting. So taken with the other questions, this would have let me find out what I really want to know: What proportion of organizations have Tech Comms professionals but don't use them for UI text? Who writes UI text in their place? Why this happens? It's possible (feasible is another matter) that hundreds of companies have tech authors who don't work on user interfaces because they've empirically discovered that someone else, say the Marketing Copywriter, is better at it. And once we've all finished laughing, I'll point out that I've met plenty of tech authors who just aren't used to thinking about users at the point of need in the way UI text and embedded user assistance require. If you've got what I regard, perhaps unfairly, as the bad kind of tech author - the old-school kind with the thousand-page pdf and the grammar obsession - if you've got one of those then you probably are better off getting the UX folk or the copywriters to do your UI text. At the very least, they'll derive terminology from user research.

    Read the article

  • Constant game speed independent of variable FPS in OpenGL with GLUT?

    - by Nazgulled
    I've been reading Koen Witters detailed article about different game loop solutions but I'm having some problems implementing the last one with GLUT, which is the recommended one. After reading a couple of articles, tutorials and code from other people on how to achieve a constant game speed, I think that what I currently have implemented (I'll post the code below) is what Koen Witters called Game Speed dependent on Variable FPS, the second on his article. First, through my searching experience, there's a couple of people that probably have the knowledge to help out on this but don't know what GLUT is and I'm going to try and explain (feel free to correct me) the relevant functions for my problem of this OpenGL toolkit. Skip this section if you know what GLUT is and how to play with it. GLUT Toolkit: GLUT is an OpenGL toolkit and helps with common tasks in OpenGL. The glutDisplayFunc(renderScene) takes a pointer to a renderScene() function callback, which will be responsible for rendering everything. The renderScene() function will only be called once after the callback registration. The glutTimerFunc(TIMER_MILLISECONDS, processAnimationTimer, 0) takes the number of milliseconds to pass before calling the callback processAnimationTimer(). The last argument is just a value to pass to the timer callback. The processAnimationTimer() will not be called each TIMER_MILLISECONDS but just once. The glutPostRedisplay() function requests GLUT to render a new frame so we need call this every time we change something in the scene. The glutIdleFunc(renderScene) could be used to register a callback to renderScene() (this does not make glutDisplayFunc() irrelevant) but this function should be avoided because the idle callback is continuously called when events are not being received, increasing the CPU load. The glutGet(GLUT_ELAPSED_TIME) function returns the number of milliseconds since glutInit was called (or first call to glutGet(GLUT_ELAPSED_TIME)). That's the timer we have with GLUT. I know there are better alternatives for high resolution timers, but let's keep with this one for now. I think this is enough information on how GLUT renders frames so people that didn't know about it could also pitch in this question to try and help if they fell like it. Current Implementation: Now, I'm not sure I have correctly implemented the second solution proposed by Koen, Game Speed dependent on Variable FPS. The relevant code for that goes like this: #define TICKS_PER_SECOND 30 #define MOVEMENT_SPEED 2.0f const int TIMER_MILLISECONDS = 1000 / TICKS_PER_SECOND; int previousTime; int currentTime; int elapsedTime; void renderScene(void) { (...) // Setup the camera position and looking point SceneCamera.LookAt(); // Do all drawing below... (...) } void processAnimationTimer(int value) { // setups the timer to be called again glutTimerFunc(TIMER_MILLISECONDS, processAnimationTimer, 0); // Get the time when the previous frame was rendered previousTime = currentTime; // Get the current time (in milliseconds) and calculate the elapsed time currentTime = glutGet(GLUT_ELAPSED_TIME); elapsedTime = currentTime - previousTime; /* Multiply the camera direction vector by constant speed then by the elapsed time (in seconds) and then move the camera */ SceneCamera.Move(cameraDirection * MOVEMENT_SPEED * (elapsedTime / 1000.0f)); // Requests to render a new frame (this will call my renderScene() once) glutPostRedisplay(); } void main(int argc, char **argv) { glutInit(&argc, argv); (...) glutDisplayFunc(renderScene); (...) // Setup the timer to be called one first time glutTimerFunc(TIMER_MILLISECONDS, processAnimationTimer, 0); // Read the current time since glutInit was called currentTime = glutGet(GLUT_ELAPSED_TIME); glutMainLoop(); } This implementation doesn't fell right. It works in the sense that helps the game speed to be constant dependent on the FPS. So that moving from point A to point B takes the same time no matter the high/low framerate. However, I believe I'm limiting the game framerate with this approach. Each frame will only be rendered when the time callback is called, that means the framerate will be roughly around TICKS_PER_SECOND frames per second. This doesn't feel right, you shouldn't limit your powerful hardware, it's wrong. It's my understanding though, that I still need to calculate the elapsedTime. Just because I'm telling GLUT to call the timer callback every TIMER_MILLISECONDS, it doesn't mean it will always do that on time. I'm not sure how can I fix this and to be completely honest, I have no idea what is the game loop in GLUT, you know, the while( game_is_running ) loop in Koen's article. But it's my understanding that GLUT is event-driven and that game loop starts when I call glutMainLoop() (which never returns), yes? I thought I could register an idle callback with glutIdleFunc() and use that as replacement of glutTimerFunc(), only rendering when necessary (instead of all the time as usual) but when I tested this with an empty callback (like void gameLoop() {}) and it was basically doing nothing, only a black screen, the CPU spiked to 25% and remained there until I killed the game and it went back to normal. So I don't think that's the path to follow. Using glutTimerFunc() is definitely not a good approach to perform all movements/animations based on that, as I'm limiting my game to a constant FPS, not cool. Or maybe I'm using it wrong and my implementation is not right? How exactly can I have a constant game speed with variable FPS? More exactly, how do I correctly implement Koen's Constant Game Speed with Maximum FPS solution (the fourth one on his article) with GLUT? Maybe this is not possible at all with GLUT? If not, what are my alternatives? What is the best approach to this problem (constant game speed) with GLUT? I originally posted this question on Stack Overflow before being pointed out about this site. The following is a different approach I tried after creating the question in SO, so I'm posting it here too. Another Approach: I've been experimenting and here's what I was able to achieve now. Instead of calculating the elapsed time on a timed function (which limits my game's framerate) I'm now doing it in renderScene(). Whenever changes to the scene happen I call glutPostRedisplay() (ie: camera moving, some object animation, etc...) which will make a call to renderScene(). I can use the elapsed time in this function to move my camera for instance. My code has now turned into this: int previousTime; int currentTime; int elapsedTime; void renderScene(void) { (...) // Setup the camera position and looking point SceneCamera.LookAt(); // Do all drawing below... (...) } void renderScene(void) { (...) // Get the time when the previous frame was rendered previousTime = currentTime; // Get the current time (in milliseconds) and calculate the elapsed time currentTime = glutGet(GLUT_ELAPSED_TIME); elapsedTime = currentTime - previousTime; /* Multiply the camera direction vector by constant speed then by the elapsed time (in seconds) and then move the camera */ SceneCamera.Move(cameraDirection * MOVEMENT_SPEED * (elapsedTime / 1000.0f)); // Setup the camera position and looking point SceneCamera.LookAt(); // All drawing code goes inside this function drawCompleteScene(); glutSwapBuffers(); /* Redraw the frame ONLY if the user is moving the camera (similar code will be needed to redraw the frame for other events) */ if(!IsTupleEmpty(cameraDirection)) { glutPostRedisplay(); } } void main(int argc, char **argv) { glutInit(&argc, argv); (...) glutDisplayFunc(renderScene); (...) currentTime = glutGet(GLUT_ELAPSED_TIME); glutMainLoop(); } Conclusion, it's working, or so it seems. If I don't move the camera, the CPU usage is low, nothing is being rendered (for testing purposes I only have a grid extending for 4000.0f, while zFar is set to 1000.0f). When I start moving the camera the scene starts redrawing itself. If I keep pressing the move keys, the CPU usage will increase; this is normal behavior. It drops back when I stop moving. Unless I'm missing something, it seems like a good approach for now. I did find this interesting article on iDevGames and this implementation is probably affected by the problem described on that article. What's your thoughts on that? Please note that I'm just doing this for fun, I have no intentions of creating some game to distribute or something like that, not in the near future at least. If I did, I would probably go with something else besides GLUT. But since I'm using GLUT, and other than the problem described on iDevGames, do you think this latest implementation is sufficient for GLUT? The only real issue I can think of right now is that I'll need to keep calling glutPostRedisplay() every time the scene changes something and keep calling it until there's nothing new to redraw. A little complexity added to the code for a better cause, I think. What do you think?

