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  • Should we use temporary variables for the returned values of functions?

    - by totymedli
    I thought about this: Is there a performance difference in these two practices: Store the return value of a function in a temporary variable than give that variable as a parameter to another function. Put the function into the other function. Specification Assuming all classes and functions are written correctly. Case 1. ClassA a = function1(); ClassB b = function2(a); function3(b); Case 2. function3(function2(function1())); I know there aren't a big difference with only one run, but supposed that we could run this a lot of times in a loop, I created some tests. Test #include <iostream> #include <ctime> #include <math.h> using namespace std; int main() { clock_t start = clock(); clock_t ends = clock(); // Case 1. start = clock(); for (int i=0; i<10000000; i++) { double a = cos(1); double b = pow(a, 2); sqrt(b); } ends = clock(); cout << (double) (ends - start) / CLOCKS_PER_SEC << endl; // Case 2. start = clock(); for (int i=0; i<10000000; i++) sqrt(pow(cos(1),2)); ends = clock(); cout << (double) (ends - start) / CLOCKS_PER_SEC << endl; return 0; } Results Case 1 = 6.375 Case 2 = 0.031 Why is the first one is much slower, and if the second one is faster why dont we always write code that way? Anyway does the second pratice has a name? I also wondered what happens if I create the variables outside the for loop in the first case, but the result was the same. Why?

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  • Tips for XNA WP7 Developers

    - by Michael B. McLaughlin
    There are several things any XNA developer should know/consider when coming to the Windows Phone 7 platform. This post assumes you are familiar with the XNA Framework and with the changes between XNA 3.1 and XNA 4.0. It’s not exhaustive; it’s simply a list of things I’ve gathered over time. I may come back and add to it over time, and I’m happy to add anything anyone else has experienced or learned as well. Display · The screen is either 800x480 or 480x800. · But you aren’t required to use only those resolutions. · The hardware scaler on the phone will scale up from 240x240. · One dimension will be capped at 800 and the other at 480; which depends on your code, but you cannot have, e.g., an 800x600 back buffer – that will be created as 800x480. · The hardware scaler will not normally change aspect ratio, though, so no unintended stretching. · Any dimension (width, height, or both) below 240 will be adjusted to 240 (without any aspect ratio adjustment such that, e.g. 200x240 will be treated as 240x240). · Dimensions below 240 will be honored in terms of calculating whether to use portrait or landscape. · If dimensions are exactly equal or if height is greater than width then game will be in portrait. · If width is greater than height, the game will be in landscape. · Landscape games will automatically flip if the user turns the phone 180°; no code required. · Default landscape is top = left. In other words a user holding a phone who starts a landscape game will see the first image presented so that the “top” of the screen is along the right edge of his/her phone, such that the natural behavior would be to turn the phone 90° so that the top of the phone will be held in the user’s left hand and the bottom would be held in the user’s right hand. · The status bar (where the clock, battery power, etc., are found) is hidden when the Game-derived class sets GraphicsDeviceManager.IsFullScreen = true. It is shown when IsFullScreen = false. The default value is false (i.e. the status bar is shown). · You should have a good reason for hiding the status bar. Users find it helpful to know what time it is, how much charge their battery has left, and whether or not their phone is in service range. This is especially true for casual games that you expect someone to play for a few minutes at a time, e.g. while waiting for some event to start, for a phone call to come in, or for a train, bus, or subway to arrive. · In portrait mode, the status bar occupies 32 pixels of space. This means that a game with a back buffer of 480x800 will be scaled down to occupy approximately 461x768 screen pixels. Setting the back buffer to 480x768 (or some resolution with the same 0.625 aspect ratio) will avoid this scaling. · In landscape mode, the status bar occupies 72 pixels of space. This means that a game with a back buffer of 800x480 will be scaled down to occupy approximately 728x437 screen pixels. Setting the back buffer to 728x480 (or some resolution with the same 1.51666667 aspect ratio) will avoid this scaling. Input · Touch input is scaled with screen size. · So if your back buffer is 600x360, a tap in the bottom right corner will come in as (599,359). You don’t need to do anything special to get this automatic scaling of touch behavior. · If you do not use full area of the screen, any touch input outside the area you use will still register as a touch input. For example, if you set a portrait resolution of 240x240, it would be scaled up to occupy a 480x480 area, centered in the screen. If you touch anywhere above this area, you will get a touch input of (X,0) where X is a number from 0 to 239 (in accordance with your 240 pixel wide back buffer). Any touch below this area will give a touch input of (X,239). · If you keep the status bar visible, touches within its area will not be passed to your game. · In general, a screen measurement is the diagonal. So a 3.5” screen is 3.5” long from the bottom right corner to the top left corner. With an aspect ratio of 0.6 (480/800 = 0.6), this means that a phone with a 3.5” screen is only approximately 1.8” wide by 3” tall. So there are approximately 267 pixels in an inch on a 3.5” screen. · Again, this time in metric! 3.5 inches is approximately 8.89 cm. So an 8.89 cm screen is 8.89 cm long from the bottom right corner to the top left corner. With an aspect ratio of 0.6, this means that a phone with an 8.89 cm screen is only approximately 4.57 cm wide by 7.62 cm tall. So there are approximately 105 pixels in a centimeter on an 8.89 cm screen. · Think about the size of your finger tip. If you do not have large hands, think about the size of the fingertip of someone with large hands. Consider that when you are sizing your touch input. Especially consider that when you are spacing two touch targets near one another. You need to judge it for yourself, but items that are next to each other and are each 100x100 should be fine when it comes to selecting items individually. Smaller targets than that are ok provided that you leave space between them. · You want your users to have a pleasant experience. Making touch controls too small or too close to one another will make them nervous about whether they will touch the right target. Take this into account when you plan out your game initially. If possible, do some quick size mockups on an actual phone using colored rectangles that you position and size where you plan to have your game controls. Adjust as necessary. · People do not have transparent hands! Nor are their hands the size of a mouse pointer icon. Consider leaving a dedicated space for input rather than forcing the user to cover up to one-third of the screen with a finger just to play the game. · Another benefit of designing your controls to use a dedicated area is that you’re less likely to have players moving their finger(s) so frantically that they accidentally hit the back button, start button, or search button (many phones have one or more of these on the screen itself – it’s easy to hit one by accident and really annoying if you hit, e.g., the search button and then quickly tap back only to find out that the game didn’t save your progress such that you just wasted all the time you spent playing). · People do not like doing somersaults in order to move something forward with accelerometer-based controls. Test your accelerometer-based controls extensively and get a lot of feedback. Very well-known games from noted publishers have created really bad accelerometer controls and been virtually unplayable as a result. Also be wary of exceptions and other possible failures that the documentation warns about. · When done properly, the accelerometer can add a nice touch to your game (see, e.g. ilomilo where the accelerometer was used to move the background; it added a nice touch without frustrating the user; I also think CarniVale does direct accelerometer controls very well). However, if done poorly, it will make your game an abomination unto the Marketplace. Days, weeks, perhaps even months of development time that you will never get back. I won’t name names; you can search the marketplace for games with terrible reviews and you’ll find them. Graphics · The maximum frame rate is 30 frames per second. This was set as a compromise between battery life and quality. · At least one model of phone is known to have a screen refresh rate that is between 59 and 60 hertz. Because of this, using a fixed time step with a target frame rate of 30 will cause a slight internal delay to build up as the framework is forced to wait slightly for the next refresh. Eventually the delay will get to the point where a draw is skipped in order to recover from the delay. (See Nick's comment below for clarification.) · To deal with that delay, you can either stay with a fixed time step and set the frame rate slightly lower or else you can go to a variable time step and make sure to adjust all of your update data (e.g. player movement distance) to take into account the elapsed time from the last update. A variable time step makes your update logic slightly more complicated but will avoid frame skips entirely. · Currently there are no custom shaders. This might change in the future (there is no hardware limitation preventing it; it simply wasn’t a feature that could be implemented in the time available before launch). · There are five built-in shaders. You can create a lot of nice effects with the built-in shaders. · There is more power on the CPU than there is on the GPU so things you might typically off-load to the GPU will instead make sense to do on the CPU side. · This is a phone. It is not a PC. It is not an Xbox 360. The emulator runs on a PC and uses the full power of your PC. It is very good for testing your code for bugs and doing early prototyping and layout. You should not use it to measure performance. Use actual phone hardware instead. · There are many phone models, each of which has slightly different performance levels for I/O, screen blitting, CPU performance, etc. Do not take your game right to the performance limit on your phone since for some other phones you might be crossing their limits and leaving players with a bad experience. Leave a cushion to account for hardware differences. · Smaller screened phones will have slightly more dots per inch (dpi). Larger screened phones will have slightly less. Either way, the dpi will be much higher than the typical 96 found on most computer screens. Make sure that whoever is doing art for your game takes this into account. · Screens are only required to have 16 bit color (65,536 colors). This is common among smart phones. Using gradients on a 16 bit display can produce an ugly artifact known as banding. Banding is when, rather than a smooth transition from one color to another, you instead see distinct lines. Be careful to avoid this when possible. Banding can be avoided through careful art creation. Its effects can be minimized and even unnoticeable when the texture in question is always moving. You should be careful not to rely on “looks good on my phone” since some phones do have 32-bit displays and thus you’ll find yourself wondering why you’re getting bad reviews that complain about the graphics. Avoid gradients; if you can’t, make sure they are 16-bit safe. Audio · Never rely on sounds as your sole signal to the player that something is happening in the game. They might have the sound off. They might be playing somewhere loud. Etc. · You have to provide controls to disable sound & music. These should be separate. · On at least one model of phone, the volume control API currently has no effect. Players can adjust sound with their hardware volume buttons, but in game selectors simply won’t work. As such, it may not be worth the effort of providing anything beyond on/off switches for sound and music. · MediaPlayer.GameHasControl will return true when a game is hooked up to a PC running Zune. When Zune is running, any attempts to do anything (beyond check GameHasControl) with MediaPlayer will cause an exception to be thrown. If this exception is thrown, catch it and disable music. Exceptions take time to propagate; you don’t want one popping up in every single run of your game’s Update method. · Remember that players can already be listening to music or using the FM radio. In this case GameHasControl will be false and you should handle this appropriately. You can, alternately, ask the player for permission to stop their current music and play your music instead, but the (current) requirement that you restore their music when done is very hard (if not impossible) to deal with. · You can still play sound effects even when the game doesn’t have control of the music, but don’t think this is a backdoor to playing music. Your game will fail certification if your “sound effect” seems to be more like music in scope and length.

