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  • C Number to Text problem with ones and tens..

    - by Joegabb
    #include<stdio.h> #include<conio.h> main() { int ones,tens,ventoteen, myloop = 0; long num2,cents2,centeens,cents1,thousands,hundreds; double num; do{ printf("Enter a number: "); scanf("%lf",&num); if(num<=10000 || num>=0) { if (num==0) { printf("\t\tZero"); } num=(num*100); num2= (long)num; thousands=num2/100000; num2=num2%100000; hundreds=num2/10000; num2=num2%10000; if ((num2>=1100) || (num2<=1900)) { tens=0; ones=0; ventoteen=num2%1000; } else { tens=num2/1000; num2=num2%1000; ones=num2/100; num2=num2%100; } if((num2>=11) && (num2<=19)) { cents1=0; cents2=0; centeens=num2%10; } else { cents1=num2/10; num2=num2%10; cents2=num2/1; } if (thousands == 1) printf("One thousand "); else if (thousands == 2) printf("Two thousand "); else if (thousands == 3) printf("Three Thousand "); else if (thousands == 4) printf("Four thousand "); else if (thousands == 5) printf("Five Thousand "); else if (thousands == 6) printf("Six thousand "); else if (thousands == 7) printf("Seven Thousand "); else if (thousands == 8) printf("Eight thousand "); else if (thousands == 9) printf("Nine Thousand "); else {} if (hundreds == 1) printf("one hundred "); else if (hundreds == 2) printf("two hundred "); else if (hundreds == 3) printf("three hundred "); else if (hundreds == 4) printf("four hundred "); else if (hundreds == 5) printf("five hundred "); else if (hundreds == 6) printf("six hundred "); else if (hundreds == 7) printf("seven hundred "); else if (hundreds == 8) printf("eight hundred "); else if (hundreds == 9) printf("nine hundred "); else {} switch(ventoteen) { case 1: printf("eleven ");break; case 2: printf("twelve ");break; case 3: printf("thirteen ");break; case 4: printf("fourteen ");break; case 5: printf("fifteen ");break; case 6: printf("sixteen ");break; case 7: printf("seventeen ");break; case 8: printf("eighteen ");break; case 9: printf("nineteen ");break; } switch(tens) { case 1: printf("ten ");break; case 2: printf("twenty ");break; case 3: printf("thirty ");break; case 4: printf("forty ");break; case 5: printf("fifty ");break; case 6: printf("sixty ");break; case 7: printf("seventy ");break; case 8: printf("eighty ");break; case 9: printf("ninety ");break; } switch(ones) { case 1: printf("one ");break; case 2: printf("two ");break; case 3: printf("three ");break; case 4: printf("four ");break; case 5: printf("five ");break; case 6: printf("six ");break; case 7: printf("seven ");break; case 8: printf("eight ");break; case 9: printf("nine ");break; } switch(cents1) { case 1: printf("and ten centavos ");break; case 2: printf("and twenty centavos ");break; case 3: printf("and thirty centavos ");break; case 4: printf("and fourty centavos ");break; case 5: printf("and fifty centavos ");break; case 6: printf("and sixty centavos ");break; case 7: printf("and seventy centavos ");break; case 8: printf("and eighty centavos ");break; case 9: printf("and ninety centavos ");break; } switch(centeens) { case 1: printf("and eleven centavos ");break; case 2: printf("and twelve centavos ");break; case 3: printf("and thirteen centavos ");break; case 4: printf("and fourteen centavos ");break; case 5: printf("and fifteen centavos ");break; case 6: printf("and sixteen centavos ");break; case 7: printf("and seventeen centavos ");break; case 8: printf("and eighteen centavos ");break; case 9: printf("and nineteen centavos ");break; } switch(cents2) { case 1: printf("and one centavos ");break; case 2: printf("and two centavos ");break; case 3: printf("and three centavos ");break; case 4: printf("and four centavos ");break; case 5: printf("and five centavos ");break; case 6: printf("and six centavos ");break; case 7: printf("and seven centavos ");break; case 8: printf("and eight centavos ");break; case 9: printf("and nine centavos ");break; } } getch(); }while(myloop == 0); return 0; } my code is working fine but the problem is when i input 1 - 90 nothing appears but when i input 100 the output would be fine and that is "One Hundred" and so as 1000 the output would be "One Thousand". thanks for the help..

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  • 256 Windows Azure Worker Roles, Windows Kinect and a 90's Text-Based Ray-Tracer

    - by Alan Smith
    For a couple of years I have been demoing a simple render farm hosted in Windows Azure using worker roles and the Azure Storage service. At the start of the presentation I deploy an Azure application that uses 16 worker roles to render a 1,500 frame 3D ray-traced animation. At the end of the presentation, when the animation was complete, I would play the animation delete the Azure deployment. The standing joke with the audience was that it was that it was a “$2 demo”, as the compute charges for running the 16 instances for an hour was $1.92, factor in the bandwidth charges and it’s a couple of dollars. The point of the demo is that it highlights one of the great benefits of cloud computing, you pay for what you use, and if you need massive compute power for a short period of time using Windows Azure can work out very cost effective. The “$2 demo” was great for presenting at user groups and conferences in that it could be deployed to Azure, used to render an animation, and then removed in a one hour session. I have always had the idea of doing something a bit more impressive with the demo, and scaling it from a “$2 demo” to a “$30 demo”. The challenge was to create a visually appealing animation in high definition format and keep the demo time down to one hour.  This article will take a run through how I achieved this. Ray Tracing Ray tracing, a technique for generating high quality photorealistic images, gained popularity in the 90’s with companies like Pixar creating feature length computer animations, and also the emergence of shareware text-based ray tracers that could run on a home PC. In order to render a ray traced image, the ray of light that would pass from the view point must be tracked until it intersects with an object. At the intersection, the color, reflectiveness, transparency, and refractive index of the object are used to calculate if the ray will be reflected or refracted. Each pixel may require thousands of calculations to determine what color it will be in the rendered image. Pin-Board Toys Having very little artistic talent and a basic understanding of maths I decided to focus on an animation that could be modeled fairly easily and would look visually impressive. I’ve always liked the pin-board desktop toys that become popular in the 80’s and when I was working as a 3D animator back in the 90’s I always had the idea of creating a 3D ray-traced animation of a pin-board, but never found the energy to do it. Even if I had a go at it, the render time to produce an animation that would look respectable on a 486 would have been measured in months. PolyRay Back in 1995 I landed my first real job, after spending three years being a beach-ski-climbing-paragliding-bum, and was employed to create 3D ray-traced animations for a CD-ROM that school kids would use to learn physics. I had got into the strange and wonderful world of text-based ray tracing, and was using a shareware ray-tracer called PolyRay. PolyRay takes a text file describing a scene as input and, after a few hours processing on a 486, produced a high quality ray-traced image. The following is an example of a basic PolyRay scene file. background Midnight_Blue   static define matte surface { ambient 0.1 diffuse 0.7 } define matte_white texture { matte { color white } } define matte_black texture { matte { color dark_slate_gray } } define position_cylindrical 3 define lookup_sawtooth 1 define light_wood <0.6, 0.24, 0.1> define median_wood <0.3, 0.12, 0.03> define dark_wood <0.05, 0.01, 0.005>     define wooden texture { noise surface { ambient 0.2  diffuse 0.7  specular white, 0.5 microfacet Reitz 10 position_fn position_cylindrical position_scale 1  lookup_fn lookup_sawtooth octaves 1 turbulence 1 color_map( [0.0, 0.2, light_wood, light_wood] [0.2, 0.3, light_wood, median_wood] [0.3, 0.4, median_wood, light_wood] [0.4, 0.7, light_wood, light_wood] [0.7, 0.8, light_wood, median_wood] [0.8, 0.9, median_wood, light_wood] [0.9, 1.0, light_wood, dark_wood]) } } define glass texture { surface { ambient 0 diffuse 0 specular 0.2 reflection white, 0.1 transmission white, 1, 1.5 }} define shiny surface { ambient 0.1 diffuse 0.6 specular white, 0.6 microfacet Phong 7  } define steely_blue texture { shiny { color black } } define chrome texture { surface { color white ambient 0.0 diffuse 0.2 specular 0.4 microfacet Phong 10 reflection 0.8 } }   viewpoint {     from <4.000, -1.000, 1.000> at <0.000, 0.000, 0.000> up <0, 1, 0> angle 60     resolution 640, 480 aspect 1.6 image_format 0 }       light <-10, 30, 20> light <-10, 30, -20>   object { disc <0, -2, 0>, <0, 1, 0>, 30 wooden }   object { sphere <0.000, 0.000, 0.000>, 1.00 chrome } object { cylinder <0.000, 0.000, 0.000>, <0.000, 0.000, -4.000>, 0.50 chrome }   After setting up the background and defining colors and textures, the viewpoint is specified. The “camera” is located at a point in 3D space, and it looks towards another point. The angle, image resolution, and aspect ratio are specified. Two lights are present in the image at defined coordinates. The three objects in the image are a wooden disc to represent a table top, and a sphere and cylinder that intersect to form a pin that will be used for the pin board toy in the final animation. When the image is rendered, the following image is produced. The pins are modeled with a chrome surface, so they reflect the environment around them. Note that the scale of the pin shaft is not correct, this will be fixed later. Modeling the Pin Board The frame of the pin-board is made up of three boxes, and six cylinders, the front box is modeled using a clear, slightly reflective solid, with the same refractive index of glass. The other shapes are modeled as metal. object { box <-5.5, -1.5, 1>, <5.5, 5.5, 1.2> glass } object { box <-5.5, -1.5, -0.04>, <5.5, 5.5, -0.09> steely_blue } object { box <-5.5, -1.5, -0.52>, <5.5, 5.5, -0.59> steely_blue } object { cylinder <-5.2, -1.2, 1.4>, <-5.2, -1.2, -0.74>, 0.2 steely_blue } object { cylinder <5.2, -1.2, 1.4>, <5.2, -1.2, -0.74>, 0.2 steely_blue } object { cylinder <-5.2, 5.2, 1.4>, <-5.2, 5.2, -0.74>, 0.2 steely_blue } object { cylinder <5.2, 5.2, 1.4>, <5.2, 5.2, -0.74>, 0.2 steely_blue } object { cylinder <0, -1.2, 1.4>, <0, -1.2, -0.74>, 0.2 steely_blue } object { cylinder <0, 5.2, 1.4>, <0, 5.2, -0.74>, 0.2 steely_blue }   In order to create the matrix of pins that make up the pin board I used a basic console application with a few nested loops to create two intersecting matrixes of pins, which models the layout used in the pin boards. The resulting image is shown below. The pin board contains 11,481 pins, with the scene file containing 23,709 lines of code. For the complete animation 2,000 scene files will be created, which is over 47 million lines of code. Each pin in the pin-board will slide out a specific distance when an object is pressed into the back of the board. This is easily modeled by setting the Z coordinate of the pin to a specific value. In order to set all of the pins in the pin-board to the correct position, a bitmap image can be used. The position of the pin can be set based on the color of the pixel at the appropriate position in the image. When the Windows Azure logo is used to set the Z coordinate of the pins, the following image is generated. The challenge now was to make a cool animation. The Azure Logo is fine, but it is static. Using a normal video to animate the pins would not work; the colors in the video would not be the same as the depth of the objects from the camera. In order to simulate the pin board accurately a series of frames from a depth camera could be used. Windows Kinect The Kenect controllers for the X-Box 360 and Windows feature a depth camera. The Kinect SDK for Windows provides a programming interface for Kenect, providing easy access for .NET developers to the Kinect sensors. The Kinect Explorer provided with the Kinect SDK is a great starting point for exploring Kinect from a developers perspective. Both the X-Box 360 Kinect and the Windows Kinect will work with the Kinect SDK, the Windows Kinect is required for commercial applications, but the X-Box Kinect can be used for hobby projects. The Windows Kinect has the advantage of providing a mode to allow depth capture with objects closer to the camera, which makes for a more accurate depth image for setting the pin positions. Creating a Depth Field Animation The depth field animation used to set the positions of the pin in the pin board was created using a modified version of the Kinect Explorer sample application. In order to simulate the pin board accurately, a small section of the depth range from the depth sensor will be used. Any part of the object in front of the depth range will result in a white pixel; anything behind the depth range will be black. Within the depth range the pixels in the image will be set to RGB values from 0,0,0 to 255,255,255. A screen shot of the modified Kinect Explorer application is shown below. The Kinect Explorer sample application was modified to include slider controls that are used to set the depth range that forms the image from the depth stream. This allows the fine tuning of the depth image that is required for simulating the position of the pins in the pin board. The Kinect Explorer was also modified to record a series of images from the depth camera and save them as a sequence JPEG files that will be used to animate the pins in the animation the Start and Stop buttons are used to start and stop the image recording. En example of one of the depth images is shown below. Once a series of 2,000 depth images has been captured, the task of creating the animation can begin. Rendering a Test Frame In order to test the creation of frames and get an approximation of the time required to render each frame a test frame was rendered on-premise using PolyRay. The output of the rendering process is shown below. The test frame contained 23,629 primitive shapes, most of which are the spheres and cylinders that are used for the 11,800 or so pins in the pin board. The 1280x720 image contains 921,600 pixels, but as anti-aliasing was used the number of rays that were calculated was 4,235,777, with 3,478,754,073 object boundaries checked. The test frame of the pin board with the depth field image applied is shown below. The tracing time for the test frame was 4 minutes 27 seconds, which means rendering the2,000 frames in the animation would take over 148 hours, or a little over 6 days. Although this is much faster that an old 486, waiting almost a week to see the results of an animation would make it challenging for animators to create, view, and refine their animations. It would be much better if the animation could be rendered in less than one hour. Windows Azure Worker Roles The cost of creating an on-premise render farm to render animations increases in proportion to the number of servers. The table below shows the cost of servers for creating a render farm, assuming a cost of $500 per server. Number of Servers Cost 1 $500 16 $8,000 256 $128,000   As well as the cost of the servers, there would be additional costs for networking, racks etc. Hosting an environment of 256 servers on-premise would require a server room with cooling, and some pretty hefty power cabling. The Windows Azure compute services provide worker roles, which are ideal for performing processor intensive compute tasks. With the scalability available in Windows Azure a job that takes 256 hours to complete could be perfumed using different numbers of worker roles. The time and cost of using 1, 16 or 256 worker roles is shown below. Number of Worker Roles Render Time Cost 1 256 hours $30.72 16 16 hours $30.72 256 1 hour $30.72   Using worker roles in Windows Azure provides the same cost for the 256 hour job, irrespective of the number of worker roles used. Provided the compute task can be broken down into many small units, and the worker role compute power can be used effectively, it makes sense to scale the application so that the task is completed quickly, making the results available in a timely fashion. The task of rendering 2,000 frames in an animation is one that can easily be broken down into 2,000 individual pieces, which can be performed by a number of worker roles. Creating a Render Farm in Windows Azure The architecture of the render farm is shown in the following diagram. The render farm is a hybrid application with the following components: ·         On-Premise o   Windows Kinect – Used combined with the Kinect Explorer to create a stream of depth images. o   Animation Creator – This application uses the depth images from the Kinect sensor to create scene description files for PolyRay. These files are then uploaded to the jobs blob container, and job messages added to the jobs queue. o   Process Monitor – This application queries the role instance lifecycle table and displays statistics about the render farm environment and render process. o   Image Downloader – This application polls the image queue and downloads the rendered animation files once they are complete. ·         Windows Azure o   Azure Storage – Queues and blobs are used for the scene description files and completed frames. A table is used to store the statistics about the rendering environment.   The architecture of each worker role is shown below.   The worker role is configured to use local storage, which provides file storage on the worker role instance that can be use by the applications to render the image and transform the format of the image. The service definition for the worker role with the local storage configuration highlighted is shown below. <?xml version="1.0" encoding="utf-8"?> <ServiceDefinition name="CloudRay" >   <WorkerRole name="CloudRayWorkerRole" vmsize="Small">     <Imports>     </Imports>     <ConfigurationSettings>       <Setting name="DataConnectionString" />     </ConfigurationSettings>     <LocalResources>       <LocalStorage name="RayFolder" cleanOnRoleRecycle="true" />     </LocalResources>   </WorkerRole> </ServiceDefinition>     The two executable programs, PolyRay.exe and DTA.exe are included in the Azure project, with Copy Always set as the property. PolyRay will take the scene description file and render it to a Truevision TGA file. As the TGA format has not seen much use since the mid 90’s it is converted to a JPG image using Dave's Targa Animator, another shareware application from the 90’s. Each worker roll will use the following process to render the animation frames. 1.       The worker process polls the job queue, if a job is available the scene description file is downloaded from blob storage to local storage. 2.       PolyRay.exe is started in a process with the appropriate command line arguments to render the image as a TGA file. 3.       DTA.exe is started in a process with the appropriate command line arguments convert the TGA file to a JPG file. 4.       The JPG file is uploaded from local storage to the images blob container. 5.       A message is placed on the images queue to indicate a new image is available for download. 6.       The job message is deleted from the job queue. 7.       The role instance lifecycle table is updated with statistics on the number of frames rendered by the worker role instance, and the CPU time used. The code for this is shown below. public override void Run() {     // Set environment variables     string polyRayPath = Path.Combine(Environment.GetEnvironmentVariable("RoleRoot"), PolyRayLocation);     string dtaPath = Path.Combine(Environment.GetEnvironmentVariable("RoleRoot"), DTALocation);       LocalResource rayStorage = RoleEnvironment.GetLocalResource("RayFolder");     string localStorageRootPath = rayStorage.RootPath;       JobQueue jobQueue = new JobQueue("renderjobs");     JobQueue downloadQueue = new JobQueue("renderimagedownloadjobs");     CloudRayBlob sceneBlob = new CloudRayBlob("scenes");     CloudRayBlob imageBlob = new CloudRayBlob("images");     RoleLifecycleDataSource roleLifecycleDataSource = new RoleLifecycleDataSource();       Frames = 0;       while (true)     {         // Get the render job from the queue         CloudQueueMessage jobMsg = jobQueue.Get();           if (jobMsg != null)         {             // Get the file details             string sceneFile = jobMsg.AsString;             string tgaFile = sceneFile.Replace(".pi", ".tga");             string jpgFile = sceneFile.Replace(".pi", ".jpg");               string sceneFilePath = Path.Combine(localStorageRootPath, sceneFile);             string tgaFilePath = Path.Combine(localStorageRootPath, tgaFile);             string jpgFilePath = Path.Combine(localStorageRootPath, jpgFile);               // Copy the scene file to local storage             sceneBlob.DownloadFile(sceneFilePath);               // Run the ray tracer.             string polyrayArguments =                 string.Format("\"{0}\" -o \"{1}\" -a 2", sceneFilePath, tgaFilePath);             Process polyRayProcess = new Process();             polyRayProcess.StartInfo.FileName =                 Path.Combine(Environment.GetEnvironmentVariable("RoleRoot"), polyRayPath);             polyRayProcess.StartInfo.Arguments = polyrayArguments;             polyRayProcess.Start();             polyRayProcess.WaitForExit();               // Convert the image             string dtaArguments =                 string.Format(" {0} /FJ /P{1}", tgaFilePath, Path.GetDirectoryName (jpgFilePath));             Process dtaProcess = new Process();             dtaProcess.StartInfo.FileName =                 Path.Combine(Environment.GetEnvironmentVariable("RoleRoot"), dtaPath);             dtaProcess.StartInfo.Arguments = dtaArguments;             dtaProcess.Start();             dtaProcess.WaitForExit();               // Upload the image to blob storage             imageBlob.UploadFile(jpgFilePath);               // Add a download job.             downloadQueue.Add(jpgFile);               // Delete the render job message             jobQueue.Delete(jobMsg);               Frames++;         }         else         {             Thread.Sleep(1000);         }           // Log the worker role activity.         roleLifecycleDataSource.Alive             ("CloudRayWorker", RoleLifecycleDataSource.RoleLifecycleId, Frames);     } }     Monitoring Worker Role Instance Lifecycle In order to get more accurate statistics about the lifecycle of the worker role instances used to render the animation data was tracked in an Azure storage table. The following class was used to track the worker role lifecycles in Azure storage.   public class RoleLifecycle : TableServiceEntity {     public string ServerName { get; set; }     public string Status { get; set; }     public DateTime StartTime { get; set; }     public DateTime EndTime { get; set; }     public long SecondsRunning { get; set; }     public DateTime LastActiveTime { get; set; }     public int Frames { get; set; }     public string Comment { get; set; }       public RoleLifecycle()     {     }       public RoleLifecycle(string roleName)     {         PartitionKey = roleName;         RowKey = Utils.GetAscendingRowKey();         Status = "Started";         StartTime = DateTime.UtcNow;         LastActiveTime = StartTime;         EndTime = StartTime;         SecondsRunning = 0;         Frames = 0;     } }     A new instance of this class is created and added to the storage table when the role starts. It is then updated each time the worker renders a frame to record the total number of frames rendered and the total processing time. These statistics are used be the monitoring application to determine the effectiveness of use of resources in the render farm. Rendering the Animation The Azure solution was deployed to Windows Azure with the service configuration set to 16 worker role instances. This allows for the application to be tested in the cloud environment, and the performance of the application determined. When I demo the application at conferences and user groups I often start with 16 instances, and then scale up the application to the full 256 instances. The configuration to run 16 instances is shown below. <?xml version="1.0" encoding="utf-8"?> <ServiceConfiguration serviceName="CloudRay" xmlns="http://schemas.microsoft.com/ServiceHosting/2008/10/ServiceConfiguration" osFamily="1" osVersion="*">   <Role name="CloudRayWorkerRole">     <Instances count="16" />     <ConfigurationSettings>       <Setting name="DataConnectionString"         value="DefaultEndpointsProtocol=https;AccountName=cloudraydata;AccountKey=..." />     </ConfigurationSettings>   </Role> </ServiceConfiguration>     About six minutes after deploying the application the first worker roles become active and start to render the first frames of the animation. The CloudRay Monitor application displays an icon for each worker role instance, with a number indicating the number of frames that the worker role has rendered. The statistics on the left show the number of active worker roles and statistics about the render process. The render time is the time since the first worker role became active; the CPU time is the total amount of processing time used by all worker role instances to render the frames.   Five minutes after the first worker role became active the last of the 16 worker roles activated. By this time the first seven worker roles had each rendered one frame of the animation.   With 16 worker roles u and running it can be seen that one hour and 45 minutes CPU time has been used to render 32 frames with a render time of just under 10 minutes.     At this rate it would take over 10 hours to render the 2,000 frames of the full animation. In order to complete the animation in under an hour more processing power will be required. Scaling the render farm from 16 instances to 256 instances is easy using the new management portal. The slider is set to 256 instances, and the configuration saved. We do not need to re-deploy the application, and the 16 instances that are up and running will not be affected. Alternatively, the configuration file for the Azure service could be modified to specify 256 instances.   <?xml version="1.0" encoding="utf-8"?> <ServiceConfiguration serviceName="CloudRay" xmlns="http://schemas.microsoft.com/ServiceHosting/2008/10/ServiceConfiguration" osFamily="1" osVersion="*">   <Role name="CloudRayWorkerRole">     <Instances count="256" />     <ConfigurationSettings>       <Setting name="DataConnectionString"         value="DefaultEndpointsProtocol=https;AccountName=cloudraydata;AccountKey=..." />     </ConfigurationSettings>   </Role> </ServiceConfiguration>     Six minutes after the new configuration has been applied 75 new worker roles have activated and are processing their first frames.   Five minutes later the full configuration of 256 worker roles is up and running. We can see that the average rate of frame rendering has increased from 3 to 12 frames per minute, and that over 17 hours of CPU time has been utilized in 23 minutes. In this test the time to provision 140 worker roles was about 11 minutes, which works out at about one every five seconds.   We are now half way through the rendering, with 1,000 frames complete. This has utilized just under three days of CPU time in a little over 35 minutes.   The animation is now complete, with 2,000 frames rendered in a little over 52 minutes. The CPU time used by the 256 worker roles is 6 days, 7 hours and 22 minutes with an average frame rate of 38 frames per minute. The rendering of the last 1,000 frames took 16 minutes 27 seconds, which works out at a rendering rate of 60 frames per minute. The frame counts in the server instances indicate that the use of a queue to distribute the workload has been very effective in distributing the load across the 256 worker role instances. The first 16 instances that were deployed first have rendered between 11 and 13 frames each, whilst the 240 instances that were added when the application was scaled have rendered between 6 and 9 frames each.   Completed Animation I’ve uploaded the completed animation to YouTube, a low resolution preview is shown below. Pin Board Animation Created using Windows Kinect and 256 Windows Azure Worker Roles   The animation can be viewed in 1280x720 resolution at the following link: http://www.youtube.com/watch?v=n5jy6bvSxWc Effective Use of Resources According to the CloudRay monitor statistics the animation took 6 days, 7 hours and 22 minutes CPU to render, this works out at 152 hours of compute time, rounded up to the nearest hour. As the usage for the worker role instances are billed for the full hour, it may have been possible to render the animation using fewer than 256 worker roles. When deciding the optimal usage of resources, the time required to provision and start the worker roles must also be considered. In the demo I started with 16 worker roles, and then scaled the application to 256 worker roles. It would have been more optimal to start the application with maybe 200 worker roles, and utilized the full hour that I was being billed for. This would, however, have prevented showing the ease of scalability of the application. The new management portal displays the CPU usage across the worker roles in the deployment. The average CPU usage across all instances is 93.27%, with over 99% used when all the instances are up and running. This shows that the worker role resources are being used very effectively. Grid Computing Scenarios Although I am using this scenario for a hobby project, there are many scenarios where a large amount of compute power is required for a short period of time. Windows Azure provides a great platform for developing these types of grid computing applications, and can work out very cost effective. ·         Windows Azure can provide massive compute power, on demand, in a matter of minutes. ·         The use of queues to manage the load balancing of jobs between role instances is a simple and effective solution. ·         Using a cloud-computing platform like Windows Azure allows proof-of-concept scenarios to be tested and evaluated on a very low budget. ·         No charges for inbound data transfer makes the uploading of large data sets to Windows Azure Storage services cost effective. (Transaction charges still apply.) Tips for using Windows Azure for Grid Computing Scenarios I found the implementation of a render farm using Windows Azure a fairly simple scenario to implement. I was impressed by ease of scalability that Azure provides, and by the short time that the application took to scale from 16 to 256 worker role instances. In this case it was around 13 minutes, in other tests it took between 10 and 20 minutes. The following tips may be useful when implementing a grid computing project in Windows Azure. ·         Using an Azure Storage queue to load-balance the units of work across multiple worker roles is simple and very effective. The design I have used in this scenario could easily scale to many thousands of worker role instances. ·         Windows Azure accounts are typically limited to 20 cores. If you need to use more than this, a call to support and a credit card check will be required. ·         Be aware of how the billing model works. You will be charged for worker role instances for the full clock our in which the instance is deployed. Schedule the workload to start just after the clock hour has started. ·         Monitor the utilization of the resources you are provisioning, ensure that you are not paying for worker roles that are idle. ·         If you are deploying third party applications to worker roles, you may well run into licensing issues. Purchasing software licenses on a per-processor basis when using hundreds of processors for a short time period would not be cost effective. ·         Third party software may also require installation onto the worker roles, which can be accomplished using start-up tasks. Bear in mind that adding a startup task and possible re-boot will add to the time required for the worker role instance to start and activate. An alternative may be to use a prepared VM and use VM roles. ·         Consider using the Windows Azure Autoscaling Application Block (WASABi) to autoscale the worker roles in your application. When using a large number of worker roles, the utilization must be carefully monitored, if the scaling algorithms are not optimal it could get very expensive!

