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• Can Google Employees See My Saved Google Chrome Passwords?

  - by Jason Fitzpatrick

  Storing your passwords in your web browser seems like a great time saver, but are the passwords secure and inaccessible to others (even employees of the browser company) when squirreled away? Today’s Question & Answer session comes to us courtesy of SuperUser—a subdivision of Stack Exchange, a community-driven grouping of Q&A web sites. The Question SuperUser reader MMA is curious if Google employees have (or could have) access to the passwords he stores in Google Chrome: I understand that we are really tempted to save our passwords in Google Chrome. The likely benefit is two fold, You don’t need to (memorize and) input those long and cryptic passwords. These are available wherever you are once you log in to your Google account. The last point sparked my doubt. Since the password is available anywhere, the storage must in some central location, and this should be at Google. Now, my simple question is, can a Google employee see my passwords? Searching over the Internet revealed several articles/messages. Do you save passwords in Chrome? Maybe you should reconsider: Talks about your passwords being stolen by someone who has access to your computer account. Nothing mentioned about the central storage security and vulnerability. There is even a response from Chrome browser security tech lead about the first issue. Chrome’s insane password security strategy: Mostly along the same line. You can steal password from somebody if you have access to the computer account. How to Steal Passwords Saved in Google Chrome in 5 Simple Steps: Teaches you how to actually perform the act mentioned in the previous two when you have access to somebody else’s account. There are many more (including this one at this site), mostly along the same line, points, counter-points, huge debates. I refrain from mentioning them here, simply carry a search if you want to find them. Coming back to my original query, can a Google employee see my password? Since I can view the password using a simple button, definitely they can be unhashed (decrypted) even if encrypted. This is very different from the passwords saved in Unix-like OS’s where the saved password can never be seen in plain text. They use a one-way encryption algorithm to encrypt your passwords. This encrypted password is then stored in the passwd or shadow file. When you attempt to login, the password you type in is encrypted again and compared with the entry in the file that stores your passwords. If they match, it must be the same password, and you are allowed access. Thus, a superuser can change my password, can block my account, but he can never see my password. So are his concerns well founded or will a little insight dispel his worry? The Answer SuperUser contributor Zeel helps put his mind at ease: Short answer: No* Passwords stored on your local machine can be decrypted by Chrome, as long as your OS user account is logged in. And then you can view those in plain text. At first this seems horrible, but how did you think auto-fill worked? When that password field gets filled in, Chrome must insert the real password into the HTML form element – or else the page wouldn’t work right, and you could not submit the form. And if the connection to the website is not over HTTPS, the plain text is then sent over the internet. In other words, if chrome can’t get the plain text passwords, then they are totally useless. A one way hash is no good, because we need to use them. Now the passwords are in fact encrypted, the only way to get them back to plain text is to have the decryption key. That key is your Google password, or a secondary key you can set up. When you sign into Chrome and sync the Google servers will transmit the encrypted passwords, settings, bookmarks, auto-fill, etc, to your local machine. Here Chrome will decrypt the information and be able to use it. On Google’s end all that info is stored in its encrpyted state, and they do not have the key to decrypt it. Your account password is checked against a hash to log in to Google, and even if you let chrome remember it, that encrypted version is hidden in the same bundle as the other passwords, impossible to access. So an employee could probably grab a dump of the encrypted data, but it wouldn’t do them any good, since they would have no way to use it.* So no, Google employees can not** access your passwords, since they are encrypted on their servers. * However, do not forget that any system that can be accessed by an authorized user can be accessed by an unauthorized user. Some systems are easier to break than other, but none are fail-proof. . . That being said, I think I will trust Google and the millions they spend on security systems, over any other password storage solution. And heck, I’m a wimpy nerd, it would be easier to beat the passwords out of me than break Google’s encryption. ** I am also assuming that there isn’t a person who just happens to work for Google gaining access to your local machine. In that case you are screwed, but employment at Google isn’t actually a factor any more. Moral: Hit Win + L before leaving machine. While we agree with zeel that it’s a pretty safe bet (as long as your computer is not compromised) that your passwords are in fact safe while stored in Chrome, we prefer to encrypt all our logins and passwords in a LastPass vault. Have something to add to the explanation? Sound off in the the comments. Want to read more answers from other tech-savvy Stack Exchange users? Check out the full discussion thread here.     

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• Scheduling thread tiles with C++ AMP

  - by Daniel Moth

  This post assumes you are totally comfortable with, what some of us call, the simple model of C++ AMP, i.e. you could write your own matrix multiplication. We are now ready to explore the tiled model, which builds on top of the non-tiled one. Tiling the extent We know that when we pass a grid (which is just an extent under the covers) to the parallel_for_each call, it determines the number of threads to schedule and their index values (including dimensionality). For the single-, two-, and three- dimensional cases you can go a step further and subdivide the threads into what we call tiles of threads (others may call them thread groups). So here is a single-dimensional example: extent<1> e(20); // 20 units in a single dimension with indices from 0-19 grid<1> g(e);      // same as extent tiled_grid<4> tg = g.tile<4>(); …on the 3rd line we subdivided the single-dimensional space into 5 single-dimensional tiles each having 4 elements, and we captured that result in a concurrency::tiled_grid (a new class in amp.h). Let's move on swiftly to another example, in pictures, this time 2-dimensional: So we start on the left with a grid of a 2-dimensional extent which has 8*6=48 threads. We then have two different examples of tiling. In the first case, in the middle, we subdivide the 48 threads into tiles where each has 4*3=12 threads, hence we have 2*2=4 tiles. In the second example, on the right, we subdivide the original input into tiles where each has 2*2=4 threads, hence we have 4*3=12 tiles. Notice how you can play with the tile size and achieve different number of tiles. The numbers you pick must be such that the original total number of threads (in our example 48), remains the same, and every tile must have the same size. Of course, you still have no clue why you would do that, but stick with me. First, we should see how we can use this tiled_grid, since the parallel_for_each function that we know expects a grid. Tiled parallel_for_each and tiled_index It turns out that we have additional overloads of parallel_for_each that accept a tiled_grid instead of a grid. However, those overloads, also expect that the lambda you pass in accepts a concurrency::tiled_index (new in amp.h), not an index<N>. So how is a tiled_index different to an index? A tiled_index object, can have only 1 or 2 or 3 dimensions (matching exactly the tiled_grid), and consists of 4 index objects that are accessible via properties: global, local, tile_origin, and tile. The global index is the same as the index we know and love: the global thread ID. The local index is the local thread ID within the tile. The tile_origin index returns the global index of the thread that is at position 0,0 of this tile, and the tile index is the position of the tile in relation to the overall grid. Confused? Here is an example accompanied by a picture that hopefully clarifies things: array_view<int, 2> data(8, 6, p_my_data); parallel_for_each(data.grid.tile<2,2>(), [=] (tiled_index<2,2> t_idx) restrict(direct3d) { /* todo */ }); Given the code above and the picture on the right, what are the values of each of the 4 index objects that the t_idx variables exposes, when the lambda is executed by T (highlighted in the picture on the right)? If you can't work it out yourselves, the solution follows: t_idx.global       = index<2> (6,3) t_idx.local          = index<2> (0,1) t_idx.tile_origin = index<2> (6,2) t_idx.tile             = index<2> (3,1) Don't move on until you are comfortable with this… the picture really helps, so use it. Tiled Matrix Multiplication Example – part 1 Let's paste here the C++ AMP matrix multiplication example, bolding the lines we are going to change (can you guess what the changes will be?) 01: void MatrixMultiplyTiled_Part1(vector<float>& vC, const vector<float>& vA, const vector<float>& vB, int M, int N, int W) 02: { 03: 04: array_view<const float,2> a(M, W, vA); 05: array_view<const float,2> b(W, N, vB); 06: array_view<writeonly<float>,2> c(M, N, vC); 07: parallel_for_each(c.grid, 08: [=](index<2> idx) restrict(direct3d) { 09: 10: int row = idx[0]; int col = idx[1]; 11: float sum = 0.0f; 12: for(int i = 0; i < W; i++) 13: sum += a(row, i) * b(i, col); 14: c[idx] = sum; 15: }); 16: } To turn this into a tiled example, first we need to decide our tile size. Let's say we want each tile to be 16*16 (which assumes that we'll have at least 256 threads to process, and that c.grid.extent.size() is divisible by 256, and moreover that c.grid.extent[0] and c.grid.extent[1] are divisible by 16). So we insert at line 03 the tile size (which must be a compile time constant). 03: static const int TS = 16; ...then we need to tile the grid to have tiles where each one has 16*16 threads, so we change line 07 to be as follows 07: parallel_for_each(c.grid.tile<TS,TS>(), ...that means that our index now has to be a tiled_index with the same characteristics as the tiled_grid, so we change line 08 08: [=](tiled_index<TS, TS> t_idx) restrict(direct3d) { ...which means, without changing our core algorithm, we need to be using the global index that the tiled_index gives us access to, so we insert line 09 as follows 09: index<2> idx = t_idx.global; ...and now this code just works and it is tiled! Closing thoughts on part 1 The process we followed just shows the mechanical transformation that can take place from the simple model to the tiled model (think of this as step 1). In fact, when we wrote the matrix multiplication example originally, the compiler was doing this mechanical transformation under the covers for us (and it has additional smarts to deal with the cases where the total number of threads scheduled cannot be divisible by the tile size). The point is that the thread scheduling is always tiled, even when you use the non-tiled model. But with this mechanical transformation, we haven't gained anything… Hint: our goal with explicitly using the tiled model is to gain even more performance. In the next post, we'll evolve this further (beyond what the compiler can automatically do for us, in this first release), so you can see the full usage of the tiled model and its benefits… Comments about this post by Daniel Moth welcome at the original blog.

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• Master-slave vs. peer-to-peer archictecture: benefits and problems

  - by Ashok_Ora

  Normal 0 false false false EN-US X-NONE X-NONE Almost two decades ago, I was a member of a database development team that introduced adaptive locking. Locking, the most popular concurrency control technique in database systems, is pessimistic. Locking ensures that two or more conflicting operations on the same data item don’t “trample” on each other’s toes, resulting in data corruption. In a nutshell, here’s the issue we were trying to address. In everyday life, traffic lights serve the same purpose. They ensure that traffic flows smoothly and when everyone follows the rules, there are no accidents at intersections. As I mentioned earlier, the problem with typical locking protocols is that they are pessimistic. Regardless of whether there is another conflicting operation in the system or not, you have to hold a lock! Acquiring and releasing locks can be quite expensive, depending on how many objects the transaction touches. Every transaction has to pay this penalty. To use the earlier traffic light analogy, if you have ever waited at a red light in the middle of nowhere with no one on the road, wondering why you need to wait when there’s clearly no danger of a collision, you know what I mean. The adaptive locking scheme that we invented was able to minimize the number of locks that a transaction held, by detecting whether there were one or more transactions that needed conflicting eyou could get by without holding any lock at all. In many “well-behaved” workloads, there are few conflicts, so this optimization is a huge win. If, on the other hand, there are many concurrent, conflicting requests, the algorithm gracefully degrades to the “normal” behavior with minimal cost. We were able to reduce the number of lock requests per TPC-B transaction from 178 requests down to 2! Wow! This is a dramatic improvement in concurrency as well as transaction latency. The lesson from this exercise was that if you can identify the common scenario and optimize for that case so that only the uncommon scenarios are more expensive, you can make dramatic improvements in performance without sacrificing correctness. So how does this relate to the architecture and design of some of the modern NoSQL systems? NoSQL systems can be broadly classified as master-slave sharded, or peer-to-peer sharded systems. NoSQL systems with a peer-to-peer architecture have an interesting way of handling changes. Whenever an item is changed, the client (or an intermediary) propagates the changes synchronously or asynchronously to multiple copies (for availability) of the data. Since the change can be propagated asynchronously, during some interval in time, it will be the case that some copies have received the update, and others haven’t. What happens if someone tries to read the item during this interval? The client in a peer-to-peer system will fetch the same item from multiple copies and compare them to each other. If they’re all the same, then every copy that was queried has the same (and up-to-date) value of the data item, so all’s good. If not, then the system provides a mechanism to reconcile the discrepancy and to update stale copies. So what’s the problem with this? There are two major issues: First, IT’S HORRIBLY PESSIMISTIC because, in the common case, it is unlikely that the same data item will be updated and read from different locations at around the same time! For every read operation, you have to read from multiple copies. That’s a pretty expensive, especially if the data are stored in multiple geographically separate locations and network latencies are high. Second, if the copies are not all the same, the application has to reconcile the differences and propagate the correct value to the out-dated copies. This means that the application program has to handle discrepancies in the different versions of the data item and resolve the issue (which can further add to cost and operation latency). Resolving discrepancies is only one part of the problem. What if the same data item was updated independently on two different nodes (copies)? In that case, due to the asynchronous nature of change propagation, you might land up with different versions of the data item in different copies. In this case, the application program also has to resolve conflicts and then propagate the correct value to the copies that are out-dated or have incorrect versions. This can get really complicated. My hunch is that there are many peer-to-peer-based applications that don’t handle this correctly, and worse, don’t even know it. Imagine have 100s of millions of records in your database – how can you tell whether a particular data item is incorrect or out of date? And what price are you willing to pay for ensuring that the data can be trusted? Multiple network messages per read request? Discrepancy and conflict resolution logic in the application, and potentially, additional messages? All this overhead, when all you were trying to do was to read a data item. Wouldn’t it be simpler to avoid this problem in the first place? Master-slave architectures like the Oracle NoSQL Database handles this very elegantly. A change to a data item is always sent to the master copy. Consequently, the master copy always has the most current and authoritative version of the data item. The master is also responsible for propagating the change to the other copies (for availability and read scalability). Client drivers are aware of master copies and replicas, and client drivers are also aware of the “currency” of a replica. In other words, each NoSQL Database client knows how stale a replica is. This vastly simplifies the job of the application developer. If the application needs the most current version of the data item, the client driver will automatically route the request to the master copy. If the application is willing to tolerate some staleness of data (e.g. a version that is no more than 1 second out of date), the client can easily determine which replica (or set of replicas) can satisfy the request, and route the request to the most efficient copy. This results in a dramatic simplification in application logic and also minimizes network requests (the driver will only send the request to exactl the right replica, not many). So, back to my original point. A well designed and well architected system minimizes or eliminates unnecessary overhead and avoids pessimistic algorithms wherever possible in order to deliver a highly efficient and high performance system. If you’ve every programmed an Oracle NoSQL Database application, you’ll know the difference! /* 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-fareast-font-family:"Times New Roman"; mso-fareast-theme-font:minor-fareast; mso-hansi-font-family:Calibri; mso-hansi-theme-font:minor-latin;}