    Read the article

  • Bluetooth RFCOMM / SDP connection to a RS232 adapter in android

    - by ThePosey
    Hello All, I am trying to use the Bluetooth Chat sample API app that google provides to connect to a bluetooth RS232 adapter hooked up to another device. Here is the app for reference: http://developer.android.com/resources/samples/BluetoothChat/index.html And here is the spec sheet for the RS232 connector just for reference: http://serialio.com/download/Docs/BlueSnap-guide-4.77_Commands.pdf Well the problem is that when I go to connect to the device with: mmSocket.connect(); (BluetoothSocket::connect()) I always get an IOException error thrown by the connect() method. When I do a toString on the exception I get "Service discovery failed". My question is mostly what are the cases that would cause an IOException to get thrown in the connect method? I know those are in the source somewhere but I don't know exactly how the java layer that you write apps in and the C/C++ layer that contains the actual stacks interface. I know that it uses the bluez bluetooth stack which is written in C/C++ but not sure how that ties into the java layer which is what I would think is throwing the exception. Any help on pointing me to where I can try to dissect this issue would be incredible. Also just to note I am able to pair with the RS232 adapter just fine but I am never able to actually connect. Here is the logcat output for more reference: I/ActivityManager( 1018): Displayed activity com.example.android.BluetoothChat/.DeviceListActivity: 326 ms (total 326 ms) E/BluetoothService.cpp( 1018): stopDiscoveryNative: D-Bus error in StopDiscovery: org.bluez.Error.Failed (Invalid discovery session) D/BluetoothChat( 1729): onActivityResult -1 D/BluetoothChatService( 1729): connect to: 00:06:66:03:0C:51 D/BluetoothChatService( 1729): setState() STATE_LISTEN - STATE_CONNECTING E/BluetoothChat( 1729): + ON RESUME + I/BluetoothChat( 1729): MESSAGE_STATE_CHANGE: STATE_CONNECTING I/BluetoothChatService( 1729): BEGIN mConnectThread E/BluetoothService.cpp( 1018): stopDiscoveryNative: D-Bus error in StopDiscovery: org.bluez.Error.Failed (Invalid discovery session) E/BluetoothEventLoop.cpp( 1018): event_filter: Received signal org.bluez.Device:PropertyChanged from /org/bluez/1498/hci0/dev_00_06_66_03_0C_51 I/BluetoothChatService( 1729): CONNECTION FAIL TOSTRING: java.io.IOException: Service discovery failed D/BluetoothChatService( 1729): setState() STATE_CONNECTING - STATE_LISTEN D/BluetoothChatService( 1729): start D/BluetoothChatService( 1729): setState() STATE_LISTEN - STATE_LISTEN I/BluetoothChat( 1729): MESSAGE_STATE_CHANGE: STATE_LISTEN V/BluetoothEventRedirector( 1080): Received android.bleutooth.device.action.UUID I/NotificationService( 1018): enqueueToast pkg=com.example.android.BluetoothChat callback=android.app.ITransientNotification$Stub$Proxy@446327c8 duration=0 I/BluetoothChat( 1729): MESSAGE_STATE_CHANGE: STATE_LISTEN E/BluetoothEventLoop.cpp( 1018): event_filter: Received signal org.bluez.Device:PropertyChanged from /org/bluez/1498/hci0/dev_00_06_66_03_0C_51 V/BluetoothEventRedirector( 1080): Received android.bleutooth.device.action.UUID The device I'm trying to connect to is the 00:06:66:03:0C:51 which I can scan for and apparently pair with just fine. The below is merged from a similar question which was successfully resolved by the selected answer here: How can one connect to an rfcomm device other than another phone in Android? The Android API provides examples of using listenUsingRfcommWithServiceRecord() to set up a socket and createRfcommSocketToServiceRecord() to connect to that socket. I'm trying to connect to an embedded device with a BlueSMiRF Gold chip. My working Python code (using the PyBluez library), which I'd like to port to Android, is as follows: sock = bluetooth.BluetoothSocket(proto=bluetooth.RFCOMM) sock.connect((device_addr, 1)) return sock.makefile() ...so the service to connect to is simply defined as channel 1, without any SDP lookup. As the only documented mechanism I see in the Android API does SDP lookup of a UUID, I'm slightly at a loss. Using "sdptool browse" from my Linux host comes up empty, so I surmise that the chip in question simply lacks SDP support.

    Read the article

  • Fixing predicated NSFetchedResultsController/NSFetchRequest performance with SQLite backend?