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  • Recreating OMS instances in a HA environment when instances on all nodes are lost

    - by rnigam
    Oracle highly recommends deploying EM in a HA environment. The best practices for HA deployments, backup and housekeeping of your Enterprise Manager environment are documented in the Oracle Enterprise Manager Advanced Configuration Guide. It is imperative that there is a good disaster recovery plan in place for your EM deployment. In this post I want to talk about a customer who failed to do the correct planning and housekeeping for EM and landed in a situation where we the all the OMSes were nearly blown away had we not jumped to help. We recently hit an issue at a customer site where we had a two node OMS setup of the Enterprise Manager and a RAC Database being used as the EM repository. An accidental delete of the OMS oracle home left us with a single node deployment. While we were trying to figure out a possible path to recover the first node, the second node was rebooted under a maintenance window. What followed was a complete site outage as the Admin and managed servers would not start on either of the nodes. In my situation there were - No backups of the Oracle Homes from any node - No OMS Configuration snapshots (created using the “emctl exportconfig oms” command) and the instance home was completely lost on node 1 which also had the Admin Server  We did however have: - A copy of the emkey.ora that I found under the OMS_ORACLE_HOME/ of the second node (NOTE: it is a bad practice to have your emkey present under the OMS Oracle home directory on the same server as the OMS. The backup of the emkey should be maintained on some other server. In this case however it was a savior in my situation since there were no backups - The oms oracle home on the second node but missing a number of files and had a number of changes done to the files in the home. There were a number of attempts to start the server by modifying various files based on the Weblogic server logs to have atleast node up and running but all of them failed. Here is how you can recover from this scenario: Follow these steps: STEP 1: Check status of emkey.ora Check whether the emkey exists is present in the EM repository or not. Run the following command: $OMS_ORACLE_HOME/bin/emctl status emkey If the output is something like this below then you are good to go and the key is present in the repository ./emctl status emkey Oracle Enterprise Manager 11g Release 1 Grid Control Copyright (c) 1996, 2010 Oracle Corporation. All rights reserved. Enter Enterprise Manager Root (SYSMAN) Password : The EMKey is configured properly. Here are the messages that you might see as the emctl status emkey output depending upon whether the EM Admin Server is up and if the key is configured properly: Case1:  AdminServer is up, emkey is proper in CredStore & not in repos. This is same as the output of the command shown above:The EMKey is configured properly Case 2: AdminServer is up, emkey is proper in CredStore & exists in repos:The EMKey is configured properly, but is not secure. Secure the EMKey by running "emctl config emkey -remove_from_repos".Case 3: AdminServer is down or emkey is corrupted in CredStore) & (emkey exists in repos): The EMKey exists in the Management Repository, but is not configured properly or is corrupted in the credential store.Configure the EMKey by running "emctl config emkey -copy_to_credstore".Case 4: (AdminServer is down or emkey is corrupted in CredStore) & (emkey does not exist in repos): The EMKey is not configured properly or is corrupted in the credential store and does not exist in the Management Repository. To correct the problem:1) Get the backed up emkey.ora file.2) Configure the emkey by running "emctl config emkey -copy_to_credstore_from_file". If not the key was not secured properly, we will have to be put in the repository before proceeding. Look at the next step 2 for doing this There may be cases (like mine) where running emctl may give errors like the following: $OMS_ORACLE_HOME/bin/emctl status emkey Exception in thread “Main Thread” java.lang.NoClassDefFoundError: oracle/security/pki/OracleWallet At oracle.sysman.emctl.config.oms.EMKeyCmds.main (EMKeyCmds.java:658) Just move to the next step to put the key back in the repository STEP 2: Put emkey.ora back in the repository Skip this step if your emkey.ora is present in the repository. If not, you need to put the key back in the repository See if you can run the following command (with sample output): $OMS_ORACLE_HOME/bin/emctl config emkey –copy_to_repos Oracle Enterprise Manager 11g Release 1 Grid Control Copyright (c) 1996, 2010 Oracle Corporation. All rights reserved. The EMKey has been copied to the Management Repository. This operation will cause the EMKey to become unsecure. After the required operation has been completed, secure the EMKey by running "emctl config emkey -remove_from_repos". Typically the key is present under $OMS_ORACLE_HOME/sysman/config directory before being removed after the install as a best practice. If you hit any errors while running emctl commands like the one mentioned in step 1, jump to step 3 and we will take care of the emkey.ora in Step 5 STEP 3: Get the port information Check for the existing port information in the emd.properties file under EM_INSTANCE_DIRECTORY (typically gc_inst directory right above the Middleware home where you have deployed em. For eg. /u01/app/oracle/product/gc_inst in case your oms home is /u01/app/oracle/product/Middleware/oms11g) In my case I got the information from the emgc.properties present in the gc_inst on the second node. If you can run emctl you may want to try the following command as well $OMS_ORACLE_HOME/bin/emctl status oms –details Note this information as this will be used in the next step STEP 4: Perform cleanup on Node 1 Note the oracle home of the Weblogic and OMS, get the list of applied patches in the homes (using opatch lsinventory command), take a backup copy of the home just in case we need it and then de-install/remove oracle homes, update inventory and cleanup processes on the first node STEP 5: Perform Software Only Installation of OMS on Node 1 Perform Weblogic 10.3.2 installation exactly under the same location as present in the earlier installation. Perform software only installation of the OMS using the following command. This will not run any configuration assistants and bypass all user interface validations runInstaller –noconfig -validationaswarnings Select the “Additional OMS” option while performing the installation. Provide the same path for OMS and Instance directories like the previous installation Use the port information collected in Step 3 while performing the installation. Once the installation is complete run the allroot.sh script to complete the binary deployment STEP 6: Apply one-off patches At this point you can apply any patches to the OMS Oracle Home previously. You only need to run opatch to install the patch in the home and not required to run the SQLs STEP 7: Copy EM key This step is only required if you were not able to use emctl command to put the emkey back into the EM repository in STEP 2 Copy the emkey.ora file of the old installation you have under $OMS_ORACLE_HOME/sysman/config directory of the newly installed OMS STEP 8: Configure Grid Control Domain Run the following command to configure the EM domain and OMS. Note that you need to use a different GC Domain name than what you used earlier. For example I have used GCDOMAIN11 as the new domain name when my previous domain name was GCDOMAIN $OMS_ORACLE_HOME/bin/omsca new –AS_USERNAME weblogic –EM_DOMAIN_NAME GCDOMAIN11 –NM_USER nodemanager -nostart This command as shown below will prompt for a number of inputs like Admin Server hostname, port, password, etc. Verify if the defaults shown are correct by pressing enter or provide a new value STEP 9: Run Add-ON Configuration Assistant After this step run the following add-on configuration assistant. This was used in my case to configure the virtualization add-on $OMS_ORACLE_HOME/addonca -oui -omsonly -name vt -install gc STEP 10: Start the OMS Now start the OMS using $OMS_ORACLE_HOME/bin/emctl start oms In a multi-node setup like mine you would either have a software load balancer or DNS round robin (using a virtual host name that resolves to one of multiple OMS hostnames) being used for load balancing. Secure the OMS against the SLB or DNS virtual hostname using the following $ OMS_HOME/bin/emctl secure oms -host slb.example.com -secure_port 1159 -slb_port 1159 -slb_console_port 443 STEP 11: Configure the Agent From the $AGENT_ORACLE_HOME/bin run the ./agentca –f At this point you should have your OMS on node 1 fully re-covered. Clean up node 2 and use the normal Additional OMS installation process documented in the official installation guide to add the additional OMS on node 2 Summary It took us nearly a little over two days to completely recover the environment with some other non-EM related issues that hit us along the way as well. In the end a situation like this could have been completely avoided had the proper housekeeping and backup of the Enterprise Manager Deployment been done in the first place. This is going to a topic that we cover in the next post. In the meantime please do refer to the Oracle Enterprise Manager Advanced Configuration Guide for planning your EM installation, backup and housekeeping procedures. This can be found here: http://download.oracle.com/docs/cd/E11857_01/index.htm Thanks This post would not have been possible without Raj Aggarwal, Prasad Chebrolu and Ravikumar Basa who helped to recover the environment and provided all the support we needed