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• Who could ask for more with LESS CSS? (Part 3 of 3&ndash;Clrizr)

  - by ToString(theory);

  Welcome back!  In the first two posts in this series, I covered some of the awesome features in CSS precompilers such as SASS and LESS, as well as how to get an initial project setup up and running in ASP.Net MVC 4. In this post, I will cover an actual advanced example of using LESS in a project, and show some of the great productivity features we gain from its usage. Introduction In the first post, I mentioned two subjects that I will be using in this example – constants, and color functions.  I’ve always enjoyed using online color scheme utilities such as Adobe Kuler or Color Scheme Designer to come up with a scheme based off of one primary color.  Using these tools, and requesting a complementary scheme you can get a couple of shades of your primary color, and a couple of shades of a complementary/accent color to display. Because there is no way in regular css to do color operations or store variables, there was no way to accomplish something like defining a primary color, and have a site theme cascade off of that.  However with tools such as LESS, that impossibility becomes a reality!  So, if you haven’t guessed it by now, this post is on the creation of a plugin/module/less file to drop into your project, plugin one color, and have your primary theme cascade from it.  I only went through the trouble of creating a module for getting Complementary colors.  However, it wouldn’t be too much trouble to go through other options such as Triad or Monochromatic to get a module that you could use off of that. Step 1 – Analysis I decided to mimic Adobe Kuler’s Complementary theme algorithm as I liked its simplicity and aesthetics.  Color Scheme Designer is great, but I do believe it can give you too many color options, which can lead to chaos and overload.  The first thing I had to check was if the complementary values for the color schemes were actually hues rotated by 180 degrees at all times – they aren’t.  Apparently Adobe applies some variance to the complementary colors to get colors that are actually more aesthetically appealing to users.  So, I opened up Excel and began to plot complementary hues based on rotation in increments of 10: Long story short, I completed the same calculations for Hue, Saturation, and Lightness.  For Hue, I only had to record the Complementary hue values, however for saturation and lightness, I had to record the values for ALL of the shades.  Since the functions were too complicated to put into LESS since they aren’t constant/linear, but rather interval functions, I instead opted to extrapolate the HSL values using the trendline function for each major interval, onto intervals of spacing 1. For example, using the hue extraction, I got the following values: Interval Function 0-60 60-140 140-270 270-360 Saturation and Lightness were much worse, but in the end, I finally had functions for all of the intervals, and then went the route of just grabbing each shades value in intervals of 1.  Step 2 – Mapping I declared variable names for each of these sections as something that shouldn’t ever conflict with a variable someone would define in their own file.  After I had each of the values, I extracted the values and put them into files of their own for hue variables, saturation variables, and lightness variables…  Example: /*HUE CONVERSIONS*/@clrizr-hue-source-0deg: 133.43;@clrizr-hue-source-1deg: 135.601;@clrizr-hue-source-2deg: 137.772;@clrizr-hue-source-3deg: 139.943;@clrizr-hue-source-4deg: 142.114;.../*SATURATION CONVERSIONS*/@clrizr-saturation-s2SV0px: 0;@clrizr-saturation-s2SV1px: 0;@clrizr-saturation-s2SV2px: 0;@clrizr-saturation-s2SV3px: 0;@clrizr-saturation-s2SV4px: 0;.../*LIGHTNESS CONVERSIONS*/@clrizr-lightness-s2LV0px: 30;@clrizr-lightness-s2LV1px: 31;@clrizr-lightness-s2LV2px: 32;@clrizr-lightness-s2LV3px: 33;@clrizr-lightness-s2LV4px: 34;...   In the end, I have 973 lines of mapping/conversion from source HSL to shade HSL for two extra primary shades, and two complementary shades. The last bit of the work was the file to compose each of the shades from these mappings. Step 3 – Clrizr Mapper The final step was the hardest to overcome as I was still trying to understand LESS to its fullest extent.  Imports As mentioned previously, I had separated the HSL mappings into different files, so the first necessary step is to import those for use into the Clrizr plugin: @import url("hue.less");@import url("saturation.less");@import url("lightness.less"); Extract Component Values For Each Shade Next, I extracted the necessary information for each shade HSL before shade composition: @clrizr-input-saturation: 1px+floor(saturation(@clrizr-input))-1;@clrizr-input-lightness: 1px+floor(lightness(@clrizr-input))-1; @clrizr-complementary-hue: formatstring("clrizr-hue-source-{0}", ceil(hue(@clrizr-input))); @clrizr-primary-2-saturation: formatstring("clrizr-saturation-s2SV{0}",@clrizr-input-saturation);@clrizr-primary-1-saturation: formatstring("clrizr-saturation-s1SV{0}",@clrizr-input-saturation);@clrizr-complementary-1-saturation: formatstring("clrizr-saturation-c1SV{0}",@clrizr-input-saturation); @clrizr-primary-2-lightness: formatstring("clrizr-lightness-s2LV{0}",@clrizr-input-lightness);@clrizr-primary-1-lightness: formatstring("clrizr-lightness-s1LV{0}",@clrizr-input-lightness);@clrizr-complementary-1-lightness: formatstring("clrizr-lightness-c1LV{0}",@clrizr-input-lightness); Here, you can see a couple of odd things…  On the first line, I am using operations to add units to the saturation and lightness.  This is due to some limitations in the operations that would give me saturation or lightness in %, which can’t be in a variable name.  So, I use first add 1px to it, which casts the result of the following functions as px instead of %, and then at the end, I remove that pixel.  You can also see here the formatstring method which is exactly what it sounds like – something like String.Format(string str, params object[] obj). Get Primary & Complementary Shades Now that I have components for each of the different shades, I can now compose them into each of their pieces.  For this, I use the @@ operator which will look for a variable with the name specified in a string, and then call that variable: @clrizr-primary-2: hsl(hue(@clrizr-input), @@clrizr-primary-2-saturation, @@clrizr-primary-2-lightness);@clrizr-primary-1: hsl(hue(@clrizr-input), @@clrizr-primary-1-saturation, @@clrizr-primary-1-lightness);@clrizr-primary: @clrizr-input;@clrizr-complementary-1: hsl(@@clrizr-complementary-hue, @@clrizr-complementary-1-saturation, @@clrizr-complementary-1-lightness);@clrizr-complementary-2: hsl(@@clrizr-complementary-hue, saturation(@clrizr-input), lightness(@clrizr-input)); That’s is it, for the most part.  These variables now hold the theme for the one input color – @clrizr-input.  However, I have one last addition… Perceptive Luminance Well, after I got the colors, I decided I wanted to also get the best font color that would go on top of it.  Black or white depending on light or dark color.  Now I couldn’t just go with checking the lightness, as that is half the story.  You see, the human eye doesn’t see ALL colors equally well but rather has more cells for interpreting green light compared to blue or red.  So, using the ratio, we can calculate the perceptive luminance of each of the shades, and get the font color that best matches it! @clrizr-perceptive-luminance-ps2: round(1 - ( (0.299 * red(@clrizr-primary-2) ) + ( 0.587 * green(@clrizr-primary-2) ) + (0.114 * blue(@clrizr-primary-2)))/255)*255;@clrizr-perceptive-luminance-ps1: round(1 - ( (0.299 * red(@clrizr-primary-1) ) + ( 0.587 * green(@clrizr-primary-1) ) + (0.114 * blue(@clrizr-primary-1)))/255)*255;@clrizr-perceptive-luminance-ps: round(1 - ( (0.299 * red(@clrizr-primary) ) + ( 0.587 * green(@clrizr-primary) ) + (0.114 * blue(@clrizr-primary)))/255)*255;@clrizr-perceptive-luminance-pc1: round(1 - ( (0.299 * red(@clrizr-complementary-1)) + ( 0.587 * green(@clrizr-complementary-1)) + (0.114 * blue(@clrizr-complementary-1)))/255)*255;@clrizr-perceptive-luminance-pc2: round(1 - ( (0.299 * red(@clrizr-complementary-2)) + ( 0.587 * green(@clrizr-complementary-2)) + (0.114 * blue(@clrizr-complementary-2)))/255)*255; @clrizr-col-font-on-primary-2: rgb(@clrizr-perceptive-luminance-ps2, @clrizr-perceptive-luminance-ps2, @clrizr-perceptive-luminance-ps2);@clrizr-col-font-on-primary-1: rgb(@clrizr-perceptive-luminance-ps1, @clrizr-perceptive-luminance-ps1, @clrizr-perceptive-luminance-ps1);@clrizr-col-font-on-primary: rgb(@clrizr-perceptive-luminance-ps, @clrizr-perceptive-luminance-ps, @clrizr-perceptive-luminance-ps);@clrizr-col-font-on-complementary-1: rgb(@clrizr-perceptive-luminance-pc1, @clrizr-perceptive-luminance-pc1, @clrizr-perceptive-luminance-pc1);@clrizr-col-font-on-complementary-2: rgb(@clrizr-perceptive-luminance-pc2, @clrizr-perceptive-luminance-pc2, @clrizr-perceptive-luminance-pc2); Conclusion That’s it!  I have posted a project on clrizr.codePlex.com for this, and included a testing page for you to test out how it works.  Feel free to use it in your own project, and if you have any questions, comments or suggestions, please feel free to leave them here as a comment, or on the contact page!

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• CodePlex Daily Summary for Monday, August 18, 2014

  CodePlex Daily Summary for Monday, August 18, 2014Popular ReleasesMagick.NET: Magick.NET 7.0.0.0001: Magick.NET linked with ImageMagick 7-Beta.CMake Tools for Visual Studio: CMake Tools for Visual Studio 1.2: This release adds the following new features and bug fixes from CMake Tools for Visual Studio 1.1: Added support for CMake 3.0. Added support for word completion. Added IntelliSense support for the CMAKEHOSTSYSTEM_INFORMATION command. Fixed syntax highlighting for tokens beginning with escape sequences. Fixed issue uninstalling CMake Tools for Visual Studio after Visual Studio has been uninstalled.GW2 Personal Assistant Overlay: GW2 Personal Assistant Overlay 1.1: Overview1.1 is the second 'stable' release of the GW2 Personal Assistant Overlay. This version includes just a couple of very minor features and some minor bug fixes. For details regarding installation, setup, and general use, see Documentation. Note: If you were using a previous version, you will probably want to copy over the following user settings files: GW2PAO.DungeonSettings.xml GW2PAO.EventSettings.xml GW2PAO.WvWSettings.xml GW2PAO.ZoneCompletionSettings.xml New FeaturesAdded new "No...WallSwitch: WallSwitch 1.2.5: Version 1.2.5 Changes: Added support for sequential order in collage mode. Added option to display multiple images per switch in collage mode. Fixed bug where border width wasn't being loaded properly, and was reverting to default values. Fixed bug where sequential order was repeating images on multiple monitors. Decreased likelihood of random images being repeated.OpenCppCoverage: OpenCppCoverage 0.9.1: - Add Jenkins support. - Command line argument can be placed inside a config file. If you do not have Visual Studio C++ 2013 you need to download redistributable packages: http://www.microsoft.com/en-us/download/details.aspx?id=40784Easy Backup Windows Service: Release 2.0 with CU: Fix log error when "To" directory not exist in fyle system. Force run program as administrator by default. Add 'everyday' schedule element. Update solution to VS 2013.Easy Backup Application: Release 2.0 with CU: Fix log error when "To" directory not exist in fyle system. Fix app location initialization. Force run program as administrator by default. Update solution to VS 2013.TEBookConverter: 1.5: Added: Turkish and French translations Added: A few interface changes Removed: SkinDynamulet: Dynamulet v0.1: DynamoDB Transaction Server v0.1Console parallel nunit tests runner: ConsoleUnitTestsRunner 1.03: bugfixingFluentx: Fluentx v1.5.3: Added few more extension methods.fastJSON: v2.1.2: 2.1.2 - bug fix circular referencesJPush.NET: JPush Server SDK 1.2.1 (For JPush V3): Assembly: 1.2.1.24728 JPush REST API Version: v3 JPush Documentation Reference .NET framework: v4.0 or above. Sample: class: JPushClientV3 2014 Augest 15th.SEToolbox: SEToolbox 01.043.008 Release 1: Changed ship/station names to use new DisplayName instead of Beacon/Antenna. Fixed issue with updated SE binaries 01.043.018 using new Voxel Material definitions.Google .Net API: Drive.Sample: Google .NET Client API – Drive.SampleInstructions for the Google .NET Client API – Drive.Sample</h2> http://code.google.com/p/google-api-dotnet-client/source/browse/?repo=samples#hg%2FDrive.SampleBrowse Source, or main file http://code.google.com/p/google-api-dotnet-client/source/browse/Drive.Sample/Program.cs?repo=samplesProgram.cs <h3>1. Checkout Instructions</h3> <p><b>Prerequisites:</b> Install Visual Studio, and <a href="http://mercurial.selenic.com/">Mercurial</a>.</p> ...FineUI - jQuery / ExtJS based ASP.NET Controls: FineUI v4.1.1: -??Form??????????????（???-5929）。 -?TemplateField??ExpandOnDoubleClick、ExpandOnEnter、ExpandToSelectRow????（LZOM-5932）。 -BodyPadding???????，??“5”“5 10”，???????????“5px”“5px 10px”。 -??TriggerBox?EnableEdit=false????，??????????????（Jango_Jing-5450）。 -???????????DataKeyNames???????????（yygy-6002）。 -????????????????????????（Gnid-6018）。 -??PageManager???AutoSizePanelID????，??????????????????（yygy-6008）。 -?FState???????????????，????????????????（????-5925）。 -??????OnClientClick???return?????????（FineU...DNN CMS Platform: 07.03.02: Major Highlights Fixed backwards compatibility issue with 3rd party control panels Fixed issue in the drag and drop functionality of the File Uploader in IE 11 and Safari Fixed issue where users were able to create pages with the same name Fixed issue that affected older versions of DNN that do not include the maxAllowedContentLength during upgrade Fixed issue that stopped some skins from being upgraded to newer versions Fixed issue that randomly showed an unexpected error during us...WordMat: WordMat for Mac: WordMat for Mac has a few limitations compared to the Windows version - Graph is not supported (Gnuplot, GeoGebra and Excel works) - Units are not supported yet (Coming up) The Mac version is yet as tested as the windows version.MFCMAPI: August 2014 Release: Build: 15.0.0.1042 Full release notes at SGriffin's blog. If you just want to run the MFCMAPI or MrMAPI, get the executables. If you want to debug them, get the symbol files and the source. The 64 bit builds will only work on a machine with Outlook 2010/2013 64 bit installed. All other machines should use the 32 bit builds, regardless of the operating system. Facebook BadgeEWSEditor: EwsEditor 1.10 Release: • Export and import of items as a full fidelity steam works - without proxy classes! - I used raw EWS POSTs. • Turned off word wrap for EWS request field in EWS POST windows. • Several windows with scrolling texts boxes were limiting content to 32k - I removed this restriction. • Split server timezone info off to separate menu item from the timezone info windows so that the timezone info window could be used without logging into a mailbox. • Lots of updates to the TimeZone window. • UserAgen...New Projectsballmon: ballmonExchange Database Recovery With and Without Log Files is Possible: This segments giving an overview of Exchange Server transaction log files. It describes process how users can recover their database with & without log filesFabs.Net: Ego tatmini ve gelisme amaçli yaptigim bir projedir.JacoChat: JacoChat is a simple chatting interface that uses my personal webserver as a "wall" for people to chat on.ManagedWin32: ManagedWin32 is a library that exposes the Win32 API to .NET applications.Open XML Extensions: The project provides additions to the Open XML SDK and related projects (e.g., PowerTools for Open XML), starting with MemoryStreams for Open XML Documents.orntic: Project for insurace companyTBOX: The Treasure Box Library: TBOX is a mutli-platform c library for unix, windows, mac, ios, android, etc. It includes asio, stream, container, algorithm, xml and other library modules.WeatherTS: Typescript weather application.?????@/????: ??????????????:????,????,????,???????,????????,??????:????????,?????! ?????????: ????????????????????,????????:??、??、???,?????????????????????! ????-??: ??????????????，????，???????????????。