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  • Send raw data to USB parallel port after upgrading to 11.10

    - by zaphod
    I have a laser cutter connected via a generic USB to parallel adapter. The laser cutter speaks HPGL, as it happens, but since this is a laser cutter and not a plotter, I usually want to generate the HPGL myself, since I care about the ordering, speed, and direction of cuts and so on. In previous versions of Ubuntu, I was able to print to the cutter by copying an HPGL file directly to the corresponding USB "lp" device. For example: cp foo.plt /dev/usblp1 Well, I just upgraded to Ubuntu 11.10 oneiric, and I can't find any "lp" devices in /dev anymore. D'oh! What's the preferred way to send raw data to a parallel port in Ubuntu? I've tried System Settings Printing + Add, hoping that I might be able to associate my device with some kind of "raw printer" driver and print to it with a command like lp -d LaserCutter foo.plt But my USB to parallel adapter doesn't seem to show up in the list. What I do see are my HP Color LaserJet, two USB-to-serial adapters, "Enter URI", and "Network Printer". Meanwhile, over in /dev, I do see /dev/ttyUSB0 and /dev/ttyUSB1 devices for the 2 USB-to-serial adapters. I don't see anything obvious corresponding to the HP printer (which was /dev/usblp0 prior to the upgrade), except for generic USB stuff. For example, sudo find /dev | grep lp produces no output. I do seem to be able to print to the HP printer just fine, though. The printer setup GUI gives it a device URI starting with "hp:" which isn't much help for the parallel adapter. The CUPS administrator's guide makes it sound like I might need to feed it a device URI of the form parallel:/dev/SOMETHING, but of course if I had a /dev/SOMETHING I'd probably just go on writing to it directly. Here's what dmesg says after I disconnect and reconnect the device from the USB port: [ 924.722906] usb 1-1.1.4: USB disconnect, device number 7 [ 959.993002] usb 1-1.1.4: new full speed USB device number 8 using ehci_hcd And here's how it shows up in lsusb -v: Bus 001 Device 008: ID 1a86:7584 QinHeng Electronics CH340S Device Descriptor: bLength 18 bDescriptorType 1 bcdUSB 1.10 bDeviceClass 0 (Defined at Interface level) bDeviceSubClass 0 bDeviceProtocol 0 bMaxPacketSize0 8 idVendor 0x1a86 QinHeng Electronics idProduct 0x7584 CH340S bcdDevice 2.52 iManufacturer 0 iProduct 2 USB2.0-Print iSerial 0 bNumConfigurations 1 Configuration Descriptor: bLength 9 bDescriptorType 2 wTotalLength 32 bNumInterfaces 1 bConfigurationValue 1 iConfiguration 0 bmAttributes 0x80 (Bus Powered) MaxPower 96mA Interface Descriptor: bLength 9 bDescriptorType 4 bInterfaceNumber 0 bAlternateSetting 0 bNumEndpoints 2 bInterfaceClass 7 Printer bInterfaceSubClass 1 Printer bInterfaceProtocol 2 Bidirectional iInterface 0 Endpoint Descriptor: bLength 7 bDescriptorType 5 bEndpointAddress 0x82 EP 2 IN bmAttributes 2 Transfer Type Bulk Synch Type None Usage Type Data wMaxPacketSize 0x0020 1x 32 bytes bInterval 0 Endpoint Descriptor: bLength 7 bDescriptorType 5 bEndpointAddress 0x02 EP 2 OUT bmAttributes 2 Transfer Type Bulk Synch Type None Usage Type Data wMaxPacketSize 0x0020 1x 32 bytes bInterval 0 Device Status: 0x0000 (Bus Powered)

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  • Diving into OpenStack Network Architecture - Part 2 - Basic Use Cases

