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  • Black Screen: How to set Projection/View Matrix

    - by Lisa
    I have a Windows Phone 8 C#/XAML with DirectX component project. I'm rendering some particles, but each particle is a rectangle versus a square (as I've set the vertices to be positions equally offset from each other). I used an Identity matrix in the view and projection matrix. I decided to add the windows aspect ratio to prevent the rectangles. But now I get a black screen. None of the particles are rendered now. I don't know what's wrong with my matrices. Can anyone see the problem? These are the default matrices in Microsoft's project example. View Matrix: XMVECTOR eye = XMVectorSet(0.0f, 0.7f, 1.5f, 0.0f); XMVECTOR at = XMVectorSet(0.0f, -0.1f, 0.0f, 0.0f); XMVECTOR up = XMVectorSet(0.0f, 1.0f, 0.0f, 0.0f); XMStoreFloat4x4(&m_constantBufferData.view, XMMatrixTranspose(XMMatrixLookAtRH(eye, at, up))); Projection Matrix: void CubeRenderer::CreateWindowSizeDependentResources() { Direct3DBase::CreateWindowSizeDependentResources(); float aspectRatio = m_windowBounds.Width / m_windowBounds.Height; float fovAngleY = 70.0f * XM_PI / 180.0f; if (aspectRatio < 1.0f) { fovAngleY /= aspectRatio; } XMStoreFloat4x4(&m_constantBufferData.projection, XMMatrixTranspose(XMMatrixPerspectiveFovRH(fovAngleY, aspectRatio, 0.01f, 100.0f))); } I've tried modifying them to use cocos2dx's WP8 example. XMMATRIX identityMatrix = XMMatrixIdentity(); float fovy = 60.0f; float aspect = m_windowBounds.Width / m_windowBounds.Height; float zNear = 0.1f; float zFar = 100.0f; float xmin, xmax, ymin, ymax; ymax = zNear * tanf(fovy * XM_PI / 360); ymin = -ymax; xmin = ymin * aspect; xmax = ymax * aspect; XMMATRIX tmpMatrix = XMMatrixPerspectiveOffCenterRH(xmin, xmax, ymin, ymax, zNear, zFar); XMMATRIX projectionMatrix = XMMatrixMultiply(tmpMatrix, identityMatrix); // View Matrix float fEyeX = m_windowBounds.Width * 0.5f; float fEyeY = m_windowBounds.Height * 0.5f; float fEyeZ = m_windowBounds.Height / 1.1566f; float fLookAtX = m_windowBounds.Width * 0.5f; float fLookAtY = m_windowBounds.Height * 0.5f; float fLookAtZ = 0.0f; float fUpX = 0.0f; float fUpY = 1.0f; float fUpZ = 0.0f; XMMATRIX tmpMatrix2 = XMMatrixLookAtRH(XMVectorSet(fEyeX,fEyeY,fEyeZ,0.f), XMVectorSet(fLookAtX,fLookAtY,fLookAtZ,0.f), XMVectorSet(fUpX,fUpY,fUpZ,0.f)); XMMATRIX viewMatrix = XMMatrixMultiply(tmpMatrix2, identityMatrix); XMStoreFloat4x4(&m_constantBufferData.view, viewMatrix); Vertex Shader cbuffer ModelViewProjectionConstantBuffer : register(b0) { //matrix model; matrix view; matrix projection; }; struct VertexInputType { float4 position : POSITION; float2 tex : TEXCOORD0; float4 color : COLOR; }; struct PixelInputType { float4 position : SV_POSITION; float2 tex : TEXCOORD0; float4 color : COLOR; }; PixelInputType main(VertexInputType input) { PixelInputType output; // Change the position vector to be 4 units for proper matrix calculations. input.position.w = 1.0f; //===================================== // TODO: ADDED for testing input.position.z = 0.0f; //===================================== // Calculate the position of the vertex against the world, view, and projection matrices. //output.position = mul(input.position, model); output.position = mul(input.position, view); output.position = mul(output.position, projection); // Store the texture coordinates for the pixel shader. output.tex = input.tex; // Store the particle color for the pixel shader. output.color = input.color; return output; } Before I render the shader, I set the view/projection matrices into the constant buffer void ParticleRenderer::SetShaderParameters() { ViewProjectionConstantBuffer* dataPtr; D3D11_MAPPED_SUBRESOURCE mappedResource; DX::ThrowIfFailed(m_d3dContext->Map(m_constantBuffer.Get(), 0, D3D11_MAP_WRITE_DISCARD, 0, &mappedResource)); dataPtr = (ViewProjectionConstantBuffer*)mappedResource.pData; dataPtr->view = m_constantBufferData.view; dataPtr->projection = m_constantBufferData.projection; m_d3dContext->Unmap(m_constantBuffer.Get(), 0); // Now set the constant buffer in the vertex shader with the updated values. m_d3dContext->VSSetConstantBuffers(0, 1, m_constantBuffer.GetAddressOf() ); // Set shader texture resource in the pixel shader. m_d3dContext->PSSetShaderResources(0, 1, &m_textureView); } Nothing, black screen... I tried so many different look at, eye, and up vectors. I tried transposing the matrices. I've set the particle center position to always be (0, 0, 0), I tried different positions too, just to make sure they're not being rendered offscreen.