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  • Video Card needs additional mounting support

    - by Sean
    We are creating a fairly decent system with crossfire. The issue we are having is apparently the video cards are physically to heavy to be supported purely by the motherboard and case mounts. Whenever we have the case horizontal everything works fine however when we bring the case upright the computer stops working. Relevant Info: Video Cards (2) - Sapphire 5870 (http://www.newegg.com/Product/Product.aspx?Item=N82E16814102856) MotherBoard - Asus M4A79T Deluxe Case - CoolerMaster 922 HAF To me it looks as if an additional support bracket that could hold up the end of the card would be required (included with the viideo cards!!!) Does anyone have any experience with this.

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  • Sharing Bandwidth and Prioritizing Realtime Traffic via HTB, Which Scenario Works Better?

    - by Mecki
    I would like to add some kind of traffic management to our Internet line. After reading a lot of documentation, I think HFSC is too complicated for me (I don't understand all the curves stuff, I'm afraid I will never get it right), CBQ is not recommend, and basically HTB is the way to go for most people. Our internal network has three "segments" and I'd like to share bandwidth more or less equally between those (at least in the beginning). Further I must prioritize traffic according to at least three kinds of traffic (realtime traffic, standard traffic, and bulk traffic). The bandwidth sharing is not as important as the fact that realtime traffic should always be treated as premium traffic whenever possible, but of course no other traffic class may starve either. The question is, what makes more sense and also guarantees better realtime throughput: Creating one class per segment, each having the same rate (priority doesn't matter for classes that are no leaves according to HTB developer) and each of these classes has three sub-classes (leaves) for the 3 priority levels (with different priorities and different rates). Having one class per priority level on top, each having a different rate (again priority won't matter) and each having 3 sub-classes, one per segment, whereas all 3 in the realtime class have highest prio, lowest prio in the bulk class, and so on. I'll try to make this more clear with the following ASCII art image: Case 1: root --+--> Segment A | +--> High Prio | +--> Normal Prio | +--> Low Prio | +--> Segment B | +--> High Prio | +--> Normal Prio | +--> Low Prio | +--> Segment C +--> High Prio +--> Normal Prio +--> Low Prio Case 2: root --+--> High Prio | +--> Segment A | +--> Segment B | +--> Segment C | +--> Normal Prio | +--> Segment A | +--> Segment B | +--> Segment C | +--> Low Prio +--> Segment A +--> Segment B +--> Segment C Case 1 Seems like the way most people would do it, but unless I don't read the HTB implementation details correctly, Case 2 may offer better prioritizing. The HTB manual says, that if a class has hit its rate, it may borrow from its parent and when borrowing, classes with higher priority always get bandwidth offered first. However, it also says that classes having bandwidth available on a lower tree-level are always preferred to those on a higher tree level, regardless of priority. Let's assume the following situation: Segment C is not sending any traffic. Segment A is only sending realtime traffic, as fast as it can (enough to saturate the link alone) and Segment B is only sending bulk traffic, as fast as it can (again, enough to saturate the full link alone). What will happen? Case 1: Segment A-High Prio and Segment B-Low Prio both have packets to send, since A-High Prio has the higher priority, it will always be scheduled first, till it hits its rate. Now it tries to borrow from Segment A, but since Segment A is on a higher level and Segment B-Low Prio has not yet hit its rate, this class is now served first, till it also hits the rate and wants to borrow from Segment B. Once both have hit their rates, both are on the same level again and now Segment A-High Prio is going to win again, until it hits the rate of Segment A. Now it tries to borrow from root (which has plenty of traffic spare, as Segment C is not using any of its guaranteed traffic), but again, it has to wait for Segment B-Low Prio to also reach the root level. Once that happens, priority is taken into account again and this time Segment A-High Prio will get all the bandwidth left over from Segment C. Case 2: High Prio-Segment A and Low Prio-Segment B both have packets to send, again High Prio-Segment A is going to win as it has the higher priority. Once it hits its rate, it tries to borrow from High Prio, which has bandwidth spare, but being on a higher level, it has to wait for Low Prio-Segment B again to also hit its rate. Once both have hit their rate and both have to borrow, High Prio-Segment A will win again until it hits the rate of the High Prio class. Once that happens, it tries to borrow from root, which has again plenty of bandwidth left (all bandwidth of Normal Prio is unused at the moment), but it has to wait again until Low Prio-Segment B hits the rate limit of the Low Prio class and also tries to borrow from root. Finally both classes try to borrow from root, priority is taken into account, and High Prio-Segment A gets all bandwidth root has left over. Both cases seem sub-optimal, as either way realtime traffic sometimes has to wait for bulk traffic, even though there is plenty of bandwidth left it could borrow. However, in case 2 it seems like the realtime traffic has to wait less than in case 1, since it only has to wait till the bulk traffic rate is hit, which is most likely less than the rate of a whole segment (and in case 1 that is the rate it has to wait for). Or am I totally wrong here? I thought about even simpler setups, using a priority qdisc. But priority queues have the big problem that they cause starvation if they are not somehow limited. Starvation is not acceptable. Of course one can put a TBF (Token Bucket Filter) into each priority class to limit the rate and thus avoid starvation, but when doing so, a single priority class cannot saturate the link on its own any longer, even if all other priority classes are empty, the TBF will prevent that from happening. And this is also sub-optimal, since why wouldn't a class get 100% of the line's bandwidth if no other class needs any of it at the moment? Any comments or ideas regarding this setup? It seems so hard to do using standard tc qdiscs. As a programmer it was such an easy task if I could simply write my own scheduler (which I'm not allowed to do).