    - by Ronen Kofman
      rkofman Normal rkofman 4 138 2014-06-05T03:38:00Z 2014-06-05T05:04:00Z 3 2735 15596 Oracle Corporation 129 36 18295 12.00 Clean Clean false false false false EN-US X-NONE HE /* 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-qformat:yes; mso-style-parent:""; mso-padding-alt:0in 5.4pt 0in 5.4pt; mso-para-margin-top:0in; mso-para-margin-right:0in; mso-para-margin-bottom:10.0pt; mso-para-margin-left:0in; line-height:115%; mso-pagination:widow-orphan; font-size:11.0pt; font-family:"Calibri","sans-serif"; mso-ascii-font-family:Calibri; mso-ascii-theme-font:minor-latin; mso-hansi-font-family:Calibri; mso-hansi-theme-font:minor-latin; mso-bidi-font-family:Arial; mso-bidi-theme-font:minor-bidi; mso-bidi-language:AR-SA;} In the previous post we reviewed several network components including Open vSwitch, Network Namespaces, Linux Bridges and veth pairs. In this post we will take three simple use cases and see how those basic components come together to create a complete SDN solution in OpenStack. With those three use cases we will review almost the entire network setup and see how all the pieces work together. The use cases we will use are: 1.       Create network – what happens when we create network and how can we create multiple isolated networks 2.       Launch a VM – once we have networks we can launch VMs and connect them to networks. 3.       DHCP request from a VM – OpenStack can automatically assign IP addresses to VMs. This is done through local DHCP service controlled by OpenStack Neutron. We will see how this service runs and how does a DHCP request and response look like. In this post we will show connectivity, we will see how packets get from point A to point B. We first focus on how a configured deployment looks like and only later we will discuss how and when the configuration is created. Personally I found it very valuable to see the actual interfaces and how they connect to each other through examples and hands on experiments. After the end game is clear and we know how the connectivity works, in a later post, we will take a step back and explain how Neutron configures the components to be able to provide such connectivity.  We are going to get pretty technical shortly and I recommend trying these examples on your own deployment or using the Oracle OpenStack Tech Preview. Understanding these three use cases thoroughly and how to look at them will be very helpful when trying to debug a deployment in case something does not work. Use case #1: Create Network Create network is a simple operation it can be performed from the GUI or command line. When we create a network in OpenStack the network is only available to the tenant who created it or it could be defined as “shared” and then it can be used by all tenants. A network can have multiple subnets but for this demonstration purpose and for simplicity we will assume that each network has exactly one subnet. Creating a network from the command line will look like this: # neutron net-create net1 Created a new network: +---------------------------+--------------------------------------+ | Field                     | Value                                | +---------------------------+--------------------------------------+ | admin_state_up            | True                                 | | id                        | 5f833617-6179-4797-b7c0-7d420d84040c | | name                      | net1                                 | | provider:network_type     | vlan                                 | | provider:physical_network | default                              | | provider:segmentation_id  | 1000                                 | | shared                    | False                                | | status                    | ACTIVE                               | | subnets                   |                                      | | tenant_id                 | 9796e5145ee546508939cd49ad59d51f     | +---------------------------+--------------------------------------+ Creating a subnet for this network will look like this: # neutron subnet-create net1 10.10.10.0/24 Created a new subnet: +------------------+------------------------------------------------+ | Field            | Value                                          | +------------------+------------------------------------------------+ | allocation_pools | {"start": "10.10.10.2", "end": "10.10.10.254"} | | cidr             | 10.10.10.0/24                                  | | dns_nameservers  |                                                | | enable_dhcp      | True                                           | | gateway_ip       | 10.10.10.1                                     | | host_routes      |                                                | | id               | 2d7a0a58-0674-439a-ad23-d6471aaae9bc           | | ip_version       | 4                                              | | name             |                                                | | network_id       | 5f833617-6179-4797-b7c0-7d420d84040c           | | tenant_id        | 9796e5145ee546508939cd49ad59d51f               | +------------------+------------------------------------------------+ We now have a network and a subnet, on the network topology view this looks like this: Now let’s dive in and see what happened under the hood. Looking at the control node we will discover that a new namespace was created: # ip netns list qdhcp-5f833617-6179-4797-b7c0-7d420d84040c   The name of the namespace is qdhcp-<network id> (see above), let’s look into the namespace and see what’s in it: # ip netns exec qdhcp-5f833617-6179-4797-b7c0-7d420d84040c ip addr 1: lo: <LOOPBACK,UP,LOWER_UP> mtu 65536 qdisc noqueue state UNKNOWN     link/loopback 00:00:00:00:00:00 brd 00:00:00:00:00:00     inet 127.0.0.1/8 scope host lo     inet6 ::1/128 scope host        valid_lft forever preferred_lft forever 12: tap26c9b807-7c: <BROADCAST,UP,LOWER_UP> mtu 1500 qdisc noqueue state UNKNOWN     link/ether fa:16:3e:1d:5c:81 brd ff:ff:ff:ff:ff:ff     inet 10.10.10.3/24 brd 10.10.10.255 scope global tap26c9b807-7c     inet6 fe80::f816:3eff:fe1d:5c81/64 scope link        valid_lft forever preferred_lft forever   We see two interfaces in the namespace, one is the loopback and the other one is an interface called “tap26c9b807-7c”. This interface has the IP address of 10.10.10.3 and it will also serve dhcp requests in a way we will see later. Let’s trace the connectivity of the “tap26c9b807-7c” interface from the namespace.  First stop is OVS, we see that the interface connects to bridge  “br-int” on OVS: # ovs-vsctl show 8a069c7c-ea05-4375-93e2-b9fc9e4b3ca1     Bridge "br-eth2"         Port "br-eth2"             Interface "br-eth2"                 type: internal         Port "eth2"             Interface "eth2"         Port "phy-br-eth2"             Interface "phy-br-eth2"     Bridge br-ex         Port br-ex             Interface br-ex                 type: internal     Bridge br-int         Port "int-br-eth2"             Interface "int-br-eth2"         Port "tap26c9b807-7c"             tag: 1             Interface "tap26c9b807-7c"                 type: internal         Port br-int             Interface br-int                 type: internal     ovs_version: "1.11.0"   In the picture above we have a veth pair which has two ends called “int-br-eth2” and "phy-br-eth2", this veth pair is used to connect two bridge in OVS "br-eth2" and "br-int". In the previous post we explained how to check the veth connectivity using the ethtool command. It shows that the two are indeed a pair: # ethtool -S int-br-eth2 NIC statistics:      peer_ifindex: 10 . .   #ip link . . 10: phy-br-eth2: <BROADCAST,MULTICAST,UP,LOWER_UP> mtu 1500 qdisc pfifo_fast state UP qlen 1000 . . Note that “phy-br-eth2” is connected to a bridge called "br-eth2" and one of this bridge's interfaces is the physical link eth2. This means that the network which we have just created has created a namespace which is connected to the physical interface eth2. eth2 is the “VM network” the physical interface where all the virtual machines connect to where all the VMs are connected. About network isolation: OpenStack supports creation of multiple isolated networks and can use several mechanisms to isolate the networks from one another. The isolation mechanism can be VLANs, VxLANs or GRE tunnels, this is configured as part of the initial setup in our deployment we use VLANs. When using VLAN tagging as an isolation mechanism a VLAN tag is allocated by Neutron from a pre-defined VLAN tags pool and assigned to the newly created network. By provisioning VLAN tags to the networks Neutron allows creation of multiple isolated networks on the same physical link.  The big difference between this and other platforms is that the user does not have to deal with allocating and managing VLANs to networks. The VLAN allocation and provisioning is handled by Neutron which keeps track of the VLAN tags, and responsible for allocating and reclaiming VLAN tags. In the example above net1 has the VLAN tag 1000, this means that whenever a VM is created and connected to this network the packets from that VM will have to be tagged with VLAN tag 1000 to go on this particular network. This is true for namespace as well, if we would like to connect a namespace to a particular network we have to make sure that the packets to and from the namespace are correctly tagged when they reach the VM network. In the example above we see that the namespace interface “tap26c9b807-7c” has vlan tag 1 assigned to it, if we examine OVS we see that it has flows which modify VLAN tag 1 to VLAN tag 1000 when a packet goes to the VM network on eth2 and vice versa. We can see this using the dump-flows command on OVS for packets going to the VM network we see the modification done on br-eth2: #  ovs-ofctl dump-flows br-eth2 NXST_FLOW reply (xid=0x4):  cookie=0x0, duration=18669.401s, table=0, n_packets=857, n_bytes=163350, idle_age=25, priority=4,in_port=2,dl_vlan=1 actions=mod_vlan_vid:1000,NORMAL  cookie=0x0, duration=165108.226s, table=0, n_packets=14, n_bytes=1000, idle_age=5343, hard_age=65534, priority=2,in_port=2 actions=drop  cookie=0x0, duration=165109.813s, table=0, n_packets=1671, n_bytes=213304, idle_age=25, hard_age=65534, priority=1 actions=NORMAL   For packets coming from the interface to the namespace we see the following modification: #  ovs-ofctl dump-flows br-int NXST_FLOW reply (xid=0x4):  cookie=0x0, duration=18690.876s, table=0, n_packets=1610, n_bytes=210752, idle_age=1, priority=3,in_port=1,dl_vlan=1000 actions=mod_vlan_vid:1,NORMAL  cookie=0x0, duration=165130.01s, table=0, n_packets=75, n_bytes=3686, idle_age=4212, hard_age=65534, priority=2,in_port=1 actions=drop  cookie=0x0, duration=165131.96s, table=0, n_packets=863, n_bytes=160727, idle_age=1, hard_age=65534, priority=1 actions=NORMAL   To summarize we can see that when a user creates a network Neutron creates a namespace and this namespace is connected through OVS to the “VM network”. OVS also takes care of tagging the packets from the namespace to the VM network with the correct VLAN tag and knows to modify the VLAN for packets coming from VM network to the namespace. Now let’s see what happens when a VM is launched and how it is connected to the “VM network”. Use case #2: Launch a VM Launching a VM can be done from Horizon or from the command line this is how we do it from Horizon: Attach the network: And Launch Once the virtual machine is up and running we can see the associated IP using the nova list command : # nova list +--------------------------------------+--------------+--------+------------+-------------+-----------------+ | ID                                   | Name         | Status | Task State | Power State | Networks        | +--------------------------------------+--------------+--------+------------+-------------+-----------------+ | 3707ac87-4f5d-4349-b7ed-3a673f55e5e1 | Oracle Linux | ACTIVE | None       | Running     | net1=10.10.10.2 | +--------------------------------------+--------------+--------+------------+-------------+-----------------+ The nova list command shows us that the VM is running and that the IP 10.10.10.2 is assigned to this VM. Let’s trace the connectivity from the VM to VM network on eth2 starting with the VM definition file. The configuration files of the VM including the virtual disk(s), in case of ephemeral storage, are stored on the compute node at/var/lib/nova/instances/<instance-id>/. Looking into the VM definition file ,libvirt.xml,  we see that the VM is connected to an interface called “tap53903a95-82” which is connected to a Linux bridge called “qbr53903a95-82”: <interface type="bridge">       <mac address="fa:16:3e:fe:c7:87"/>       <source bridge="qbr53903a95-82"/>       <target dev="tap53903a95-82"/>     </interface>   Looking at the bridge using the brctl show command we see this: # brctl show bridge name     bridge id               STP enabled     interfaces qbr53903a95-82          8000.7e7f3282b836       no              qvb53903a95-82                                                         tap53903a95-82    The bridge has two interfaces, one connected to the VM (“tap53903a95-82 “) and another one ( “qvb53903a95-82”) connected to “br-int” bridge on OVS: # ovs-vsctl show 83c42f80-77e9-46c8-8560-7697d76de51c     Bridge "br-eth2"         Port "br-eth2"             Interface "br-eth2"                 type: internal         Port "eth2"             Interface "eth2"         Port "phy-br-eth2"             Interface "phy-br-eth2"     Bridge br-int         Port br-int             Interface br-int                 type: internal         Port "int-br-eth2"             Interface "int-br-eth2"         Port "qvo53903a95-82"             tag: 3             Interface "qvo53903a95-82"     ovs_version: "1.11.0"   As we showed earlier “br-int” is connected to “br-eth2” on OVS using the veth pair int-br-eth2,phy-br-eth2 and br-eth2 is connected to the physical interface eth2. The whole flow end to end looks like this: VM è tap53903a95-82 (virtual interface)è qbr53903a95-82 (Linux bridge) è qvb53903a95-82 (interface connected from Linux bridge to OVS bridge br-int) è int-br-eth2 (veth one end) è phy-br-eth2 (veth the other end) è eth2 physical interface. The purpose of the Linux Bridge connecting to the VM is to allow security group enforcement with iptables. Security groups are enforced at the edge point which are the interface of the VM, since iptables nnot be applied to OVS bridges we use Linux bridge to apply them. In the future we hope to see this Linux Bridge going away rules.  VLAN tags: As we discussed in the first use case net1 is using VLAN tag 1000, looking at OVS above we see that qvo41f1ebcf-7c is tagged with VLAN tag 3. The modification from VLAN tag 3 to 1000 as we go to the physical network is done by OVS  as part of the packet flow of br-eth2 in the same way we showed before. To summarize, when a VM is launched it is connected to the VM network through a chain of elements as described here. During the packet from VM to the network and back the VLAN tag is modified. Use case #3: Serving a DHCP request coming from the virtual machine In the previous use cases we have shown that both the namespace called dhcp-<some id> and the VM end up connecting to the physical interface eth2  on their respective nodes, both will tag their packets with VLAN tag 1000.We saw that the namespace has an interface with IP of 10.10.10.3. Since the VM and the namespace are connected to each other and have interfaces on the same subnet they can ping each other, in this picture we see a ping from the VM which was assigned 10.10.10.2 to the namespace: The fact that they are connected and can ping each other can become very handy when something doesn’t work right and we need to isolate the problem. In such case knowing that we should be able to ping from the VM to the namespace and back can be used to trace the disconnect using tcpdump or other monitoring tools. To serve DHCP requests coming from VMs on the network Neutron uses a Linux tool called “dnsmasq”,this is a lightweight DNS and DHCP service you can read more about it here. If we look at the dnsmasq on the control node with the ps command we see this: dnsmasq --no-hosts --no-resolv --strict-order --bind-interfaces --interface=tap26c9b807-7c --except-interface=lo --pid-file=/var/lib/neutron/dhcp/5f833617-6179-4797-b7c0-7d420d84040c/pid --dhcp-hostsfile=/var/lib/neutron/dhcp/5f833617-6179-4797-b7c0-7d420d84040c/host --dhcp-optsfile=/var/lib/neutron/dhcp/5f833617-6179-4797-b7c0-7d420d84040c/opts --leasefile-ro --dhcp-range=tag0,10.10.10.0,static,120s --dhcp-lease-max=256 --conf-file= --domain=openstacklocal The service connects to the tap interface in the namespace (“--interface=tap26c9b807-7c”), If we look at the hosts file we see this: # cat  /var/lib/neutron/dhcp/5f833617-6179-4797-b7c0-7d420d84040c/host fa:16:3e:fe:c7:87,host-10-10-10-2.openstacklocal,10.10.10.2   If you look at the console output above you can see the MAC address fa:16:3e:fe:c7:87 which is the VM MAC. This MAC address is mapped to IP 10.10.10.2 and so when a DHCP request comes with this MAC dnsmasq will return the 10.10.10.2.If we look into the namespace at the time we initiate a DHCP request from the VM (this can be done by simply restarting the network service in the VM) we see the following: # ip netns exec qdhcp-5f833617-6179-4797-b7c0-7d420d84040c tcpdump -n 19:27:12.191280 IP 0.0.0.0.bootpc > 255.255.255.255.bootps: BOOTP/DHCP, Request from fa:16:3e:fe:c7:87, length 310 19:27:12.191666 IP 10.10.10.3.bootps > 10.10.10.2.bootpc: BOOTP/DHCP, Reply, length 325   To summarize, the DHCP service is handled by dnsmasq which is configured by Neutron to listen to the interface in the DHCP namespace. Neutron also configures dnsmasq with the combination of MAC and IP so when a DHCP request comes along it will receive the assigned IP. Summary In this post we relied on the components described in the previous post and saw how network connectivity is achieved using three simple use cases. These use cases gave a good view of the entire network stack and helped understand how an end to end connection is being made between a VM on a compute node and the DHCP namespace on the control node. One conclusion we can draw from what we saw here is that if we launch a VM and it is able to perform a DHCP request and receive a correct IP then there is reason to believe that the network is working as expected. We saw that a packet has to travel through a long list of components before reaching its destination and if it has done so successfully this means that many components are functioning properly. In the next post we will look at some more sophisticated services Neutron supports and see how they work. We will see that while there are some more components involved for the most part the concepts are the same. @RonenKofman

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  • Failed to start up after upgrading software

    - by Landy
    I asked this question in SuperUser one hour ago, then I know this community so I moved the question here... I've been running Ubuntu 10.10 in a physical x86-64 machine. Today Update Manager reminded me that there are some updates to install and I confirmed the action. I should had read the update list but I didn't. I can only remember there is an update about cups. After the upgrading, Update Manager requires a restart and I confirmed too. But after the restart, the computer can't start up. There are errors in the console. Begin: Running /scripts/init-premount ... done. Begin: Mounting root file system ... Begin: Running /scripts/local-top ... done. [xxx]usb 1-8: new high speed USB device using ehci_hcd and address 3 [xxx]usb 2-1: new full speed USB device using ohci_hcd and address 2 [xxx]hub 2-1:1.0: USB hub found [xxx]hub 2-1:1.0: 4 ports detected [xxx]usb 2-1.1: new low speed USB device using ohci_hcd and address 3 Gave up waiting for root device. Common probles: - Boot args (cat /proc/cmdline) - Check rootdelay=(did the system wait long enough) - Check root= (did the system wait for the right device?) - Missing modules (cat /proc/modules; ls /dev) FATAL: Could not load /lib/modules/2.6.35-22-generic/modules.dep: No such file or directory FATAL: Could not load /lib/modules/2.6.35-22-generic/modules.dep: No such file or directory ALERT! /dev/sda1 does not exist. Dropping to a shell! BusyBox v1.15.3 (Ubuntu 1:1.15.3-1ubuntu5) built-in shell(ash) Enter 'help' for a list of built-in commands. (initramfs)[cursor is here] At the moment, I can't input anything in the console. The keyboard doesn't work at all. What's wrong? How can I check boot args or "root=" as suggested? How can I fix this issue? Thanks. =============== PS1: the /dev/sda1 is type ext4 (rw,nosuid,nodev) PS2: the /dev/sda1 can be mounted and accessed successfully under SUSE 11 SP1 x64. PS3: From this link, I think the keyboard doesn't work because the USB driver is not loaded at that time.

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  • Writing or extending existing emacs packages: is it worth or should I move to Netbeans/Eclipse?

    - by Andrea
    I'm finishing my master degree course in CS and I've almost become addicted to Emacs. I've used it to write in C, Latex, Java, JSP,XML, CommonLisp, Ada and other languages no other editor supported, like AMPL. I'd like to improve the packages I've been using the most or create new ones, but, in practice, I find that the implementation of Emacs leaves a lot to be desired. There are a lot of poorly-featured/poorly-maintained packages with either overlapping functionalities or obscure incompatibilities, and Elisp just seems to foster the situation by lacking the common features modern lisps have. In contrast Eclipse and Netbeans are actively improved and it does seem they can be effective for non-mainstream languages. I tried Hibachi for Ada in Eclipse and it worked well, there's CUPS for Lisp in Eclipse and LambdaBeans built using NetBeans components. On the other hand those plugins seem to be less active than their Emacs' counterparts, for example Hibachi was archived last year. What's your opinion on this? Which editor should I write extension for? EDIT: To answer Larry Coleman (see comment below): I like Emacs as a user because it is efficient both for me and the computer I'm using. It's fast and the textual interface (i.e. minibuffer) allows for quick interaction. It's solid and packages are usually small and easy to manage. If I need to correct or remove something I usually just have to change a row in my .emacs or an elisp file, or delete a directory. Eclipse plugins rely on a more complicated process that screwed my Eclipse configuration a couple of times, forcing me to do a clean reinstall. Emacs works as long as I use the basic packages. If I need something more complicated the situation gets pretty hairy. As a "power user" I think that the best I can hope for is to write a severely crippled version of the extensions I'd actually like to have; in other words, that it's not worth the trouble. I'd like to write extensions for the things I'd like to have automated in Emacs, for example project support with automated tag-table update on file writing. There are a few projects on this that lack integration, documentation, extensibility and so forth. The best one is probably CEDET, for which I believe the Greenspun's 10th rule can be applied. EDIT: To comment Larry Coleman's answer I'm pretty sure I can pick elisp programming but the extensions I have in mind don't exist yet despite their relative simplicity and the effort more knowledgeable people poured into related projects.This makes me wonder whether it is so because of the way emacs is developed, i.e. people tend to write their own little extensions without coordination, or its implementation, its extension language not being able to keep up with the growing complexity.

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  • Failed to start up after upgrading software in ubuntu 10.10

    - by Landy
    I asked this question in SuperUser one hour ago, then I know this community so I moved the question here... I've been running Ubuntu 10.10 in a physical x86-64 machine. Today Update Manager reminded me that there are some updates to install and I confirmed the action. I should had read the update list but I didn't. I can only remember there is an update about cups. After the upgrading, Update Manager requires a restart and I confirmed too. But after the restart, the computer can't start up. There are errors in the console. Begin: Running /scripts/init-premount ... done. Begin: Mounting root file system ... Begin: Running /scripts/local-top ... done. [xxx]usb 1-8: new high speed USB device using ehci_hcd and address 3 [xxx]usb 2-1: new full speed USB device using ohci_hcd and address 2 [xxx]hub 2-1:1.0: USB hub found [xxx]hub 2-1:1.0: 4 ports detected [xxx]usb 2-1.1: new low speed USB device using ohci_hcd and address 3 Gave up waiting for root device. Common probles: - Boot args (cat /proc/cmdline) - Check rootdelay=(did the system wait long enough) - Check root= (did the system wait for the right device?) - Missing modules (cat /proc/modules; ls /dev) FATAL: Could not load /lib/modules/2.6.35-22-generic/modules.dep: No such file or directory FATAL: Could not load /lib/modules/2.6.35-22-generic/modules.dep: No such file or directory ALERT! /dev/sda1 does not exist. Dropping to a shell! BusyBox v1.15.3 (Ubuntu 1:1.15.3-1ubuntu5) built-in shell(ash) Enter 'help' for a list of built-in commands. (initramfs)[cursor is here] At the moment, I can't input anything in the console. The keyboard doesn't work at all. What's wrong? How can I check boot args or "root=" as suggested? How can I fix this issue? Thanks. =============== PS1: the /dev/sda1 is type ext4 (rw,nosuid,nodev) PS2: the /dev/sda1 can be mounted and accessed successfully under SUSE 11 SP1 x64. PS3: From this link, I think the keyboard doesn't work because the USB driver is not loaded at that time.