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  • SQL SERVER – Backing Up and Recovering the Tail End of a Transaction Log – Notes from the Field #042

    - by Pinal Dave
    [Notes from Pinal]: The biggest challenge which people face is not taking backup, but the biggest challenge is to restore a backup successfully. I have seen so many different examples where users have failed to restore their database because they made some mistake while they take backup and were not aware of the same. Tail Log backup was such an issue in earlier version of SQL Server but in the latest version of SQL Server, Microsoft team has fixed the confusion with additional information on the backup and restore screen itself. Now they have additional information, there are a few more people confused as they have no clue about this. Previously they did not find this as a issue and now they are finding tail log as a new learning. Linchpin People are database coaches and wellness experts for a data driven world. In this 42nd episode of the Notes from the Fields series database expert Tim Radney (partner at Linchpin People) explains in a very simple words, Backing Up and Recovering the Tail End of a Transaction Log. Many times when restoring a database over an existing database SQL Server will warn you about needing to make a tail end of the log backup. This might be your reminder that you have to choose to overwrite the database or could be your reminder that you are about to write over and lose any transactions since the last transaction log backup. You might be asking yourself “What is the tail end of the transaction log”. The tail end of the transaction log is simply any committed transactions that have occurred since the last transaction log backup. This is a very crucial part of a recovery strategy if you are lucky enough to be able to capture this part of the log. Most organizations have chosen to accept some amount of data loss. You might be shaking your head at this statement however if your organization is taking transaction logs backup every 15 minutes, then your potential risk of data loss is up to 15 minutes. Depending on the extent of the issue causing you to have to perform a restore, you may or may not have access to the transaction log (LDF) to be able to back up those vital transactions. For example, if the storage array or disk that holds your transaction log file becomes corrupt or damaged then you wouldn’t be able to recover the tail end of the log. If you do have access to the physical log file then you can still back up the tail end of the log. In 2013 I presented a session at the PASS Summit called “The Ultimate Tail Log Backup and Restore” and have been invited back this year to present it again. During this session I demonstrate how you can back up the tail end of the log even after the data file becomes corrupt. In my demonstration I set my database offline and then delete the data file (MDF). The database can’t become more corrupt than that. I attempt to bring the database back online to change the state to RECOVERY PENDING and then backup the tail end of the log. I can do this by specifying WITH NO_TRUNCATE. Using NO_TRUNCATE is equivalent to specifying both COPY_ONLY and CONTINUE_AFTER_ERROR. It as its name says, does not try to truncate the log. This is a great demo however how could I achieve backing up the tail end of the log if the failure destroys my entire instance of SQL and all I had was the LDF file? During my demonstration I also demonstrate that I can attach the log file to a database on another instance and then back up the tail end of the log. If I am performing proper backups then my most recent full, differential and log files should be on a server other than the one that crashed. I am able to achieve this task by creating new database with the same name as the failed database. I then set the database offline, delete my data file and overwrite the log with my good log file. I attempt to bring the database back online and then backup the log with NO_TRUNCATE just like in the first example. I encourage each of you to view my blog post and watch the video demonstration on how to perform these tasks. I really hope that none of you ever have to perform this in production, however it is a really good idea to know how to do this just in case. It really isn’t a matter of “IF” you will have to perform a restore of a production system but more of a “WHEN”. Being able to recover the tail end of the log in these sever cases could be the difference of having to notify all your business customers of data loss or not. If you want me to take a look at your server and its settings, or if your server is facing any issue we can Fix Your SQL Server. Note: Tim has also written an excellent book on SQL Backup and Recovery, a must have for everyone. Reference: Pinal Dave (http://blog.sqlauthority.com)Filed under: Notes from the Field, PostADay, SQL, SQL Authority, SQL Performance, SQL Query, SQL Server, SQL Tips and Tricks, T SQL