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• The code works but when using printf it gives me a weird answer. Help please [closed]

  - by user71458

  //Programmer-William Chen //Seventh Period Computer Science II //Problem Statement - First get the elapsed times and the program will find the //split times for the user to see. // //Algorithm- First the programmer makes the prototype and calls them in the //main function. The programmer then asks the user to input lap time data. //Secondly, you convert the splits into seconds and subtract them so you can //find the splits. Then the average is all the lap time's in seconds. Finally, //the programmer printf all the results for the user to see. #include <iostream> #include <stdlib.h> #include <math.h> #include <conio.h> #include <stdio.h> using namespace std; void thisgetsElapsedTimes( int &m1, int &m2, int &m3, int &m4, int &m5, int &s1, int &s2, int &s3, int &s4, int &s5); //this is prototype void thisconvertstoseconds ( int &m1, int &m2, int &m3, int &m4, int &m5, int &s1, int &s2, int &s3, int &s4, int &s5, int &split1, int &split2, int &split3, int &split4, int &split5);//this too void thisfindsSplits(int &m1, int &m2, int &m3, int &m4, int &m5, int &split1, int &split2, int &split3, int &split4, int &split5, int &split6, int &split7, int &split8, int &split9, int &split10);// this is part of prototype void thisisthesecondconversation (int &split1M, int &split2M, int &split3M, int &split4M, int &split5M, int &split1S,int &split2S, int &split3S, int &split4S, int &split5S, int &split1, int &split2, int &split3, int &split4, int &split5);//this gets a value void thisfindstheaverage(double &average, int &split1, int &split2, int &split3, int &split4, int &split5);//and this void thisprintsstuff( int &split1M, int &split2M, int &split3M, int &split4M, int &split5M, int &split1S, int &split2S, int &split3S, int &split4S, int &split5S, double &average); //this prints int main(int argc, char *argv[]) { int m1, m2, m3, m4, m5, s1, s2, s3, s4, s5, split1, split2, split3, split4, split5, split1M, split2M, split3M, split4M, split5M, split1S, split2S, split3S, split4S, split5S; int split6, split7, split8, split9, split10; double average; char thistakescolon; thisgetsElapsedTimes ( m1, m2, m3, m4, m5, s1, s2, s3, s4, s5); thisconvertstoseconds ( m1, m2, m3, m4, m5, s1, s2, s3, s4, s5, split1, split2, split3, split4, split5); thisfindsSplits ( m1, m2, m3, m4, m5, split1, split2, split3, split4, split5, split6, split7, split8, split9, split10); thisisthesecondconversation ( split1M, split2M, split3M, split4M, split5M, split1S, split2S, split3S, split4S, split5S, split1, split2, split3, split4, split5); thisfindstheaverage ( average, split1, split2, split3, split4, split5); thisprintsstuff ( split1M, split2M, split3M, split4M, split5M, split1S, split2S, split3S, split4S, split5S, average); // these are calling statements and they call from the main function to the other functions. system("PAUSE"); return 0; } void thisgetsElapsedTimes(int &m1, int &m2, int &m3, int &m4, int &m5, int &s1, int &s2, int &s3, int &s4, int &s5) { char thistakescolon; cout << "Enter the elapsed time:" << endl; cout << " Kilometer 1 "; cin m1 thistakescolon s1; cout << " Kilometer 2 "; cin m2 thistakescolon s2; cout << " Kilometer 3 " ; cin m3 thistakescolon s3; cout << " Kilometer 4 "; cin m4 thistakescolon s4; cout << " Kilometer 5 "; cin m5 thistakescolon s5; // this gets the data required to get the results needed for the user to see // . } void thisconvertstoseconds (int &m1, int &m2, int &m3, int &m4, int &m5, int &s1, int &s2, int &s3, int &s4, int &s5, int &split1, int &split2, int &split3, int &split4, int &split5) { split1 = (m1 * 60) + s1;//this converts for minutes to seconds for m1 split2 = (m2 * 60) + s2;//this converts for minutes to seconds for m2 split3 = (m3 * 60) + s3;//this converts for minutes to seconds for m3 split4 = (m4 * 60) + s4;//this converts for minutes to seconds for m4 split5 = (m5 * 60) + s5;//this converts for minutes to seconds for m5 } void thisfindsSplits (int &m1, int &m2, int &m3, int &m4, int &m5,int &split1, int &split2, int &split3, int &split4, int &split5, int &split6, int &split7, int &split8, int &split9, int &split10)//this is function heading { split6 = split1; //this is split for the first lap. split7 = split2 - split1;//this is split for the second lap. split8 = split3 - split2;//this is split for the third lap. split9 = split4 - split3;//this is split for the fourth lap. split10 = split5 - split4;//this is split for the fifth lap. } void thisfindstheaverage(double &average, int &split1, int &split2, int &split3, int &split4, int &split5) { average = (split1 + split2 + split3 + split4 + split5)/5; // this finds the average from all the splits in seconds } void thisisthesecondconversation (int &split1M, int &split2M, int &split3M, int &split4M, int &split5M, int &split1S,int &split2S, int &split3S, int &split4S, int &split5S, int &split1, int &split2, int &split3, int &split4, int &split5) { split1M = split1 * 60; //this finds the split times split1S = split1M - split1 * 60; //then this finds split2M = split2 * 60; //and all of this split2S = split2M - split2 * 60; //does basically split3M = split3 * 60; //the same thing split3S = split3M - split3 * 60; //all of it split4M = split4 * 60; //it's also a split4S = split4M - split4 * 60; //function split5M = split5 * 60; //and it finds the splits split5S = split5M - split5 * 60; //for each lap. } void thisprintsstuff (int &split1M, int &split2M, int &split3M, int &split4M, int &split5M, int &split1S, int &split2S, int &split3S, int &split4S, int &split5S, double &average)// this is function heading { printf("\n kilometer 1 %d" , ":02%d",'split1M','split1S'); printf("\n kilometer 2 %d" , ":02%d",'split2M','split2S'); printf("\n kilometer 3 %d" , ":02%d",'split3M','split3S'); printf("\n kilometer 4 %d" , ":02%d",'split4M','split4S'); printf("\n kilometer 5 %d" , ":02%d",'split5M','split5S'); printf("\n your average pace is ",'average',"per kilometer \n", "William Chen\n"); // this printf so the programmer // can allow the user to see // the results from the data gathered. }

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• Data Source Connection Pool Sizing

  - by Steve Felts

  Normal 0 false false false EN-US X-NONE X-NONE MicrosoftInternetExplorer4 /* 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:0in; mso-para-margin-bottom:.0001pt; mso-pagination:widow-orphan; font-size:10.0pt; font-family:"Times New Roman","serif";} One of the most time-consuming procedures of a database application is establishing a connection. The connection pooling of the data source can be used to minimize this overhead.  That argues for using the data source instead of accessing the database driver directly. Configuring the size of the pool in the data source is somewhere between an art and science – this article will try to move it closer to science.  From the beginning, WLS data source has had an initial capacity and a maximum capacity configuration values.  When the system starts up and when it shrinks, initial capacity is used.  The pool can grow to maximum capacity.  Customers found that they might want to set the initial capacity to 0 (more on that later) but didn’t want the pool to shrink to 0.  In WLS 10.3.6, we added minimum capacity to specify the lower limit to which a pool will shrink.  If minimum capacity is not set, it defaults to the initial capacity for upward compatibility.   We also did some work on the shrinking in release 10.3.4 to reduce thrashing; the algorithm that used to shrink to the maximum of the currently used connections or the initial capacity (basically the unused connections were all released) was changed to shrink by half of the unused connections. The simple approach to sizing the pool is to set the initial/minimum capacity to the maximum capacity.  Doing this creates all connections at startup, avoiding creating connections on demand and the pool is stable.  However, there are a number of reasons not to take this simple approach. When WLS is booted, the deployment of the data source includes synchronously creating the connections.  The more connections that are configured in initial capacity, the longer the boot time for WLS (there have been several projects for parallel boot in WLS but none that are available).  Related to creating a lot of connections at boot time is the problem of logon storms (the database gets too much work at one time).   WLS has a solution for that by setting the login delay seconds on the pool but that also increases the boot time. There are a number of cases where it is desirable to set the initial capacity to 0.  By doing that, the overhead of creating connections is deferred out of the boot and the database doesn’t need to be available.  An application may not want WLS to automatically connect to the database until it is actually needed, such as for some code/warm failover configurations. There are a number of cases where minimum capacity should be less than maximum capacity.  Connections are generally expensive to keep around.  They cause state to be kept on both the client and the server, and the state on the backend may be heavy (for example, a process).  Depending on the vendor, connection usage may cost money.  If work load is not constant, then database connections can be freed up by shrinking the pool when connections are not in use.  When using Active GridLink, connections can be created as needed according to runtime load balancing (RLB) percentages instead of by connection load balancing (CLB) during data source deployment. Shrinking is an effective technique for clearing the pool when connections are not in use.  In addition to the obvious reason that there times where the workload is lighter,  there are some configurations where the database and/or firewall conspire to make long-unused or too-old connections no longer viable.  There are also some data source features where the connection has state and cannot be used again unless the state matches the request.  Examples of this are identity based pooling where the connection has a particular owner and XA affinity where the connection is associated with a particular RAC node.  At this point, WLS does not re-purpose (discard/replace) connections and shrinking is a way to get rid of the unused existing connection and get a new one with the correct state when needed. So far, the discussion has focused on the relationship of initial, minimum, and maximum capacity.  Computing the maximum size requires some knowledge about the application and the current number of simultaneously active users, web sessions, batch programs, or whatever access patterns are common.  The applications should be written to only reserve and close connections as needed but multiple statements, if needed, should be done in one reservation (don’t get/close more often than necessary).  This means that the size of the pool is likely to be significantly smaller then the number of users.   If possible, you can pick a size and see how it performs under simulated or real load.  There is a high-water mark statistic (ActiveConnectionsHighCount) that tracks the maximum connections concurrently used.  In general, you want the size to be big enough so that you never run out of connections but no bigger.   It will need to deal with spikes in usage, which is where shrinking after the spike is important.  Of course, the database capacity also has a big influence on the decision since it’s important not to overload the database machine.  Planning also needs to happen if you are running in a Multi-Data Source or Active GridLink configuration and expect that the remaining nodes will take over the connections when one of the nodes in the cluster goes down.  For XA affinity, additional headroom is also recommended.  In summary, setting initial and maximum capacity to be the same may be simple but there are many other factors that may be important in making the decision about sizing.