    - by Jaanus
    I have a series of NSFetchedResultsControllers powering some table views, and their performance on device was abysmal, on the order of seconds. Since it all runs on main thread, it's blocking my app at startup, which is not great. I investigated and turns out the predicate is the problem: NSPredicate *somePredicate = [NSPredicate predicateWithFormat:@"ANY somethings == %@", something]; [fetchRequest setPredicate:somePredicate]; I.e the fetch entity, call it "things", has a many-to-many relation with entity "something". This predicate is a filter that limits the results to only things that have a relation with a particular "something". When I removed the predicate for testing, fetch time (the initial performFetch: call) dropped (for some extreme cases) from 4 seconds to around 100ms or less, which is acceptable. I am troubled by this, though, as it negates a lot of the benefit I was hoping to gain with Core Data and NSFRC, which otherwise seems like a powerful tool. So, my question is, how can I optimize this performance? Am I using the predicate wrong? Should I modify the model/schema somehow? And what other ways there are to fix this? Is this kind of degraded performance to be expected? (There are on the order of hundreds of <1KB objects.) EDIT WITH DETAILS: Here's the code: [fetchRequest setFetchLimit:200]; NSLog(@"before fetch"); BOOL success = [frc performFetch:&error]; if (!success) { NSLog(@"Fetch request error: %@", error); } NSLog(@"after fetch"); Updated logs (previously, I had some application inefficiencies degrading the performance here. These are the updated logs that should be as close to optimal as you can get under my current environment): 2010-02-05 12:45:22.138 Special Ppl[429:207] before fetch 2010-02-05 12:45:22.144 Special Ppl[429:207] CoreData: sql: SELECT DISTINCT 0, t0.Z_PK, t0.Z_OPT, <model fields> FROM ZTHING t0 LEFT OUTER JOIN Z_1THINGS t1 ON t0.Z_PK = t1.Z_2THINGS WHERE t1.Z_1SOMETHINGS = ? ORDER BY t0.ZID DESC LIMIT 200 2010-02-05 12:45:22.663 Special Ppl[429:207] CoreData: annotation: sql connection fetch time: 0.5094s 2010-02-05 12:45:22.668 Special Ppl[429:207] CoreData: annotation: total fetch execution time: 0.5240s for 198 rows. 2010-02-05 12:45:22.706 Special Ppl[429:207] after fetch If I do the same fetch without predicate (by commenting out the two lines in the beginning of the question): 2010-02-05 12:44:10.398 Special Ppl[414:207] before fetch 2010-02-05 12:44:10.405 Special Ppl[414:207] CoreData: sql: SELECT 0, t0.Z_PK, t0.Z_OPT, <model fields> FROM ZTHING t0 ORDER BY t0.ZID DESC LIMIT 200 2010-02-05 12:44:10.426 Special Ppl[414:207] CoreData: annotation: sql connection fetch time: 0.0125s 2010-02-05 12:44:10.431 Special Ppl[414:207] CoreData: annotation: total fetch execution time: 0.0262s for 200 rows. 2010-02-05 12:44:10.457 Special Ppl[414:207] after fetch 20-fold difference in times. 500ms is not that great, and there does not seem to be a way to do it in background thread or otherwise optimize that I can think of. (Apart from going to a binary store where this becomes a non-issue, so I might do that. Binary store performance is consistently ~100ms for the above 200-object predicated query.) (I nested another question here previously, which I now moved away).

    Read the article

  • Forms Authentication works on dev server but not production server (same SQL db)

    - by Desmond
    Hi, I have the same problem as a previously solved question however, this solution did not help me. I have posted the previous question and answer below: http://stackoverflow.com/questions/2215963/forms-authentication-works-on-dev-server-but-not-production-server-same-sql-db/2963985#2963985 Question: I've never had this problem before, I'm at a total loss. I have a SQL Server 2008 database with ASP.NET Forms Authentication, profiles and roles created and is functional on the development workstation. I can login using the created users without problem. I back up the database on the development computer and restore it on the production server. I xcopy the DLLs and ASP.NET files to the server. I make the necessary changes in the web.config, changing the SQL connection strings to point to the production server database and upload it. I've made sure to generate a machine key and it is the same on both the development web.config and the production web.config. And yet, when I try to login on the production server, the same user that I'm able to login successfully with on the development computer, fails on the production server. There is other content in the database, the schema generated by FluentNHibernate. This content is able to be queried successfully on both development and production servers. This is mind boggling, I believe I've verified everything, but obviously it is still not working and I must have missed something. Please, any ideas? Answer: I ran into a problem with similar symptoms at one point by forgetting to set the applicationName attribute in the web.config under the membership providers element. Users are associated to a specific application. Since I didn't set the applicationName, it defaulted to the application path (something like "/MyApplication"). When it was moved to production, the path changed (for example to "/WebSiteFolder/SomeSubFolder /MyApplication"), so the application name defaulted to the new production path and an association could not be made to the original user accounts that were set up in development. Could your issues possibly be the same as mine? I have this already in my web.config but still get the issue. Any ideas? <membership> <providers> <clear/> <add name="AspNetSqlMembershipProvider" type="System.Web.Security.SqlMembershipProvider, System.Web, Version=2.0.0.0, Culture=neutral, PublicKeyToken=b03f5f7f11d50a3a" connectionStringName="ApplicationServices" enablePasswordRetrieval="false" enablePasswordReset="true" requiresQuestionAndAnswer="false" requiresUniqueEmail="false" passwordFormat="Hashed" maxInvalidPasswordAttempts="5" minRequiredPasswordLength="6" minRequiredNonalphanumericCharacters="0" passwordAttemptWindow="10" passwordStrengthRegularExpression="" applicationName="/"/> </providers> </membership> <profile> <providers> <clear/> <add name="AspNetSqlProfileProvider" type="System.Web.Profile.SqlProfileProvider, System.Web, Version=2.0.0.0, Culture=neutral, PublicKeyToken=b03f5f7f11d50a3a" connectionStringName="ApplicationServices" applicationName="/"/> </providers> </profile> <roleManager enabled="false"> <providers> <clear/> <add connectionStringName="ApplicationServices" applicationName="/" name="AspNetSqlRoleProvider" type="System.Web.Security.SqlRoleProvider, System.Web, Version=2.0.0.0, Culture=neutral, PublicKeyToken=b03f5f7f11d50a3a"/> <add applicationName="/" name="AspNetWindowsTokenRoleProvider" type="System.Web.Security.WindowsTokenRoleProvider, System.Web, Version=2.0.0.0, Culture=neutral, PublicKeyToken=b03f5f7f11d50a3a"/> </providers> </roleManager> Any help is greatly appriciated.