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  • C#/.NET &ndash; Finding an Item&rsquo;s Index in IEnumerable&lt;T&gt;

    - by James Michael Hare
    Sorry for the long blogging hiatus.  First it was, of course, the holidays hustle and bustle, then my brother and his wife gave birth to their son, so I’ve been away from my blogging for two weeks. Background: Finding an item’s index in List<T> is easy… Many times in our day to day programming activities, we want to find the index of an item in a collection.  Now, if we have a List<T> and we’re looking for the item itself this is trivial: 1: // assume have a list of ints: 2: var list = new List<int> { 1, 13, 42, 64, 121, 77, 5, 99, 132 }; 3:  4: // can find the exact item using IndexOf() 5: var pos = list.IndexOf(64); This will return the position of the item if it’s found, or –1 if not.  It’s easy to see how this works for primitive types where equality is well defined.  For complex types, however, it will attempt to compare them using EqualityComparer<T>.Default which, in a nutshell, relies on the object’s Equals() method. So what if we want to search for a condition instead of equality?  That’s also easy in a List<T> with the FindIndex() method: 1: // assume have a list of ints: 2: var list = new List<int> { 1, 13, 42, 64, 121, 77, 5, 99, 132 }; 3:  4: // finds index of first even number or -1 if not found. 5: var pos = list.FindIndex(i => i % 2 == 0);   Problem: Finding an item’s index in IEnumerable<T> is not so easy... This is all well and good for lists, but what if we want to do the same thing for IEnumerable<T>?  A collection of IEnumerable<T> has no indexing, so there’s no direct method to find an item’s index.  LINQ, as powerful as it is, gives us many tools to get us this information, but not in one step.  As with almost any problem involving collections, there are several ways to accomplish the same goal.  And once again as with almost any problem involving collections, the choice of the solution somewhat depends on the situation. So let’s look at a few possible alternatives.  I’m going to express each of these as extension methods for simplicity and consistency. Solution: The TakeWhile() and Count() combo One of the things you can do is to perform a TakeWhile() on the list as long as your find condition is not true, and then do a Count() of the items it took.  The only downside to this method is that if the item is not in the list, the index will be the full Count() of items, and not –1.  So if you don’t know the size of the list beforehand, this can be confusing. 1: // a collection of extra extension methods off IEnumerable<T> 2: public static class EnumerableExtensions 3: { 4: // Finds an item in the collection, similar to List<T>.FindIndex() 5: public static int FindIndex<T>(this IEnumerable<T> list, Predicate<T> finder) 6: { 7: // note if item not found, result is length and not -1! 8: return list.TakeWhile(i => !finder(i)).Count(); 9: } 10: } Personally, I don’t like switching the paradigm of not found away from –1, so this is one of my least favorites.  Solution: Select with index Many people don’t realize that there is an alternative form of the LINQ Select() method that will provide you an index of the item being selected: 1: list.Select( (item,index) => do something here with the item and/or index... ) This can come in handy, but must be treated with care.  This is because the index provided is only as pertains to the result of previous operations (if any).  For example: 1: // assume have a list of ints: 2: var list = new List<int> { 1, 13, 42, 64, 121, 77, 5, 99, 132 }; 3:  4: // you'd hope this would give you the indexes of the even numbers 5: // which would be 2, 3, 8, but in reality it gives you 0, 1, 2 6: list.Where(item => item % 2 == 0).Select((item,index) => index); The reason the example gives you the collection { 0, 1, 2 } is because the where clause passes over any items that are odd, and therefore only the even items are given to the select and only they are given indexes. Conversely, we can’t select the index and then test the item in a Where() clause, because then the Where() clause would be operating on the index and not the item! So, what we have to do is to select the item and index and put them together in an anonymous type.  It looks ugly, but it works: 1: // extensions defined on IEnumerable<T> 2: public static class EnumerableExtensions 3: { 4: // finds an item in a collection, similar to List<T>.FindIndex() 5: public static int FindIndex<T>(this IEnumerable<T> list, Predicate<T> finder) 6: { 7: // if you don't name the anonymous properties they are the variable names 8: return list.Select((item, index) => new { item, index }) 9: .Where(p => finder(p.item)) 10: .Select(p => p.index + 1) 11: .FirstOrDefault() - 1; 12: } 13: }     So let’s look at this, because i know it’s convoluted: First Select() joins the items and their indexes into an anonymous type. Where() filters that list to only the ones matching the predicate. Second Select() picks the index of the matches and adds 1 – this is to distinguish between not found and first item. FirstOrDefault() returns the first item found from the previous clauses or default (zero) if not found. Subtract one so that not found (zero) will be –1, and first item (one) will be zero. The bad thing is, this is ugly as hell and creates anonymous objects for each item tested until it finds the match.  This concerns me a bit but we’ll defer judgment until compare the relative performances below. Solution: Convert ToList() and use FindIndex() This solution is easy enough.  We know any IEnumerable<T> can be converted to List<T> using the LINQ extension method ToList(), so we can easily convert the collection to a list and then just use the FindIndex() method baked into List<T>. 1: // a collection of extension methods for IEnumerable<T> 2: public static class EnumerableExtensions 3: { 4: // find the index of an item in the collection similar to List<T>.FindIndex() 5: public static int FindIndex<T>(this IEnumerable<T> list, Predicate<T> finder) 6: { 7: return list.ToList().FindIndex(finder); 8: } 9: } This solution is simplicity itself!  It is very concise and elegant and you need not worry about anyone misinterpreting what it’s trying to do (as opposed to the more convoluted LINQ methods above). But the main thing I’m concerned about here is the performance hit to allocate the List<T> in the ToList() call, but once again we’ll explore that in a second. Solution: Roll your own FindIndex() for IEnumerable<T> Of course, you can always roll your own FindIndex() method for IEnumerable<T>.  It would be a very simple for loop which scans for the item and counts as it goes.  There’s many ways to do this, but one such way might look like: 1: // extension methods for IEnumerable<T> 2: public static class EnumerableExtensions 3: { 4: // Finds an item matching a predicate in the enumeration, much like List<T>.FindIndex() 5: public static int FindIndex<T>(this IEnumerable<T> list, Predicate<T> finder) 6: { 7: int index = 0; 8: foreach (var item in list) 9: { 10: if (finder(item)) 11: { 12: return index; 13: } 14:  15: index++; 16: } 17:  18: return -1; 19: } 20: } Well, it’s not quite simplicity, and those less familiar with LINQ may prefer it since it doesn’t include all of the lambdas and behind the scenes iterators that come with deferred execution.  But does having this long, blown out method really gain us much in performance? Comparison of Proposed Solutions So we’ve now seen four solutions, let’s analyze their collective performance.  I took each of the four methods described above and run them over 100,000 iterations of lists of size 10, 100, 1000, and 10000 and here’s the performance results.  Then I looked for targets at the begining of the list (best case), middle of the list (the average case) and not in the list (worst case as must scan all of the list). Each of the times below is the average time in milliseconds for one execution as computer over the 100,000 iterations: Searches Matching First Item (Best Case)   10 100 1000 10000 TakeWhile 0.0003 0.0003 0.0003 0.0003 Select 0.0005 0.0005 0.0005 0.0005 ToList 0.0002 0.0003 0.0013 0.0121 Manual 0.0001 0.0001 0.0001 0.0001   Searches Matching Middle Item (Average Case)   10 100 1000 10000 TakeWhile 0.0004 0.0020 0.0191 0.1889 Select 0.0008 0.0042 0.0387 0.3802 ToList 0.0002 0.0007 0.0057 0.0562 Manual 0.0002 0.0013 0.0129 0.1255   Searches Where Not Found (Worst Case)   10 100 1000 10000 TakeWhile 0.0006 0.0039 0.0381 0.3770 Select 0.0012 0.0081 0.0758 0.7583 ToList 0.0002 0.0012 0.0100 0.0996 Manual 0.0003 0.0026 0.0253 0.2514   Notice something interesting here, you’d think the “roll your own” loop would be the most efficient, but it only wins when the item is first (or very close to it) regardless of list size.  In almost all other cases though and in particular the average case and worst case, the ToList()/FindIndex() combo wins for performance, even though it is creating some temporary memory to hold the List<T>.  If you examine the algorithm, the reason why is most likely because once it’s in a ToList() form, internally FindIndex() scans the internal array which is much more efficient to iterate over.  Thus, it takes a one time performance hit (not including any GC impact) to create the List<T> but after that the performance is much better. Summary If you’re concerned about too many throw-away objects, you can always roll your own FindIndex() method, but for sheer simplicity and overall performance, using the ToList()/FindIndex() combo performs best on nearly all list sizes in the average and worst cases.    Technorati Tags: C#,.NET,Litte Wonders,BlackRabbitCoder,Software,LINQ,List

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  • Set-Cookie Headers getting stripped in ASP.NET HttpHandlers