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  • Send raw data to USB parallel port after upgrading to 11.10 oneiric

    - by zaphod
    I have a laser cutter connected via a generic USB to parallel adapter. The laser cutter speaks HPGL, as it happens, but since this is a laser cutter and not a plotter, I usually want to generate the HPGL myself, since I care about the ordering, speed, and direction of cuts and so on. In previous versions of Ubuntu, I was able to print to the cutter by copying an HPGL file directly to the corresponding USB "lp" device. For example: cp foo.plt /dev/usblp1 Well, I just upgraded to Ubuntu 11.10 oneiric, and I can't find any "lp" devices in /dev anymore. D'oh! What's the preferred way to send raw data to a parallel port in Ubuntu? I've tried System Settings Printing + Add, hoping that I might be able to associate my device with some kind of "raw printer" driver and print to it with a command like lp -d LaserCutter foo.plt But my USB to parallel adapter doesn't seem to show up in the list. What I do see are my HP Color LaserJet, two USB-to-serial adapters, "Enter URI", and "Network Printer". Meanwhile, over in /dev, I do see /dev/ttyUSB0 and /dev/ttyUSB1 devices for the 2 USB-to-serial adapters. I don't see anything obvious corresponding to the HP printer (which was /dev/usblp0 prior to the upgrade), except for generic USB stuff. For example, sudo find /dev | grep lp produces no output. I do seem to be able to print to the HP printer just fine, though. The printer setup GUI gives it a device URI starting with "hp:" which isn't much help for the parallel adapter. The CUPS administrator's guide makes it sound like I might need to feed it a device URI of the form parallel:/dev/SOMETHING, but of course if I had a /dev/SOMETHING I'd probably just go on writing to it directly. Here's what dmesg says after I disconnect and reconnect the device from the USB port: [ 924.722906] usb 1-1.1.4: USB disconnect, device number 7 [ 959.993002] usb 1-1.1.4: new full speed USB device number 8 using ehci_hcd And here's how it shows up in lsusb -v: Bus 001 Device 008: ID 1a86:7584 QinHeng Electronics CH340S Device Descriptor: bLength 18 bDescriptorType 1 bcdUSB 1.10 bDeviceClass 0 (Defined at Interface level) bDeviceSubClass 0 bDeviceProtocol 0 bMaxPacketSize0 8 idVendor 0x1a86 QinHeng Electronics idProduct 0x7584 CH340S bcdDevice 2.52 iManufacturer 0 iProduct 2 USB2.0-Print iSerial 0 bNumConfigurations 1 Configuration Descriptor: bLength 9 bDescriptorType 2 wTotalLength 32 bNumInterfaces 1 bConfigurationValue 1 iConfiguration 0 bmAttributes 0x80 (Bus Powered) MaxPower 96mA Interface Descriptor: bLength 9 bDescriptorType 4 bInterfaceNumber 0 bAlternateSetting 0 bNumEndpoints 2 bInterfaceClass 7 Printer bInterfaceSubClass 1 Printer bInterfaceProtocol 2 Bidirectional iInterface 0 Endpoint Descriptor: bLength 7 bDescriptorType 5 bEndpointAddress 0x82 EP 2 IN bmAttributes 2 Transfer Type Bulk Synch Type None Usage Type Data wMaxPacketSize 0x0020 1x 32 bytes bInterval 0 Endpoint Descriptor: bLength 7 bDescriptorType 5 bEndpointAddress 0x02 EP 2 OUT bmAttributes 2 Transfer Type Bulk Synch Type None Usage Type Data wMaxPacketSize 0x0020 1x 32 bytes bInterval 0 Device Status: 0x0000 (Bus Powered)

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  • Nginx HTTPS redirects causing loop

    - by Ben Chiappetta
    I've been banging my head against the wall trying to figure this out, so if anyone can help I'd appreciate it. My Nginx conf has three different redirect loops, haven't been able to get any of the three to work right. The three problem areas are: Redirecting memcache directory to SSL Redirecting accounts directory to SSL Redirecting SSL to www if non-www nginx.conf: user nginx; worker_processes 1; error_log /var/log/nginx/error.log warn; pid /var/run/nginx.pid; events { worker_connections 1024; } http { include /etc/nginx/mime.types; default_type application/octet-stream; log_format main '$remote_addr - $remote_user [$time_local] "$request" ' '$status $body_bytes_sent "$http_referer" ' '"$http_user_agent" "$http_x_forwarded_for"'; access_log /var/log/nginx/access.log main; error_log /var/log/nginx/error.log notice; sendfile on; #tcp_nopush on; keepalive_timeout 65; proxy_set_header X-Url-Scheme $scheme; #gzip on; rewrite_log on; include /etc/nginx/conf.d/*.conf; } conf.d/default.conf: server { listen 80; server_name <redacted>.net; rewrite ^(.*) http://www.<redacted>.net$1; } server { listen 80; server_name www.<redacted>.net; set_real_ip_from 192.168.30.4; set_real_ip_from 192.168.30.5; set_real_ip_from 192.168.30.10; real_ip_header X-Forwarded-For; #charset koi8-r; access_log /var/log/nginx/host.access.log main; root /var/www/html; index index.php index.html index.htm; location =/memcache { rewrite ^/(.*)$ https://$server_name$request_uri? permanent; } location /accounts { rewrite ^/(.*)$ https://$server_name$request_uri? permanent; } #error_page 404 /404.html; # redirect server error pages to the static page /50x.html # error_page 500 502 503 504 /50x.html; location = /50x.html { } # pass the PHP scripts to FastCGI server listening on 127.0.0.1:9000 # location ~ \.php$ { fastcgi_pass 127.0.0.1:9000; fastcgi_index index.php; fastcgi_param SCRIPT_FILENAME $document_root$fastcgi_script_name; include /etc/nginx/fastcgi_params; try_files $uri = 404; } # deny access to .htaccess files, if Apache's document root # concurs with nginx's one # location ~ /\.ht { deny all; } } conf.d/ssl.conf: # HTTPS server # server { listen 443; server_name <redacted>.net; rewrite ^(.*) https://www.<redacted>.net$1; } server { listen 443 default_server ssl; server_name www.<redacted>.net; set_real_ip_from 192.168.30.4; set_real_ip_from 192.168.30.5; set_real_ip_from 192.168.30.10; real_ip_header X-Forwarded-For; proxy_set_header X-Forwarded_Proto https; proxy_set_header Host $host; proxy_redirect off; proxy_max_temp_file_size 0; proxy_set_header X-Forwarded-Ssl on; set $https_enabled on; ssl_certificate <redacted>.crt; ssl_certificate_key <redacted>.key; ssl_session_timeout 5m; ssl_protocols SSLv2 SSLv3 TLSv1; ssl_ciphers HIGH:!aNULL:!MD5; ssl_prefer_server_ciphers on; root /var/www/html; index index.php index.html index.htm; location /memcache { auth_basic "Restricted"; auth_basic_user_file $document_root/memcache/.htpasswd; } location ~ \.php$ { fastcgi_pass 127.0.0.1:9000; fastcgi_index index.php; fastcgi_param SCRIPT_FILENAME $document_root$fastcgi_script_name; fastcgi_param HTTPS on; include /etc/nginx/fastcgi_params; try_files $uri = 404; } }

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  • 5x5 matrix multiplication in C

    - by Rick
    I am stuck on this problem in my homework. I've made it this far and am sure the problem is in my three for loops. The question directly says to use 3 for loops so I know this is probably just a logic error. #include<stdio.h> void matMult(int A[][5],int B[][5],int C[][5]); int printMat_5x5(int A[5][5]); int main() { int A[5][5] = {{1,2,3,4,6}, {6,1,5,3,8}, {2,6,4,9,9}, {1,3,8,3,4}, {5,7,8,2,5}}; int B[5][5] = {{3,5,0,8,7}, {2,2,4,8,3}, {0,2,5,1,2}, {1,4,0,5,1}, {3,4,8,2,3}}; int C[5][5] = {0}; matMult(A,B,C); printMat_5x5(A); printf("\n"); printMat_5x5(B); printf("\n"); printMat_5x5(C); return 0; } void matMult(int A[][5], int B[][5], int C[][5]) { int i; int j; int k; for(i = 0; i <= 2; i++) { for(j = 0; j <= 4; j++) { for(k = 0; k <= 3; k++) { C[i][j] += A[i][k] * B[k][j]; } } } } int printMat_5x5(int A[5][5]){ int i; int j; for (i = 0;i < 5;i++) { for(j = 0;j < 5;j++) { printf("%2d",A[i][j]); } printf("\n"); } } EDIT: Here is the question, sorry for not posting it the first time. (2) Write a C function to multiply two five by five matrices. The prototype should read void matMult(int a[][5],int b[][5],int c[][5]); The resulting matrix product (a times b) is returned in the two dimensional array c (the third parameter of the function). Program your solution using three nested for loops (each generating the counter values 0, 1, 2, 3, 4) That is, DO NOT code specific formulas for the 5 by 5 case in the problem, but make your code general so it can be easily changed to compute the product of larger square matrices. Write a main program to test your function using the arrays a: 1 2 3 4 6 6 1 5 3 8 2 6 4 9 9 1 3 8 3 4 5 7 8 2 5 b: 3 5 0 8 7 2 2 4 8 3 0 2 5 1 2 1 4 0 5 1 3 4 8 2 3 Print your matrices in a neat format using a C function created for printing five by five matrices. Print all three matrices. Generate your test arrays in your main program using the C array initialization feature. enter code here

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  • Python: Improving long cumulative sum

    - by Bo102010
    I have a program that operates on a large set of experimental data. The data is stored as a list of objects that are instances of a class with the following attributes: time_point - the time of the sample cluster - the name of the cluster of nodes from which the sample was taken code - the name of the node from which the sample was taken qty1 = the value of the sample for the first quantity qty2 = the value of the sample for the second quantity I need to derive some values from the data set, grouped in three ways - once for the sample as a whole, once for each cluster of nodes, and once for each node. The values I need to derive depend on the (time sorted) cumulative sums of qty1 and qty2: the maximum value of the element-wise sum of the cumulative sums of qty1 and qty2, the time point at which that maximum value occurred, and the values of qty1 and qty2 at that time point. I came up with the following solution: dataset.sort(key=operator.attrgetter('time_point')) # For the whole set sys_qty1 = 0 sys_qty2 = 0 sys_combo = 0 sys_max = 0 # For the cluster grouping cluster_qty1 = defaultdict(int) cluster_qty2 = defaultdict(int) cluster_combo = defaultdict(int) cluster_max = defaultdict(int) cluster_peak = defaultdict(int) # For the node grouping node_qty1 = defaultdict(int) node_qty2 = defaultdict(int) node_combo = defaultdict(int) node_max = defaultdict(int) node_peak = defaultdict(int) for t in dataset: # For the whole system ###################################################### sys_qty1 += t.qty1 sys_qty2 += t.qty2 sys_combo = sys_qty1 + sys_qty2 if sys_combo > sys_max: sys_max = sys_combo # The Peak class is to record the time point and the cumulative quantities system_peak = Peak(time_point=t.time_point, qty1=sys_qty1, qty2=sys_qty2) # For the cluster grouping ################################################## cluster_qty1[t.cluster] += t.qty1 cluster_qty2[t.cluster] += t.qty2 cluster_combo[t.cluster] = cluster_qty1[t.cluster] + cluster_qty2[t.cluster] if cluster_combo[t.cluster] > cluster_max[t.cluster]: cluster_max[t.cluster] = cluster_combo[t.cluster] cluster_peak[t.cluster] = Peak(time_point=t.time_point, qty1=cluster_qty1[t.cluster], qty2=cluster_qty2[t.cluster]) # For the node grouping ##################################################### node_qty1[t.node] += t.qty1 node_qty2[t.node] += t.qty2 node_combo[t.node] = node_qty1[t.node] + node_qty2[t.node] if node_combo[t.node] > node_max[t.node]: node_max[t.node] = node_combo[t.node] node_peak[t.node] = Peak(time_point=t.time_point, qty1=node_qty1[t.node], qty2=node_qty2[t.node]) This produces the correct output, but I'm wondering if it can be made more readable/Pythonic, and/or faster/more scalable. The above is attractive in that it only loops through the (large) dataset once, but unattractive in that I've essentially copied/pasted three copies of the same algorithm. To avoid the copy/paste issues of the above, I tried this also: def find_peaks(level, dataset): def grouping(object, attr_name): if attr_name == 'system': return attr_name else: return object.__dict__[attrname] cuml_qty1 = defaultdict(int) cuml_qty2 = defaultdict(int) cuml_combo = defaultdict(int) level_max = defaultdict(int) level_peak = defaultdict(int) for t in dataset: cuml_qty1[grouping(t, level)] += t.qty1 cuml_qty2[grouping(t, level)] += t.qty2 cuml_combo[grouping(t, level)] = (cuml_qty1[grouping(t, level)] + cuml_qty2[grouping(t, level)]) if cuml_combo[grouping(t, level)] > level_max[grouping(t, level)]: level_max[grouping(t, level)] = cuml_combo[grouping(t, level)] level_peak[grouping(t, level)] = Peak(time_point=t.time_point, qty1=node_qty1[grouping(t, level)], qty2=node_qty2[grouping(t, level)]) return level_peak system_peak = find_peaks('system', dataset) cluster_peak = find_peaks('cluster', dataset) node_peak = find_peaks('node', dataset) For the (non-grouped) system-level calculations, I also came up with this, which is pretty: dataset.sort(key=operator.attrgetter('time_point')) def cuml_sum(seq): rseq = [] t = 0 for i in seq: t += i rseq.append(t) return rseq time_get = operator.attrgetter('time_point') q1_get = operator.attrgetter('qty1') q2_get = operator.attrgetter('qty2') timeline = [time_get(t) for t in dataset] cuml_qty1 = cuml_sum([q1_get(t) for t in dataset]) cuml_qty2 = cuml_sum([q2_get(t) for t in dataset]) cuml_combo = [q1 + q2 for q1, q2 in zip(cuml_qty1, cuml_qty2)] combo_max = max(cuml_combo) time_max = timeline.index(combo_max) q1_at_max = cuml_qty1.index(time_max) q2_at_max = cuml_qty2.index(time_max) However, despite this version's cool use of list comprehensions and zip(), it loops through the dataset three times just for the system-level calculations, and I can't think of a good way to do the cluster-level and node-level calaculations without doing something slow like: timeline = defaultdict(int) cuml_qty1 = defaultdict(int) #...etc. for c in cluster_list: timeline[c] = [time_get(t) for t in dataset if t.cluster == c] cuml_qty1[c] = [q1_get(t) for t in dataset if t.cluster == c] #...etc. Does anyone here at Stack Overflow have suggestions for improvements? The first snippet above runs well for my initial dataset (on the order of a million records), but later datasets will have more records and clusters/nodes, so scalability is a concern. This is my first non-trivial use of Python, and I want to make sure I'm taking proper advantage of the language (this is replacing a very convoluted set of SQL queries, and earlier versions of the Python version were essentially very ineffecient straight transalations of what that did). I don't normally do much programming, so I may be missing something elementary. Many thanks!

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  • Minimum-Waste Print Job Grouping Algorithm?

    - by Matt Mc
    I work at a publishing house and I am setting up one of our presses for "ganging", in other words, printing multiple jobs simultaneously. Given that different print jobs can have different quantities, and anywhere from 1 to 20 jobs might need to be considered at a time, the problem would be to determine which jobs to group together to minimize waste (waste coming from over-printing on smaller-quantity jobs in a given set, that is). Given the following stable data: All jobs are equal in terms of spatial size--placement on paper doesn't come into consideration. There are three "lanes", meaning that three jobs can be printed simultaneously. Ideally, each lane has one job. Part of the problem is minimizing how many lanes each job is run on. If necessary, one job could be run on two lanes, with a second job on the third lane. The "grouping" waste from a given set of jobs (let's say the quantities of them are x, y and z) would be the highest number minus the two lower numbers. So if x is the higher number, the grouping waste would be (x - y) + (x - z). Otherwise stated, waste is produced by printing job Y and Z (in excess of their quantities) up to the quantity of X. The grouping waste would be a qualifier for the given set, meaning it could not exceed a certain quantity or the job would simply be printed alone. So the question is stated: how to determine which sets of jobs are grouped together, out of any given number of jobs, based on the qualifiers of 1) Three similar quantities OR 2) Two quantities where one is approximately double the other, AND with the aim of minimal total grouping waste across the various sets. (Edit) Quantity Information: Typical job quantities can be from 150 to 350 on foreign languages, or 500 to 1000 on English print runs. This data can be used to set up some scenarios for an algorithm. For example, let's say you had 5 jobs: 1000, 500, 500, 450, 250 By looking at it, I can see a couple of answers. Obviously (1000/500/500) is not efficient as you'll have a grouping waste of 1000. (500/500/450) is better as you'll have a waste of 50, but then you run (1000) and (250) alone. But you could also run (1000/500) with 1000 on two lanes, (500/250) with 500 on two lanes and then (450) alone. In terms of trade-offs for lane minimization vs. wastage, we could say that any grouping waste over 200 is excessive. (End Edit) ...Needless to say, quite a problem. (For me.) I am a moderately skilled programmer but I do not have much familiarity with algorithms and I am not fully studied in the mathematics of the area. I'm I/P writing a sort of brute-force program that simply tries all options, neglecting any option tree that seems to have excessive grouping waste. However, I can't help but hope there's an easier and more efficient method. I've looked at various websites trying to find out more about algorithms in general and have been slogging my way through the symbology, but it's slow going. Unfortunately, Wikipedia's articles on the subject are very cross-dependent and it's difficult to find an "in". The only thing I've been able to really find would seem to be a definition of the rough type of algorithm I need: "Exclusive Distance Clustering", one-dimensionally speaking. I did look at what seems to be the popularly referred-to algorithm on this site, the Bin Packing one, but I was unable to see exactly how it would work with my problem. Any help is appreciated. :)

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  • Explain the Peak and Flag Algorithm