    Read the article

  • MySQL Query GROUP_CONCAT Over Multiple Rows

    - by PeteGO
    I'm getting name and address data out of generic question / answer data to create some kind of normalised reporting database. The query I've got uses group_concat and works for individual sets of questions but not for multiple sets. I've tried to simplify what I'm doing by using just forename and surname and just 3 records, 2 for 1 person and 1 for another. In reality though there are more than 300,000 records. Example of results with qs.Id = 1. QuestionSetId Forename Surname ------------------------------------------------------- 1 Bob Jones Example of results with qs.Id IN (1, 2, 3). QuestionSetId Forename Surname ------------------------------------------------------- 3 Bob,Bob,Frank Jones,Jones,Smith What I would like to see for qs.Id IN (1, 2, 3). QuestionSetId Forename Surname ------------------------------------------------------- 1 Bob Jones 2 Bob Jones 3 Frank Smith So how can I make the 2nd example return a separate row for each set of name and address information? I realise the current way the data is stored is "questionable" but I cannot change the way the data is stored. I can get sets of individual answers but not sure how to combine the others. My simplified Schema that I cannot change: CREATE TABLE StaticQuestion ( Id INT NOT NULL, StaticText VARCHAR(500) NOT NULL); CREATE TABLE Question ( Id INT NOT NULL, Text VARCHAR(500) NOT NULL); CREATE TABLE StaticQuestionQuestionLink ( Id INT NOT NULL, StaticQuestionId INT NOT NULL, QuestionId INT NOT NULL, DateEffective DATETIME NOT NULL); CREATE TABLE Answer ( Id INT NOT NULL, Text VARCHAR(500) NOT NULL); CREATE TABLE QuestionSet ( Id INT NOT NULL, DateEffective DATETIME NOT NULL); CREATE TABLE QuestionAnswerLink ( Id INT NOT NULL, QuestionSetId INT NOT NULL, QuestionId INT NOT NULL, AnswerId INT NOT NULL, StaticQuestionId INT NOT NULL); Some example data for only forename and surname. INSERT INTO StaticQuestion (Id, StaticText) VALUES (1, 'FirstName'), (2, 'LastName'); INSERT INTO Question (Id, Text) VALUES (1, 'What is your first name?'), (2, 'What is your forename?'), (3, 'What is your Surname?'); INSERT INTO StaticQuestionQuestionLink (Id, StaticQuestionId, QuestionId, DateEffective) VALUES (1, 1, 1, '2001-01-01'), (2, 1, 2, '2008-08-08'), (3, 2, 3, '2001-01-01'); INSERT INTO Answer (Id, Text) VALUES (1, 'Bob'), (2, 'Jones'), (3, 'Bob'), (4, 'Jones'), (5, 'Frank'), (6, 'Smith'); INSERT INTO QuestionSet (Id, DateEffective) VALUES (1, '2002-03-25'), (2, '2009-05-05'), (3, '2009-08-06'); INSERT INTO QuestionAnswerLink (Id, QuestionSetId, QuestionId, AnswerId, StaticQuestionId) VALUES (1, 1, 1, 1, 1), (2, 1, 3, 2, 2), (3, 2, 2, 3, 1), (4, 2, 3, 4, 2), (5, 3, 2, 5, 1), (6, 3, 3, 6, 2); Just in case SQLFiddle is down here are the 3 queries from the examples I've linked to: 1: - working query but only on 1 set of data. SELECT MAX(QuestionSetId) AS QuestionSetId, GROUP_CONCAT(Forename) AS Forename, GROUP_CONCAT(Surname) AS Surname FROM (SELECT x.QuestionSetId, CASE x.StaticQuestionId WHEN 1 THEN Text END AS Forename, CASE x.StaticQuestionId WHEN 2 THEN Text END AS Surname FROM (SELECT (SELECT link.StaticQuestionId FROM StaticQuestionQuestionLink link WHERE link.Id = qa.QuestionId AND link.DateEffective <= qs.DateEffective AND link.StaticQuestionId IN (1, 2) ORDER BY link.DateEffective DESC LIMIT 1) AS StaticQuestionId, a.Text, qa.QuestionSetId FROM QuestionSet qs INNER JOIN QuestionAnswerLink qa ON qs.Id = qa.QuestionSetId INNER JOIN Answer a ON qa.AnswerId = a.Id WHERE qs.Id IN (1)) x) y 2: - working query but undesired results on multiple sets of data. SELECT MAX(QuestionSetId) AS QuestionSetId, GROUP_CONCAT(Forename) AS Forename, GROUP_CONCAT(Surname) AS Surname FROM (SELECT x.QuestionSetId, CASE x.StaticQuestionId WHEN 1 THEN Text END AS Forename, CASE x.StaticQuestionId WHEN 2 THEN Text END AS Surname FROM (SELECT (SELECT link.StaticQuestionId FROM StaticQuestionQuestionLink link WHERE link.Id = qa.QuestionId AND link.DateEffective <= qs.DateEffective AND link.StaticQuestionId IN (1, 2) ORDER BY link.DateEffective DESC LIMIT 1) AS StaticQuestionId, a.Text, qa.QuestionSetId FROM QuestionSet qs INNER JOIN QuestionAnswerLink qa ON qs.Id = qa.QuestionSetId INNER JOIN Answer a ON qa.AnswerId = a.Id WHERE qs.Id IN (1, 2, 3)) x) y 3: - working query on multiple sets of data only on 1 field (answer) though. SELECT qs.Id AS QuestionSet, a.Text AS Answer FROM QuestionSet qs INNER JOIN QuestionAnswerLink qalink ON qs.Id = qalink.QuestionSetId INNER JOIN StaticQuestionQuestionLink sqqlink ON qalink.QuestionId = sqqlink.QuestionId INNER JOIN Answer a ON qalink.AnswerId = a.Id WHERE sqqlink.StaticQuestionId = 1 /* FirstName */ AND sqqlink.DateEffective = (SELECT DateEffective FROM StaticQuestionQuestionLink WHERE StaticQuestionId = 1 AND DateEffective <= qs.DateEffective ORDER BY DateEffective DESC LIMIT 1)

    Read the article

  • SQLAlchemy session management in long-running process

    - by codeape
    Scenario: A .NET-based application server (Wonderware IAS/System Platform) hosts automation objects that communicate with various equipment on the factory floor. CPython is hosted inside this application server (using Python for .NET). The automation objects have scripting functionality built-in (using a custom, .NET-based language). These scripts call Python functions. The Python functions are part of a system to track Work-In-Progress on the factory floor. The purpose of the system is to track the produced widgets along the process, ensure that the widgets go through the process in the correct order, and check that certain conditions are met along the process. The widget production history and widget state is stored in a relational database, this is where SQLAlchemy plays its part. For example, when a widget passes a scanner, the automation software triggers the following script (written in the application server's custom scripting language): ' wiget_id and scanner_id provided by automation object ' ExecFunction() takes care of calling a CPython function retval = ExecFunction("WidgetScanned", widget_id, scanner_id); ' if the python function raises an Exception, ErrorOccured will be true ' in this case, any errors should cause the production line to stop. if (retval.ErrorOccured) then ProductionLine.Running = False; InformationBoard.DisplayText = "ERROR: " + retval.Exception.Message; InformationBoard.SoundAlarm = True end if; The script calls the WidgetScanned python function: # pywip/functions.py from pywip.database import session from pywip.model import Widget, WidgetHistoryItem from pywip import validation, StatusMessage from datetime import datetime def WidgetScanned(widget_id, scanner_id): widget = session.query(Widget).get(widget_id) validation.validate_widget_passed_scanner(widget, scanner) # raises exception on error widget.history.append(WidgetHistoryItem(timestamp=datetime.now(), action=u"SCANNED", scanner_id=scanner_id)) widget.last_scanner = scanner_id widget.last_update = datetime.now() return StatusMessage("OK") # ... there are a dozen similar functions My question is: How do I best manage SQLAlchemy sessions in this scenario? The application server is a long-running process, typically running months between restarts. The application server is single-threaded. Currently, I do it the following way: I apply a decorator to the functions I make avaliable to the application server: # pywip/iasfunctions.py from pywip import functions def ias_session_handling(func): def _ias_session_handling(*args, **kwargs): try: retval = func(*args, **kwargs) session.commit() return retval except: session.rollback() raise return _ias_session_handling # ... actually I populate this module with decorated versions of all the functions in pywip.functions dynamically WidgetScanned = ias_session_handling(functions.WidgetScanned) Question: Is the decorator above suitable for handling sessions in a long-running process? Should I call session.remove()? The SQLAlchemy session object is a scoped session: # pywip/database.py from sqlalchemy.orm import scoped_session, sessionmaker session = scoped_session(sessionmaker()) I want to keep the session management out of the basic functions. For two reasons: There is another family of functions, sequence functions. The sequence functions call several of the basic functions. One sequence function should equal one database transaction. I need to be able to use the library from other environments. a) From a TurboGears web application. In that case, session management is done by TurboGears. b) From an IPython shell. In that case, commit/rollback will be explicit. (I am truly sorry for the long question. But I felt I needed to explain the scenario. Perhaps not necessary?)