    - by Rick Strahl
    Yikes, I ran into a real bummer of an edge case yesterday in one of my older low level handler implementations (for West Wind Web Connection in this case). Basically this handler is a connector for a backend Web framework that creates self contained HTTP output. An ASP.NET Handler captures the full output, and then shoves the result down the ASP.NET Response object pipeline writing out the content into the Response.OutputStream and seperately sending the HttpHeaders in the Response.Headers collection. The headers turned out to be the problem and specifically Http Cookies, which for some reason ended up getting stripped out in some scenarios. My handler works like this: Basically the HTTP response from the backend app would return a full set of HTTP headers plus the content. The ASP.NET handler would read the headers one at a time and then dump them out via Response.AppendHeader(). But I found that in some situations Set-Cookie headers sent along were simply stripped inside of the Http Handler. After a bunch of back and forth with some folks from Microsoft (thanks Damien and Levi!) I managed to pin this down to a very narrow edge scenario. It's easiest to demonstrate the problem with a simple example HttpHandler implementation. The following simulates the very much simplified output generation process that fails in my handler. Specifically I have a couple of headers including a Set-Cookie header and some output that gets written into the Response object.using System.Web; namespace wwThreads { public class Handler : IHttpHandler { /* NOTE: * * Run as a web.config set handler (see entry below) * * Best way is to look at the HTTP Headers in Fiddler * or Chrome/FireBug/IE tools and look for the * WWHTREADSID cookie in the outgoing Response headers * ( If the cookie is not there you see the problem! ) */ public void ProcessRequest(HttpContext context) { HttpRequest request = context.Request; HttpResponse response = context.Response; // If ClearHeaders is used Set-Cookie header gets removed! // if commented header is sent... response.ClearHeaders(); response.ClearContent(); // Demonstrate that other headers make it response.AppendHeader("RequestId", "asdasdasd"); // This cookie gets removed when ClearHeaders above is called // When ClearHEaders is omitted above the cookie renders response.AppendHeader("Set-Cookie", "WWTHREADSID=ThisIsThEValue; path=/"); // *** This always works, even when explicit // Set-Cookie above fails and ClearHeaders is called //response.Cookies.Add(new HttpCookie("WWTHREADSID", "ThisIsTheValue")); response.Write(@"Output was created.<hr/> Check output with Fiddler or HTTP Proxy to see whether cookie was sent."); } public bool IsReusable { get { return false; } } } } In order to see the problem behavior this code has to be inside of an HttpHandler, and specifically in a handler defined in web.config with: <add name=".ck_handler" path="handler.ck" verb="*" type="wwThreads.Handler" preCondition="integratedMode" /> Note: Oddly enough this problem manifests only when configured through web.config, not in an ASHX handler, nor if you paste that same code into an ASPX page or MVC controller. What's the problem exactly? The code above simulates the more complex code in my live handler that picks up the HTTP response from the backend application and then peels out the headers and sends them one at a time via Response.AppendHeader. One of the headers in my app can be one or more Set-Cookie. I found that the Set-Cookie headers were not making it into the Response headers output. Here's the Chrome Http Inspector trace: Notice, no Set-Cookie header in the Response headers! Now, running the very same request after removing the call to Response.ClearHeaders() command, the cookie header shows up just fine: As you might expect it took a while to track this down. At first I thought my backend was not sending the headers but after closer checks I found that indeed the headers were set in the backend HTTP response, and they were indeed getting set via Response.AppendHeader() in the handler code. Yet, no cookie in the output. In the simulated example the problem is this line:response.AppendHeader("Set-Cookie", "WWTHREADSID=ThisIsThEValue; path=/"); which in my live code is more dynamic ( ie. AppendHeader(token[0],token[1[]) )as it parses through the headers. Bizzaro Land: Response.ClearHeaders() causes Cookie to get stripped Now, here is where it really gets bizarre: The problem occurs only if: Response.ClearHeaders() was called before headers are added It only occurs in Http Handlers declared in web.config Clearly this is an edge of an edge case but of course - knowing my relationship with Mr. Murphy - I ended up running smack into this problem. So in the code above if you remove the call to ClearHeaders(), the cookie gets set!  Add it back in and the cookie is not there. If I run the above code in an ASHX handler it works. If I paste the same code (with a Response.End()) into an ASPX page, or MVC controller it all works. Only in the HttpHandler configured through Web.config does it fail! Cue the Twilight Zone Music. Workarounds As is often the case the fix for this once you know the problem is not too difficult. The difficulty lies in tracking inconsistencies like this down. Luckily there are a few simple workarounds for the Cookie issue. Don't use AppendHeader for Cookies The easiest and obvious solution to this problem is simply not use Response.AppendHeader() to set Cookies. Duh! Under normal circumstances in application level code there's rarely a reason to write out a cookie like this:response.AppendHeader("Set-Cookie", "WWTHREADSID=ThisIsThEValue; path=/"); but rather create the cookie using the Response.Cookies collection:response.Cookies.Add(new HttpCookie("WWTHREADSID", "ThisIsTheValue")); Unfortunately, in my case where I dynamically read headers from the original output and then dynamically  write header key value pairs back  programmatically into the Response.Headers collection, I actually don't look at each header specifically so in my case the cookie is just another header. My first thought was to simply trap for the Set-Cookie header and then parse out the cookie and create a Cookie object instead. But given that cookies can have a lot of different options this is not exactly trivial, plus I don't really want to fuck around with cookie values which can be notoriously brittle. Don't use Response.ClearHeaders() The real mystery in all this is why calling Response.ClearHeaders() prevents a cookie value later written with Response.AppendHeader() to fail. I fired up Reflector and took a quick look at System.Web and HttpResponse.ClearHeaders. There's all sorts of resetting going on but nothing that seems to indicate that headers should be removed later on in the request. The code in ClearHeaders() does access the HttpWorkerRequest, which is the low level interface directly into IIS, and so I suspect it's actually IIS that's stripping the headers and not ASP.NET, but it's hard to know. Somebody from Microsoft and the IIS team would have to comment on that. In my application it's probably safe to simply skip ClearHeaders() in my handler. The ClearHeaders/ClearContent was mainly for safety but after reviewing my code there really should never be a reason that headers would be set prior to this method firing. However, if for whatever reason headers do need to be cleared, it's easy enough to manually clear the headers out:private void RemoveHeaders(HttpResponse response) { List<string> headers = new List<string>(); foreach (string header in response.Headers) { headers.Add(header); } foreach (string header in headers) { response.Headers.Remove(header); } response.Cookies.Clear(); } Now I can replace the call the Response.ClearHeaders() and I don't get the funky side-effects from Response.ClearHeaders(). Summary I realize this is a total edge case as this occurs only in HttpHandlers that are manually configured. It looks like you'll never run into this in any of the higher level ASP.NET frameworks or even in ASHX handlers - only web.config defined handlers - which is really, really odd. After all those frameworks use the same underlying ASP.NET architecture. Hopefully somebody from Microsoft has an idea what crazy dependency was triggered here to make this fail. IAC, there are workarounds to this should you run into it, although I bet when you do run into it, it'll likely take a bit of time to find the problem or even this post in a search because it's not easily to correlate the problem to the solution. It's quite possible that more than cookies are affected by this behavior. Searching for a solution I read a few other accounts where headers like Referer were mysteriously disappearing, and it's possible that something similar is happening in those cases. Again, extreme edge case, but I'm writing this up here as documentation for myself and possibly some others that might have run into this. © Rick Strahl, West Wind Technologies, 2005-2012Posted in ASP.NET   IIS7   Tweet !function(d,s,id){var js,fjs=d.getElementsByTagName(s)[0];if(!d.getElementById(id)){js=d.createElement(s);js.id=id;js.src="//platform.twitter.com/widgets.js";fjs.parentNode.insertBefore(js,fjs);}}(document,"script","twitter-wjs"); (function() { var po = document.createElement('script'); po.type = 'text/javascript'; po.async = true; po.src = 'https://apis.google.com/js/plusone.js'; var s = document.getElementsByTagName('script')[0]; s.parentNode.insertBefore(po, s); })();

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  • How to Use the Signature Editor in Outlook 2013