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  • Pirates, Treasure Chests and Architectural Mapping

    Pirate 1: Why do pirates create treasure maps? Pirate 2: I do not know.Pirate 1: So they can find their gold. Yes, that was a bad joke, but it does illustrate a point. Pirates are known for drawing treasure maps to their most prized possession. These documents detail the decisions pirates made in order to hide and find their chests of gold. The map allows them to trace the steps they took originally to hide their treasure so that they may return. As software engineers, programmers, and architects we need to treat software implementations much like our treasure chest. Why is software like a treasure chest? It cost money, time,  and resources to develop (Usually) It can make or save money, time, and resources (Hopefully) If we operate under the assumption that software is like a treasure chest then wouldn’t make sense to document the steps, rationale, concerns, and decisions about how it was designed? Pirates are notorious for documenting where they hide their treasure.  Shouldn’t we as creators of software do the same? By documenting our design decisions and rationale behind them will help others be able to understand and maintain implemented systems. This can only be done if the design decisions are correctly mapped to its corresponding implementation. This allows for architectural decisions to be traced from the conceptual model, architectural design and finally to the implementation. Mapping gives software professional a method to trace the reason why specific areas of code were developed verses other options. Just like the pirates we need to able to trace our steps from the start of a project to its implementation,  so that we will understand why specific choices were chosen. The traceability of a software implementation that actually maps back to its originating design decisions is invaluable for ensuring that architectural drifting and erosion does not take place. The drifting and erosion is prevented by allowing others to understand the rational of why an implementation was created in a specific manor or methodology The process of mapping distinct design concerns/decisions to the location of its implemented is called traceability. In this context traceability is defined as method for connecting distinctive software artifacts. This process allows architectural design models and decisions to be directly connected with its physical implementation. The process of mapping architectural design concerns to a software implementation can be very complex. However, most design decision can be placed in  a few generalized categories. Commonly Mapped Design Decisions Design Rationale Components and Connectors Interfaces Behaviors/Properties Design rational is one of the hardest categories to map directly to an implementation. Typically this rational is mapped or document in code via comments. These comments consist of general design decisions and reasoning because they do not directly refer to a specific part of an application. They typically focus more on the higher level concerns. Components and connectors can directly be mapped to architectural concerns. Typically concerns subdivide an application in to distinct functional areas. These functional areas then can map directly back to their originating concerns.Interfaces can be mapped back to design concerns in one of two ways. Interfaces that pertain to specific function definitions can be directly mapped back to its originating concern(s). However, more complicated interfaces require additional analysis to ensure that the proper mappings are created. Depending on the complexity some Behaviors\Properties can be translated directly into a generic implementation structure that is ready for business logic. In addition, some behaviors can be translated directly in to an actual implementation depending on the complexity and architectural tools used. Mapping design concerns to an implementation is a lot of work to maintain, but is doable. In order to ensure that concerns are mapped correctly and that an implementation correctly reflects its design concerns then one of two standard approaches are usually used. All Changes Come From ArchitectureBy forcing all application changes to come through the architectural model prior to implementation then the existing mappings will be used to locate where in the implementation changes need to occur. Allow Changes From Implementation Or Architecture By allowing changes to come from the implementation and/or the architecture then the other area must be kept in sync. This methodology is more complex compared to the previous approach.  One reason to justify the added complexity for an application is due to the fact that this approach tends to detect and prevent architectural drift and erosion. Additionally, this approach is usually maintained via software because of the complexity. Reference:Taylor, R. N., Medvidovic, N., & Dashofy, E. M. (2009). Software architecture: Foundations, theory, and practice Hoboken, NJ: John Wiley & Sons  