    - by Isaac Levin
    EDIT Just was pointed that the requirements state peaks cannot be ends of Arrays. So I ran across this site http://codility.com/ Which gives you programming problems and gives you certificates if you can solve them in 2 hours. The very first question is one I have seen before, typically called the Peaks and Flags question. If you are not familiar A non-empty zero-indexed array A consisting of N integers is given. A peak is an array element which is larger than its neighbours. More precisely, it is an index P such that 0 < P < N - 1 and A[P - 1] < A[P] A[P + 1] . For example, the following array A: A[0] = 1 A[1] = 5 A[2] = 3 A[3] = 4 A[4] = 3 A[5] = 4 A[6] = 1 A[7] = 2 A[8] = 3 A[9] = 4 A[10] = 6 A[11] = 2 has exactly four peaks: elements 1, 3, 5 and 10. You are going on a trip to a range of mountains whose relative heights are represented by array A. You have to choose how many flags you should take with you. The goal is to set the maximum number of flags on the peaks, according to certain rules. Flags can only be set on peaks. What's more, if you take K flags, then the distance between any two flags should be greater than or equal to K. The distance between indices P and Q is the absolute value |P - Q|. For example, given the mountain range represented by array A, above, with N = 12, if you take: two flags, you can set them on peaks 1 and 5; three flags, you can set them on peaks 1, 5 and 10; four flags, you can set only three flags, on peaks 1, 5 and 10. You can therefore set a maximum of three flags in this case. Write a function that, given a non-empty zero-indexed array A of N integers, returns the maximum number of flags that can be set on the peaks of the array. For example, given the array above the function should return 3, as explained above. Assume that: N is an integer within the range [1..100,000]; each element of array A is an integer within the range [0..1,000,000,000]. Complexity: expected worst-case time complexity is O(N); expected worst-case space complexity is O(N), beyond input storage (not counting the storage required for input arguments). Elements of input arrays can be modified. So this makes sense, but I failed it using this code public int GetFlags(int[] A) { List<int> peakList = new List<int>(); for (int i = 0; i <= A.Length - 1; i++) { if ((A[i] > A[i + 1] && A[i] > A[i - 1])) { peakList.Add(i); } } List<int> flagList = new List<int>(); int distance = peakList.Count; flagList.Add(peakList[0]); for (int i = 1, j = 0, max = peakList.Count; i < max; i++) { if (Math.Abs(Convert.ToDecimal(peakList[j]) - Convert.ToDecimal(peakList[i])) >= distance) { flagList.Add(peakList[i]); j = i; } } return flagList.Count; } EDIT int[] A = new int[] { 7, 10, 4, 5, 7, 4, 6, 1, 4, 3, 3, 7 }; The correct answer is 3, but my application says 2 This I do not get, since there are 4 peaks (indices 1,4,6,8) and from that, you should be able to place a flag at 2 of the peaks (1 and 6) Am I missing something here? Obviously my assumption is that the beginning or end of an Array can be a peak, is this not the case? If this needs to go in Stack Exchange Programmers, I will move it, but thought dialog here would be helpful. EDIT

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  • Duplication of menu items with ViewPager and Fragments

    - by Julian
    I'm building an Android Application (minimum SDK Level 10, Gingerbread 2.3.3) with some Fragments in a ViewPager. I'm using ActionBarSherlock to create an ActionBar and android-viewpagertabs to add tabs to the ViewPager just like in the Market client. I have one global menu item that I want to be shown on every tab/fragment. On the first of the three tabs I want to have two additional menu items. But now two strange things happen: First if I start the app, everything seems to be fine, I can see all three menu items on the first page and only one item if i swipe to the second and third tab. But if I swipe back to the second tab from the third one, I can see all three items again which shouldn't happen. If I swipe back to the first and then again to the second tab, everything is fine again. The other strange thing is that every time I rotate the device, the menu items from the fragment are added again, even though they are already in the menu. Code of the FragmentActivity that displays the ViewPager and its tabs: public class MainActivity extends FragmentActivity { public static final String TAG = "MainActivity"; private ActionBar actionBar; private Adapter adapter; private ViewPager viewPager; private ViewPagerTabs tabs; @Override public void onCreate(Bundle savedInstanceState) { super.onCreate(savedInstanceState); setContentView(R.layout.volksempfaenger); actionBar = getSupportActionBar(); adapter = new Adapter(getSupportFragmentManager()); adapter.addFragment(getString(R.string.title_tab_subscriptions), SubscriptionGridFragment.class); // adding more fragments here viewPager = (ViewPager) findViewById(R.id.viewpager); viewPager.setAdapter(adapter); tabs = (ViewPagerTabs) findViewById(R.id.tabs); tabs.setViewPager(viewPager); } public static class Adapter extends FragmentPagerAdapter implements ViewPagerTabProvider { private FragmentManager fragmentManager; private ArrayList<Class<? extends Fragment>> fragments; private ArrayList<String> titles; public Adapter(FragmentManager fm) { super(fm); fragmentManager = fm; fragments = new ArrayList<Class<? extends Fragment>>(); titles = new ArrayList<String>(); } public void addFragment(String title, Class<? extends Fragment> fragment) { titles.add(title); fragments.add(fragment); } @Override public int getCount() { return fragments.size(); } public String getTitle(int position) { return titles.get(position); } @Override public Fragment getItem(int position) { try { return fragments.get(position).newInstance(); } catch (InstantiationException e) { Log.wtf(TAG, e); } catch (IllegalAccessException e) { Log.wtf(TAG, e); } return null; } @Override public Object instantiateItem(View container, int position) { FragmentTransaction fragmentTransaction = fragmentManager .beginTransaction(); Fragment f = getItem(position); fragmentTransaction.add(container.getId(), f); fragmentTransaction.commit(); return f; } } @Override public boolean onCreateOptionsMenu(Menu menu) { BaseActivity.addGlobalMenu(this, menu); return true; } @Override public boolean onOptionsItemSelected(MenuItem item) { return BaseActivity.handleGlobalMenu(this, item); } } Code of the fragment that shall have its own menu items: public class SubscriptionGridFragment extends Fragment { private GridView subscriptionList; private SubscriptionListAdapter adapter; @Override public void onCreate(Bundle savedInstanceState) { super.onCreate(savedInstanceState); setHasOptionsMenu(true); } // ... @Override public void onCreateOptionsMenu(Menu menu, MenuInflater inflater) { inflater.inflate(R.menu.subscription_list, menu); } @Override public boolean onOptionsItemSelected(MenuItem item) { // ... } }

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  • What can I do to improve a project if there is a no-listening situation. Developers vs Management

    - by NazGul
    Hi all, I hope that I'm not the only one and I can get a answer from someone with more experience than me, so I can think cleaner and I don't get depressed with this developer's life. I'm working as developer for a small company three years now. In that three years I'm working in the same project and sincerely, I think this project could be used as a CASE STUDY because it has all the situations that cannot happen in a project and that makes a project fails. To begin with, and I believe you've already noticed, the project has 3 years already (develoment only) and is still unfinished, because in every meeting there is a "new priority" ,or a "new problem" to be solve or a "new feature" to be add. So, first problem is no target set. How can you know when something is finished if you don't know what you want? I understand Management, because they see an oportunity and try to get that, but I don't understand how can they not see (or hear us) that they'll lose all they already have and what they'll eventually get. Second, there is no team group. My team consists of three people, a Senior Developer, a DBA and, finally, I for all the work (support, testing, new features, bug fixing, meeting, projet management of clients, etc) aka Junior Developer. The first (senior developer), does not perform any tests on his changes, so, most of the time, his changes give us problems (us = me, since I'm the one who will fix it). The second (DBA) is an uncompromising person and you can not talk to him, believe me, I tried! In his view, everything he does is fantastic... even if it is the most complicated to make it... And he does everything he wants, even if we need that only for 5 months later and would help some extra-hand to do the things we have to do for now. As you can see, there is very hard to work with no help... Third, there is no testings. Every... I repeat, Every release of the project, the customers wants to kill us, because there is a lot of bugs. Management? They say that they want tests before the release. Us? We say the same. Time? No time. Management? There is always some time to open the application and click in some buttons. Us? Try to explain that it is not so simple. Management doesn't care... end of story. Actually, must of the bugs could be avoid with a rigorous work... Some people just want to do the show to the Management. "Did you ask for this? Cool, it's done. Bugs? The Do-all-the-work guy will solve." Unfortunally for me, sometimes the Do-all-the-work also has to finish it. And to makes this all better, I'm the person who will listen the complaints from the customers. Cool, huh? I know, everyone makes mistakes. But there is mistakes and mistakes... To complete, in the Management view, "the problem is the lack of an individual project management", because we cannot do all the stuff they ask, even if there is no PM for the project itself. And ask us to work overtime without any reward... I do say all this stuff to the management and others members, but by telling this, the I'm the bad guy, the guy who is complain when everything is going well... but we need to work overtime... sigh What can I do to make it works? Anyone has a situation like this, what did you do? I hope you could understand my problem, my English is a little rusty. Thanks.

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  • Selecting unique records in XSLT/XPath

    - by Daniel I-S
    I have to select only unique records from an XML document, in the context of an <xsl:for-each> loop. I am limited by Visual Studio to using XSL 1.0. <availList> <item> <schDate>2010-06-24</schDate> <schFrmTime>10:00:00</schFrmTime> <schToTime>13:00:00</schToTime> <variousOtherElements></variousOtherElements> </item> <item> <schDate>2010-06-24</schDate> <schFrmTime>10:00:00</schFrmTime> <schToTime>13:00:00</schToTime> <variousOtherElements></variousOtherElements> </item> <item> <schDate>2010-06-25</schDate> <schFrmTime>10:00:00</schFrmTime> <schToTime>12:00:00</schToTime> <variousOtherElements></variousOtherElements> </item> <item> <schDate>2010-06-26</schDate> <schFrmTime>13:00:00</schFrmTime> <schToTime>14:00:00</schToTime> <variousOtherElements></variousOtherElements> </item> <item> <schDate>2010-06-26</schDate> <schFrmTime>10:00:00</schFrmTime> <schToTime>12:00:00</schToTime> <variousOtherElements></variousOtherElements> </item> </availList> The uniqueness must be based on the value of the three child elements: schDate, schFrmTime and schToTime. If two item elements have the same values for all three child elements, they are duplicates. In the above XML, items one and two are duplicates. The rest are unique. As indicated above, each item contains other elements that we do not wish to include in the comparison. 'Uniqueness' should be a factor of those three elements, and those alone. I have attempted to accomplish this through the following: availList/item[not(schDate = preceding:: schDate and schFrmTime = preceding:: schFrmTime and schToTime = preceding:: schToTime)] The idea behind this is to select records where there is no preceding element with the same schDate, schFrmTime and schToTime. However, its output is missing the last item. This is because my XPath is actually excluding items where all of the child element values are matched within the entire preceding document. No single item matches all of the last item's child elements - but because each element's value is individually present in another item, the last item gets excluded. I could get the correct result by comparing all child values as a concatenated string to the same concatenated values for each preceding item. Does anybody know of a way I could do this?

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  • Ajax Control Toolkit July 2011 Release and the New HTML Editor Extender

    - by Stephen Walther
    I’m happy to announce the July 2011 release of the Ajax Control Toolkit which includes important bug fixes and a completely new HTML Editor Extender control. You can download the July 2011 Release by visiting the Ajax Control Toolkit CodePlex site at: http://AjaxControlToolkit.CodePlex.com Using the New HTML Editor Extender Control You can use the new HTML Editor Extender to extend any standard ASP.NET TextBox control so that it supports rich formatting such as bold, italics, bulleted lists, numbered lists, typefaces and different foreground and background colors. The following code illustrates how you can extend a standard ASP.NET TextBox control with the HtmlEditorExtender: <%@ Page Language="C#" AutoEventWireup="true" CodeBehind="Simple.aspx.cs" Inherits="WebApplication1.Simple" %> <%@ Register TagPrefix="asp" Namespace="AjaxControlToolkit" Assembly="AjaxControlToolkit" %> <html xmlns="http://www.w3.org/1999/xhtml"> <head runat="server"> <title>Simple</title> </head> <body> <form id="form1" runat="server"> <asp:ToolkitScriptManager runat="Server" /> <asp:TextBox ID="txtComments" TextMode="MultiLine" Columns="60" Rows="8" runat="server" /> <asp:HtmlEditorExtender TargetControlID="txtComments" runat="server" /> </form> </body> </html> This page has the following three controls: ToolkitScriptManager – The ToolkitScriptManager renders all of the scripts required by the Ajax Control Toolkit. TextBox – The TextBox control is a standard ASP.NET TextBox which is set to display multiple lines (a TextArea instead of an Input element). HtmlEditorExtender – The HtmlEditorExtender is set to extend the TextBox control. You can use the standard TextBox Text property to read the rich text entered into the TextBox control on the server. Lightweight and HTML5 The HTML Editor Extender works on all modern browsers including the most recent versions of Mozilla Firefox (Firefox 5), Google Chrome (Chrome 12), and Apple Safari (Safari 5). Furthermore, the HTML Editor Extender is compatible with Microsoft Internet Explorer 6 and newer. The HTML Editor Extender is very lightweight. It takes advantage of the HTML5 ContentEditable attribute so it does not require an iframe or complex browser workarounds. If you select View Source in your browser while using the HTML Editor Extender, we hope that you will be pleasantly surprised by how little markup and script is generated by the HTML Editor Extender. Customizable Toolbar Buttons Depending on the web application that you are building, you will want to display different toolbar buttons with the HTML Editor Extender. One of the design goals of the HTML Editor Extender was to make it very easy for you to customize the toolbar buttons. Imagine, for example, that you want to use the HTML Editor Extender when accepting comments on blog posts. In that case, you might want to restrict the type of formatting that a user can display. You might want to enable a user to format text as bold or italic but you do not want the user to make any other formatting changes. The following page illustrates how you can customize the HTML Editor Extender toolbar: <%@ Page Language="C#" AutoEventWireup="true" CodeBehind="CustomToolbar.aspx.cs" Inherits="WebApplication1.CustomToolbar" %> <%@ Register TagPrefix="asp" Namespace="AjaxControlToolkit" Assembly="AjaxControlToolkit" %> <html> <head runat="server"> <title>Custom Toolbar</title> </head> <body> <form id="form1" runat="server"> <asp:ToolkitScriptManager Runat="server" /> <asp:TextBox ID="txtComments" TextMode="MultiLine" Columns="50" Rows="10" Text="Hello <b>world!</b>" Runat="server" /> <asp:HtmlEditorExtender TargetControlID="txtComments" runat="server"> <Toolbar> <asp:Bold /> <asp:Italic /> </Toolbar> </asp:HtmlEditorExtender> </form> </body> </html> Notice that the HTML Editor Extender in the page above has a Toolbar subtag. You can list the toolbar buttons which you want to appear within the subtag. In the case above, only Bold and Italic buttons are displayed. Here is a complete list of the Toolbar buttons currently supported by the HTML Editor Extender: Undo Redo Bold Italic Underline StrikeThrough Subscript Superscript JustifyLeft JustifyCenter JustifyRight JustifyFull InsertOrderedList InsertUnorderedList CreateLink UnLink RemoveFormat SelectAll UnSelect Delete Cut Copy Paste BackgroundColorSelector ForeColorSelector FontNameSelector FontSizeSelector Indent Outdent InsertHorizontalRule HorizontalSeparator Of course the HTML Editor Extender was designed to be extensible. You can create your own buttons and add them to the control. Compatible with the AntiXSS Library When using the HTML Editor Extender on a public facing website, we strongly recommend that you use the HTML Editor Extender with the AntiXSS Library. If you allow users to submit arbitrary HTML, and you don’t take any action to strip out malicious markup, then you are opening your website to Cross-Site Scripting Attacks (XSS attacks). The HTML Editor Extender uses the Provider Model to support different Sanitizer Providers. The July 2011 release of the Ajax Control Toolkit ships with a single Sanitizer Provider which uses the AntiXSS library (see http://AntiXss.CodePlex.com ). A Sanitizer Provider is responsible for sanitizing HTML markup by removing any malicious elements, attributes, and attribute values. For example, the AntiXss Sanitizer Provider will take the following block of HTML: <b><a href=""javascript:doEvil()"">Visit Grandma</a></b> <script>doEvil()</script> And return the following sanitized block of HTML: <b><a href="">Visit Grandma</a></b> Notice that the JavaScript href and <SCRIPT> tag are both stripped out. Be aware that there are a depressingly large number of ways to sneak evil markup into your HTML. You definitely want a Sanitizer as a safety net. Before you can use the AntiXSS Sanitizer Provider, you must add three assemblies to your web application: AntiXSSLibrary.dll, HtmlSanitizationLibrary.dll, and SanitizerProviders.dll. All three assemblies are included with the CodePlex download of the Ajax Control Toolkit in the SanitizerProviders folder. Here’s how you modify your web.config file to use the AntiXSS Sanitizer Provider: <configuration> <configSections> <sectionGroup name="system.web"> <section name="sanitizer" requirePermission="false" type="AjaxControlToolkit.Sanitizer.ProviderSanitizerSection, AjaxControlToolkit"/> </sectionGroup> </configSections> <system.web> <compilation targetFramework="4.0" debug="true"/> <sanitizer defaultProvider="AntiXssSanitizerProvider"> <providers> <add name="AntiXssSanitizerProvider" type="AjaxControlToolkit.Sanitizer.AntiXssSanitizerProvider"></add> </providers> </sanitizer> </system.web> </configuration> You can detect whether the HTML Editor Extender is using the AntiXSS Sanitizer Provider by checking the HtmlEditorExtender SanitizerProvider property like this: if (MyHtmlEditorExtender.SanitizerProvider == null) { throw new Exception("Please enable the AntiXss Sanitizer!"); } When the SanitizerProvider property has the value null, you know that a Sanitizer Provider has not been configured in the web.config file. Because the AntiXSS library requires Full Trust, you cannot use the AntiXSS Sanitizer Provider with most shared website hosting providers. Because most shared hosting providers only support Medium Trust and not Full Trust, we do not recommend using the HTML Editor Extender with a public website hosted with a shared hosting provider. Why a New HTML Editor Control? The Ajax Control Toolkit now includes two HTML Editor controls. Why did we introduce a new HTML Editor control when there was already an existing HTML Editor? We think you will like the new HTML Editor much more than the previous one. We had several goals with the new HTML Editor Extender: Lightweight – We wanted to leverage HTML5 to create a lightweight HTML Editor. The new HTML Editor generates much less markup and script than the previous HTML Editor. Secure – We wanted to make it easy to integrate the AntiXSS library with the HTML Editor. If you are creating a public facing website, we strongly recommend that you use the AntiXSS Provider. Customizable – We wanted to make it easy for users to customize the toolbar buttons displayed by the HTML Editor. Compatibility – We wanted to ensure that the HTML Editor will work with the latest versions of the most popular browsers (including Internet Explorer 6 and higher). The old HTML Editor control is still included in the Ajax Control Toolkit and continues to live in the AjaxControlToolkit.HTMLEditor namespace. We have not modified the control and you can continue to use the control in the same way as you have used it in the past. However, we hope that you will consider migrating to the new HTML Editor Extender for the reasons listed above. Summary We’ve introduced a new Ajax Control Toolkit control with this release. I want to thank the developers and testers on the Superexpert team for the huge amount of work which they put into this control. It was a non-trivial task to build an entirely new control which has the complexity of the HTML Editor in less than 6 weeks. Please let us know what you think! We want to hear your feedback. If you discover issues with the new HTML Editor Extender control, or you have questions about the control, or you have ideas for how it can be improved, then please post them to this blog. Tomorrow starts a new sprint