    - by Lori Kaufman
    The Signature Editor in Outlook 2013 allows you to create a custom signature from text, graphics, or business cards. We will show you how to use the various features of the Signature Editor to customize your signatures. To open the Signature Editor, click the File tab and select Options on the left side of the Account Information screen. Then, click Mail on the left side of the Options dialog box and click the Signatures button. For more details, refer to one of the articles mentioned above. Changing the font for your signature is pretty self-explanatory. Select the text for which you want to change the font and select the desired font from the drop-down list. You can also set the justification (left, center, right) for each line of text separately. The drop-down list that reads Automatic by default allows you to change the color of the selected text. Click OK to accept your changes and close the Signatures and Stationery dialog box. To see your signature in an email, click Mail on the Navigation Bar. Click New Email on the Home tab. The Message window displays and your default signature is inserted into the body of the email. NOTE: You shouldn’t use fonts that are not common in your signatures. In order for the recipient to see your signature as you intended, the font you choose also needs to be installed on the recipient’s computer. If the font is not installed, the recipient would see a different font, the wrong characters, or even placeholder characters, which are empty square boxes. Close the Message window using the File tab or the X button in the upper, right corner of the Message window. You can save it as a draft if you want, but it’s not necessary. If you decide to use a font that is not common, a better way to do so would be to create a signature as an image, or logo. Create your image or logo in an image editing program making it the exact size you want to use in your signature. Save the image in a file size as small as possible. The .jpg format works well for pictures, the .png format works well for detailed graphics, and the .gif format works well for simple graphics. The .gif format generally produces the smallest files. To insert an image in your signature, open the Signatures and Stationery dialog box again. Either delete the text currently in the editor, if any, or create a new signature. Then, click the image button on the editor’s toolbar. On the Insert Picture dialog box, navigate to the location of your image, select the file, and click Insert. If you want to insert an image from the web, you must enter the full URL for the image in the File name edit box (instead of the local image filename). For example, http://www.somedomain.com/images/signaturepic.gif. If you want to link to the image at the specified URL, you must also select Link to File from the Insert drop-down list to maintain the URL reference. The image is inserted into the Edit signature box. Click OK to accept your changes and close the Signatures and Stationery dialog box. Create a new email message again. You’ll notice the image you inserted into the signature displays in the body of the message. Close the Message window using the File tab or the X button in the upper, right corner of the Message window. You may want to put a link to a webpage or an email link in your signature. To do this, open the Signatures and Stationery dialog box again. Enter the text to display for the link, highlight the text, and click the Hyperlink button on the editor’s toolbar. On the Insert Hyperlink dialog box, select the type of link from the list on the left and enter the webpage, email, or other type of address in the Address edit box. You can change the text that will display in the signature for the link in the Text to display edit box. Click OK to accept your changes and close the dialog box. The link displays in the editor with the default blue, underlined text. Click OK to accept your changes and close the Signatures and Stationery dialog box. Here’s an example of an email message with a link in the signature. Close the Message window using the File tab or the X button in the upper, right corner of the Message window. You can also insert your contact information into your signature as a Business Card. To do so, click Business Card on the editor’s toolbar. On the Insert Business Card dialog box, select the contact you want to insert as a Business Card. Select a size for the Business Card image from the Size drop-down list. Click OK. The Business Card image displays in the Signature Editor. Click OK to accept your changes and close the Signatures and Stationery dialog box. When you insert a Business Card into your signature, the Business Card image displays in the body of the email message and a .vcf file containing your contact information is attached to the email. This .vcf file can be imported into programs like Outlook that support this format. Close the Message window using the File tab or the X button in the upper, right corner of the Message window. You can also insert your Business Card into your signature without the image or without the .vcf file attached. If you want to provide recipients your contact info in a .vcf file, but don’t want to attach it to every email, you can upload the .vcf file to a location on the internet and add a link to the file, such as “Get my vCard,” in your signature. NOTE: If you want to edit your business card, such as applying a different template to it, you must select a different View other than People for your Contacts folder so you can open the full contact editing window.     

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  • Can I override fonts installed by ttf-mscorefonts-installer, prefer Liberation fonts?

    - by conner_bw
    I had to apt-get install ttf-mscorefonts-installer on Ubuntu 12.04/12.10. The short version is I need to pipe PDF files out of an application that requires these fonts for certain glyphs. The problem, after running this command, is that the fonts in my web browser (and some java apps) are now "ugly." Obviously this is a subjective opinion but it is the one I hold. I want the old fonts back for most cases (Liberation, DejaVu, Ubuntu, ...). I'm not sure how best to describe this but here's an example: Example CSS in Webbrowser font-family: Verdana,Arial,sans-serif; Without ttf-mscorefonts-installer (Case 1): $ fc-match Verdana LiberationSans-Regular.ttf: "Liberation Sans" "Regular" $ fc-match Arial LiberationSans-Regular.ttf: "Liberation Sans" "Regular" $ fc-match sans-serif LiberationSans-Regular.ttf: "Liberation Sans" "Regular"` With ttf-mscorefonts-installer (Case 2): $ fc-match Verdana Verdana.ttf: "Verdana" "Normal" $ fc-match Arial Arial.ttf: "Arial" "Normal" $ fc-match sans-serif LiberationSans-Regular.ttf: "Liberation Sans" "Regular"` I want (Case 1). Optionally, I want the fonts in (Case 2) not to look "ugly" IE. they are more jagged, less smooth than their free alternatives in my web browsers. Is this possible?

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  • Arbitrary Rotation about a Sphere

    - by Der
    I'm coding a mechanic which allows a user to move around the surface of a sphere. The position on the sphere is currently stored as theta and phi, where theta is the angle between the z-axis and the xz projection of the current position (i.e. rotation about the y axis), and phi is the angle from the y-axis to the position. I explained that poorly, but it is essentially theta = yaw, phi = pitch Vector3 position = new Vector3(0,0,1); position.X = (float)Math.Sin(phi) * (float)Math.Sin(theta); position.Y = (float)Math.Sin(phi) * (float)Math.Cos(theta); position.Z = (float)Math.Cos(phi); position *= r; I believe this is accurate, however I could be wrong. I need to be able to move in an arbitrary pseudo two dimensional direction around the surface of a sphere at the origin of world space with radius r. For example, holding W should move around the sphere in an upwards direction relative to the orientation of the player. I believe I should be using a Quaternion to represent the position/orientation on the sphere, but I can't think of the correct way of doing it. Spherical geometry is not my strong suit. Essentially, I need to fill the following block: public void Move(Direction dir) { switch (dir) { case Direction.Left: // update quaternion to rotate left break; case Direction.Right: // update quaternion to rotate right break; case Direction.Up: // update quaternion to rotate upward break; case Direction.Down: // update quaternion to rotate downward break; } }

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  • Google respond differently to two identical nginx setups and 200 codes; any ideas?

    - by Yuji Tomita
    I'm rather confused... I have a linode.com VPS which has been cloned recently, so the settings are the same between nginx servers. One lives on a dev subdomain, one on a www. I'm trying to run a google experiment on my live server, which claims: Web server rejects utm_expid. Your server doesn't support added query arguments in URLs. My logs show on the dev server where it works: 74.125.186.32 - - [13/Sep/2012:13:33:45 -0700] "GET /product/iphone-case/?utm_expid=25706866-0 HTTP/1.1" 200 12521 "-" "Google_Analytics_Content_Experiments 74.125.186.32 - - [13/Sep/2012:13:33:45 -0700] "GET /product/iphone-case/?ab_reviews=True&utm_expid=25706866-0 HTTP/1.1" 200 14679 "-" "Google_Analytics_Content_Experiments My production server shows google making a second request. 74.125.186.41 - - [13/Sep/2012:13:34:49 -0700] "GET /product/iphone-case/?ab_reviews=on&utm_expid=25706866-1 HTTP/1.1" 200 12104 "-" "Google_Analytics_Content_Experiments 74.125.186.41 - - [13/Sep/2012:13:34:49 -0700] "GET /product/iphone-case/?utm_expid=25706866-1 HTTP/1.1" 200 12122 "-" "Google_Analytics_Content_Experiments 74.125.186.41 - - [13/Sep/2012:13:34:49 -0700] "GET /product/iphone-case/ <--- A second request for some reason. HTTP/1.1" 200 12522 "-" "Google_Analytics_Content_Experiments I'm not sure how google determines why it needs to send a second request without the querystring. The original request has clearly sent a 200 OK status response. Does anybody have any suggestions where to look next? The HTML (compared by diff) on the two pages is exactly the same.