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• A* PathFinding Poor Performance

  - by RedShft

  After debugging for a few hours, the algorithm seems to be working. Right now to check if it works i'm checking the end node position to the currentNode position when the while loop quits. So far the values look correct. The problem is, the farther I get from the NPC, who is current stationary, the worse the performance gets. It gets to a point where the game is unplayable less than 10 fps. My current PathGraph is 2500 nodes, which I believe is pretty small, right? Any ideas on how to improve performance? struct Node { bool walkable; //Whether this node is blocked or open vect2 position; //The tile's position on the map in pixels int xIndex, yIndex; //The index values of the tile in the array Node*[4] connections; //An array of pointers to nodes this current node connects to Node* parent; int gScore; int hScore; int fScore; } class AStar { private: SList!Node openList; SList!Node closedList; //Node*[4] connections; //The connections of the current node; Node currentNode; //The current node being processed Node[] Path; //The path found; const int connectionCost = 10; Node start, end; ////////////////////////////////////////////////////////// void AddToList(ref SList!Node list, ref Node node ) { list.insert( node ); } void RemoveFrom(ref SList!Node list, ref Node node ) { foreach( elem; list ) { if( node.xIndex == elem.xIndex && node.yIndex == elem.yIndex ) { auto a = find( list[] , elem ); list.linearRemove( take(a, 1 ) ); } } } bool IsInList( SList!Node list, ref Node node ) { foreach( elem; list ) { if( node.xIndex == elem.xIndex && node.yIndex == elem.yIndex ) return true; } return false; } void ClearList( SList!Node list ) { list.clear; } void SetParentNode( ref Node parent, ref Node child ) { child.parent = &parent; } void SetStartAndEndNode( vect2 vStart, vect2 vEnd, Node[] PathGraph ) { int startXIndex, startYIndex; int endXIndex, endYIndex; startXIndex = cast(int)( vStart.x / 32 ); startYIndex = cast(int)( vStart.y / 32 ); endXIndex = cast(int)( vEnd.x / 32 ); endYIndex = cast(int)( vEnd.y / 32 ); foreach( node; PathGraph ) { if( node.xIndex == startXIndex && node.yIndex == startYIndex ) { start = node; } if( node.xIndex == endXIndex && node.yIndex == endYIndex ) { end = node; } } } void SetStartScores( ref Node start ) { start.gScore = 0; start.hScore = CalculateHScore( start, end ); start.fScore = CalculateFScore( start ); } Node GetLowestFScore() { Node lowest; lowest.fScore = 10000; foreach( elem; openList ) { if( elem.fScore < lowest.fScore ) lowest = elem; } return lowest; } //This function current sets the program into an infinite loop //I still need to debug to figure out why the parent nodes aren't correct void GeneratePath() { while( currentNode.position != start.position ) { Path ~= currentNode; currentNode = *currentNode.parent; } } void ReversePath() { Node[] temp; for(int i = Path.length - 1; i >= 0; i-- ) { temp ~= Path[i]; } Path = temp.dup; } public: //@FIXME It seems to find the path, but now performance is terrible void FindPath( vect2 vStart, vect2 vEnd, Node[] PathGraph ) { openList.clear; closedList.clear; SetStartAndEndNode( vStart, vEnd, PathGraph ); SetStartScores( start ); AddToList( openList, start ); while( currentNode.position != end.position ) { currentNode = GetLowestFScore(); if( currentNode.position == end.position ) break; else { RemoveFrom( openList, currentNode ); AddToList( closedList, currentNode ); for( int i = 0; i < currentNode.connections.length; i++ ) { if( currentNode.connections[i] is null ) continue; else { if( IsInList( closedList, *currentNode.connections[i] ) && currentNode.gScore < currentNode.connections[i].gScore ) { currentNode.connections[i].gScore = currentNode.gScore + connectionCost; currentNode.connections[i].hScore = abs( currentNode.connections[i].xIndex - end.xIndex ) + abs( currentNode.connections[i].yIndex - end.yIndex ); currentNode.connections[i].fScore = currentNode.connections[i].gScore + currentNode.connections[i].hScore; currentNode.connections[i].parent = &currentNode; } else if( IsInList( openList, *currentNode.connections[i] ) && currentNode.gScore < currentNode.connections[i].gScore ) { currentNode.connections[i].gScore = currentNode.gScore + connectionCost; currentNode.connections[i].hScore = abs( currentNode.connections[i].xIndex - end.xIndex ) + abs( currentNode.connections[i].yIndex - end.yIndex ); currentNode.connections[i].fScore = currentNode.connections[i].gScore + currentNode.connections[i].hScore; currentNode.connections[i].parent = &currentNode; } else { currentNode.connections[i].gScore = currentNode.gScore + connectionCost; currentNode.connections[i].hScore = abs( currentNode.connections[i].xIndex - end.xIndex ) + abs( currentNode.connections[i].yIndex - end.yIndex ); currentNode.connections[i].fScore = currentNode.connections[i].gScore + currentNode.connections[i].hScore; currentNode.connections[i].parent = &currentNode; AddToList( openList, *currentNode.connections[i] ); } } } } } writeln( "Current Node Position: ", currentNode.position ); writeln( "End Node Position: ", end.position ); if( currentNode.position == end.position ) { writeln( "Current Node Parent: ", currentNode.parent ); //GeneratePath(); //ReversePath(); } } Node[] GetPath() { return Path; } } This is my first attempt at A* so any help would be greatly appreciated.

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• setting up bind to work with nsupdate (SERVFAIL)

  - by funny_ha_ha

  I'm trying to update my DNS-Server dynamically using nsupdate. Prerequisite I'm using Debian 6 on my DNS-Server and Debian 4 on my client. I created a public/private key pair using: dnssec-keygen -C -a HMAC-MD5 -b 512 -n USER sub.example.com. I then edited my named.conf.local to contain my public key and the new zone i wish to update. It now looks like this (note: I also tried allow-update { any; }; without success): zone "example.com" { type master; file "/etc/bind/primary/example.com"; notify yes; allow-update { none; }; allow-query { any; }; }; zone "sub.example.com" { type master; file "/etc/bind/primary/sub.example.com"; notify yes; allow-update { key "sub.example.com."; }; allow-query { any; }; }; key sub.example.com. { algorithm HMAC-MD5; secret "xxxx xxxx"; }; Next, I copied the private key file (key.private) to another server I want to update the zone from. I also created a textfile (update) on this server which contained the update information (note: I tried toying around with this stuff too. no success): server example.com zone sub.example.com update add sub.example.com. 86400 A 10.10.10.1 show send Now I'm trying to update the zone using: nsupdate -k key.private -v update The Problem Said command gives me the following output: Outgoing update query: ;; ->>HEADER<<- opcode: UPDATE, status: NOERROR, id: 0 ;; flags: ; ZONE: 0, PREREQ: 0, UPDATE: 0, ADDITIONAL: 0 ;; ZONE SECTION: ;sub.example.com. IN SOA ;; UPDATE SECTION: sub.example.com. 86400 IN A 10.10.10.1 update failed: SERVFAIL named debug Level 3 gives me the following information when I issue the nsupdate command on the remote server (note: I obfuscated the client IP): 06-Aug-2012 14:51:33.977 client X.X.X.X#33182: new TCP connection 06-Aug-2012 14:51:33.977 client X.X.X.X#33182: replace 06-Aug-2012 14:51:33.978 clientmgr @0x2ada3c7ee760: createclients 06-Aug-2012 14:51:33.978 clientmgr @0x2ada3c7ee760: recycle 06-Aug-2012 14:51:33.978 client @0x2ada475f1120: accept 06-Aug-2012 14:51:33.978 client X.X.X.X#33182: read 06-Aug-2012 14:51:33.978 client X.X.X.X#33182: TCP request 06-Aug-2012 14:51:33.978 client X.X.X.X#33182: request has valid signature 06-Aug-2012 14:51:33.978 client X.X.X.X#33182: recursion not available 06-Aug-2012 14:51:33.978 client X.X.X.X#33182: update 06-Aug-2012 14:51:33.978 client X.X.X.X#33182: send 06-Aug-2012 14:51:33.978 client X.X.X.X#33182: sendto 06-Aug-2012 14:51:33.979 client X.X.X.X#33182: senddone 06-Aug-2012 14:51:33.979 client X.X.X.X#33182: next 06-Aug-2012 14:51:33.979 client X.X.X.X#33182: endrequest 06-Aug-2012 14:51:33.979 client X.X.X.X#33182: read 06-Aug-2012 14:51:33.986 client X.X.X.X#33182: next 06-Aug-2012 14:51:33.986 client X.X.X.X#33182: request failed: end of file 06-Aug-2012 14:51:33.986 client X.X.X.X#33182: endrequest 06-Aug-2012 14:51:33.986 client X.X.X.X#33182: closetcp But it doesn't do anything. The zone isn't updated, nor does my nsupdate change anything. I'm not sure if the file /etc/bind/primary/sub.example.com should exist prior to the first update or not. I tried it without the file, with an empty file and with a pre-configured zone file. Without success. The sparse information I found on the net pointed me towards file and folder permissions regarding the bind working directory, so I changed the permissions of both /etc/bind and /var/cache/bind (which is the home dir of my "bind" user). I'm not a 100% sure if the permissions are correct.. but it looks good to me: ls -lah /var/cache/bind/ total 224K drwxrwxr-x 2 bind bind 4.0K Aug 6 03:13 . drwxr-xr-x 12 root root 4.0K Jul 21 11:27 .. -rw-r--r-- 1 bind bind 211K Aug 6 03:21 named.run ls -lah /etc/bind/ total 72K drwxr-sr-x 3 bind bind 4.0K Aug 6 14:41 . drwxr-xr-x 87 root root 4.0K Jul 30 01:24 .. -rw------- 1 bind bind 125 Aug 6 02:54 key.public -rw------- 1 bind bind 156 Aug 6 02:54 key.private -rw-r--r-- 1 bind bind 2.5K Aug 6 03:07 bind.keys -rw-r--r-- 1 bind bind 237 Aug 6 03:07 db.0 -rw-r--r-- 1 bind bind 271 Aug 6 03:07 db.127 -rw-r--r-- 1 bind bind 237 Aug 6 03:07 db.255 -rw-r--r-- 1 bind bind 353 Aug 6 03:07 db.empty -rw-r--r-- 1 bind bind 270 Aug 6 03:07 db.local -rw-r--r-- 1 bind bind 3.0K Aug 6 03:07 db.root -rw-r--r-- 1 bind bind 493 Aug 6 03:32 named.conf -rw-r--r-- 1 bind bind 490 Aug 6 03:07 named.conf.default-zones -rw-r--r-- 1 bind bind 1.2K Aug 6 14:18 named.conf.local -rw-r--r-- 1 bind bind 666 Jul 29 22:51 named.conf.options drwxr-sr-x 2 bind bind 4.0K Aug 6 03:57 primary/ -rw-r----- 1 root bind 77 Mar 19 02:57 rndc.key -rw-r--r-- 1 bind bind 1.3K Aug 6 03:07 zones.rfc1918 ls -lah /etc/bind/primary/ total 20K drwxr-sr-x 2 bind bind 4.0K Aug 6 03:57 . drwxr-sr-x 3 bind bind 4.0K Aug 6 14:41 .. -rw-r--r-- 1 bind bind 356 Jul 30 00:45 example.com

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• Problem with RAID5 (mdadm) - disk detached