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  • RHEL hangs after starting virt-who succesfully

    - by Nick
    Idea #1: Is there a way to REPAIR an RHEL 6.2 installation? During the start-up procedure, after a recent forced reboot, my Linux machine (RHEL 6.2) hangs right after successfully starting virt-who. I can use login screens (Alt + F2/F3...) in text mode. I am clueless -- how can I find out what is the next step in the startup sequence? That step is most likely what is causing it to hang. These are the last lines saved to /var/log/boot.log: Starting RPC idmapd: [60G[[0;32m OK [0;39m] Starting cups: [60G[[0;32m OK [0;39m] Starting acpi daemon: [60G[[0;32m OK [0;39m] Starting HAL daemon: [60G[[0;32m OK [0;39m] Starting PC/SC smart card daemon (pcscd): [60G[[0;32m OK [0;39m] Retrigger failed udev events[60G[[0;32m OK [0;39m] Loading autofs4: [60G[[0;32m OK [0;39m] Starting automount: [60G[[0;32m OK [0;39m] Enabling Bluetooth devices: Starting sshd: [60G[[0;32m OK [0;39m] Starting ntpd: [60G[[0;32m OK [0;39m] Starting mysqld: [60G[[0;32m OK [0;39m] Starting postfix: [60G[[0;32m OK [0;39m] Starting abrt daemon: [60G[[0;32m OK [0;39m] Starting ksm: [60G[[0;32m OK [0;39m] Starting ksmtuned: [60G[[0;32m OK [0;39m] Starting Qpid AMQP daemon: [60G[[0;32m OK [0;39m] Starting crond: [60G[[0;32m OK [0;39m] Starting atd: [60G[[0;32m OK [0;39m] Starting libvirtd daemon: [60G[[0;32m OK [0;39m] Starting rhsmcertd 240 1440[60G[[0;32m OK [0;39m] Starting virt-who: [60G[[0;32m OK [0;39m]

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  • SQL SERVER – Guest Posts – Feodor Georgiev – The Context of Our Database Environment – Going Beyond the Internal SQL Server Waits – Wait Type – Day 21 of 28

    - by pinaldave
    This guest post is submitted by Feodor. Feodor Georgiev is a SQL Server database specialist with extensive experience of thinking both within and outside the box. He has wide experience of different systems and solutions in the fields of architecture, scalability, performance, etc. Feodor has experience with SQL Server 2000 and later versions, and is certified in SQL Server 2008. In this article Feodor explains the server-client-server process, and concentrated on the mutual waits between client and SQL Server. This is essential in grasping the concept of waits in a ‘global’ application plan. Recently I was asked to write a blog post about the wait statistics in SQL Server and since I had been thinking about writing it for quite some time now, here it is. It is a wide-spread idea that the wait statistics in SQL Server will tell you everything about your performance. Well, almost. Or should I say – barely. The reason for this is that SQL Server is always a part of a bigger system – there are always other players in the game: whether it is a client application, web service, any other kind of data import/export process and so on. In short, the SQL Server surroundings look like this: This means that SQL Server, aside from its internal waits, also depends on external waits and settings. As we can see in the picture above, SQL Server needs to have an interface in order to communicate with the surrounding clients over the network. For this communication, SQL Server uses protocol interfaces. I will not go into detail about which protocols are best, but you can read this article. Also, review the information about the TDS (Tabular data stream). As we all know, our system is only as fast as its slowest component. This means that when we look at our environment as a whole, the SQL Server might be a victim of external pressure, no matter how well we have tuned our database server performance. Let’s dive into an example: let’s say that we have a web server, hosting a web application which is using data from our SQL Server, hosted on another server. The network card of the web server for some reason is malfunctioning (think of a hardware failure, driver failure, or just improper setup) and does not send/receive data faster than 10Mbs. On the other end, our SQL Server will not be able to send/receive data at a faster rate either. This means that the application users will notify the support team and will say: “My data is coming very slow.” Now, let’s move on to a bit more exciting example: imagine that there is a similar setup as the example above – one web server and one database server, and the application is not using any stored procedure calls, but instead for every user request the application is sending 80kb query over the network to the SQL Server. (I really thought this does not happen in real life until I saw it one day.) So, what happens in this case? To make things worse, let’s say that the 80kb query text is submitted from the application to the SQL Server at least 100 times per minute, and as often as 300 times per minute in peak times. Here is what happens: in order for this query to reach the SQL Server, it will have to be broken into a of number network packets (according to the packet size settings) – and will travel over the network. On the other side, our SQL Server network card will receive the packets, will pass them to our network layer, the packets will get assembled, and eventually SQL Server will start processing the query – parsing, allegorizing, generating the query execution plan and so on. So far, we have already had a serious network overhead by waiting for the packets to reach our Database Engine. There will certainly be some processing overhead – until the database engine deals with the 80kb query and its 20 subqueries. The waits you see in the DMVs are actually collected from the point the query reaches the SQL Server and the packets are assembled. Let’s say that our query is processed and it finally returns 15000 rows. These rows have a certain size as well, depending on the data types returned. This means that the data will have converted to packages (depending on the network size package settings) and will have to reach the application server. There will also be waits, however, this time you will be able to see a wait type in the DMVs called ASYNC_NETWORK_IO. What this wait type indicates is that the client is not consuming the data fast enough and the network buffers are filling up. Recently Pinal Dave posted a blog on Client Statistics. What Client Statistics does is captures the physical flow characteristics of the query between the client(Management Studio, in this case) and the server and back to the client. As you see in the image, there are three categories: Query Profile Statistics, Network Statistics and Time Statistics. Number of server roundtrips–a roundtrip consists of a request sent to the server and a reply from the server to the client. For example, if your query has three select statements, and they are separated by ‘GO’ command, then there will be three different roundtrips. TDS Packets sent from the client – TDS (tabular data stream) is the language which SQL Server speaks, and in order for applications to communicate with SQL Server, they need to pack the requests in TDS packets. TDS Packets sent from the client is the number of packets sent from the client; in case the request is large, then it may need more buffers, and eventually might even need more server roundtrips. TDS packets received from server –is the TDS packets sent by the server to the client during the query execution. Bytes sent from client – is the volume of the data set to our SQL Server, measured in bytes; i.e. how big of a query we have sent to the SQL Server. This is why it is best to use stored procedures, since the reusable code (which already exists as an object in the SQL Server) will only be called as a name of procedure + parameters, and this will minimize the network pressure. Bytes received from server – is the amount of data the SQL Server has sent to the client, measured in bytes. Depending on the number of rows and the datatypes involved, this number will vary. But still, think about the network load when you request data from SQL Server. Client processing time – is the amount of time spent in milliseconds between the first received response packet and the last received response packet by the client. Wait time on server replies – is the time in milliseconds between the last request packet which left the client and the first response packet which came back from the server to the client. Total execution time – is the sum of client processing time and wait time on server replies (the SQL Server internal processing time) Here is an illustration of the Client-server communication model which should help you understand the mutual waits in a client-server environment. Keep in mind that a query with a large ‘wait time on server replies’ means the server took a long time to produce the very first row. This is usual on queries that have operators that need the entire sub-query to evaluate before they proceed (for example, sort and top operators). However, a query with a very short ‘wait time on server replies’ means that the query was able to return the first row fast. However a long ‘client processing time’ does not necessarily imply the client spent a lot of time processing and the server was blocked waiting on the client. It can simply mean that the server continued to return rows from the result and this is how long it took until the very last row was returned. The bottom line is that developers and DBAs should work together and think carefully of the resource utilization in the client-server environment. From experience I can say that so far I have seen only cases when the application developers and the Database developers are on their own and do not ask questions about the other party’s world. I would recommend using the Client Statistics tool during new development to track the performance of the queries, and also to find a synchronous way of utilizing resources between the client – server – client. Here is another example: think about similar setup as above, but add another server to the game. Let’s say that we keep our media on a separate server, and together with the data from our SQL Server we need to display some images on the webpage requested by our user. No matter how simple or complicated the logic to get the images is, if the images are 500kb each our users will get the page slowly and they will still think that there is something wrong with our data. Anyway, I don’t mean to get carried away too far from SQL Server. Instead, what I would like to say is that DBAs should also be aware of ‘the big picture’. I wrote a blog post a while back on this topic, and if you are interested, you can read it here about the big picture. And finally, here are some guidelines for monitoring the network performance and improving it: Run a trace and outline all queries that return more than 1000 rows (in Profiler you can actually filter and sort the captured trace by number of returned rows). This is not a set number; it is more of a guideline. The general thought is that no application user can consume that many rows at once. Ask yourself and your fellow-developers: ‘why?’. Monitor your network counters in Perfmon: Network Interface:Output queue length, Redirector:Network errors/sec, TCPv4: Segments retransmitted/sec and so on. Make sure to establish a good friendship with your network administrator (buy them coffee, for example J ) and get into a conversation about the network settings. Have them explain to you how the network cards are setup – are they standalone, are they ‘teamed’, what are the settings – full duplex and so on. Find some time to read a bit about networking. In this short blog post I hope I have turned your attention to ‘the big picture’ and the fact that there are other factors affecting our SQL Server, aside from its internal workings. As a further reading I would still highly recommend the Wait Stats series on this blog, also I would recommend you have the coffee break conversation with your network admin as soon as possible. This guest post is written by Feodor Georgiev. Read all the post in the Wait Types and Queue series. Reference: Pinal Dave (http://blog.SQLAuthority.com) Filed under: Pinal Dave, PostADay, Readers Contribution, SQL, SQL Authority, SQL Query, SQL Server, SQL Tips and Tricks, SQL Wait Stats, SQL Wait Types, T SQL

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  • Heaps of Trouble?

    - by Paul White NZ
    If you’re not already a regular reader of Brad Schulz’s blog, you’re missing out on some great material.  In his latest entry, he is tasked with optimizing a query run against tables that have no indexes at all.  The problem is, predictably, that performance is not very good.  The catch is that we are not allowed to create any indexes (or even new statistics) as part of our optimization efforts. In this post, I’m going to look at the problem from a slightly different angle, and present an alternative solution to the one Brad found.  Inevitably, there’s going to be some overlap between our entries, and while you don’t necessarily need to read Brad’s post before this one, I do strongly recommend that you read it at some stage; he covers some important points that I won’t cover again here. The Example We’ll use data from the AdventureWorks database, copied to temporary unindexed tables.  A script to create these structures is shown below: CREATE TABLE #Custs ( CustomerID INTEGER NOT NULL, TerritoryID INTEGER NULL, CustomerType NCHAR(1) COLLATE SQL_Latin1_General_CP1_CI_AI NOT NULL, ); GO CREATE TABLE #Prods ( ProductMainID INTEGER NOT NULL, ProductSubID INTEGER NOT NULL, ProductSubSubID INTEGER NOT NULL, Name NVARCHAR(50) COLLATE SQL_Latin1_General_CP1_CI_AI NOT NULL, ); GO CREATE TABLE #OrdHeader ( SalesOrderID INTEGER NOT NULL, OrderDate DATETIME NOT NULL, SalesOrderNumber NVARCHAR(25) COLLATE SQL_Latin1_General_CP1_CI_AI NOT NULL, CustomerID INTEGER NOT NULL, ); GO CREATE TABLE #OrdDetail ( SalesOrderID INTEGER NOT NULL, OrderQty SMALLINT NOT NULL, LineTotal NUMERIC(38,6) NOT NULL, ProductMainID INTEGER NOT NULL, ProductSubID INTEGER NOT NULL, ProductSubSubID INTEGER NOT NULL, ); GO INSERT #Custs ( CustomerID, TerritoryID, CustomerType ) SELECT C.CustomerID, C.TerritoryID, C.CustomerType FROM AdventureWorks.Sales.Customer C WITH (TABLOCK); GO INSERT #Prods ( ProductMainID, ProductSubID, ProductSubSubID, Name ) SELECT P.ProductID, P.ProductID, P.ProductID, P.Name FROM AdventureWorks.Production.Product P WITH (TABLOCK); GO INSERT #OrdHeader ( SalesOrderID, OrderDate, SalesOrderNumber, CustomerID ) SELECT H.SalesOrderID, H.OrderDate, H.SalesOrderNumber, H.CustomerID FROM AdventureWorks.Sales.SalesOrderHeader H WITH (TABLOCK); GO INSERT #OrdDetail ( SalesOrderID, OrderQty, LineTotal, ProductMainID, ProductSubID, ProductSubSubID ) SELECT D.SalesOrderID, D.OrderQty, D.LineTotal, D.ProductID, D.ProductID, D.ProductID FROM AdventureWorks.Sales.SalesOrderDetail D WITH (TABLOCK); The query itself is a simple join of the four tables: SELECT P.ProductMainID AS PID, P.Name, D.OrderQty, H.SalesOrderNumber, H.OrderDate, C.TerritoryID FROM #Prods P JOIN #OrdDetail D ON P.ProductMainID = D.ProductMainID AND P.ProductSubID = D.ProductSubID AND P.ProductSubSubID = D.ProductSubSubID JOIN #OrdHeader H ON D.SalesOrderID = H.SalesOrderID JOIN #Custs C ON H.CustomerID = C.CustomerID ORDER BY P.ProductMainID ASC OPTION (RECOMPILE, MAXDOP 1); Remember that these tables have no indexes at all, and only the single-column sampled statistics SQL Server automatically creates (assuming default settings).  The estimated query plan produced for the test query looks like this (click to enlarge): The Problem The problem here is one of cardinality estimation – the number of rows SQL Server expects to find at each step of the plan.  The lack of indexes and useful statistical information means that SQL Server does not have the information it needs to make a good estimate.  Every join in the plan shown above estimates that it will produce just a single row as output.  Brad covers the factors that lead to the low estimates in his post. In reality, the join between the #Prods and #OrdDetail tables will produce 121,317 rows.  It should not surprise you that this has rather dire consequences for the remainder of the query plan.  In particular, it makes a nonsense of the optimizer’s decision to use Nested Loops to join to the two remaining tables.  Instead of scanning the #OrdHeader and #Custs tables once (as it expected), it has to perform 121,317 full scans of each.  The query takes somewhere in the region of twenty minutes to run to completion on my development machine. A Solution At this point, you may be thinking the same thing I was: if we really are stuck with no indexes, the best we can do is to use hash joins everywhere. We can force the exclusive use of hash joins in several ways, the two most common being join and query hints.  A join hint means writing the query using the INNER HASH JOIN syntax; using a query hint involves adding OPTION (HASH JOIN) at the bottom of the query.  The difference is that using join hints also forces the order of the join, whereas the query hint gives the optimizer freedom to reorder the joins at its discretion. Adding the OPTION (HASH JOIN) hint results in this estimated plan: That produces the correct output in around seven seconds, which is quite an improvement!  As a purely practical matter, and given the rigid rules of the environment we find ourselves in, we might leave things there.  (We can improve the hashing solution a bit – I’ll come back to that later on). Faster Nested Loops It might surprise you to hear that we can beat the performance of the hash join solution shown above using nested loops joins exclusively, and without breaking the rules we have been set. The key to this part is to realize that a condition like (A = B) can be expressed as (A <= B) AND (A >= B).  Armed with this tremendous new insight, we can rewrite the join predicates like so: SELECT P.ProductMainID AS PID, P.Name, D.OrderQty, H.SalesOrderNumber, H.OrderDate, C.TerritoryID FROM #OrdDetail D JOIN #OrdHeader H ON D.SalesOrderID >= H.SalesOrderID AND D.SalesOrderID <= H.SalesOrderID JOIN #Custs C ON H.CustomerID >= C.CustomerID AND H.CustomerID <= C.CustomerID JOIN #Prods P ON P.ProductMainID >= D.ProductMainID AND P.ProductMainID <= D.ProductMainID AND P.ProductSubID = D.ProductSubID AND P.ProductSubSubID = D.ProductSubSubID ORDER BY D.ProductMainID OPTION (RECOMPILE, LOOP JOIN, MAXDOP 1, FORCE ORDER); I’ve also added LOOP JOIN and FORCE ORDER query hints to ensure that only nested loops joins are used, and that the tables are joined in the order they appear.  The new estimated execution plan is: This new query runs in under 2 seconds. Why Is It Faster? The main reason for the improvement is the appearance of the eager Index Spools, which are also known as index-on-the-fly spools.  If you read my Inside The Optimiser series you might be interested to know that the rule responsible is called JoinToIndexOnTheFly. An eager index spool consumes all rows from the table it sits above, and builds a index suitable for the join to seek on.  Taking the index spool above the #Custs table as an example, it reads all the CustomerID and TerritoryID values with a single scan of the table, and builds an index keyed on CustomerID.  The term ‘eager’ means that the spool consumes all of its input rows when it starts up.  The index is built in a work table in tempdb, has no associated statistics, and only exists until the query finishes executing. The result is that each unindexed table is only scanned once, and just for the columns necessary to build the temporary index.  From that point on, every execution of the inner side of the join is answered by a seek on the temporary index – not the base table. A second optimization is that the sort on ProductMainID (required by the ORDER BY clause) is performed early, on just the rows coming from the #OrdDetail table.  The optimizer has a good estimate for the number of rows it needs to sort at that stage – it is just the cardinality of the table itself.  The accuracy of the estimate there is important because it helps determine the memory grant given to the sort operation.  Nested loops join preserves the order of rows on its outer input, so sorting early is safe.  (Hash joins do not preserve order in this way, of course). The extra lazy spool on the #Prods branch is a further optimization that avoids executing the seek on the temporary index if the value being joined (the ‘outer reference’) hasn’t changed from the last row received on the outer input.  It takes advantage of the fact that rows are still sorted on ProductMainID, so if duplicates exist, they will arrive at the join operator one after the other. The optimizer is quite conservative about introducing index spools into a plan, because creating and dropping a temporary index is a relatively expensive operation.  It’s presence in a plan is often an indication that a useful index is missing. I want to stress that I rewrote the query in this way primarily as an educational exercise – I can’t imagine having to do something so horrible to a production system. Improving the Hash Join I promised I would return to the solution that uses hash joins.  You might be puzzled that SQL Server can create three new indexes (and perform all those nested loops iterations) faster than it can perform three hash joins.  The answer, again, is down to the poor information available to the optimizer.  Let’s look at the hash join plan again: Two of the hash joins have single-row estimates on their build inputs.  SQL Server fixes the amount of memory available for the hash table based on this cardinality estimate, so at run time the hash join very quickly runs out of memory. This results in the join spilling hash buckets to disk, and any rows from the probe input that hash to the spilled buckets also get written to disk.  The join process then continues, and may again run out of memory.  This is a recursive process, which may eventually result in SQL Server resorting to a bailout join algorithm, which is guaranteed to complete eventually, but may be very slow.  The data sizes in the example tables are not large enough to force a hash bailout, but it does result in multiple levels of hash recursion.  You can see this for yourself by tracing the Hash Warning event using the Profiler tool. The final sort in the plan also suffers from a similar problem: it receives very little memory and has to perform multiple sort passes, saving intermediate runs to disk (the Sort Warnings Profiler event can be used to confirm this).  Notice also that because hash joins don’t preserve sort order, the sort cannot be pushed down the plan toward the #OrdDetail table, as in the nested loops plan. Ok, so now we understand the problems, what can we do to fix it?  We can address the hash spilling by forcing a different order for the joins: SELECT P.ProductMainID AS PID, P.Name, D.OrderQty, H.SalesOrderNumber, H.OrderDate, C.TerritoryID FROM #Prods P JOIN #Custs C JOIN #OrdHeader H ON H.CustomerID = C.CustomerID JOIN #OrdDetail D ON D.SalesOrderID = H.SalesOrderID ON P.ProductMainID = D.ProductMainID AND P.ProductSubID = D.ProductSubID AND P.ProductSubSubID = D.ProductSubSubID ORDER BY D.ProductMainID OPTION (MAXDOP 1, HASH JOIN, FORCE ORDER); With this plan, each of the inputs to the hash joins has a good estimate, and no hash recursion occurs.  The final sort still suffers from the one-row estimate problem, and we get a single-pass sort warning as it writes rows to disk.  Even so, the query runs to completion in three or four seconds.  That’s around half the time of the previous hashing solution, but still not as fast as the nested loops trickery. Final Thoughts SQL Server’s optimizer makes cost-based decisions, so it is vital to provide it with accurate information.  We can’t really blame the performance problems highlighted here on anything other than the decision to use completely unindexed tables, and not to allow the creation of additional statistics. I should probably stress that the nested loops solution shown above is not one I would normally contemplate in the real world.  It’s there primarily for its educational and entertainment value.  I might perhaps use it to demonstrate to the sceptical that SQL Server itself is crying out for an index. Be sure to read Brad’s original post for more details.  My grateful thanks to him for granting permission to reuse some of his material. Paul White Email: [email protected] Twitter: @PaulWhiteNZ