    Read the article

  • Databinding to ObservableCollection in a different UserControl - how to preserve current selections?

    - by Dave
    Scope of question expanded on 2010-03-25 I ended up figuring out my problem, but here's a new problem that came up as a result of solving the original question, because I want to be able to award the bounty to someone!!! Once I figured out my problem, I soon found out that when the ObservableCollection updates, the databound ComboBox has its contents repopulated, but most of the selections have been blanked out. I assume that in this case, MVVM is going to make it difficult for me to remember the last selected item. I have an idea, but it seems a little nasty. I'll award the bounty to whomever comes up with a nice solution for this! Question re-written on 2010-03-24 I have two UserControls, where one is a dialog that has a TabControl, and the other is one that appears within said TabControl. I'll just call them CandyDialog and CandyNameViewer for simplicity's sake. There's also a data management class called Tracker that manages information storage, which for all intents and purposes just exposes a public property that is an ObservableCollection. I display the CandyNameViewer in CandyDialog via code behind, like this: private void CandyDialog_Loaded( object sender, RoutedEventArgs e) { _candyviewer = new CandyViewer(); _candyviewer.DataContext = _tracker; candy_tab.Content = _candyviewer; } The CandyViewer's XAML looks like this (edited for kaxaml): <Page xmlns="http://schemas.microsoft.com/winfx/2006/xaml/presentation" xmlns:x="http://schemas.microsoft.com/winfx/2006/xaml"> <Page.Resources> <DataTemplate x:Key="CandyItemTemplate"> <Grid> <Grid.ColumnDefinitions> <ColumnDefinition Width="120"></ColumnDefinition> <ColumnDefinition Width="150"></ColumnDefinition> </Grid.ColumnDefinitions> <TextBox Grid.Column="0" Text="{Binding CandyName}" Margin="3"></TextBox> <!-- just binding to DataContext ends up using InventoryItem as parent, so we need to get to the UserControl --> <ComboBox Grid.Column="1" SelectedItem="{Binding SelectedCandy, Mode=TwoWay}" ItemsSource="{Binding RelativeSource={RelativeSource FindAncestor, AncestorType={x:Type UserControl}}, Path=DataContext.CandyNames}" Margin="3"></ComboBox> </Grid> </DataTemplate> </Page.Resources> <Grid> <ListBox DockPanel.Dock="Top" ItemsSource="{Binding CandyBoxContents, Mode=TwoWay}" ItemTemplate="{StaticResource CandyItemTemplate}" /> </Grid> </Page> Now everything works fine when the controls are loaded. As long as CandyNames is populated first, and then the consumer UserControl is displayed, all of the names are there. I obviously don't get any errors in the Output Window or anything like that. The issue I have is that when the ObservableCollection is modified from the model, those changes are not reflected in the consumer UserControl! I've never had this problem before; all of my previous uses of ObservableCollection updated fine, although in those cases I wasn't databinding across assemblies. Although I am currently only adding and removing candy names to/from the ObservableCollection, at a later date I will likely also allow renaming from the model side. Is there something I did wrong? Is there a good way to actually debug this? Reed Copsey indicates here that inter-UserControl databinding is possible. Unfortunately, my favorite Bea Stollnitz article on WPF databinding debugging doesn't suggest anything that I could use for this particular problem.

    Read the article

  • Implementing a popularity algorithm in Django

    - by TheLizardKing
    I am creating a site similar to reddit and hacker news that has a database of links and votes. I am implementing hacker news' popularity algorithm and things are going pretty swimmingly until it comes to actually gathering up these links and displaying them. The algorithm is simple: Y Combinator's Hacker News: Popularity = (p - 1) / (t + 2)^1.5` Votes divided by age factor. Where` p : votes (points) from users. t : time since submission in hours. p is subtracted by 1 to negate submitter's vote. Age factor is (time since submission in hours plus two) to the power of 1.5.factor is (time since submission in hours plus two) to the power of 1.5. I asked a very similar question over yonder http://stackoverflow.com/questions/1964395/complex-ordering-in-django but instead of contemplating my options I choose one and tried to make it work because that's how I did it with PHP/MySQL but I now know Django does things a lot differently. My models look something (exactly) like this class Link(models.Model): category = models.ForeignKey(Category) user = models.ForeignKey(User) created = models.DateTimeField(auto_now_add = True) modified = models.DateTimeField(auto_now = True) fame = models.PositiveIntegerField(default = 1) title = models.CharField(max_length = 256) url = models.URLField(max_length = 2048) def __unicode__(self): return self.title class Vote(models.Model): link = models.ForeignKey(Link) user = models.ForeignKey(User) created = models.DateTimeField(auto_now_add = True) modified = models.DateTimeField(auto_now = True) karma_delta = models.SmallIntegerField() def __unicode__(self): return str(self.karma_delta) and my view: def index(request): popular_links = Link.objects.select_related().annotate(karma_total = Sum('vote__karma_delta')) return render_to_response('links/index.html', {'links': popular_links}) Now from my previous question, I am trying to implement the algorithm using the sorting function. An answer from that question seems to think I should put the algorithm in the select and sort then. I am going to paginate these results so I don't think I can do the sorting in python without grabbing everything. Any suggestions on how I could efficiently do this? EDIT This isn't working yet but I think it's a step in the right direction: from django.shortcuts import render_to_response from linkett.apps.links.models import * def index(request): popular_links = Link.objects.select_related() popular_links = popular_links.extra( select = { 'karma_total': 'SUM(vote.karma_delta)', 'popularity': '(karma_total - 1) / POW(2, 1.5)', }, order_by = ['-popularity'] ) return render_to_response('links/index.html', {'links': popular_links}) This errors out into: Caught an exception while rendering: column "karma_total" does not exist LINE 1: SELECT ((karma_total - 1) / POW(2, 1.5)) AS "popularity", (S... EDIT 2 Better error? TemplateSyntaxError: Caught an exception while rendering: missing FROM-clause entry for table "vote" LINE 1: SELECT ((vote.karma_total - 1) / POW(2, 1.5)) AS "popularity... My index.html is simply: {% block content %} {% for link in links %} karma-up {{ link.karma_total }} karma-down {{ link.title }} Posted by {{ link.user }} to {{ link.category }} at {{ link.created }} {% empty %} No Links {% endfor %} {% endblock content %} EDIT 3 So very close! Again, all these answers are great but I am concentrating on a particular one because I feel it works best for my situation. from django.db.models import Sum from django.shortcuts import render_to_response from linkett.apps.links.models import * def index(request): popular_links = Link.objects.select_related().extra( select = { 'popularity': '(SUM(links_vote.karma_delta) - 1) / POW(2, 1.5)', }, tables = ['links_link', 'links_vote'], order_by = ['-popularity'], ) return render_to_response('links/test.html', {'links': popular_links}) Running this I am presented with an error hating on my lack of group by values. Specifically: TemplateSyntaxError at / Caught an exception while rendering: column "links_link.id" must appear in the GROUP BY clause or be used in an aggregate function LINE 1: ...karma_delta) - 1) / POW(2, 1.5)) AS "popularity", "links_lin... Not sure why my links_link.id wouldn't be in my group by but I am not sure how to alter my group by, django usually does that.