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  • White Paper on Analysis Services Tabular Large-scale Solution #ssas #tabular

    - by Marco Russo (SQLBI)
    Since the first beta of Analysis Services 2012, I worked with many companies designing and implementing solutions based on Analysis Services Tabular. I am glad that Microsoft published a white paper about a case-study using one of these scenarios: An Analysis Services Case Study: Using Tabular Models in a Large-scale Commercial Solution. Alberto Ferrari is the author of the white paper and many people contributed to it. The final result is a very technical document based on a case study, which provides a level of detail that I don’t see often in other case studies (which are usually more marketing-oriented). This white paper has the following structure: Requirements (data model, capacity planning, client tool) Options considered (SQL Server Columnstore Indexes, SSAS Multidimensional, SSAS Tabular) Data Model optimizations (memory compression, query performance, scalability) Partitioning and Processing strategy for near real-time latency Hardware selection (NUMA analysis, Azure VM tests) Scalability tests (estimation of maximum users per node) If you are in charge of evaluating Tabular as analytical engine, or if you have to design your solution based on Tabular, this white paper is a must read. But if you just want to increase your knowledge of Analysis Services, you will find a lot of useful technical information. That said, my favorite quote of the document is the following one, funny but true: […] After several trials, the clear winner was a video gaming machine that one guy on the team used at home. That computer outperformed any available server, running twice as fast as the server-class machines we had in house. At that point, it was clear that the criteria for choosing the server would have to be expanded a bit, simply because it would have been impossible to convince the boss to build a cluster of gaming machines and trust it to serve our customers.  But, honestly, if a business has the flexibility to buy gaming machines (assuming the machines can handle capacity) – do this. Owen Graupman, inContact I want to write a longer discussion about how companies are adopting Tabular in scenarios where it is the hidden engine of a more complex solution (and not the classical “BI system”), because it is more frequent than you might expect (and has several advantages over many alternative approaches).

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  • Wubi shows error

    - by Quirk
    I tried installing Ubuntu 12.04 using Wubi and it just doesn't work, every time, without fail. I had the following scenarios: 1. I downloaded only wubi.exe and ran it. The wubi installer started downloading the amd64.iso using torrent. But when there were just about 40 secs left to download, it shows an error 404: File not found. 2. I downloaded the iso file seperately and put it in the same folder as the wubi.exe. Now there are two cases: a. Offline: wubi says it could not download the metalinks file and hence cannot download the iso. So I download the meta files separately and place them in the same directory. wubi shows the same error again. b. Online: wubi works in same way as in case 1. and the same problem occurs as in case 1. In short wubi doesn't recognize the already downloaded iso in the directory at all. 3. I burn the iso into a Cd and run it. The same thing occurs as in Case 2. Just in case that you know, I installed SP3 for win XP just before using wubi. While Windows is running alright, is it possible that its causing conflicts for wubi?

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  • Delaying a Foreach loop half a second

    - by Sigh-AniDe
    I have created a game that has a ghost that mimics the movement of the player after 10 seconds. The movements are stored in a list and i use a foreach loop to go through the commands. The ghost mimics the movements but it does the movements way too fast, in split second from spawn time it catches up to my current movement. How do i slow down the foreach so that it only does a command every half a second? I don't know how else to do it. Please help this is what i tried : The foreach runs inside the update method DateTime dt = DateTime.Now; foreach ( string commandDirection in ghostMovements ) { int mapX = ( int )( ghostPostition.X / scalingFactor ); int mapY = ( int )( ghostPostition.Y / scalingFactor ); // If the dt is the same as current time if ( dt == DateTime.Now ) { if ( commandDirection == "left" ) { switch ( ghostDirection ) { case ghostFacingUp: angle = 1.6f; ghostDirection = ghostFacingRight; Program.form.direction = ""; dt.AddMilliseconds( 500 );// add half a second to dt break; case ghostFacingRight: angle = 3.15f; ghostDirection = ghostFacingDown; Program.form.direction = ""; dt.AddMilliseconds( 500 ); break; case ghostFacingDown: angle = -1.6f; ghostDirection = ghostFacingLeft; Program.form.direction = ""; dt.AddMilliseconds( 500 ); break; case ghostFacingLeft: angle = 0.0f; ghostDirection = ghostFacingUp; Program.form.direction = ""; dt.AddMilliseconds( 500 ); break; } } } }

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  • Best practices: Ajax and server side scripting with stored procedures

    - by Luka Milani
    I need to rebuild an old huge website and probably to port everyting to ASP.NET and jQuery and I would like to ask for some suggestion and tips. Actually the website uses: Ajax (client site with prototype.js) ASP (vb script server side) SQL Server 2005 IIS 7 as web server This website uses hundred of stored procedures and the requests are made by an ajax call and only 1 ASP page that contain an huge select case Shortly an example: JAVASCRIPT + PROTOTYPE: var data = { action: 'NEWS', callback: 'doNews', param1: $('text_example').value, ......: ..........}; AjaxGet(data); // perform a call using another function + prototype SERVER SIDE ASP: <% ...... select case request("Action") case "NEWS" With cmmDB .ActiveConnection = Conn .CommandText = "sp_NEWS_TO_CALL_for_example" .CommandType = adCmdStoredProc Set par0DB = .CreateParameter("Param1", adVarchar, adParamInput,6) Set par1DB = .CreateParameter(".....", adInteger, adParamInput) ' ........ ' can be more parameters .Parameters.Append par0DB .Parameters.Append par1DB par0DB.Value = request("Param1") par1DB.Value = request(".....") set rs=cmmDB.execute RecodsetToJSON rs, jsa ' create JSON response using a sub End With .... %> So as you can see I have an ASP page that has a lot of CASE and this page answers to all the ajax request in the site. My question are: Instead of having many CASES is it possible to create dynamic vb code that parses the ajax request and creates dynamically the call to the desired SP (also implementing the parameters passed by JS)? What is the best approach to handle situations like this, by using the advantages of .Net + protoype or jQuery? How the big sites handle situation like this? Do they do it by creating 1 page for request? Thanks in advance for suggestion, direction and tips.

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  • Rotation and translation like in GTA 1 OpenGL

    - by user1876377
    Okay, so I have a figure in XZ plain. I want to move it forward/backward and rotate at it's own Y axis, then move forward again in the rotation's direction, like the character in GTA 1. Code so far: Init: spaceship_position = glm::vec3(0,0,0); spaceship_rotation = glm::vec3(0,0,0); spaceship_scale = glm::vec3(1, 1, 1); Draw: glm::mat4 transform = glm::scale<float>(spaceship_scale) * glm::rotate<float>(spaceship_rotation.x, 1, 0, 0) * glm::rotate<float>(spaceship_rotation.y, 0, 1, 0) * glm::rotate<float>(spaceship_rotation.z, 0, 0, 1) * glm::translate<float>(spaceship_position); drawMesh(spaceship, texture, transform); Update: switch (key.keysym.sym) { case SDLK_UP: spaceship_position.z += 0.1; break; case SDLK_DOWN: spaceship_position.z -= 0.1; break; case SDLK_LEFT: spaceship_rotation.y += 1; break; case SDLK_RIGHT: spaceship_rotation.y -= 1; break; } So this only moves on the Z axis, but how can I move the object on both Z and X axis where the object is facing?

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  • Sql-server-2008 client Access license

    - by thushya
    Hi, case 1 : i have one user makes 10 connection from single computer, maximum number of connection at a given time = 10, what is the number CAL i need here ? case 2 : i have 10 users have access to only 1 computer, 10 user connect from single computer - maximum connection at any given time = 1, what is the number CAL i need here ? case 3 : i have 10 users using 10 computers, all 10 are making total of 5 connection maximum in any given time, what is the number of CAL i need here ? Thanks.

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  • Parsing mathematical experssions with two values that have parenthesis and minus signs

    - by user45921
    I'm trying to parse equations like these which only has two values or the square root of a certain value from a text file: 100+100 -100-100 -(100)+(-100) sqrt(100) by the minues signs, parenthesis and the operator symbol in the middle and the square root, and I have no idea how to start off... I've got the file part done and the simple calculation parts except that I couldnt get my program to solve the equations in the above. #include <stdio.h> #include <string.h> #include <stdlib.h> #include <math.h> main(){ FILE *fp; char buff[255], sym,sym2,del1,del2,del3,del4; double num1, num2; int ret; fp = fopen("input.txt","r"); while(fgets(buff,sizeof(buff),fp)!=NULL){ char *tok = buff; sscanf(tok,"%lf%c%lf",&num1,&sym,&num2); switch(sym){ case '+': printf("%lf\n", num1+num2); break; case '-': printf("%lf\n", num1-num2); break; case '*': printf("%lf\n", num1*num2); break; case '/': printf("%lf\n", num1/num2); break; default: printf("The input value is not correct\n"); break; } } fclose(fp); } that is what have I written for the other basic operations without parenthesis and the minus sign for the second value and it works great for the simple ones. I'm using a switch method to calculate the add, sub, mul and divide but I'm not sure how to properly use the sscanf function (if I am not using it properly) or if there is another way using a function like strtok to properly parse the parenthesis and the minus signs. Any kind help?