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  • Building a Windows Phone 7 Twitter Application using Silverlight

    - by ScottGu
    On Monday I had the opportunity to present the MIX 2010 Day 1 Keynote in Las Vegas (you can watch a video of it here).  In the keynote I announced the release of the Silverlight 4 Release Candidate (we’ll ship the final release of it next month) and the VS 2010 RC tools for Silverlight 4.  I also had the chance to talk for the first time about how Silverlight and XNA can now be used to build Windows Phone 7 applications. During my talk I did two quick Windows Phone 7 coding demos using Silverlight – a quick “Hello World” application and a “Twitter” data-snacking application.  Both applications were easy to build and only took a few minutes to create on stage.  Below are the steps you can follow yourself to build them on your own machines as well. [Note: In addition to blogging, I am also now using Twitter for quick updates and to share links. Follow me at: twitter.com/scottgu] Building a “Hello World” Windows Phone 7 Application First make sure you’ve installed the Windows Phone Developer Tools CTP – this includes the Visual Studio 2010 Express for Windows Phone development tool (which will be free forever and is the only thing you need to develop and build Windows Phone 7 applications) as well as an add-on to the VS 2010 RC that enables phone development within the full VS 2010 as well. After you’ve downloaded and installed the Windows Phone Developer Tools CTP, launch the Visual Studio 2010 Express for Windows Phone that it installs or launch the VS 2010 RC (if you have it already installed), and then choose “File”->”New Project.”  Here, you’ll find the usual list of project template types along with a new category: “Silverlight for Windows Phone”. The first CTP offers two application project templates. The first is the “Windows Phone Application” template - this is what we’ll use for this example. The second is the “Windows Phone List Application” template - which provides the basic layout for a master-details phone application: After creating a new project, you’ll get a view of the design surface and markup. Notice that the design surface shows the phone UI, letting you easily see how your application will look while you develop. For those familiar with Visual Studio, you’ll also find the familiar ToolBox, Solution Explorer and Properties pane. For our HelloWorld application, we’ll start out by adding a TextBox and a Button from the Toolbox. Notice that you get the same design experience as you do for Silverlight on the web or desktop. You can easily resize, position and align your controls on the design surface. Changing properties is easy with the Properties pane. We’ll change the name of the TextBox that we added to username and change the page title text to “Hello world.” We’ll then write some code by double-clicking on the button and create an event handler in the code-behind file (MainPage.xaml.cs). We’ll start out by changing the title text of the application. The project template included this title as a TextBlock with the name textBlockListTitle (note that the current name incorrectly includes the word “list”; that will be fixed for the final release.)  As we write code against it we get intellisense showing the members available.  Below we’ll set the Text property of the title TextBlock to “Hello “ + the Text property of the TextBox username: We now have all the code necessary for a Hello World application.  We have two choices when it comes to deploying and running the application. We can either deploy to an actual device itself or use the built-in phone emulator: Because the phone emulator is actually the phone operating system running in a virtual machine, we’ll get the same experience developing in the emulator as on the device. For this sample, we’ll just press F5 to start the application with debugging using the emulator.  Once the phone operating system loads, the emulator will run the new “Hello world” application exactly as it would on the device: Notice that we can change several settings of the emulator experience with the emulator toolbar – which is a floating toolbar on the top right.  This includes the ability to re-size/zoom the emulator and two rotate buttons.  Zoom lets us zoom into even the smallest detail of the application: The orientation buttons allow us easily see what the application looks like in landscape mode (orientation change support is just built into the default template): Note that the emulator can be reused across F5 debug sessions - that means that we don’t have to start the emulator for every deployment. We’ve added a dialog that will help you from accidentally shutting down the emulator if you want to reuse it.  Launching an application on an already running emulator should only take ~3 seconds to deploy and run. Within our Hello World application we’ll click the “username” textbox to give it focus.  This will cause the software input panel (SIP) to open up automatically.  We can either type a message or – since we are using the emulator – just type in text.  Note that the emulator works with Windows 7 multi-touch so, if you have a touchscreen, you can see how interaction will feel on a device just by pressing the screen. We’ll enter “MIX 10” in the textbox and then click the button – this will cause the title to update to be “Hello MIX 10”: We provide the same Visual Studio experience when developing for the phone as other .NET applications. This means that we can set a breakpoint within the button event handler, press the button again and have it break within the debugger: Building a “Twitter” Windows Phone 7 Application using Silverlight Rather than just stop with “Hello World” let’s keep going and evolve it to be a basic Twitter client application. We’ll return to the design surface and add a ListBox, using the snaplines within the designer to fit it to the device screen and make the best use of phone screen real estate.  We’ll also rename the Button “Lookup”: We’ll then return to the Button event handler in Main.xaml.cs, and remove the original “Hello World” line of code and take advantage of the WebClient networking class to asynchronously download a Twitter feed. This takes three lines of code in total: (1) declaring and creating the WebClient, (2) attaching an event handler and then (3) calling the asynchronous DownloadStringAsync method. In the DownloadStringAsync call, we’ll pass a Twitter Uri plus a query string which pulls the text from the “username” TextBox. This feed will pull down the respective user’s most frequent posts in an XML format. When the call completes, the DownloadStringCompleted event is fired and our generated event handler twitter_DownloadStringCompleted will be called: The result returned from the Twitter call will come back in an XML based format.  To parse this we’ll use LINQ to XML. LINQ to XML lets us create simple queries for accessing data in an xml feed. To use this library, we’ll first need to add a reference to the assembly (right click on the References folder in the solution explorer and choose “Add Reference): We’ll then add a “using System.Xml.Linq” namespace reference at the top of the code-behind file at the top of Main.xaml.cs file: We’ll then add a simple helper class called TwitterItem to our project. TwitterItem has three string members – UserName, Message and ImageSource: We’ll then implement the twitter_DownloadStringCompleted event handler and use LINQ to XML to parse the returned XML string from Twitter.  What the query is doing is pulling out the three key pieces of information for each Twitter post from the username we passed as the query string. These are the ImageSource for their profile image, the Message of their tweet and their UserName. For each Tweet in the XML, we are creating a new TwitterItem in the IEnumerable<XElement> returned by the Linq query.  We then assign the generated TwitterItem sequence to the ListBox’s ItemsSource property: We’ll then do one more step to complete the application. In the Main.xaml file, we’ll add an ItemTemplate to the ListBox. For the demo, I used a simple template that uses databinding to show the user’s profile image, their tweet and their username. <ListBox Height="521" HorizonalAlignment="Left" Margin="0,131,0,0" Name="listBox1" VerticalAlignment="Top" Width="476"> <ListBox.ItemTemplate> <DataTemplate> <StackPanel Orientation="Horizontal" Height="132"> <Image Source="{Binding ImageSource}" Height="73" Width="73" VerticalAlignment="Top" Margin="0,10,8,0"/> <StackPanel Width="370"> <TextBlock Text="{Binding UserName}" Foreground="#FFC8AB14" FontSize="28" /> <TextBlock Text="{Binding Message}" TextWrapping="Wrap" FontSize="24" /> </StackPanel> </StackPanel> </DataTemplate> </ListBox.ItemTemplate> </ListBox> Now, pressing F5 again, we are able to reuse the emulator and re-run the application. Once the application has launched, we can type in a Twitter username and press the  Button to see the results. Try my Twitter user name (scottgu) and you’ll get back a result of TwitterItems in the Listbox: Try using the mouse (or if you have a touchscreen device your finger) to scroll the items in the Listbox – you should find that they move very fast within the emulator.  This is because the emulator is hardware accelerated – and so gives you the same fast performance that you get on the actual phone hardware. Summary Silverlight and the VS 2010 Tools for Windows Phone (and the corresponding Expression Blend Tools for Windows Phone) make building Windows Phone applications both really easy and fun.  At MIX this week a number of great partners (including Netflix, FourSquare, Seesmic, Shazaam, Major League Soccer, Graphic.ly, Associated Press, Jackson Fish and more) showed off some killer application prototypes they’ve built over the last few weeks.  You can watch my full day 1 keynote to see them in action. I think they start to show some of the promise and potential of using Silverlight with Windows Phone 7.  I’ll be doing more blog posts in the weeks and months ahead that cover that more. Hope this helps, Scott

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  • [GEEK SCHOOL] Network Security 1: Securing User Accounts and Passwords in Windows

    - by Matt Klein
    This How-To Geek School class is intended for people who want to learn more about security when using Windows operating systems. You will learn many principles that will help you have a more secure computing experience and will get the chance to use all the important security tools and features that are bundled with Windows. Obviously, we will share everything you need to know about using them effectively. In this first lesson, we will talk about password security; the different ways of logging into Windows and how secure they are. In the proceeding lesson, we will explain where Windows stores all the user names and passwords you enter while working in this operating systems, how safe they are, and how to manage this data. Moving on in the series, we will talk about User Account Control, its role in improving the security of your system, and how to use Windows Defender in order to protect your system from malware. Then, we will talk about the Windows Firewall, how to use it in order to manage the apps that get access to the network and the Internet, and how to create your own filtering rules. After that, we will discuss the SmartScreen Filter – a security feature that gets more and more attention from Microsoft and is now widely used in its Windows 8.x operating systems. Moving on, we will discuss ways to keep your software and apps up-to-date, why this is important and which tools you can use to automate this process as much as possible. Last but not least, we will discuss the Action Center and its role in keeping you informed about what’s going on with your system and share several tips and tricks about how to stay safe when using your computer and the Internet. Let’s get started by discussing everyone’s favorite subject: passwords. The Types of Passwords Found in Windows In Windows 7, you have only local user accounts, which may or may not have a password. For example, you can easily set a blank password for any user account, even if that one is an administrator. The only exception to this rule are business networks where domain policies force all user accounts to use a non-blank password. In Windows 8.x, you have both local accounts and Microsoft accounts. If you would like to learn more about them, don’t hesitate to read the lesson on User Accounts, Groups, Permissions & Their Role in Sharing, in our Windows Networking series. Microsoft accounts are obliged to use a non-blank password due to the fact that a Microsoft account gives you access to Microsoft services. Using a blank password would mean exposing yourself to lots of problems. Local accounts in Windows 8.1 however, can use a blank password. On top of traditional passwords, any user account can create and use a 4-digit PIN or a picture password. These concepts were introduced by Microsoft to speed up the sign in process for the Windows 8.x operating system. However, they do not replace the use of a traditional password and can be used only in conjunction with a traditional user account password. Another type of password that you encounter in Windows operating systems is the Homegroup password. In a typical home network, users can use the Homegroup to easily share resources. A Homegroup can be joined by a Windows device only by using the Homegroup password. If you would like to learn more about the Homegroup and how to use it for network sharing, don’t hesitate to read our Windows Networking series. What to Keep in Mind When Creating Passwords, PINs and Picture Passwords When creating passwords, a PIN, or a picture password for your user account, we would like you keep in mind the following recommendations: Do not use blank passwords, even on the desktop computers in your home. You never know who may gain unwanted access to them. Also, malware can run more easily as administrator because you do not have a password. Trading your security for convenience when logging in is never a good idea. When creating a password, make it at least eight characters long. Make sure that it includes a random mix of upper and lowercase letters, numbers, and symbols. Ideally, it should not be related in any way to your name, username, or company name. Make sure that your passwords do not include complete words from any dictionary. Dictionaries are the first thing crackers use to hack passwords. Do not use the same password for more than one account. All of your passwords should be unique and you should use a system like LastPass, KeePass, Roboform or something similar to keep track of them. When creating a PIN use four different digits to make things slightly harder to crack. When creating a picture password, pick a photo that has at least 10 “points of interests”. Points of interests are areas that serve as a landmark for your gestures. Use a random mixture of gesture types and sequence and make sure that you do not repeat the same gesture twice. Be aware that smudges on the screen could potentially reveal your gestures to others. The Security of Your Password vs. the PIN and the Picture Password Any kind of password can be cracked with enough effort and the appropriate tools. There is no such thing as a completely secure password. However, passwords created using only a few security principles are much harder to crack than others. If you respect the recommendations shared in the previous section of this lesson, you will end up having reasonably secure passwords. Out of all the log in methods in Windows 8.x, the PIN is the easiest to brute force because PINs are restricted to four digits and there are only 10,000 possible unique combinations available. The picture password is more secure than the PIN because it provides many more opportunities for creating unique combinations of gestures. Microsoft have compared the two login options from a security perspective in this post: Signing in with a picture password. In order to discourage brute force attacks against picture passwords and PINs, Windows defaults to your traditional text password after five failed attempts. The PIN and the picture password function only as alternative login methods to Windows 8.x. Therefore, if someone cracks them, he or she doesn’t have access to your user account password. However, that person can use all the apps installed on your Windows 8.x device, access your files, data, and so on. How to Create a PIN in Windows 8.x If you log in to a Windows 8.x device with a user account that has a non-blank password, then you can create a 4-digit PIN for it, to use it as a complementary login method. In order to create one, you need to go to “PC Settings”. If you don’t know how, then press Windows + C on your keyboard or flick from the right edge of the screen, on a touch-enabled device, then press “Settings”. The Settings charm is now open. Click or tap the link that says “Change PC settings”, on the bottom of the charm. In PC settings, go to Accounts and then to “Sign-in options”. Here you will find all the necessary options for changing your existing password, creating a PIN, or a picture password. To create a PIN, press the “Add” button in the PIN section. The “Create a PIN” wizard is started and you are asked to enter the password of your user account. Type it and press “OK”. Now you are asked to enter a 4-digit pin in the “Enter PIN” and “Confirm PIN” fields. The PIN has been created and you can now use it to log in to Windows. How to Create a Picture Password in Windows 8.x If you log in to a Windows 8.x device with a user account that has a non-blank password, then you can also create a picture password and use it as a complementary login method. In order to create one, you need to go to “PC settings”. In PC Settings, go to Accounts and then to “Sign-in options”. Here you will find all the necessary options for changing your existing password, creating a PIN, or a picture password. To create a picture password, press the “Add” button in the “Picture password” section. The “Create a picture password” wizard is started and you are asked to enter the password of your user account. You are shown a guide on how the picture password works. Take a few seconds to watch it and learn the gestures that can be used for your picture password. You will learn that you can create a combination of circles, straight lines, and taps. When ready, press “Choose picture”. Browse your Windows 8.x device and select the picture you want to use for your password and press “Open”. Now you can drag the picture to position it the way you want. When you like how the picture is positioned, press “Use this picture” on the left. If you are not happy with the picture, press “Choose new picture” and select a new one, as shown during the previous step. After you have confirmed that you want to use this picture, you are asked to set up your gestures for the picture password. Draw three gestures on the picture, any combination you wish. Please remember that you can use only three gestures: circles, straight lines, and taps. Once you have drawn those three gestures, you are asked to confirm. Draw the same gestures one more time. If everything goes well, you are informed that you have created your picture password and that you can use it the next time you sign in to Windows. If you don’t confirm the gestures correctly, you will be asked to try again, until you draw the same gestures twice. To close the picture password wizard, press “Finish”. Where Does Windows Store Your Passwords? Are They Safe? All the passwords that you enter in Windows and save for future use are stored in the Credential Manager. This tool is a vault with the usernames and passwords that you use to log on to your computer, to other computers on the network, to apps from the Windows Store, or to websites using Internet Explorer. By storing these credentials, Windows can automatically log you the next time you access the same app, network share, or website. Everything that is stored in the Credential Manager is encrypted for your protection.