    Read the article

  • Convert extended ASCII characters to it's right presentation using Console.ReadKey() method and ConsoleKeyInfo variable

    - by mishamosher
    Readed about 30 minutes, and didn't found some specific for this in this site. Suppose the following, in C#, console application: ConsoleKeyInfo cki; cki = Console.ReadKey(true); Console.WriteLine(cki.KeyChar.ToString()); //Or Console.WriteLine(cki.KeyChar) as well Console.ReadKey(true); Now, let's put ¿ in the console entry, and asign it to cki via a Console.ReadKey(true). What will be shown isn't the ¿ symbol, the ¨ symbol is the one that's shown instead. And the same happens with many other characters. Examples: ñ shows ¤, ¡ shows -, ´ shows ï. Now, let's take the same code snipplet and add some things for a more Console.ReadLine() like behavior: string data = string.Empty; ConsoleKeyInfo cki; for (int i = 0; i < 10; i++) { cki = Console.ReadKey(true); data += cki.KeyChar; } Console.WriteLine(data); Console.ReadKey(true); The question, how to handle this by the right way, end printing the right characters that should be stored on data, not things like ¨, ¤, -, ï, etc? Please note that I want a solution that works with ConsoleKeyInfo and Console.ReadKey(), not use other variable types, or read methods. EDIT: Because ReadKey() method, that comes from Console namespace, depends on Kernel32.dll and it definetively bad handles the extended ASCII and unicode, it's not an option anymore to just find a valid conversion for what it returns. The only valid way to handle the bad behavior of ReadKey() is to use the cki.Key property that's written in cki = Console.ReadKey(true) execution and apply a switch to it, then, return the right values on dependence of what key was pressed. For example, to handle the Ñ key pressing: string data = string.Empty; ConsoleKeyInfo cki; cki = Console.ReadKey(true); switch (cki.Key) { case ConsoleKey.Oem3: if (cki.Modifiers.ToString().Contains("Shift")) //Could added handlers for Alt and Control, but not putted in here to keep the code small and simple data += "Ñ"; else data += "ñ"; break; } Console.WriteLine(data); Console.ReadKey(true); So, now the question has a wider focus... Which others functions completes it's execution with only one key pressed, and returns what's pressed (a substitute of ReadKey())? I think that there's not such substitutes, but a confirmed answer would be usefull. EDIT2: HA! Found the way, for something I used for so many times Windows 98 SE. There are the codepages, the ones responsibles for how's presented the info in the console. ReadLine() reconfigures the codepage to use properly the extended ASCII and Unicode characters. ReadKey() leaves it in EN-US default (codepage 850). Just use a codepage that prints the characters you want, and that's all. Refer to http://en.wikipedia.org/wiki/Code_page for some of them :) So, for the Ñ key press, the solution is this: Console.OutputEncoding = Encoding.GetEncoding(1252); //Also 28591 is valid for `Ñ` key, and others too string data = string.Empty; ConsoleKeyInfo cki; cki = Console.ReadKey(true); data += cki.KeyChar; Console.WriteLine(data); Console.ReadKey(true); Simple :) Now I'm wrrr with myself... how could I forget those codepages!? Question answered, so, no more about this!

    Read the article

  • Postmortem debugging with WinDBG.

    - by Drazar
    I have an WCF-service running on an server, and occasionally(1-2 times every month) it throws an COMException with the informative message ”Unknown error (0x8005008)”. When i googled for this particular error I only got threads about problems when creating virtual directories in IIS. And the source code hasn’t anything with making a virtual directory in IIS. DirectoryServiceLib.LdapProvider.Directory - CreatePost - Could not create employee for 195001010000,000000000000: System.Runtime.InteropServices.COMException (0x80005008): Unknown error (0x80005008) at System.DirectoryServices.PropertyValueCollection.PopulateList I've taken a memorydump when I catch the Exception for further analysis in WinDBG. After switching to the right thread I executed the !CLRStack command: 000000001b8ab6d8 000000007708671a [NDirectMethodFrameStandalone: 000000001b8ab6d8] Common.MemoryDump.MiniDumpWriteDump(IntPtr, Int32, IntPtr, MINIDUMP_TYPE, IntPtr, IntPtr, IntPtr) 000000001b8ab680 000007ff002808d8 DomainBoundILStubClass.IL_STUB_PInvoke(IntPtr, Int32, IntPtr, MINIDUMP_TYPE, IntPtr, IntPtr, IntPtr) 000000001b8ab780 000007ff00280812 Common.MemoryDump.CreateMiniDump(System.String) 000000001b8ab7e0 000007ff0027b218 DirectoryServiceLib.LdapProvider.Directory.CreatePost(System.String, DirectoryServiceLib.Model.Post, DirectoryServiceLib.Model.Presumptions, Services.Common.SourceEnum, System.String) 000000001b8ad6d8 000007fef8816869 [HelperMethodFrame: 000000001b8ad6d8] 000000001b8ad820 000007feec2b6c6f System.DirectoryServices.PropertyValueCollection.PopulateList() 000000001b8ad860 000007feec225f0f System.DirectoryServices.PropertyValueCollection..ctor(System.DirectoryServices.DirectoryEntry, System.String) 000000001b8ad8a0 000007feec22d023 System.DirectoryServices.PropertyCollection.get_Item(System.String) 000000001b8ad8f0 000007ff00274d34 Common.DirectoryEntryExtension.GetStringAttribute(System.String) 000000001b8ad940 000007ff0027f507 DirectoryServiceLib.LdapProvider.DirectoryPost.Copy(DirectoryServiceLib.LdapProvider.DirectoryPost) 000000001b8ad980 000007ff0027a7cf DirectoryServiceLib.LdapProvider.Directory.CreatePost(System.String, DirectoryServiceLib.Model.Post, DirectoryServiceLib.Model.Presumptions, Services.Common.SourceEnum, System.String) 000000001b8adbe0 000007ff00279532 DirectoryServiceLib.WCFDirectory.CreatePost(System.String, DirectoryServiceLib.Model.Post, DirectoryServiceLib.Model.Presumptions, Services.Common.SourceEnum, System.String) 000000001b8adc60 000007ff001f47bd DynamicClass.SyncInvokeCreatePost(System.Object, System.Object[], System.Object[]) My conclusion is that it fails when the code is calling System.DirectoryServices.PropertyCollection.get_Item(System.String). So after issuing an !CLRStack -a I get this result: 000000001b8ad8a0 000007feec22d023 System.DirectoryServices.PropertyCollection.get_Item(System.String) PARAMETERS: this = <no data> propertyName = <no data> LOCALS: <CLR reg> = 0x0000000001dcef78 <no data> My very first question is why does it display no data on the propertyname? I am kinda new on Windbg. However I executed an dumpobject on = 0x0000000001dcef78: 0:013> !do 0x0000000001dcef78 Name: System.String MethodTable: 000007fef66d6960 EEClass: 000007fef625eec8 Size: 74(0x4a) bytes File: C:\Windows\Microsoft.Net\assembly\GAC_64\mscorlib\v4.0_4.0.0.0__b77a5c561934e089\mscorlib.dll String: personalprescriptioncode Fields: MT Field Offset Type VT Attr Value Name 000007fef66dc848 40000ed 8 System.Int32 1 instance 24 m_stringLength 000007fef66db388 40000ee c System.Char 1 instance 70 m_firstChar 000007fef66d6960 40000ef 10 System.String 0 shared static Empty >> Domain:Value 0000000000174e10:00000000019d1420 000000001a886f50:00000000019d1420 << So when the source code wants to fetch the personalprescriptioncode from Active Directory(what is used for persistence layer) it fails. Looking back at the stack it is when issuing the Copy method. DirectoryServiceLib.LdapProvider.DirectoryPost.Copy(DirectoryServiceLib.LdapProvider.DirectoryPost) So looking in the sourcecode: DirectoryPost postInLimbo = DirectoryPostFactory.Instance().GetDirectoryPost(LdapConfigReader.Instance().GetConfigValue("LimboDN"), idGenPerson.ID.UserId); if (postInLimbo != null) newPost.Copy(postInLimbo); This code is looking for another post in OU=limbo with the same UserId and if it finds one it copies the attributes to the new post. In this case it does and it fails with personalprescriptioncode. I've looked in Active Directory under OU=Limbo and the post exist there with the attribute personalprescriptioncode=31243. Question 1: Why does it display no data for some of the PARAMETERS and LOCALS? Is it the GC who has cleaned up before the memorydump had been created. Question 2: Is there anymore i can do to get to the solution to this problem?