  - by poscaman

  Having these lines in /var/log/syslog Apr 18 16:53:05 Server kernel: [4487878.816036] ata4: EH in SWNCQ mode,QC:qc_active 0x1 sactive 0x1 Apr 18 16:53:05 Server kernel: [4487878.816058] ata4: SWNCQ:qc_active 0x1 defer_bits 0x0 last_issue_tag 0x0 Apr 18 16:53:05 Server kernel: [4487878.816059] dhfis 0x1 dmafis 0x1 sdbfis 0x0 Apr 18 16:53:05 Server kernel: [4487878.816093] ata4: ATA_REG 0x40 ERR_REG 0x0 Apr 18 16:53:05 Server kernel: [4487878.816108] ata4: tag : dhfis dmafis sdbfis sacitve Apr 18 16:53:05 Server kernel: [4487878.816125] ata4: tag 0x0: 1 1 0 1 Apr 18 16:53:05 Server kernel: [4487878.816150] ata4.00: exception Emask 0x0 SAct 0x1 SErr 0x0 action 0x6 frozen Apr 18 16:53:05 Server kernel: [4487878.816178] ata4.00: failed command: WRITE FPDMA QUEUED Apr 18 16:53:05 Server kernel: [4487878.816199] ata4.00: cmd 61/08:00:00:88:e0/00:00:e8:00:00/40 tag 0 ncq 4096 out Apr 18 16:53:05 Server kernel: [4487878.816200] res 40/00:00:01:4f:c2/00:00:00:00:00/00 Emask 0x4 (timeout) Apr 18 16:53:05 Server kernel: [4487878.816253] ata4.00: status: { DRDY } Apr 18 16:53:05 Server kernel: [4487878.816272] ata4: hard resetting link Apr 18 16:53:05 Server kernel: [4487878.816274] ata4: nv: skipping hardreset on occupied port Apr 18 16:53:06 Server kernel: [4487879.676029] ata4: SATA link up 3.0 Gbps (SStatus 123 SControl 300) Apr 18 16:53:07 Server kernel: [4487880.416749] ata4.00: n_sectors mismatch 3907029168 != 268435455 Apr 18 16:53:07 Server kernel: [4487880.416752] ata4.00: revalidation failed (errno=-19) Apr 18 16:53:07 Server kernel: [4487880.416773] ata4.00: limiting speed to UDMA/133:PIO2 Apr 18 16:53:11 Server kernel: [4487884.676024] ata4: hard resetting link Apr 18 16:53:11 Server kernel: [4487884.676027] ata4: nv: skipping hardreset on occupied port Apr 18 16:53:12 Server kernel: [4487885.144032] ata4: SATA link up 3.0 Gbps (SStatus 123 SControl 300) Apr 18 16:53:12 Server kernel: [4487885.240185] ata4.00: failed to IDENTIFY (INIT_DEV_PARAMS failed, err_mask=0x80) Apr 18 16:53:12 Server kernel: [4487885.240190] ata4.00: revalidation failed (errno=-5) Apr 18 16:53:12 Server kernel: [4487885.240210] ata4.00: disabled Apr 18 16:53:17 Server kernel: [4487890.144023] ata4: hard resetting link Apr 18 16:53:17 Server kernel: [4487891.024033] ata4: SATA link up 3.0 Gbps (SStatus 123 SControl 300) Apr 18 16:53:17 Server kernel: [4487891.033357] ata4.00: ATA-8: WDC WD20EARS-00S8B1, 80.00A80, max UDMA/133 Apr 18 16:53:17 Server kernel: [4487891.033360] ata4.00: 3907029168 sectors, multi 1: LBA48 NCQ (depth 31/32) Apr 18 16:53:17 Server kernel: [4487891.048347] ata4.00: configured for UDMA/133 Apr 18 16:53:17 Server kernel: [4487891.048361] sd 3:0:0:0: [sdc] Result: hostbyte=DID_OK driverbyte=DRIVER_SENSE Apr 18 16:53:17 Server kernel: [4487891.048365] sd 3:0:0:0: [sdc] Sense Key : Aborted Command [current] [descriptor] Apr 18 16:53:17 Server kernel: [4487891.048369] Descriptor sense data with sense descriptors (in hex): Apr 18 16:53:17 Server kernel: [4487891.048371] 72 0b 00 00 00 00 00 0c 00 0a 80 00 00 00 00 00 Apr 18 16:53:17 Server kernel: [4487891.048378] 00 00 00 00 Apr 18 16:53:17 Server kernel: [4487891.048382] sd 3:0:0:0: [sdc] Add. Sense: No additional sense information Apr 18 16:53:17 Server kernel: [4487891.048385] sd 3:0:0:0: [sdc] CDB: Write(10): 2a 00 e8 e0 88 00 00 00 08 00 Apr 18 16:53:17 Server kernel: [4487891.048393] end_request: I/O error, dev sdc, sector 3907028992 Apr 18 16:53:17 Server kernel: [4487891.048420] sd 3:0:0:0: rejecting I/O to offline device Apr 18 16:53:17 Server kernel: [4487891.048440] sd 3:0:0:0: rejecting I/O to offline device Apr 18 16:53:17 Server kernel: [4487891.048458] end_request: I/O error, dev sdc, sector 3907028992 Apr 18 16:53:17 Server kernel: [4487891.048477] md: super_written gets error=-5, uptodate=0 Apr 18 16:53:17 Server kernel: [4487891.048482] raid5: Disk failure on sdc, disabling device. Apr 18 16:53:17 Server kernel: [4487891.048483] raid5: Operation continuing on 3 devices. Apr 18 16:53:17 Server kernel: [4487891.048525] ata4: EH complete Apr 18 16:53:17 Server kernel: [4487891.048554] sd 3:0:0:0: rejecting I/O to offline device Apr 18 16:53:17 Server kernel: [4487891.048576] sd 3:0:0:0: rejecting I/O to offline device Apr 18 16:53:17 Server kernel: [4487891.048596] sd 3:0:0:0: rejecting I/O to offline device Apr 18 16:53:17 Server kernel: [4487891.048615] sd 3:0:0:0: [sdc] READ CAPACITY(16) failed Apr 18 16:53:17 Server kernel: [4487891.048617] sd 3:0:0:0: [sdc] Result: hostbyte=DID_NO_CONNECT driverbyte=DRIVER_OK Apr 18 16:53:17 Server kernel: [4487891.048620] sd 3:0:0:0: [sdc] Sense not available. Apr 18 16:53:17 Server kernel: [4487891.048624] sd 3:0:0:0: rejecting I/O to offline device Apr 18 16:53:17 Server kernel: [4487891.048643] sd 3:0:0:0: rejecting I/O to offline device Apr 18 16:53:17 Server kernel: [4487891.048663] sd 3:0:0:0: rejecting I/O to offline device Apr 18 16:53:17 Server kernel: [4487891.048681] sd 3:0:0:0: [sdc] READ CAPACITY failed Apr 18 16:53:17 Server kernel: [4487891.048683] sd 3:0:0:0: [sdc] Result: hostbyte=DID_NO_CONNECT driverbyte=DRIVER_OK Apr 18 16:53:17 Server kernel: [4487891.048685] sd 3:0:0:0: [sdc] Sense not available. Apr 18 16:53:17 Server kernel: [4487891.048689] sd 3:0:0:0: rejecting I/O to offline device Apr 18 16:53:17 Server kernel: [4487891.048709] sd 3:0:0:0: rejecting I/O to offline device Apr 18 16:53:17 Server kernel: [4487891.048800] sd 3:0:0:0: rejecting I/O to offline device Apr 18 16:53:17 Server kernel: [4487891.048860] sd 3:0:0:0: rejecting I/O to offline device Apr 18 16:53:17 Server kernel: [4487891.049028] sd 3:0:0:0: [sdc] Asking for cache data failed Apr 18 16:53:17 Server kernel: [4487891.049048] sd 3:0:0:0: [sdc] Assuming drive cache: write through Apr 18 16:53:17 Server kernel: [4487891.049071] sdc: detected capacity change from 2000398934016 to 0 Apr 18 16:53:17 Server kernel: [4487891.049080] ata4.00: detaching (SCSI 3:0:0:0) Apr 18 16:53:18 Server kernel: [4487891.061149] sd 3:0:0:0: [sdc] Stopping disk Apr 18 16:53:18 Server kernel: [4487891.485492] RAID5 conf printout: Apr 18 16:53:18 Server kernel: [4487891.485496] --- rd:4 wd:3 Apr 18 16:53:18 Server kernel: [4487891.485500] disk 0, o:1, dev:sdb Apr 18 16:53:18 Server kernel: [4487891.485502] disk 1, o:0, dev:sdc Apr 18 16:53:18 Server kernel: [4487891.485504] disk 2, o:1, dev:sdd Apr 18 16:53:18 Server kernel: [4487891.485506] disk 3, o:1, dev:sde Apr 18 16:53:18 Server kernel: [4487891.497014] RAID5 conf printout: Apr 18 16:53:18 Server kernel: [4487891.497016] --- rd:4 wd:3 Apr 18 16:53:18 Server kernel: [4487891.497018] disk 0, o:1, dev:sdb Apr 18 16:53:18 Server kernel: [4487891.497019] disk 2, o:1, dev:sdd Apr 18 16:53:18 Server kernel: [4487891.497021] disk 3, o:1, dev:sde Apr 18 16:53:18 Server kernel: [4487891.838719] scsi 3:0:0:0: Direct-Access ATA WDC WD20EARS-00S 80.0 PQ: 0 ANSI: 5 Apr 18 16:53:18 Server kernel: [4487891.838886] sd 3:0:0:0: Attached scsi generic sg3 type 0 Apr 18 16:53:18 Server kernel: [4487891.838911] sd 3:0:0:0: [sdf] 3907029168 512-byte logical blocks: (2.00 TB/1.81 TiB) Apr 18 16:53:18 Server kernel: [4487891.838964] sd 3:0:0:0: [sdf] Write Protect is off Apr 18 16:53:18 Server kernel: [4487891.838967] sd 3:0:0:0: [sdf] Mode Sense: 00 3a 00 00 Apr 18 16:53:18 Server kernel: [4487891.838988] sd 3:0:0:0: [sdf] Write cache: enabled, read cache: enabled, doesn't support DPO or FUA Apr 18 16:53:20 Server kernel: [4487891.839147] sdf: unknown partition table Apr 18 16:53:20 Server kernel: [4487893.130026] sd 3:0:0:0: [sdf] Attached SCSI disk Right now, i'm unable to do anything on /dev/sdc. Is there any way to try to re-attach it? I don't want to power-down the server unless absolutely necessary System: Debian Stable 2.6.32-5-amd64 mdadm version 3.1.4-1+8efb9d1 cat /proc/mdstat Personalities : [raid6] [raid5] [raid4] md0 : active raid5 sdb[0] sdc[4](F) sde[3] sdd[2] 5860543488 blocks level 5, 64k chunk, algorithm 2 [4/3] [U_UU] unused devices: <none> mdadm --examine --scan ARRAY /dev/md0 UUID=1a7744b5:912ec7af:f82a9565:e3b453b4

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• How do I sign my certificate using the root certificate

  - by Asif Alam

  I am using certificate based authentication between my server and client. I have generated Root Certificate. My client at the time of installation will generate a new Certificate and use the Root Certificate to sign it. I need to use Windows API. Cannot use any windows tools like makecert. Till now I have been able to Install the Root certificate in store. Below code X509Certificate2 ^ certificate = gcnew X509Certificate2("C:\\rootcert.pfx","test123"); X509Store ^ store = gcnew X509Store( "teststore",StoreLocation::CurrentUser ); store->Open( OpenFlags::ReadWrite ); store->Add( certificate ); store->Close(); Then open the installed root certificate to get the context GetRootCertKeyInfo(){ HCERTSTORE hCertStore; PCCERT_CONTEXT pSignerCertContext=NULL; DWORD dwSize = NULL; CRYPT_KEY_PROV_INFO* pKeyInfo = NULL; DWORD dwKeySpec; if ( !( hCertStore = CertOpenStore(CERT_STORE_PROV_SYSTEM, 0, NULL, CERT_SYSTEM_STORE_CURRENT_USER,L"teststore"))) { _tprintf(_T("Error 0x%x\n"), GetLastError()); } pSignerCertContext = CertFindCertificateInStore(hCertStore,MY_ENCODING_TYPE,0,CERT_FIND_ANY,NULL,NULL); if(NULL == pSignerCertContext) { _tprintf(_T("Error 0x%x\n"), GetLastError()); } if(!(CertGetCertificateContextProperty( pSignerCertContext, CERT_KEY_PROV_INFO_PROP_ID, NULL, &dwSize))) { _tprintf(_T("Error 0x%x\n"), GetLastError()); } if(pKeyInfo) free(pKeyInfo); if(!(pKeyInfo = (CRYPT_KEY_PROV_INFO*)malloc(dwSize))) { _tprintf(_T("Error 0x%x\n"), GetLastError()); } if(!(CertGetCertificateContextProperty( pSignerCertContext, CERT_KEY_PROV_INFO_PROP_ID, pKeyInfo, &dwSize))) { _tprintf(_T("Error 0x%x\n"), GetLastError()); } return pKeyInfo; } Then finally created the certificate and signed with the pKeyInfo // Acquire key container if (!CryptAcquireContext(&hCryptProv, _T("trykeycon"), NULL, PROV_RSA_FULL, CRYPT_MACHINE_KEYSET)) { _tprintf(_T("Error 0x%x\n"), GetLastError()); // Try to create a new key container _tprintf(_T("CryptAcquireContext... ")); if (!CryptAcquireContext(&hCryptProv, _T("trykeycon"), NULL, PROV_RSA_FULL, CRYPT_NEWKEYSET | CRYPT_MACHINE_KEYSET)) { _tprintf(_T("Error 0x%x\n"), GetLastError()); return 0; } else { _tprintf(_T("Success\n")); } } else { _tprintf(_T("Success\n")); } // Generate new key pair _tprintf(_T("CryptGenKey... ")); if (!CryptGenKey(hCryptProv, AT_SIGNATURE, 0x08000000 /*RSA-2048-BIT_KEY*/, &hKey)) { _tprintf(_T("Error 0x%x\n"), GetLastError()); return 0; } else { _tprintf(_T("Success\n")); } //some code CERT_NAME_BLOB SubjectIssuerBlob; memset(&SubjectIssuerBlob, 0, sizeof(SubjectIssuerBlob)); SubjectIssuerBlob.cbData = cbEncoded; SubjectIssuerBlob.pbData = pbEncoded; // Prepare algorithm structure for self-signed certificate CRYPT_ALGORITHM_IDENTIFIER SignatureAlgorithm; memset(&SignatureAlgorithm, 0, sizeof(SignatureAlgorithm)); SignatureAlgorithm.pszObjId = szOID_RSA_SHA1RSA; // Prepare Expiration date for self-signed certificate SYSTEMTIME EndTime; GetSystemTime(&EndTime); EndTime.wYear += 5; // Create self-signed certificate _tprintf(_T("CertCreateSelfSignCertificate... ")); CRYPT_KEY_PROV_INFO* aKeyInfo; aKeyInfo = GetRootCertKeyInfo(); pCertContext = CertCreateSelfSignCertificate(NULL, &SubjectIssuerBlob, 0, aKeyInfo, &SignatureAlgorithm, 0, &EndTime, 0); With the above code I am able to create the certificate but it does not looks be signed by the root certificate. I am unable to figure what I did is right or not.. Any help with be greatly appreciated.. Thanks Asif

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• Rendering a random generated maze in WinForms.NET