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  • What are the industry metrics for average spend on dev hardware and software? [on hold]

    - by RationalGeek
    I'm trying to budget for my dev shop and compare our budget items to industry expectations. I'm hoping to find some information on what percentage of a dev's salary is generally spent on tooling, both hardware and software. Where can I find such information? If instead there is a source that looks at raw dollars that is useful, too. I can extrapolate what I need from that. NOTE: Your anecdotal evidence from your own job will not be very helpful. I'm looking for industry average statistics from a credible source. EDIT: I'm reluctant to even keep this question going based on the passionate negative responses of commenters, but I do think this is valuable information (assuming anyone will care to answer) so let me make one attempt to clarify why I'm looking for this information, and then leave it at that. I'm not sure why understanding and validating my motives is a necessary step to providing the information, but apparently that is the case, so I will do my best. Firstly, let me respond to the idea that us "management types" shouldn't use these types of metrics to evaluate budgets. I agree in part. Ideally, you should spend whatever is necessary on developers in order to keep them fully happy and productive. And this is true of all employees. However, companies operate in a world of limited resources, and every dollar spent in one area means a dollar not spent in another. So it is not enough to simply say "I need to spend $10,000 per developer next year" without having some way to justify that position. One way to help justify it is to compare yourself against the industry. If it is the case that on average a software shops spends 5% (making up that number) of their total development budget (salaries being the large portion of the other 95%, for arguments sake), and I'm only spending 3%, it helps in the justification process. So, it is not my intent to use this information to limit what I spend on developers, but rather to arm myself with the necessary justification to spend what I need to spend on developers to give them the best tools I can. I have been a developer for many years and I understand the need for proper tooling. Next, let's examine the idea that even considering the relationship between a spend on developer salaries and developer tooling is ludicrous and should be banned from budgetary thinking. As Jimmy Hoffa put it in their comment, it's like saying "I'm going to spend no more than 10% of median employee salary on light bulbs and coffee from now on.". Well, yes, it is like saying that, and from a budgeting perspective, this is a useful way to look at things. If you know that, on average, an employee consumes X dollars of coffee a year, then you can project a coffee budget based on that. And you can compare it to an industry metric to understand where you fall: do you spend more on coffee than other companies or less? Why might this be? If you are a coffee supply manager, that seems like a useful thought process. The same seems to hold true for developers. Now, on to the idea that I need to compare "apples to apples" and only look at other shops that are in the same place geographically, the same business, the same application architecture, and the same development frameworks. I guess if I could find such a statistic that said "a shop that is exactly identical to yours spends X on developer tooling" it would be wonderful. But there is plenty of value in an average statistic. Here's an analogy: let's say you are working on a household budget and need to decide how much to spend on groceries. Is it enough to know that the average consumer spends 15% on groceries and therefore decide that you will budget exactly 15%? No. You have to tweak your budget based on your individual needs and situation. But the generalized statistic does help in this evaluation. You can know if your budget is grossly off from what others are doing, and this can help you figure out why this is. So, I will concede the point that it would be better to find statistics that align to my shop, though I think any statistics I could find would be useful for what I'm doing. In that light, let's say that my shop is mostly focused on ASP.NET web applications. That doesn't map perfectly to reality because large enterprises have very heterogenous IT environments. But if I was going to pick one technology that is our focus that would be it. But, if you were to point me at some statistics that are related to a Linux shop doing embedded Java applications, I would still find it useful as a point of comparison. SUMMARY: Let me try to rephrase my question. I'm trying to find industry metrics on how much dev shops spend on developer tooling, both hardware and software. I don't so much care whether it is expressed as a percentage of total budget or as X dollars per dev or as Y percentage of salary. Any metric would be useful. If there are metrics that are specific to ASP.NET dev shops in the Northeast US, all the better, but I would be happy to find anything.