    Read the article

  • TripleDES in Perl/PHP/ColdFusion

    - by Seidr
    Recently a problem arose regarding hooking up an API with a payment processor who were requesting a string to be encrypted to be used as a token, using the TripleDES standard. Our Applications run using ColdFusion, which has an Encrypt tag - that supports TripleDES - however the result we were getting back was not what the payment processor expected. First of all, here is the resulting token the payment processor were expecting. AYOF+kRtg239Mnyc8QIarw== And below is the snippet of ColdFusion we were using, and the resulting string. <!--- Coldfusion Crypt (here be monsters) ---> <cfset theKey="123412341234123412341234"> <cfset theString = "username=test123"> <cfset strEncodedEnc = Encrypt(theString, theKey, "DESEDE", "Base64")> <!--- resulting string(strEncodedEnc): tc/Jb7E9w+HpU2Yvn5dA7ILGmyNTQM0h ---> As you can see, this was not returning the string we were hoping for. Seeking a solution, we ditched ColdFusion for this process and attempted to reproduce the token in PHP. Now I'm aware that various languages implement encryption in different ways - for example in the past managing encryption between a C# application and PHP back-end, I've had to play about with padding in order to get the two to talk, but my experience has been that PHP generally behaves when it comes to encryption standards. Anyway, on to the PHP source we tried, and the resulting string. /* PHP Circus (here be Elephants) */ $theKey="123412341234123412341234"; $theString="username=test123"; $strEncodedEnc=base64_encode(mcrypt_ecb (MCRYPT_3DES, $theKey, $theString, MCRYPT_ENCRYPT)); /* resulting string(strEncodedEnc): sfiSu4mVggia8Ysw98x0uw== */ As you can plainly see, we've got another string that differs from both the string expected by the payment processor AND the one produced by ColdFusion. Cue head-against-wall integration techniques. After many to-and-fro communications with the payment processor (lots and lots of reps stating 'we can't help with coding issues, you must be doing it incorrectly, read the manual') we were finally escalated to someone with more than a couple of brain-cells to rub together, who was able to step back and actually look at and diagnose the issue. He agreed, our CF and PHP attempts were not resulting in the correct string. After a quick search, he also agreed that it was not neccesarily our source, but rather how the two languages implemented their vision of the TripleDES standard. Coming into the office this morning, we were met by an email with a snippet of source code, in Perl. This is was the code they were directly using on their end to produce the expected token. #!/usr/bin/perl # Perl Crypt Calamity (here be...something) use strict; use CGI; use MIME::Base64; use Crypt::TripleDES; my $cgi = CGI->new(); my $param = $cgi->Vars(); $param->{key} = "123412341234123412341234"; $param->{string} = "username=test123"; my $des = Crypt::TripleDES->new(); my $enc = $des->encrypt3($param->{string}, $param->{key}); $enc = encode_base64($enc); $enc =~ s/\n//gs; # resulting string (enc): AYOF+kRtg239Mnyc8QIarw== So, there we have it. Three languages, three implementations of what they quote in the documentation as TripleDES Standard Encryption, and three totally different resulting strings. My question is, from your experience of these three languages and their implementations of the TripleDES algorithm, have you been able to get any two of them to give the same response, and if so what tweaks to the code did you have to make in order to come to the result? I understand this is a very drawn out question, but I wanted to give clear and precise setting for each stage of testing that we had to perform. I'll also be performing some more investigatory work on this subject later, and will post any findings that I come up with to this question, so that others may avoid this headache.