  - by Claus Jørgensen

  Hi I'm trying to create a maze-generator, and for this I have implemented the Randomized Prim's Algorithm in C#. However, the result of the generation is invalid. I can't figure out if it's my rendering, or the implementation that's invalid. So for starters, I'd like to have someone take a look at the implementation: maze is a matrix of cells. var cell = maze[0, 0]; cell.Connected = true; var walls = new HashSet<MazeWall>(cell.Walls); while (walls.Count > 0) { var randomWall = walls.GetRandom(); var randomCell = randomWall.A.Connected ? randomWall.B : randomWall.A; if (!randomCell.Connected) { randomWall.IsPassage = true; randomCell.Connected = true; foreach (var wall in randomCell.Walls) walls.Add(wall); } walls.Remove(randomWall); } Here's a example on the rendered result: Edit Ok, lets have a look at the rendering part then: private void MazePanel_Paint(object sender, PaintEventArgs e) { int size = 20; int cellSize = 10; MazeCell[,] maze = RandomizedPrimsGenerator.Generate(size); mazePanel.Size = new Size( size * cellSize + 1, size * cellSize + 1 ); e.Graphics.DrawRectangle(Pens.Blue, 0, 0, size * cellSize, size * cellSize ); for (int y = 0; y < size; y++) for (int x = 0; x < size; x++) { foreach(var wall in maze[x, y].Walls.Where(w => !w.IsPassage)) { if (wall.Direction == MazeWallOrientation.Horisontal) { e.Graphics.DrawLine(Pens.Blue, x * cellSize, y * cellSize, x * cellSize + cellSize, y * cellSize ); } else { e.Graphics.DrawLine(Pens.Blue, x * cellSize, y * cellSize, x * cellSize, y * cellSize + cellSize ); } } } } And I guess, to understand this we need to see the MazeCell and MazeWall class: namespace MazeGenerator.Maze { class MazeCell { public int Column { get; set; } public int Row { get; set; } public bool Connected { get; set; } private List<MazeWall> walls = new List<MazeWall>(); public List<MazeWall> Walls { get { return walls; } set { walls = value; } } public MazeCell() { this.Connected = false; } public void AddWall(MazeCell b) { walls.Add(new MazeWall(this, b)); } } enum MazeWallOrientation { Horisontal, Vertical, Undefined } class MazeWall : IEquatable<MazeWall> { public IEnumerable<MazeCell> Cells { get { yield return CellA; yield return CellB; } } public MazeCell CellA { get; set; } public MazeCell CellB { get; set; } public bool IsPassage { get; set; } public MazeWallOrientation Direction { get { if (CellA.Column == CellB.Column) { return MazeWallOrientation.Horisontal; } else if (CellA.Row == CellB.Row) { return MazeWallOrientation.Vertical; } else { return MazeWallOrientation.Undefined; } } } public MazeWall(MazeCell a, MazeCell b) { this.CellA = a; this.CellB = b; a.Walls.Add(this); b.Walls.Add(this); IsPassage = false; } #region IEquatable<MazeWall> Members public bool Equals(MazeWall other) { return (this.CellA == other.CellA) && (this.CellB == other.CellB); } #endregion } }

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• backtracking in haskell

  - by dmindreader

  I have to traverse a matrix and say how many "characteristic areas" of each type it has. A characteristic area is defined as a zone where elements of value n or n are adjacent. For example, given the matrix: 0 1 2 2 0 1 1 2 0 3 0 0 There's a single characteristic area of type 1 which is equal to the original matrix: 0 1 2 2 0 1 1 2 0 3 0 0 There are two characteristic areas of type 2: 0 0 2 2 0 0 0 0 0 0 0 2 0 0 0 0 0 0 0 0 0 3 0 0 And one characteristic area of type 3: 0 0 0 0 0 0 0 0 0 3 0 0 So, for the function call: countAreas [[0,1,2,2],[0,1,1,2],[0,3,0,0]] The result should be [1,2,1] I haven't defined countAreas yet, I'm stuck with my visit function when it has no more possible squares in which to move it gets stuck and doesn't make the proper recursive call. I'm new to functional programming and I'm still scratching my head about how to implement a backtracking algorithm here. Take a look at my code, what can I do to change it? move_right :: (Int,Int) -> [[Int]] -> Int -> Bool move_right (i,j) mat cond | (j + 1) < number_of_columns mat && consult (i,j+1) mat /= cond = True | otherwise = False move_left :: (Int,Int) -> [[Int]] -> Int -> Bool move_left (i,j) mat cond | (j - 1) >= 0 && consult (i,j-1) mat /= cond = True | otherwise = False move_up :: (Int,Int) -> [[Int]] -> Int -> Bool move_up (i,j) mat cond | (i - 1) >= 0 && consult (i-1,j) mat /= cond = True | otherwise = False move_down :: (Int,Int) -> [[Int]] -> Int -> Bool move_down (i,j) mat cond | (i + 1) < number_of_rows mat && consult (i+1,j) mat /= cond = True | otherwise = False imp :: (Int,Int) -> Int imp (i,j) = i number_of_rows :: [[Int]] -> Int number_of_rows i = length i number_of_columns :: [[Int]] -> Int number_of_columns (x:xs) = length x consult :: (Int,Int) -> [[Int]] -> Int consult (i,j) l = (l !! i) !! j visited :: (Int,Int) -> [(Int,Int)] -> Bool visited x y = elem x y add :: (Int,Int) -> [(Int,Int)] -> [(Int,Int)] add x y = x:y visit :: (Int,Int) -> [(Int,Int)] -> [[Int]] -> Int -> [(Int,Int)] visit (i,j) vis mat cond | move_right (i,j) mat cond && not (visited (i,j+1) vis) = visit (i,j+1) (add (i,j+1) vis) mat cond | move_down (i,j) mat cond && not (visited (i+1,j) vis) = visit (i+1,j) (add (i+1,j) vis) mat cond | move_left (i,j) mat cond && not (visited (i,j-1) vis) = visit (i,j-1) (add (i,j-1) vis) mat cond | move_up (i,j) mat cond && not (visited (i-1,j) vis) = visit (i-1,j) (add (i-1,j) vis) mat cond | otherwise = vis

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• Getting RINGING response on SIP UAC without sending it from the other UAC

  - by TacB0sS

  Hi, I hope this would be my last question about this SIP subject, I have managed to overcome the last issue I had by asking a friend to help me from a remote computer, I'm able to connect between the computers, but here is the thing, according to all the examples I saw, the Callee should invoke the Ringing response, but in my application case I didn't implement it yet, but I still receive on the Caller UAC a Ringing response, this is the SIP messages that are on the caller end: Outgoing Request 5: INVITE sip:[email protected] SIP/2.0 Contact: "Client 310" <sip:[email protected]> From: "Client 310" <sip:[email protected]> Max-Forwards: 32 CSeq: 2 INVITE Call-ID: [email protected] Allow: INVITE,CANCEL,ACK,BYE,OPTIONS Content-Type: application/sdp Proxy-Authorization: Digest username="310",nonce="012afffb",realm="asterisk",uri="sip:[email protected]",algorithm=MD5,response="d19ca5b98450b4be7bd4045edb8a3a2f" Via: SIP/2.0/UDP hostName.hn:5060 To: "Client 320" <sip:[email protected]>;tag=as5a8fa200 Content-Length: 257 v=0 o=310 7108915969559970847 7108915969559970847 IN IP4 xxx.xxx.x.xxx s=- i=Nu-Art Software - TacB0sS VoIP information c=IN IP4 xxx.xxx.x.xxx m=audio 3312 RTP/AVP 0 8 101 a=rtpmap:0 PCMU/8000 a=rtpmap:8 PCMA/8000 a=rtpmap:101 telephone-event/8000 Incoming Response 6: SIP/2.0 100 Trying Via: SIP/2.0/UDP hostName.hn:5060;branch=f8d171d3278788df9e03eb9cf3acba70-xxx.xxx.x.xxx-2-invite-hostName.hn-5060333732;received=79.181.6.233 From: "Client 310" <sip:[email protected]> To: "Client 320" <sip:[email protected]>;tag=as5a8fa200 Call-ID: [email protected] CSeq: 2 INVITE User-Agent: Freeswitch 1.2.3 Allow: INVITE,ACK,CANCEL,OPTIONS,BYE,REFER,SUBSCRIBE,NOTIFY,INFO Supported: replaces Contact: <sip:[email protected]> Content-Length: 0 Incoming Response 7: SIP/2.0 180 Ringing Via: SIP/2.0/UDP hostName.hn:5060;branch=f8d171d3278788df9e03eb9cf3acba70-xxx.xxx.x.xxx-2-invite-hostName.hn-5060333732;received=79.181.6.233 From: "Client 310" <sip:[email protected]> To: "Client 320" <sip:[email protected]>;tag=as5a8fa200 Call-ID: [email protected] CSeq: 2 INVITE User-Agent: Freeswitch 1.2.3 Allow: INVITE,ACK,CANCEL,OPTIONS,BYE,REFER,SUBSCRIBE,NOTIFY,INFO Supported: replaces Contact: <sip:[email protected]> Content-Length: 0 Call to: [email protected] is Ringing Incoming Response 8: SIP/2.0 183 Session Progress Via: SIP/2.0/UDP hostName.hn:5060;branch=f8d171d3278788df9e03eb9cf3acba70-xxx.xxx.x.xxx-2-invite-hostName.hn-5060333732;received=79.181.6.233 From: "Client 310" <sip:[email protected]> To: "Client 320" <sip:[email protected]>;tag=as5a8fa200 Call-ID: [email protected] CSeq: 2 INVITE User-Agent: Freeswitch 1.2.3 Allow: INVITE,ACK,CANCEL,OPTIONS,BYE,REFER,SUBSCRIBE,NOTIFY,INFO Supported: replaces Contact: <sip:[email protected]> Content-Type: application/sdp Content-Length: 264 v=0 o=root 27669 27669 IN IP4 yy.yy.yy.yy s=session c=IN IP4 yy.yy.yy.yy t=0 0 m=audio 10914 RTP/AVP 0 8 101 a=rtpmap:0 PCMU/8000 a=rtpmap:8 PCMA/8000 a=rtpmap:101 telephone-event/8000 a=fmtp:101 0-16 a=silenceSupp:off - - - - a=ptime:20 a=sendrecv Incoming Response 9: SIP/2.0 503 Service Unavailable Via: SIP/2.0/UDP hostName.hn:5060;branch=f8d171d3278788df9e03eb9cf3acba70-xxx.xxx.x.xxx-2-invite-hostName.hn-5060333732;received=79.181.6.233 From: "Client 310" <sip:[email protected]> To: "Client 320" <sip:[email protected]>;tag=as5a8fa200 Call-ID: [email protected] CSeq: 2 INVITE User-Agent: Freeswitch 1.2.3 Allow: INVITE,ACK,CANCEL,OPTIONS,BYE,REFER,SUBSCRIBE,NOTIFY,INFO Supported: replaces Content-Length: 0 I do not respond to the invite, that is why all this is happening, but why am I getting a ringing if I'm not the one sending it. Thanks, Adam.

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• Boost tuple + transform

  - by JH

  Is it possible to do the following. Say my boost tuple has <String, int> I would like to use std::transform + mem_fun to insert only the String element in a corresponding vector. Is it possible or are we required to use a loop and push_back(get<0) Ie the following doesn't like to compile... (unknown types...) result.resize(storage.size()) std::transform(storage.begin(), storage.end(), result.begin(), std::mem_fun(&boost::get<0>)); Here is an example (trying one of the comments): #include <boost/tuple/tuple.hpp> #include <vector> #include <string> #include <algorithm> int main(int argc, char**argv) { std::vector< boost::tuple<std::string, int> > storage; std::vector< std::string> result; result.resize(storage.size()); std::transform(storage.begin(), storage.end(), result.begin(), &boost::get<0, boost::tuple<std::string, int> >); return 0; } Output: g++ test.cpp /usr/include/boost/tuple/detail/tuple_basic.hpp: In instantiation of `boost::tuples::cons<boost::tuples::tuple<std::string, int, boost::tuples::null_type, boost::tuples::null_type, boost::tuples::null_type, boost::tuples::null_type, boost::tuples::null_type, boost::tuples::null_type, boost::tuples::null_type, boost::tuples::null_type>, TT>': /usr/include/boost/tuple/detail/tuple_basic.hpp:151: instantiated from `boost::tuples::element<0, boost::tuples::cons<boost::tuples::tuple<std::string, int, boost::tuples::null_type, boost::tuples::null_type, boost::tuples::null_type, boost::tuples::null_type, boost::tuples::null_type, boost::tuples::null_type, boost::tuples::null_type, boost::tuples::null_type>, TT> >' test.cpp:14: instantiated from here /usr/include/boost/tuple/detail/tuple_basic.hpp:329: error: `boost::tuples::cons<HT, TT>::tail' has incomplete type /usr/include/boost/tuple/detail/tuple_basic.hpp:329: error: invalid use of template type parameter test.cpp: In function `int main(int, char**)': test.cpp:14: error: no matching function for call to `transform(__gnu_cxx::__normal_iterator<boost::tuples::tuple<std::string, int, boost::tuples::null_type, boost::tuples::null_type, boost::tuples::null_type, boost::tuples::null_type, boost::tuples::null_type, boost::tuples::null_type, boost::tuples::null_type, boost::tuples::null_type>*, std::vector<boost::tuples::tuple<std::string, int, boost::tuples::null_type, boost::tuples::null_type, boost::tuples::null_type, boost::tuples::null_type, boost::tuples::null_type, boost::tuples::null_type, boost::tuples::null_type, boost::tuples::null_type>, std::allocator<boost::tuples::tuple<std::string, int, boost::tuples::null_type, boost::tuples::null_type, boost::tuples::null_type, boost::tuples::null_type, boost::tuples::null_type, boost::tuples::null_type, boost::tuples::null_type, boost::tuples::null_type> > > >, __gnu_cxx::__normal_iterator<boost::tuples::tuple<std::string, int, boost::tuples::null_type, boost::tuples::null_type, boost::tuples::null_type, boost::tuples::null_type, boost::tuples::null_type, boost::tuples::null_type, boost::tuples::null_type, boost::tuples::null_type>*, std::vector<boost::tuples::tuple<std::string, int, boost::tuples::null_type, boost::tuples::null_type, boost::tuples::null_type, boost::tuples::null_type, boost::tuples::null_type, boost::tuples::null_type, boost::tuples::null_type, boost::tuples::null_type>, std::allocator<boost::tuples::tuple<std::string, int, boost::tuples::null_type, boost::tuples::null_type, boost::tuples::null_type, boost::tuples::null_type, boost::tuples::null_type, boost::tuples::null_type, boost::tuples::null_type, boost::tuples::null_type> > > >, __gnu_cxx::__normal_iterator<std::string*, std::vector<std::string, std::allocator<std::string> > >, <unresolved overloaded function type>)'