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  • Database Partitioning and Multiple Data Source Considerations

    - by Jeffrey McDaniel
    With the release of P6 Reporting Database 3.0 partitioning was added as a feature to help with performance and data management.  Careful investigation of requirements should be conducting prior to installation to help improve overall performance throughout the lifecycle of the data warehouse, preventing future maintenance that would result in data loss. Before installation try to determine how many data sources and partitions will be required along with the ranges.  In P6 Reporting Database 3.0 any adjustments outside of defaults must be made in the scripts and changes will require new ETL runs for each data source.  Considerations: 1. Standard Edition or Enterprise Edition of Oracle Database.   If you aren't using Oracle Enterprise Edition Database; the partitioning feature is not available. Multiple Data sources are only supported on Enterprise Edition of Oracle   Database. 2. Number of Data source Ids for partitioning during configuration.   This setting will specify how many partitions will be allocated for tables containing data source information.  This setting requires some evaluation prior to installation as       there are repercussions if you don't estimate correctly.   For example, if you configured the software for only 2 data sources and the partition setting was set to 2, however along came a 3rd data source.  The necessary steps to  accommodate this change are as follows: a) By default, 3 partitions are configured in the Reporting Database scripts. Edit the create_star_tables_part.sql script located in <installation directory>\star\scripts   and search for partition.  You’ll see P1, P2, P3.  Add additional partitions and sub-partitions for P4 and so on. These will appear in several areas.  (See P6 Reporting Database 3.0 Installation and Configuration guide for more information on this and how to adjust partition ranges). b) Run starETL -r.  This will recreate each table with the new partition key.  The effect of this step is that all tables data will be lost except for history related tables.   c) Run starETL for each of the 3 data sources (with the data source # (starETL.bat "-s2" -as defined in P6 Reporting Database 3.0 Installation and Configuration guide) The best strategy for this setting is to overestimate based on possible growth.  If during implementation it is deemed that there are atleast 2 data sources with possibility for growth, it is a better idea to set this setting to 4 or 5, allowing room for the future and preventing a ‘start over’ scenario. 3. The Number of Partitions and the Number of Months per Partitions are not specific to multi-data source.  These settings work in accordance to a sub partition of larger tables with regard to time related data.  These settings are dataset specific for optimization.  The number of months per partition is self explanatory, optimally the smaller the partition, the better query performance so if the dataset has an extremely large number of spread/history records, a lower number of months is optimal.  Working in accordance with this setting is the number of partitions, this will determine how many "buckets" will be created per the number of months setting.  For example, if you kept the default for # of partitions of 3, and select 2 months for each partitions you would end up with: -1st partition, 2 months -2nd partition, 2 months -3rd partition, all the remaining records Therefore with records to this setting, it is important to analyze your source db spread ranges and history settings when determining the proper number of months per partition and number of partitions to optimize performance.  Also be aware the DBA will need to monitor when these partition ranges will fill up and when additional partitions will need to be added.  If you get to the final range partition and there are no additional range partitions all data will be included into the last partition. 

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  • How to move a rectangle properly?