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  • Squid configuration for proxy server

    - by Ian Rob
    I have a server with 10 ip's that I want to give access to some friends via authentication but I'm stuck on squid's config file. Let's say I have these ip's available on my server: 212.77.23.10 212.77.1.10 68.44.82.112 And I want to allocate each one of them to a different user like so: 212.77.23.10 goes to user manilodisan using password 123456 212.77.1.10 goes to user manilodisan1 using password 123456 68.44.82.112 goes to user manilodisan2 using password 123456 I managed to add the passwords and authentication works ok but how do I do to restrict one user to one of the available ip's? I have a basic setup from different bits I found over the internet but nothing seems to work. Here's my squid.conf (all comments are removed to make it lighter): acl ip1 myip 212.77.23.10 acl ip2 myip 212.77.1.10 tcp_outgoing_address 212.77.23.10 ip1 tcp_outgoing_address 212.77.1.10 ip2 http_port 8888 visible_hostname weezie auth_param basic program /usr/lib/squid/ncsa_auth /etc/squid/squid-passwd acl ncsa_users proxy_auth REQUIRED http_access allow ncsa_users acl all src 0.0.0.0/0.0.0.0 acl manager proto cache_object acl localhost src 127.0.0.1/255.255.255.255 acl to_localhost dst 127.0.0.0/8 acl SSL_ports port 443 # https acl SSL_ports port 563 # snews acl SSL_ports port 873 # rsync acl Safe_ports port 80 # http acl Safe_ports port 21 # ftp acl Safe_ports port 443 # https acl Safe_ports port 70 # gopher acl Safe_ports port 210 # wais acl Safe_ports port 1025-65535 # unregistered ports acl Safe_ports port 280 # http-mgmt acl Safe_ports port 488 # gss-http acl Safe_ports port 591 # filemaker acl Safe_ports port 777 # multiling http acl Safe_ports port 631 # cups acl Safe_ports port 873 # rsync acl Safe_ports port 901 # SWAT acl purge method PURGE acl CONNECT method CONNECT http_access allow manager localhost http_access deny manager http_access allow purge localhost http_access deny purge http_access deny !Safe_ports http_access deny CONNECT !SSL_ports http_access allow localhost http_access deny all icp_access allow all hierarchy_stoplist cgi-bin ? access_log /var/log/squid/access.log squid acl QUERY urlpath_regex cgi-bin \? cache deny QUERY refresh_pattern ^ftp: 1440 20% 10080 refresh_pattern ^gopher: 1440 0% 1440 refresh_pattern . 0 20% 4320 acl apache rep_header Server ^Apache broken_vary_encoding allow apache extension_methods REPORT MERGE MKACTIVITY CHECKOUT hosts_file /etc/hosts forwarded_for off coredump_dir /var/spool/squid

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  • Uploading and Importing CSV file to SQL Server in ASP.NET WebForms

    - by Vincent Maverick Durano
    Few weeks ago I was working with a small internal project  that involves importing CSV file to Sql Server database and thought I'd share the simple implementation that I did on the project. In this post I will demonstrate how to upload and import CSV file to SQL Server database. As some may have already know, importing CSV file to SQL Server is easy and simple but difficulties arise when the CSV file contains, many columns with different data types. Basically, the provider cannot differentiate data types between the columns or the rows, blindly it will consider them as a data type based on first few rows and leave all the data which does not match the data type. To overcome this problem, I used schema.ini file to define the data type of the CSV file and allow the provider to read that and recognize the exact data types of each column. Now what is schema.ini? Taken from the documentation: The Schema.ini is a information file, used to define the data structure and format of each column that contains data in the CSV file. If schema.ini file exists in the directory, Microsoft.Jet.OLEDB provider automatically reads it and recognizes the data type information of each column in the CSV file. Thus, the provider intelligently avoids the misinterpretation of data types before inserting the data into the database. For more information see: http://msdn.microsoft.com/en-us/library/ms709353%28VS.85%29.aspx Points to remember before creating schema.ini:   1. The schema information file, must always named as 'schema.ini'.   2. The schema.ini file must be kept in the same directory where the CSV file exists.   3. The schema.ini file must be created before reading the CSV file.   4. The first line of the schema.ini, must the name of the CSV file, followed by the properties of the CSV file, and then the properties of the each column in the CSV file. Here's an example of how the schema looked like: [Employee.csv] ColNameHeader=False Format=CSVDelimited DateTimeFormat=dd-MMM-yyyy Col1=EmployeeID Long Col2=EmployeeFirstName Text Width 100 Col3=EmployeeLastName Text Width 50 Col4=EmployeeEmailAddress Text Width 50 To get started lets's go a head and create a simple blank database. Just for the purpose of this demo I created a database called TestDB. After creating the database then lets go a head and fire up Visual Studio and then create a new WebApplication project. Under the root application create a folder called UploadedCSVFiles and then place the schema.ini on that folder. The uploaded CSV files will be stored in this folder after the user imports the file. Now add a WebForm in the project and set up the HTML mark up and add one (1) FileUpload control one(1)Button and three (3) Label controls. After that we can now proceed with the codes for uploading and importing the CSV file to SQL Server database. Here are the full code blocks below: 1: using System; 2: using System.Data; 3: using System.Data.SqlClient; 4: using System.Data.OleDb; 5: using System.IO; 6: using System.Text; 7:   8: namespace WebApplication1 9: { 10: public partial class CSVToSQLImporting : System.Web.UI.Page 11: { 12: private string GetConnectionString() 13: { 14: return System.Configuration.ConfigurationManager.ConnectionStrings["DBConnectionString"].ConnectionString; 15: } 16: private void CreateDatabaseTable(DataTable dt, string tableName) 17: { 18:   19: string sqlQuery = string.Empty; 20: string sqlDBType = string.Empty; 21: string dataType = string.Empty; 22: int maxLength = 0; 23: StringBuilder sb = new StringBuilder(); 24:   25: sb.AppendFormat(string.Format("CREATE TABLE {0} (", tableName)); 26:   27: for (int i = 0; i < dt.Columns.Count; i++) 28: { 29: dataType = dt.Columns[i].DataType.ToString(); 30: if (dataType == "System.Int32") 31: { 32: sqlDBType = "INT"; 33: } 34: else if (dataType == "System.String") 35: { 36: sqlDBType = "NVARCHAR"; 37: maxLength = dt.Columns[i].MaxLength; 38: } 39:   40: if (maxLength > 0) 41: { 42: sb.AppendFormat(string.Format(" {0} {1} ({2}), ", dt.Columns[i].ColumnName, sqlDBType, maxLength)); 43: } 44: else 45: { 46: sb.AppendFormat(string.Format(" {0} {1}, ", dt.Columns[i].ColumnName, sqlDBType)); 47: } 48: } 49:   50: sqlQuery = sb.ToString(); 51: sqlQuery = sqlQuery.Trim().TrimEnd(','); 52: sqlQuery = sqlQuery + " )"; 53:   54: using (SqlConnection sqlConn = new SqlConnection(GetConnectionString())) 55: { 56: sqlConn.Open(); 57: SqlCommand sqlCmd = new SqlCommand(sqlQuery, sqlConn); 58: sqlCmd.ExecuteNonQuery(); 59: sqlConn.Close(); 60: } 61:   62: } 63: private void LoadDataToDatabase(string tableName, string fileFullPath, string delimeter) 64: { 65: string sqlQuery = string.Empty; 66: StringBuilder sb = new StringBuilder(); 67:   68: sb.AppendFormat(string.Format("BULK INSERT {0} ", tableName)); 69: sb.AppendFormat(string.Format(" FROM '{0}'", fileFullPath)); 70: sb.AppendFormat(string.Format(" WITH ( FIELDTERMINATOR = '{0}' , ROWTERMINATOR = '\n' )", delimeter)); 71:   72: sqlQuery = sb.ToString(); 73:   74: using (SqlConnection sqlConn = new SqlConnection(GetConnectionString())) 75: { 76: sqlConn.Open(); 77: SqlCommand sqlCmd = new SqlCommand(sqlQuery, sqlConn); 78: sqlCmd.ExecuteNonQuery(); 79: sqlConn.Close(); 80: } 81: } 82: protected void Page_Load(object sender, EventArgs e) 83: { 84:   85: } 86: protected void BTNImport_Click(object sender, EventArgs e) 87: { 88: if (FileUpload1.HasFile) 89: { 90: FileInfo fileInfo = new FileInfo(FileUpload1.PostedFile.FileName); 91: if (fileInfo.Name.Contains(".csv")) 92: { 93:   94: string fileName = fileInfo.Name.Replace(".csv", "").ToString(); 95: string csvFilePath = Server.MapPath("UploadedCSVFiles") + "\\" + fileInfo.Name; 96:   97: //Save the CSV file in the Server inside 'MyCSVFolder' 98: FileUpload1.SaveAs(csvFilePath); 99:   100: //Fetch the location of CSV file 101: string filePath = Server.MapPath("UploadedCSVFiles") + "\\"; 102: string strSql = "SELECT * FROM [" + fileInfo.Name + "]"; 103: string strCSVConnString = "Provider=Microsoft.Jet.OLEDB.4.0;Data Source=" + filePath + ";" + "Extended Properties='text;HDR=YES;'"; 104:   105: // load the data from CSV to DataTable 106:   107: OleDbDataAdapter adapter = new OleDbDataAdapter(strSql, strCSVConnString); 108: DataTable dtCSV = new DataTable(); 109: DataTable dtSchema = new DataTable(); 110:   111: adapter.FillSchema(dtCSV, SchemaType.Mapped); 112: adapter.Fill(dtCSV); 113:   114: if (dtCSV.Rows.Count > 0) 115: { 116: CreateDatabaseTable(dtCSV, fileName); 117: Label2.Text = string.Format("The table ({0}) has been successfully created to the database.", fileName); 118:   119: string fileFullPath = filePath + fileInfo.Name; 120: LoadDataToDatabase(fileName, fileFullPath, ","); 121:   122: Label1.Text = string.Format("({0}) records has been loaded to the table {1}.", dtCSV.Rows.Count, fileName); 123: } 124: else 125: { 126: LBLError.Text = "File is empty."; 127: } 128: } 129: else 130: { 131: LBLError.Text = "Unable to recognize file."; 132: } 133:   134: } 135: } 136: } 137: } The code above consists of three (3) private methods which are the GetConnectionString(), CreateDatabaseTable() and LoadDataToDatabase(). The GetConnectionString() is a method that returns a string. This method basically gets the connection string that is configured in the web.config file. The CreateDatabaseTable() is method that accepts two (2) parameters which are the DataTable and the filename. As the method name already suggested, this method automatically create a Table to the database based on the source DataTable and the filename of the CSV file. The LoadDataToDatabase() is a method that accepts three (3) parameters which are the tableName, fileFullPath and delimeter value. This method is where the actual saving or importing of data from CSV to SQL server happend. The codes at BTNImport_Click event handles the uploading of CSV file to the specified location and at the same time this is where the CreateDatabaseTable() and LoadDataToDatabase() are being called. If you notice I also added some basic trappings and validations within that event. Now to test the importing utility then let's create a simple data in a CSV format. Just for the simplicity of this demo let's create a CSV file and name it as "Employee" and add some data on it. Here's an example below: 1,VMS,Durano,[email protected] 2,Jennifer,Cortes,[email protected] 3,Xhaiden,Durano,[email protected] 4,Angel,Santos,[email protected] 5,Kier,Binks,[email protected] 6,Erika,Bird,[email protected] 7,Vianne,Durano,[email protected] 8,Lilibeth,Tree,[email protected] 9,Bon,Bolger,[email protected] 10,Brian,Jones,[email protected] Now save the newly created CSV file in some location in your hard drive. Okay let's run the application and browse the CSV file that we have just created. Take a look at the sample screen shots below: After browsing the CSV file. After clicking the Import Button Now if we look at the database that we have created earlier you'll notice that the Employee table is created with the imported data on it. See below screen shot.   That's it! I hope someone find this post useful! Technorati Tags: ASP.NET,CSV,SQL,C#,ADO.NET

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  • Inside the Concurrent Collections: ConcurrentBag

    - by Simon Cooper
    Unlike the other concurrent collections, ConcurrentBag does not really have a non-concurrent analogy. As stated in the MSDN documentation, ConcurrentBag is optimised for the situation where the same thread is both producing and consuming items from the collection. We'll see how this is the case as we take a closer look. Again, I recommend you have ConcurrentBag open in a decompiler for reference. Thread Statics ConcurrentBag makes heavy use of thread statics - static variables marked with ThreadStaticAttribute. This is a special attribute that instructs the CLR to scope any values assigned to or read from the variable to the executing thread, not globally within the AppDomain. This means that if two different threads assign two different values to the same thread static variable, one value will not overwrite the other, and each thread will see the value they assigned to the variable, separately to any other thread. This is a very useful function that allows for ConcurrentBag's concurrency properties. You can think of a thread static variable: [ThreadStatic] private static int m_Value; as doing the same as: private static Dictionary<Thread, int> m_Values; where the executing thread's identity is used to automatically set and retrieve the corresponding value in the dictionary. In .NET 4, this usage of ThreadStaticAttribute is encapsulated in the ThreadLocal class. Lists of lists ConcurrentBag, at its core, operates as a linked list of linked lists: Each outer list node is an instance of ThreadLocalList, and each inner list node is an instance of Node. Each outer ThreadLocalList is owned by a particular thread, accessible through the thread local m_locals variable: private ThreadLocal<ThreadLocalList<T>> m_locals It is important to note that, although the m_locals variable is thread-local, that only applies to accesses through that variable. The objects referenced by the thread (each instance of the ThreadLocalList object) are normal heap objects that are not specific to any thread. Thinking back to the Dictionary analogy above, if each value stored in the dictionary could be accessed by other means, then any thread could access the value belonging to other threads using that mechanism. Only reads and writes to the variable defined as thread-local are re-routed by the CLR according to the executing thread's identity. So, although m_locals is defined as thread-local, the m_headList, m_nextList and m_tailList variables aren't. This means that any thread can access all the thread local lists in the collection by doing a linear search through the outer linked list defined by these variables. Adding items So, onto the collection operations. First, adding items. This one's pretty simple. If the current thread doesn't already own an instance of ThreadLocalList, then one is created (or, if there are lists owned by threads that have stopped, it takes control of one of those). Then the item is added to the head of that thread's list. That's it. Don't worry, it'll get more complicated when we account for the other operations on the list! Taking & Peeking items This is where it gets tricky. If the current thread's list has items in it, then it peeks or removes the head item (not the tail item) from the local list and returns that. However, if the local list is empty, it has to go and steal another item from another list, belonging to a different thread. It iterates through all the thread local lists in the collection using the m_headList and m_nextList variables until it finds one that has items in it, and it steals one item from that list. Up to this point, the two threads had been operating completely independently. To steal an item from another thread's list, the stealing thread has to do it in such a way as to not step on the owning thread's toes. Recall how adding and removing items both operate on the head of the thread's linked list? That gives us an easy way out - a thread trying to steal items from another thread can pop in round the back of another thread's list using the m_tail variable, and steal an item from the back without the owning thread knowing anything about it. The owning thread can carry on completely independently, unaware that one of its items has been nicked. However, this only works when there are at least 3 items in the list, as that guarantees there will be at least one node between the owning thread performing operations on the list head and the thread stealing items from the tail - there's no chance of the two threads operating on the same node at the same time and causing a race condition. If there's less than three items in the list, then there does need to be some synchronization between the two threads. In this case, the lock on the ThreadLocalList object is used to mediate access to a thread's list when there's the possibility of contention. Thread synchronization In ConcurrentBag, this is done using several mechanisms: Operations performed by the owner thread only take out the lock when there are less than three items in the collection. With three or greater items, there won't be any conflict with a stealing thread operating on the tail of the list. If a lock isn't taken out, the owning thread sets the list's m_currentOp variable to a non-zero value for the duration of the operation. This indicates to all other threads that there is a non-locked operation currently occuring on that list. The stealing thread always takes out the lock, to prevent two threads trying to steal from the same list at the same time. After taking out the lock, the stealing thread spinwaits until m_currentOp has been set to zero before actually performing the steal. This ensures there won't be a conflict with the owning thread when the number of items in the list is on the 2-3 item borderline. If any add or remove operations are started in the meantime, and the list is below 3 items, those operations try to take out the list's lock and are blocked until the stealing thread has finished. This allows a thread to steal an item from another thread's list without corrupting it. What about synchronization in the collection as a whole? Collection synchronization Any thread that operates on the collection's global structure (accessing anything outside the thread local lists) has to take out the collection's global lock - m_globalListsLock. This single lock is sufficient when adding a new thread local list, as the items inside each thread's list are unaffected. However, what about operations (such as Count or ToArray) that need to access every item in the collection? In order to ensure a consistent view, all operations on the collection are stopped while the count or ToArray is performed. This is done by freezing the bag at the start, performing the global operation, and unfreezing at the end: The global lock is taken out, to prevent structural alterations to the collection. m_needSync is set to true. This notifies all the threads that they need to take out their list's lock irregardless of what operation they're doing. All the list locks are taken out in order. This blocks all locking operations on the lists. The freezing thread waits for all current lockless operations to finish by spinwaiting on each m_currentOp field. The global operation can then be performed while the bag is frozen, but no other operations can take place at the same time, as all other threads are blocked on a list's lock. Then, once the global operation has finished, the locks are released, m_needSync is unset, and normal concurrent operation resumes. Concurrent principles That's the essence of how ConcurrentBag operates. Each thread operates independently on its own local list, except when they have to steal items from another list. When stealing, only the stealing thread is forced to take out the lock; the owning thread only has to when there is the possibility of contention. And a global lock controls accesses to the structure of the collection outside the thread lists. Operations affecting the entire collection take out all locks in the collection to freeze the contents at a single point in time. So, what principles can we extract here? Threads operate independently Thread-static variables and ThreadLocal makes this easy. Threads operate entirely concurrently on their own structures; only when they need to grab data from another thread is there any thread contention. Minimised lock-taking Even when two threads need to operate on the same data structures (one thread stealing from another), they do so in such a way such that the probability of actually blocking on a lock is minimised; the owning thread always operates on the head of the list, and the stealing thread always operates on the tail. Management of lockless operations Any operations that don't take out a lock still have a 'hook' to force them to lock when necessary. This allows all operations on the collection to be stopped temporarily while a global snapshot is taken. Hopefully, such operations will be short-lived and infrequent. That's all the concurrent collections covered. I hope you've found it as informative and interesting as I have. Next, I'll be taking a closer look at ThreadLocal, which I came across while analyzing ConcurrentBag. As you'll see, the operation of this class deserves a much closer look.

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