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• Recognize objects in image

  - by DoomStone

  Hello I am in the process of doing a school project, where we have a robot driving on the ground in between Flamingo plates. We need to create an algorithm that can identify the locations of these plates, so we can create paths around them (We are using A Star for that). So far have we worked with AForged Library and we have created the following class, the only problem with this is that when it create the rectangles dose it not take in account that the plates are not always parallel with the camera border, and it that case will it just create a rectangle that cover the whole plate. So we need to some way find the rotation on the object, or another way to identify this. I have create an image that might help explain this Image the describe the problem: http://img683.imageshack.us/img683/9835/imagerectangle.png Any help on how I can do this would be greatly appreciated. Any other information or ideers are always welcome. public class PasteMap { private Bitmap image; private Bitmap processedImage; private Rectangle[] rectangels; public void initialize(Bitmap image) { this.image = image; } public void process() { processedImage = image; processedImage = applyFilters(processedImage); processedImage = filterWhite(processedImage); rectangels = extractRectangles(processedImage); //rectangels = filterRectangles(rectangels); processedImage = drawRectangelsToImage(processedImage, rectangels); } public Bitmap getProcessedImage { get { return processedImage; } } public Rectangle[] getRectangles { get { return rectangels; } } private Bitmap applyFilters(Bitmap image) { image = new ContrastCorrection(2).Apply(image); image = new GaussianBlur(10, 10).Apply(image); return image; } private Bitmap filterWhite(Bitmap image) { Bitmap test = new Bitmap(image.Width, image.Height); for (int width = 0; width < image.Width; width++) { for (int height = 0; height < image.Height; height++) { if (image.GetPixel(width, height).R > 200 && image.GetPixel(width, height).G > 200 && image.GetPixel(width, height).B > 200) { test.SetPixel(width, height, Color.White); } else test.SetPixel(width, height, Color.Black); } } return test; } private Rectangle[] extractRectangles(Bitmap image) { BlobCounter bc = new BlobCounter(); bc.FilterBlobs = true; bc.MinWidth = 5; bc.MinHeight = 5; // process binary image bc.ProcessImage( image ); Blob[] blobs = bc.GetObjects(image, false); // process blobs List<Rectangle> rects = new List<Rectangle>(); foreach (Blob blob in blobs) { if (blob.Area > 1000) { rects.Add(blob.Rectangle); } } return rects.ToArray(); } private Rectangle[] filterRectangles(Rectangle[] rects) { List<Rectangle> Rectangles = new List<Rectangle>(); foreach (Rectangle rect in rects) { if (rect.Width > 75 && rect.Height > 75) Rectangles.Add(rect); } return Rectangles.ToArray(); } private Bitmap drawRectangelsToImage(Bitmap image, Rectangle[] rects) { BitmapData data = image.LockBits(new Rectangle(0, 0, image.Width, image.Height), ImageLockMode.ReadWrite, PixelFormat.Format24bppRgb); foreach (Rectangle rect in rects) Drawing.FillRectangle(data, rect, Color.Red); image.UnlockBits(data); return image; } }

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• Creating a GraphicsPath from a semi-transparent bitmap

  - by Moozhe

  I want to create a GraphicsPath and a list of Points to form the outline of the non-transparent area of a bitmap. If needed, I can guarantee that each image has only one solid collection of nontransparent pixels. So for example, I should be able to record the points either clockwise or counter-clockwise along the edge of the pixels and perform a full closed loop. The speed of this algorithm is not important. However, the efficiency of the resulting points is semi-important if I can skip some points to reduce in a smaller and less complex GraphicsPath. I will list my current code below which works perfectly with most images. However, some images which are more complex end up with paths which seem to connect in the wrong order. I think I know why this occurs, but I can't come up with a solution. public static Point[] GetOutlinePoints(Bitmap image) { List<Point> outlinePoints = new List<Point>(); BitmapData bitmapData = image.LockBits(new Rectangle(0, 0, image.Width, image.Height), ImageLockMode.ReadOnly, PixelFormat.Format32bppArgb); byte[] originalBytes = new byte[image.Width * image.Height * 4]; Marshal.Copy(bitmapData.Scan0, originalBytes, 0, originalBytes.Length); for (int x = 0; x < bitmapData.Width; x++) { for (int y = 0; y < bitmapData.Height; y++) { byte alpha = originalBytes[y * bitmapData.Stride + 4 * x + 3]; if (alpha != 0) { Point p = new Point(x, y); if (!ContainsPoint(outlinePoints, p)) outlinePoints.Add(p); break; } } } for (int y = 0; y < bitmapData.Height; y++) { for (int x = bitmapData.Width - 1; x >= 0; x--) { byte alpha = originalBytes[y * bitmapData.Stride + 4 * x + 3]; if (alpha != 0) { Point p = new Point(x, y); if (!ContainsPoint(outlinePoints, p)) outlinePoints.Add(p); break; } } } for (int x = bitmapData.Width - 1; x >= 0; x--) { for (int y = bitmapData.Height - 1; y >= 0; y--) { byte alpha = originalBytes[y * bitmapData.Stride + 4 * x + 3]; if (alpha != 0) { Point p = new Point(x, y); if (!ContainsPoint(outlinePoints, p)) outlinePoints.Add(p); break; } } } for (int y = bitmapData.Height - 1; y >= 0; y--) { for (int x = 0; x < bitmapData.Width; x++) { byte alpha = originalBytes[y * bitmapData.Stride + 4 * x + 3]; if (alpha != 0) { Point p = new Point(x, y); if (!ContainsPoint(outlinePoints, p)) outlinePoints.Add(p); break; } } } // Added to close the loop outlinePoints.Add(outlinePoints[0]); image.UnlockBits(bitmapData); return outlinePoints.ToArray(); } public static bool ContainsPoint(IEnumerable<Point> points, Point value) { foreach (Point p in points) { if (p == value) return true; } return false; } And when I turn the points into a path: GraphicsPath outlinePath = new GraphicsPath(); outlinePath.AddLines(_outlinePoints); Here's an example showing what I want. The red outline should be an array of points which can be made into a GraphicsPath in order to perform hit detection, draw an outline pen, and fill it with a brush.

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• Linking LLVM JIT Code to Static LLVM Libraries?

  - by inflector

  I'm in the process of implementing a cross-platform (Mac OS X, Windows, and Linux) application which will do lots of CPU intensive analysis of financial data. The bulk of the analysis engine will be written in C++ for speed reasons, with a user-accessible scripting engine interfacing with the C++ testing engine. I want to write several scripting front-ends over time to emulate other popular software with existing large user bases. The first front will be a VisualBasic-like scripting language. I'm thinking that LLVM would be perfect for my needs. Performance is very important because of the sheer amount of data; it can take hours or days to run a single run of tests to get an answer. I believe that using LLVM will also allow me to use a single back-end solution while I implement different front-ends for different flavors of the scripting language over time. The testing engine itself will be separated from the interface and testing will even take place in a separate process with progress and results being reported to the testing management interface. Tests will consist of scripting code integrated with the testing engine code. In a previous implementation of a similar commercial testing system I wrote, I built a fast interpreter which easily interfaced with the testing library because it was written in C++ and linked directly to the testing engine library. Callbacks from scripting code to testing library objects involved translating between the formats with significant overhead. I'm imagining that with LLVM, I could implement the callbacks into C++ directly so that I could make the scripting code work almost as if it had been written in C++. Likewise, if all the code was compiled to LLVM byte-code format, it seems like the LLVM optimizers could optimize across the boundaries between the scripting language and the testing engine code that was written in C++. I don't want to have to compile the testing engine every time. Ideally, I'd like to JIT compile only the scripting code. For small tests, I'd skip some optimization passes, while for large tests, I'd perform full optimizations during the link. So is this possible? Can I precompile the testing engine to a .o object file or .a library file and then link in the scripting code using the JIT? Finally, ideally, I'd like to have the scripting code implement specific methods as subclasses for a specific C++ class. So the C++ testing engine would only see C++ objects while the JIT setup code compiled scripting code that implemented some of the methods for the objects. It seems that if I used the right name mangling algorithm it would be relatively easy to set up the LLVM generation for the scripting language to look like a C++ method call which could then be linked into the testing engine. Thus the linking stage would go in two directions, calls from the scripting language into the testing engine objects to retrieve pricing information and test state information and calls from the testing engine of methods of some particular C++ objects where the code was supplied not from C++ but from the scripting language. In summary: 1) Can I link in precompiled (either .bc, .o, or .a) files as part of the JIT compilation, code-generation process? 2) Can I link in code using that the process in 1) above in such a way that I am able to create code that acts as if it was all written in C++?

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• curl multipart/form-data help

  - by user253530

  Hi am trying to post some data on a website using CURL. The posting process has 3 steps. 1. enter a URL, submit and get to the 2nd step with some fields already completed 2. submit again, after you entered some more data and preview the form. 3. submit the final data. The problem is that after the second step, the form data looks like this POSTDATA =-----------------------------12249266671528 Content-Disposition: form-data; name="title" Filme 2010, filme 2009, filme noi, programe TV, program cinema, premiere cinema, trailere filme - CineMagia.ro -----------------------------12249266671528 Content-Disposition: form-data; name="category" 3 -----------------------------12249266671528 Content-Disposition: form-data; name="tags" filme, programe tv, program cinema -----------------------------12249266671528 Content-Disposition: form-data; name="bodytext" Filme 2010, filme 2009, filme noi, programe TV, program cinema, premiere cinema, trailere filme -----------------------------12249266671528 Content-Disposition: form-data; name="trackback" -----------------------------12249266671528 Content-Disposition: form-data; name="url" http://cinemagia.ro -----------------------------12249266671528 Content-Disposition: form-data; name="phase" 2 -----------------------------12249266671528 Content-Disposition: form-data; name="randkey" 9510520 -----------------------------12249266671528 Content-Disposition: form-data; name="id" 17753 -----------------------------12249266671528-- I am stuck trying to devise an algorithm that will generate this kind of POST data for the second step. Just to mention the URL of the form never changes. It is always: http://www.xxx.com/submit. There is only a hidden input called "phase" that changes according to the step i am currently on (phase = 1, phase = 2, phase = 3). Any help, be it either code, pseudo-code or just guidance would be greatly appreciated. My code so far: function postBlvsocialbookmarkingcom($curl,$vars) { extract($vars); $baseUrl = "http://www.blv-socialbookmarking.com/"; //step 1: login $curl->setRedirect(); $page = $curl->post ($baseUrl.'login.php?return=/index.php', array ('username' => $username, 'password' => $password, 'processlogin' => '1', 'return' => '/index.php')); if ($err = $curl->getError ()) { return $err; } //post step 1---- //get random key $page = $curl->post($baseUrl.'/submit', array()); $randomKey = explode('<input type="hidden" name="randkey" value="',$page); $randKey = explode('"',$randomKey[1]); //------------------------------------- $page = $curl->post($baseUrl.'/submit', array('url'=>$address,'phase'=>'1','randkey'=>$randKey[0],'id'=>'c_1')); if ($err = $curl->getError ()) { return $err; } //echo $page; // //post step 2 $page = $curl->post ($baseUrl.'/submit', array ('title' => $title, 'category'=>'1', 'tags' => $tags, 'bodytext' => $description, 'phase' => '2')); if ($err = $curl->getError ()) { return $err; } echo $page; //post step 3 $page = $curl->post ($baseUrl.'/submit', array ('phase' => '3')); if ($err = $curl->getError ()) { return $err; } echo $page; }

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• Mobile App Data Syncronization

  - by Matt Rogish

  Let's say I have a mobile app that uses HTML5 SQLite DB (and/or the HTML5 key-value store). Assets (media files, PDFs, etc.) are stored locally on the mobile device. Luckily enough, the mobile device is a read-only copy of the "centralized" storage, so the mobile device won't have to propagate changes upstream. However, as the server changes assets (creates new ones, modifies existing, deletes old ones) I need to propagate those changes back to the mobile app. Assume that server changes are grouped into changesets (version number n) that contain some information (added element XYZ, deleted id = 45, etc.) and that the mobile device has limited CPU/bandwidth, so most of the processing has to take place on the server. I can think of a couple of methods to do this. All have trade-offs and at this point, I'm unsure which is the right course of action... Method 1: For change set n, store the "diff" of the current n and previous n-1. When a client with version y asks if there have been any changes, send the change sets from version y up to the current version. e.g. added item 334, contents: xxx. Deleted picture 44. Deleted PDF 11. Changed 33. added picture 99. Characteristics: Diffs take up space, although in theory would be kept small. However, all diffs must be kept around indefinitely (should a v1 app have not been updated for a year, must apply v2..v100). High latency devices (mobile apps) will incur a penalty to send lots of small files (assume cannot be zipped or tarr'd up into one file) Very few server CPU resources required, as all it does is send the client a list of files "Dumb" - if I change an item in change set 3, and change it to something else in 4, the client is going to perform both actions, even though #3 is rendered moot by #4. Or, if an asset is added in #4 and removed in #5 - the client will download a file just to delete it later. Method 2: Very similar to method 1 except on the server, do some sort of a diff between the change sets represented by the app version and server version. Package that up and send that single change set to the client. Characteristics: Client-efficient: The client only has to process one file, duplicate or irrelevant changes are stripped out. Server CPU/space intensive. The change sets must be diff'd and then written out to a file that is then sent to the client. Makes diff server scalability an issue. Possibly ways to cache the results and re-use them, but in the wild there's likely to be a lot of different versions so the diff re-use has a limit Diff algorithm is complicated. The change sets must be structured in such a way that an efficient and effective diff can be performed. Method 3: Instead of keeping diffs, write out the entire versioned asset collection to a mobile-database import file. When client requests an update, send the entire database to client and have them update their assets appropriately. Characteristics: Conceptually simple -- easy to develop and deploy Very inefficient as the client database is restored every update. If only one new thing was added, the whole database is refreshed. Server space and CPU efficient. Only the latest version DB needs kept around and the server just throws the file to the client. Others?? Thoughts? Thanks!!