    Read the article

  • Disable antialiasing for a specific GDI device context

    - by Jacob Stanley
    I'm using a third party library to render an image to a GDI DC and I need to ensure that any text is rendered without any smoothing/antialiasing so that I can convert the image to a predefined palette with indexed colors. The third party library i'm using for rendering doesn't support this and just renders text as per the current windows settings for font rendering. They've also said that it's unlikely they'll add the ability to switch anti-aliasing off any time soon. The best work around I've found so far is to call the third party library in this way (error handling and prior settings checks ommitted for brevity): private static void SetFontSmoothing(bool enabled) { int pv = 0; SystemParametersInfo(Spi.SetFontSmoothing, enabled ? 1 : 0, ref pv, Spif.None); } // snip Graphics graphics = Graphics.FromImage(bitmap) IntPtr deviceContext = graphics.GetHdc(); SetFontSmoothing(false); thirdPartyComponent.Render(deviceContext); SetFontSmoothing(true); This obviously has a horrible effect on the operating system, other applications flicker from cleartype enabled to disabled and back every time I render the image. So the question is, does anyone know how I can alter the font rendering settings for a specific DC? Even if I could just make the changes process or thread specific instead of affecting the whole operating system, that would be a big step forward! (That would give me the option of farming this rendering out to a separate process- the results are written to disk after rendering anyway) EDIT: I'd like to add that I don't mind if the solution is more complex than just a few API calls. I'd even be happy with a solution that involved hooking system dlls if it was only about a days work. EDIT: Background Information The third-party library renders using a palette of about 70 colors. After the image (which is actually a map tile) is rendered to the DC, I convert each pixel from it's 32-bit color back to it's palette index and store the result as an 8bpp greyscale image. This is uploaded to the video card as a texture. During rendering, I re-apply the palette (also stored as a texture) with a pixel shader executing on the video card. This allows me to switch and fade between different palettes instantaneously instead of needing to regenerate all the required tiles. It takes between 10-60 seconds to generate and upload all the tiles for a typical view of the world. EDIT: Renamed GraphicsDevice to Graphics The class GraphicsDevice in the previous version of this question is actually System.Drawing.Graphics. I had renamed it (using GraphicsDevice = ...) because the code in question is in the namespace MyCompany.Graphics and the compiler wasn't able resolve it properly. EDIT: Success! I even managed to port the PatchIat function below to C# with the help of Marshal.GetFunctionPointerForDelegate. The .NET interop team really did a fantastic job! I'm now using the following syntax, where Patch is an extension method on System.Diagnostics.ProcessModule: module.Patch( "Gdi32.dll", "CreateFontIndirectA", (CreateFontIndirectA original) => font => { font->lfQuality = NONANTIALIASED_QUALITY; return original(font); }); private unsafe delegate IntPtr CreateFontIndirectA(LOGFONTA* lplf); private const int NONANTIALIASED_QUALITY = 3; [StructLayout(LayoutKind.Sequential)] private struct LOGFONTA { public int lfHeight; public int lfWidth; public int lfEscapement; public int lfOrientation; public int lfWeight; public byte lfItalic; public byte lfUnderline; public byte lfStrikeOut; public byte lfCharSet; public byte lfOutPrecision; public byte lfClipPrecision; public byte lfQuality; public byte lfPitchAndFamily; public unsafe fixed sbyte lfFaceName [32]; }

    Read the article

  • Technology to communicate with someone with expressive aphasia?

    - by rascher
    A family member had a stroke a few years back and now has expressive aphasia. She understands what is said to her, is cognitive of what is going on, but cannot express herself. She is able to respond to yes/no questions (do you want to go shopping? are you looking for your earrings?) She is not, however, able to read (English is not her native language and she hasn't read Hindi for decades.) I am the technologist in the family, and I intend to come up with something to help us communicate. The idea is to have some sort of picture book where she can point to what she wants. My first question: does some sort of assistive technology for people with expressive aphasia already exist? These can be hardware or software devices? If not, then such a software doesn't seem difficult to write. My initial thought is to have an interface with pictures - maybe separated by category (food, shopping) - where she can point to an individual picture to indicate what she needs. We could easily add more items with such a software, and we could have an interface where she (or we) could "flip pages". Which suggests that the best solution would use a touch screen rather than a mouse. It would be really difficult to train her to aim a mouse or find keys on a keyboard. We're thinking of maybe getting a tablet and writing some basic software. But tablets computers are expensive and fragile - I'm not sure if it would be able to stand spills or being knocked about in a nursing home. So my next question: what kind of tablet-like devices are out there which I can program on? I don't know anything about hardware, but if there is something then we could special-order it. What would be safe and durable for such a project? We could do something on an iPod or cell phone, but I feel like that interface would be too small. Finally, does anyone here have experience with this kind of assistive technology? Things I might not anticipate when designing such a system? edit I've added a (pretty hefty!) bounty. I'd kinda like to open this question up to any suggestions, comments, and experiences that people might have. This is a pretty real and important project, so while we will (are working on) a solution, any insights would be particularly helpful. Right now the plan is to mount a screen in her room. We'll either teach her to use a trackball or use a touch-screen panel, after seeing what she is able to use with a simple prototype. Then software akin to an old "hypercard" deck: ---------------------------------------------------------------- | -------------- -------------- | | | Clothes | | Food | ... | | -------------- -------------- | | | | Pic of item 1 Pic of item 2 Pic of item 3 | | | | | | | | | | Pic of item 4 Pic of item 5 Pic of item 6 | | | | | | | | | | <-Back Next-> | ---------------------------------------------------------------- commentcommentcomment!

    Read the article

  • Calculating negative fractions in Objective C

    - by Mark Reid
    I've been coding my way through Steve Kochan's Programming in Objective-C 2.0 book. I'm up to an exercise in chapter 7, ex 4, in case anyone has the book. The question posed by the exercise it will the Fraction class written work with negative fractions such as -1/2 + -2/3? Here's the implementation code in question - @implementation Fraction @synthesize numerator, denominator; -(void) print { NSLog(@"%i/%i", numerator, denominator); } -(void) setTo: (int) n over: (int) d { numerator = n; denominator = d; } -(double) convertToNum { if (denominator != 0) return (double) numerator / denominator; else return 1.0; } -(Fraction *) add: (Fraction *) f { // To add two fractions: // a/b + c/d = ((a * d) + (b * c)) / (b * d) // result will store the result of the addition Fraction *result = [[Fraction alloc] init]; int resultNum, resultDenom; resultNum = (numerator * f.denominator) + (denominator * f.numerator); resultDenom = denominator * f.denominator; [result setTo: resultNum over: resultDenom]; [result reduce]; return result; } -(Fraction *) subtract: (Fraction *) f { // To subtract two fractions: // a/b - c/d = ((a * d) - (b * c)) / (b * d) // result will store the result of the addition Fraction *result = [[Fraction alloc] init]; int resultNum, resultDenom; resultNum = numerator * f.denominator - denominator * f.numerator; resultDenom = denominator * f.denominator; [result setTo: resultNum over: resultDenom]; [result reduce]; return result; } -(Fraction *) multiply: (Fraction *) f { // To multiply two fractions // a/b * c/d = (a*c) / (b*d) // result will store the result of the addition Fraction *result = [[Fraction alloc] init]; int resultNum, resultDenom; resultNum = numerator * f.numerator; resultDenom = denominator * f.denominator; [result setTo: resultNum over: resultDenom]; [result reduce]; return result; } -(Fraction *) divide: (Fraction *) f { // To divide two fractions // a/b / c/d = (a*d) / (b*c) // result will store the result of the addition Fraction *result = [[Fraction alloc] init]; int resultNum, resultDenom; resultNum = numerator * f.denominator; resultDenom = denominator * f.numerator; [result setTo: resultNum over: resultDenom]; [result reduce]; return result; } -(void) reduce { int u = numerator; int v = denominator; int temp; while (v != 0) { temp = u % v; u = v; v = temp; } numerator /= u; denominator /= u; } @end My question to you is will it work with negative fractions and can you explain how you know? Part of the issue is I don't know how to calculate negative fractions myself so I'm not too sure how to know. Many thanks.

    Read the article

< Previous Page | 373 374 375 376 377 378 379 380 381 382 383 384  | Next Page >