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  • Why do we have to use break in switch

    - by trejder
    Who decided, and basing on what concepts, that switch construction (in many languages) has to be, like it is? Why do we have to use break in each statement? Why do we have to write something like this: switch(a) { case 1: result = 'one'; break; case 2: result = 'two'; break; default: result = 'not determined'; break; } I've noticed this construction in PHP and JS, but there are probably many other languages that uses it. If switch is an alternative of if, why we can't use the same construction for switch, as for if? I.e.: switch(a) { case 1: { result = 'one'; } case 2: { result = 'two'; } default: { result = 'not determined'; } } It is said, that break prevents execution of a blocks following current one. But, does someone really run into situation, where there was any need for execution of current block and following ones? I didn't. For me, break is always there. In every block. In every code.

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  • What is the value to checking in broken unit tests?

    - by Adam W.
    While there are ways of keeping unit tests from being executed, what is the value of checking in broken unit tests? I will use a simple example. Case sensitivity. The current code is Case Sensitive. A valid input into the method is "Cat" and it would return an enum of Animal.Cat. However, the desired functionality of the method should not be case sensitive. So if the method described was passed "cat" it could possibly return something like Animal.Null instead of Animal.Cat and the unit test would fail. Though a simple code change would make this work, a more complex issue may take weeks to fix, but identifying the bug with a unit test could be a less complex task. The application currently being analyzed has 4 years of code that "works". However, recent discussions regarding unit tests has found flaws in the code. Some just need explicit implementation documentation (ex. case sensitive or not), or code that does not execute the bug based on how it is currently called. But unit tests can be created executing specific scenarios that will cause the bug to be seen and are valid inputs. What is the value of checking in unit tests that exercise the bug until someone can get around to fixing the code? Should this unit test be flagged with ignore, priority, category etc, to determine whether a build was successful based on tests executed? Eventually the unit test should be created to execute the code once someone fixes it. On one hand it shows that identified bugs have not been fixed. On the other, there could be hundreds of failed unit tests showing up in the logs and weeding through the ones that should fail vs. failures due to a code check-in would be difficult to find.

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  • Update kernel patch for VMware Player 4.0.3

    As I stated some days ago, after upgrading to Ubuntu Precise Pangolin, aka 12.04 LTS, I had a minor obstacle with VMware products. Today, VMware offered to upgrade to Player 4.0.3 due to security-related reasons. Initially, I thought that this update might have the patch for kernel 3.2.0 integrated but sadly that is not the case. 'Hacking' the kernel patch My first intuitive try to run the existing patch against the sources of VMware Player 4.0.3 failed, as the patch by Stefano Angeleri (weltall) is originally written explicitely against Workstation 8.0.2 and Player 4.0.2. But this is nothing to worry about seriously. Just fire up your favourite editor of choice and modify the version signature for VMware Player, like so: nano patch-modules_3.2.0.sh And update line 8 for the new version: plreqver=4.0.3 Save the shell script and run it as super-user (root): sudo ./patch-modules_3.2.0.sh In case that you previously patched your VMware sources you have to remove some artifacts beforehand. Otherwise, the patch script will inform you like so: /usr/lib/vmware/modules/source/.patched found. You have already patched your sources. Exiting In that case, simply remove the 'hidden' file and run the shell script again: sudo rm /usr/lib/vmware/modules/source/.patchedsudo ./patch-modules_3.2.0.sh To finalise your installation either restart the vmware service or reboot your machine. On first start VMware will present you their EULA which you have to accept, and everything gets back to normal operation mode. Currently, I would assume that in case of VMware Workstation 8.0.3 you can follow the same steps as just described.

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  • What is the value of checking in failing unit tests?

    - by Adam W.
    While there are ways of keeping unit tests from being executed, what is the value of checking in failing unit tests? I will use a simple example: Case Sensitivity. The current code is case sensitive. A valid input into the method is "Cat" and it would return an enum of Animal.Cat. However, the desired functionality of the method should not be case sensitive. So if the method described was passed "cat" it could possibly return something like Animal.Null instead of Animal.Cat and the unit test would fail. Though a simple code change would make this work, a more complex issue may take weeks to fix, but identifying the bug with a unit test could be a less complex task. The application currently being analyzed has 4 years of code that "works". However, recent discussions regarding unit tests have found flaws in the code. Some just need explicit implementation documentation (ex. case sensitive or not), or code that does not execute the bug based on how it is currently called. But unit tests can be created executing specific scenarios that will cause the bug to be seen and are valid inputs. What is the value of checking in unit tests that exercise the bug until someone can get around to fixing the code? Should this unit test be flagged with ignore, priority, category etc, to determine whether a build was successful based on tests executed? Eventually the unit test should be created to execute the code once someone fixes it. On one hand it shows that identified bugs have not been fixed. On the other, there could be hundreds of failed unit tests showing up in the logs and weeding through the ones that should fail vs. failures due to a code check-in would be difficult to find.

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  • Dealing with Fine-Grained Cache Entries in Coherence

    - by jpurdy
    On occasion we have seen significant memory overhead when using very small cache entries. Consider the case where there is a small key (say a synthetic key stored in a long) and a small value (perhaps a number or short string). With most backing maps, each cache entry will require an instance of Map.Entry, and in the case of a LocalCache backing map (used for expiry and eviction), there is additional metadata stored (such as last access time). Given the size of this data (usually a few dozen bytes) and the granularity of Java memory allocation (often a minimum of 32 bytes per object, depending on the specific JVM implementation), it is easily possible to end up with the case where the cache entry appears to be a couple dozen bytes but ends up occupying several hundred bytes of actual heap, resulting in anywhere from a 5x to 10x increase in stated memory requirements. In most cases, this increase applies to only a few small NamedCaches, and is inconsequential -- but in some cases it might apply to one or more very large NamedCaches, in which case it may dominate memory sizing calculations. Ultimately, the requirement is to avoid the per-entry overhead, which can be done either at the application level by grouping multiple logical entries into single cache entries, or at the backing map level, again by combining multiple entries into a smaller number of larger heap objects. At the application level, it may be possible to combine objects based on parent-child or sibling relationships (basically the same requirements that would apply to using partition affinity). If there is no natural relationship, it may still be possible to combine objects, effectively using a Coherence NamedCache as a "map of maps". This forces the application to first find a collection of objects (by performing a partial hash) and then to look within that collection for the desired object. This is most naturally implemented as a collection of entry processors to avoid pulling unnecessary data back to the client (and also to encapsulate that logic within a service layer). At the backing map level, the NIO storage option keeps keys on heap, and so has limited benefit for this situation. The Elastic Data features of Coherence naturally combine entries into larger heap objects, with the caveat that only data -- and not indexes -- can be stored in Elastic Data.

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  • Particle trajectory smoothing: where to do the simulation?

    - by nkint
    I have a particle system in which I have particles that are moving to a target and the new targets are received via network. The list of new target are some noisy coordinates of a moving target stored in the server that I want to smooth in the client. For doing the smoothing and the particle I wrote a simple particle engine with standard euler integration model. So, my pseudo code is something like that: # pseudo code class Particle: def update(): # do euler motion model integration: # if the distance to the target is more than a limit # add a new force to the accelleration # seeking the target, # and add the accelleration to velocity # and velocity to the position positionHistory.push_back(position); if history.length > historySize : history.pop_front() class ParticleEngine: particleById = dict() # an associative array # where the keys are the id # and particle istances are sotred as values # this method is called each time a new tcp packet is received and parsed def setNetTarget(int id, Vec2D new_target): particleById[id].setNewTarget(new_target) # this method is called each new frame def draw(): for p in particleById.values: p.update() beginVertex(LINE_STRIP) for v in p.positionHistory: vertex(v.x, v.y) endVertex() The new target that are arriving are noisy but setting some accelleration/velocity parameters let the particle to have a smoothed trajectories. But if a particle trajectory is a circle after a while the particle position converge to the center (a normal behaviour of euler integration model). So I decided to change the simulation and use some other interpolation (spline?) or smooth method (kalman filter?) between the targets. Something like: switch( INTERPOLATION_MODEL ): case EULER_MOTION: ... case HERMITE_INTERPOLATION: ... case SPLINE_INTERPOLATION: ... case KALMAN_FILTER_SMOOTHING: ... Now my question: where to write the motion simulation / trajectory interpolation? In the Particle? So I will have some Particle subclass like ParticleEuler, ParticleSpline, ParticleKalman, etc..? Or in the particle engine?

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