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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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• Python TEA implementation

  - by Gaks

  Anybody knows proper python implementation of TEA (Tiny Encryption Algorithm)? I tried the one I've found here: http://sysadminco.com/code/python-tea/ - but it does not seem to work properly. It returns different results than other implementations in C or Java. I guess it's caused by completely different data types in python (or no data types in fact). Here's the code and an example: def encipher(v, k): y=v[0];z=v[1];sum=0;delta=0x9E3779B9;n=32 w=[0,0] while(n>0): y += (z << 4 ^ z >> 5) + z ^ sum + k[sum & 3] y &= 4294967295L # maxsize of 32-bit integer sum += delta z += (y << 4 ^ y >> 5) + y ^ sum + k[sum>>11 & 3] z &= 4294967295L n -= 1 w[0]=y; w[1]=z return w def decipher(v, k): y=v[0] z=v[1] sum=0xC6EF3720 delta=0x9E3779B9 n=32 w=[0,0] # sum = delta<<5, in general sum = delta * n while(n>0): z -= (y << 4 ^ y >> 5) + y ^ sum + k[sum>>11 & 3] z &= 4294967295L sum -= delta y -= (z << 4 ^ z >> 5) + z ^ sum + k[sum&3] y &= 4294967295L n -= 1 w[0]=y; w[1]=z return w Python example: >>> import tea >>> key = [0xbe168aa1, 0x16c498a3, 0x5e87b018, 0x56de7805] >>> v = [0xe15034c8, 0x260fd6d5] >>> res = tea.encipher(v, key) >>> "%X %X" % (res[0], res[1]) **'70D16811 F935148F'** C example: #include <unistd.h> #include <stdio.h> void encipher(unsigned long *const v,unsigned long *const w, const unsigned long *const k) { register unsigned long y=v[0],z=v[1],sum=0,delta=0x9E3779B9, a=k[0],b=k[1],c=k[2],d=k[3],n=32; while(n-->0) { sum += delta; y += (z << 4)+a ^ z+sum ^ (z >> 5)+b; z += (y << 4)+c ^ y+sum ^ (y >> 5)+d; } w[0]=y; w[1]=z; } int main() { unsigned long v[] = {0xe15034c8, 0x260fd6d5}; unsigned long key[] = {0xbe168aa1, 0x16c498a3, 0x5e87b018, 0x56de7805}; unsigned long res[2]; encipher(v, res, key); printf("%X %X\n", res[0], res[1]); return 0; } $ ./tea **D6942D68 6F87870D** Please note, that both examples were run with the same input data (v and key), but results were different. I'm pretty sure C implementation is correct - it comes from a site referenced by wikipedia (I couldn't post a link to it because I don't have enough reputation points yet - some antispam thing)

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• Managing highly repetitive code and documentation in Java

  - by polygenelubricants

  Highly repetitive code is generally a bad thing, and there are design patterns that can help minimize this. However, sometimes it's simply inevitable due to the constraints of the language itself. Take the following example from java.util.Arrays: /** * Assigns the specified long value to each element of the specified * range of the specified array of longs. The range to be filled * extends from index <tt>fromIndex</tt>, inclusive, to index * <tt>toIndex</tt>, exclusive. (If <tt>fromIndex==toIndex</tt>, the * range to be filled is empty.) * * @param a the array to be filled * @param fromIndex the index of the first element (inclusive) to be * filled with the specified value * @param toIndex the index of the last element (exclusive) to be * filled with the specified value * @param val the value to be stored in all elements of the array * @throws IllegalArgumentException if <tt>fromIndex &gt; toIndex</tt> * @throws ArrayIndexOutOfBoundsException if <tt>fromIndex &lt; 0</tt> or * <tt>toIndex &gt; a.length</tt> */ public static void fill(long[] a, int fromIndex, int toIndex, long val) { rangeCheck(a.length, fromIndex, toIndex); for (int i=fromIndex; i<toIndex; i++) a[i] = val; } The above snippet appears in the source code 8 times, with very little variation in the documentation/method signature but exactly the same method body, one for each of the root array types int[], short[], char[], byte[], boolean[], double[], float[], and Object[]. I believe that unless one resorts to reflection (which is an entirely different subject in itself), this repetition is inevitable. I understand that as a utility class, such high concentration of repetitive Java code is highly atypical, but even with the best practice, repetition does happen! Refactoring doesn't always work because it's not always possible (the obvious case is when the repetition is in the documentation). Obviously maintaining this source code is a nightmare. A slight typo in the documentation, or a minor bug in the implementation, is multiplied by however many repetitions was made. In fact, the best example happens to involve this exact class: Google Research Blog - Extra, Extra - Read All About It: Nearly All Binary Searches and Mergesorts are Broken (by Joshua Bloch, Software Engineer) The bug is a surprisingly subtle one, occurring in what many thought to be just a simple and straightforward algorithm. // int mid =(low + high) / 2; // the bug int mid = (low + high) >>> 1; // the fix The above line appears 11 times in the source code! So my questions are: How are these kinds of repetitive Java code/documentation handled in practice? How are they developed, maintained, and tested? Do you start with "the original", and make it as mature as possible, and then copy and paste as necessary and hope you didn't make a mistake? And if you did make a mistake in the original, then just fix it everywhere, unless you're comfortable with deleting the copies and repeating the whole replication process? And you apply this same process for the testing code as well? Would Java benefit from some sort of limited-use source code preprocessing for this kind of thing? Perhaps Sun has their own preprocessor to help write, maintain, document and test these kind of repetitive library code? A comment requested another example, so I pulled this one from Google Collections: com.google.common.base.Predicates lines 276-310 (AndPredicate) vs lines 312-346 (OrPredicate). The source for these two classes are identical, except for: AndPredicate vs OrPredicate (each appears 5 times in its class) "And(" vs Or(" (in the respective toString() methods) #and vs #or (in the @see Javadoc comments) true vs false (in apply; ! can be rewritten out of the expression) -1 /* all bits on */ vs 0 /* all bits off */ in hashCode() &= vs |= in hashCode()

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• Find optimal strategy and AI for the game 'Proximity'?

  - by smci

  'Proximity' is a strategy game of territorial domination similar to Othello, Go and Risk. Two players, uses a 10x12 hex grid. Game invented by Brian Cable in 2007. Seems to be a worthy game for discussing a) optimal algorithm then b) how to build an AI. Strategies are going to be probabilistic or heuristic-based, due to the randomness factor, and the insane branching factor (20^120). So it will be kind of hard to compare objectively. A compute time limit of 5s per turn seems reasonable. Game: Flash version here and many copies elsewhere on the web Rules: here Object: to have control of the most armies after all tiles have been placed. Each turn you received a randomly numbered tile (value between 1 and 20 armies) to place on any vacant board space. If this tile is adjacent to any ally tiles, it will strengthen each tile's defenses +1 (up to a max value of 20). If it is adjacent to any enemy tiles, it will take control over them if its number is higher than the number on the enemy tile. Thoughts on strategy: Here are some initial thoughts; setting the computer AI to Expert will probably teach a lot: minimizing your perimeter seems to be a good strategy, to prevent flips and minimize worst-case damage like in Go, leaving holes inside your formation is lethal, only more so with the hex grid because you can lose armies on up to 6 squares in one move low-numbered tiles are a liability, so place them away from your main territory, near the board edges and scattered. You can also use low-numbered tiles to plug holes in your formation, or make small gains along the perimeter which the opponent will not tend to bother attacking. a triangle formation of three pieces is strong since they mutually reinforce, and also reduce the perimeter Each tile can be flipped at most 6 times, i.e. when its neighbor tiles are occupied. Control of a formation can flow back and forth. Sometimes you lose part of a formation and plug any holes to render that part of the board 'dead' and lock in your territory/ prevent further losses. Low-numbered tiles are obvious-but-low-valued liabilities, but high-numbered tiles can be bigger liabilities if they get flipped (which is harder). One lucky play with a 20-army tile can cause a swing of 200 (from +100 to -100 armies). So tile placement will have both offensive and defensive considerations. Comment 1,2,4 seem to resemble a minimax strategy where we minimize the maximum expected possible loss (modified by some probabilistic consideration of the value ß the opponent can get from 1..20 i.e. a structure which can only be flipped by a ß=20 tile is 'nearly impregnable'.) I'm not clear what the implications of comments 3,5,6 are for optimal strategy. Interested in comments from Go, Chess or Othello players. (The sequel ProximityHD for XBox Live, allows 4-player -cooperative or -competitive local multiplayer increases the branching factor since you now have 5 tiles in your hand at any given time, of which you can only play one. Reinforcement of ally tiles is increased to +2 per ally.)

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• How can I read the verbose output from a Cmdlet in C# using Exchange Powershell

  - by mrkeith

  Environment: Exchange 2007 sp3 (2003 sp2 mixed mode) Visual Studio 2008, .Net 3.5 Hello, I'm working with an Exchange powershell move-mailbox cmdlet and have noted when I do so from the Exchange Management shell (using the Verbose switch) there is a ton of real-time information provided. To provide a little context, I'm attempting to create a UI application that moves mailboxes similarly to the Exchange Management Console but desire to support an input file and specific server/database destinations for each entry (and threading). Here's roughly what I have at present but I'm not sure if there is an event I need to register for or what... And to be clear, I desire to get this information in real-time so I may update my UI to reflect what's occurring in the move sequence for the appropriate user (pretty much like the native functionality offered in the Management Console). And in case you are wondering, the reason why I'm not content with the Management Console functionality is, I have an algorithm which I'm using to balance users depending on storage limit, Blackberry use, journaling, exception mailbox size etc which demands user be mapped to specific locations... and I do not desire to create many/several move groups for each common destination or to hunt for lists of users individually through the management console UI. I can not seem to find any good documentation or examples of how to tie into reading the verbose messages that are provided within the console using C# (I see value in being able to read this kind of information in many different scenarios). I've explored the Invoke and InvokeAsync methods and the StateChanged & DataReady events but none of these seem to provide the information (verbose comments) that I'm after. Any direction or examples that can be provided will be very appreciated! A code sample which is little more than how I would ordinarily call any other powershell command follows: // config to use ExMgmt shell, create runspace and open it RunspaceConfiguration rsConfig = RunspaceConfiguration.Create(); PSSnapInException snapInException = null; PSSnapInInfo info = rsConfig.AddPSSnapIn("Microsoft.Exchange.Management.PowerShell.Admin", out snapInException); if (snapInException != null) throw snapInException; Runspace runspace = RunspaceFactory.CreateRunspace(rsConfig); try { runspace.Open(); // create a pipeline and feed script text Pipeline pipeline = runspace.CreatePipeline(); string targetDatabase = @"myServer\myStorageGroup\myDB"; string mbxOwner = "[email protected]"; Command myMoveMailbox = new Command("Move-Mailbox", false, false); myMoveMailbox.Parameters.Add("Identity", mbxOwner); myMoveMailbox.Parameters.Add("TargetDatabase", targetDatabase); myMoveMailbox.Parameters.Add("Verbose"); myMoveMailbox.Parameters.Add("ValidateOnly"); myMoveMailbox.Parameters.Add("Confirm", false); pipeline.Commands.Add(myMoveMailbox); System.Collections.ObjectModel.Collection output = null; // these next few lines that are commented out are where I've tried // registering for events and calling asynchronously but this doesn't // seem to get me anywhere closer // //pipeline.StateChanged += new EventHandler(pipeline_StateChanged); //pipeline.Output.DataReady += new EventHandler(Output_DataReady); //pipeline.InvokeAsync(); //pipeline.Input.Close(); //return; tried these variations that are commented out but none seem to be useful output = pipeline.Invoke(); // Check for errors in the pipeline and throw an exception if necessary if (pipeline.Error != null && pipeline.Error.Count 0) { StringBuilder pipelineError = new StringBuilder(); pipelineError.AppendFormat("Error calling Test() Cmdlet. "); foreach (object item in pipeline.Error.ReadToEnd()) pipelineError.AppendFormat("{0}\n", item.ToString()); throw new Exception(pipelineError.ToString()); } foreach (PSObject psObject in output) { // blah, blah, blah // this is normally where I would read details about a particular PS command // but really pertains to a command once it finishes and has nothing to do with // the verbose messages that I'm after... since this part of the methods pertains // to the after-effects of a command having run, I'm suspecting I need to look to // the asynch invoke method but am not certain or knowing how. } } finally { runspace.Close(); } Thanks! Keith

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