    - by bodycountPP
    I recently started to learn OpenGL. Right now I finished the first chapter of the "OpenGL SuperBible". There were two examples. The first had the complete code and showed how to draw a simple triangle. The second example is supposed to show how to move a rectangle using SpecialKeys. The only code provided for this example was the SpecialKeys method. I still tried to implement it but I had two problems. In the previous example I declared and instaciated vVerts in the SetupRC() method. Now as it is also used in the SpecialKeys() method, I moved the declaration and instantiation to the top of the code. Is this proper c++ practice? I copied the part where vertex positions are recalculated from the book, but I had to pick the vertices for the rectangle on my own. So now every time I press a key for the first time the rectangle's upper left vertex is moved to (-0,5:-0.5). This ok because of GLfloat blockX = vVerts[0]; //Upper left X GLfloat blockY = vVerts[7]; // Upper left Y But I also think that this is the reason why my rectangle is shifted in the beginning. After the first time a key was pressed everything works just fine. Here is my complete code I hope you can help me on those two points. GLBatch squareBatch; GLShaderManager shaderManager; //Load up a triangle GLfloat vVerts[] = {-0.5f,0.5f,0.0f, 0.5f,0.5f,0.0f, 0.5f,-0.5f,0.0f, -0.5f,-0.5f,0.0f}; //Window has changed size, or has just been created. //We need to use the window dimensions to set the viewport and the projection matrix. void ChangeSize(int w, int h) { glViewport(0,0,w,h); } //Called to draw the scene. void RenderScene(void) { //Clear the window with the current clearing color glClear(GL_COLOR_BUFFER_BIT|GL_DEPTH_BUFFER_BIT|GL_STENCIL_BUFFER_BIT); GLfloat vRed[] = {1.0f,0.0f,0.0f,1.0f}; shaderManager.UseStockShader(GLT_SHADER_IDENTITY,vRed); squareBatch.Draw(); //perform the buffer swap to display the back buffer glutSwapBuffers(); } //This function does any needed initialization on the rendering context. //This is the first opportunity to do any OpenGL related Tasks. void SetupRC() { //Blue Background glClearColor(0.0f,0.0f,1.0f,1.0f); shaderManager.InitializeStockShaders(); squareBatch.Begin(GL_QUADS,4); squareBatch.CopyVertexData3f(vVerts); squareBatch.End(); } //Respond to arrow keys by moving the camera frame of reference void SpecialKeys(int key,int x,int y) { GLfloat stepSize = 0.025f; GLfloat blockSize = 0.5f; GLfloat blockX = vVerts[0]; //Upper left X GLfloat blockY = vVerts[7]; // Upper left Y if(key == GLUT_KEY_UP) { blockY += stepSize; } if(key == GLUT_KEY_DOWN){blockY -= stepSize;} if(key == GLUT_KEY_LEFT){blockX -= stepSize;} if(key == GLUT_KEY_RIGHT){blockX += stepSize;} //Recalculate vertex positions vVerts[0] = blockX; vVerts[1] = blockY - blockSize*2; vVerts[3] = blockX + blockSize * 2; vVerts[4] = blockY - blockSize *2; vVerts[6] = blockX+blockSize*2; vVerts[7] = blockY; vVerts[9] = blockX; vVerts[10] = blockY; squareBatch.CopyVertexData3f(vVerts); glutPostRedisplay(); } //Main entry point for GLUT based programs int main(int argc, char** argv) { //Sets the working directory. Not really needed gltSetWorkingDirectory(argv[0]); //Passes along the command-line parameters and initializes the GLUT library. glutInit(&argc,argv); //Tells the GLUT library what type of display mode to use, when creating the window. //Double buffered window, RGBA-Color mode,depth-buffer as part of our display, stencil buffer also available glutInitDisplayMode(GLUT_DOUBLE|GLUT_RGBA|GLUT_DEPTH|GLUT_STENCIL); //Window size glutInitWindowSize(800,600); glutCreateWindow("MoveRect"); glutReshapeFunc(ChangeSize); glutDisplayFunc(RenderScene); glutSpecialFunc(SpecialKeys); //initialize GLEW library GLenum err = glewInit(); //Check that nothing goes wrong with the driver initialization before we try and do any rendering. if(GLEW_OK != err) { fprintf(stderr,"Glew Error: %s\n",glewGetErrorString); return 1; } SetupRC(); glutMainLoop(); return 0; }

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  • PHP crashing (seg-fault) under mod_fcgi, apache

    - by Andras Gyomrey
    I've been programming a site using: Zend Framework 1.11.5 (complete MVC) PHP 5.3.6 Apache 2.2.19 CentOS 5.6 i686 virtuozzo on vps cPanel WHM 11.30.1 (build 4) Mysql 5.1.56-log Mysqli API 5.1.56 The issue started here http://stackoverflow.com/questions/6769515/php-programming-seg-fault. In brief, php is giving me random segmentation-faults. [Wed Jul 20 17:45:34 2011] [error] mod_fcgid: process /usr/local/cpanel/cgi-sys/php5(11562) exit(communication error), get unexpected signal 11 [Wed Jul 20 17:45:34 2011] [warn] [client 190.78.208.30] (104)Connection reset by peer: mod_fcgid: error reading data from FastCGI server [Wed Jul 20 17:45:34 2011] [error] [client 190.78.208.30] Premature end of script headers: index.php About extensions. When i compile php with "--enable-debug" flag, i have to disable this line: zend_extension="/usr/local/IonCube/ioncube_loader_lin_5.3.so" Otherwise, the server doesn't accept requests and i get a "The connection with the server was reset". It is possible that i have to disable eaccelerator too because of the same reason. I still don't get why apache gets running it some times and some others not: extension="eaccelerator.so" Anyway, after i get httpd running, seg-faults can occurr randomly. If i don't compile php with "--enable-debug" flag, i can get DETERMINISTICALLY a php crash: <?php class Admin_DbController extends Controller_BaseController { public function updateSqlDefinitionsAction() { $db = Zend_Registry::get('db'); $row = $db->fetchRow("SHOW CREATE TABLE 222AFI"); } } ?> BUT if i compile php with "--enable-debug" flag, it's really hard to get this error. I must add some complexity to make it crash. I have to be doing many paralell requests for a few seconds to get a crash: <?php class Admin_DbController extends Controller_BaseController { public function updateSqlDefinitionsAction() { $db = Zend_Registry::get('db'); $tableList = $db->listTables(); foreach ($tableList as $tableName){ $row = $db->fetchRow("SHOW CREATE TABLE " . $db->quoteIdentifier($tableName)); file_put_contents( DB_DEFINITIONS_PATH . '/' . $tableName . '.sql', $row['Create Table'] . ';' ); } } } ?> Please notice this is the same script, but creating DDL for all tables in database rather than for one. It seems that if php is heavy loaded (with extensions and me doing many paralell requests) it's when i get php to crash. About starting httpd with "-X": i've tried. The thing is, it is already hard to make php crash with --enable-debug. With "-X" option (which only enables one child process) i can't do parallel requests. So i haven't been able to create to proper debug backtrace: https://bugs.php.net/bugs-generating-backtrace.php My concrete question is, what do i do to get a coredump? root@GWT4 [~]# httpd -V Server version: Apache/2.2.19 (Unix) Server built: Jul 20 2011 19:18:58 Cpanel::Easy::Apache v3.4.2 rev9999 Server's Module Magic Number: 20051115:28 Server loaded: APR 1.4.5, APR-Util 1.3.12 Compiled using: APR 1.4.5, APR-Util 1.3.12 Architecture: 32-bit Server MPM: Prefork threaded: no forked: yes (variable process count) Server compiled with.... -D APACHE_MPM_DIR="server/mpm/prefork" -D APR_HAS_SENDFILE -D APR_HAS_MMAP -D APR_HAVE_IPV6 (IPv4-mapped addresses enabled) -D APR_USE_SYSVSEM_SERIALIZE -D APR_USE_PTHREAD_SERIALIZE -D SINGLE_LISTEN_UNSERIALIZED_ACCEPT -D APR_HAS_OTHER_CHILD -D AP_HAVE_RELIABLE_PIPED_LOGS -D DYNAMIC_MODULE_LIMIT=128 -D HTTPD_ROOT="/usr/local/apache" -D SUEXEC_BIN="/usr/local/apache/bin/suexec" -D DEFAULT_PIDLOG="logs/httpd.pid" -D DEFAULT_SCOREBOARD="logs/apache_runtime_status" -D DEFAULT_LOCKFILE="logs/accept.lock" -D DEFAULT_ERRORLOG="logs/error_log" -D AP_TYPES_CONFIG_FILE="conf/mime.types" -D SERVER_CONFIG_FILE="conf/httpd.conf"

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  • How do I create a Linked Server in SQL Server 2005 to a password protected Access 95 database?

    - by Brad Knowles
    I need to create a linked server with SQL Server Management Studio 2005 to an Access 95 database, which happens to be password protected at the database level. User level security has not been implemented. I cannot convert the Access database to a newer version. It is being used by a 3rd party application; so modifying it, in any way, is not allowed. I've tried using the Jet 4.0 OLE DB Provider and the ODBC OLE DB Provider. The 3rd party application creates a System DSN (with the proper database password), but I've not had any luck in using either method. If I were using a standard connection string, I think it would look something like this: Provider=Microsoft.Jet.OLEDB.4.0;Data Source='C:\Test.mdb';Jet OLEDB:Database Password=####; I'm fairly certain I need to somehow incorporate Jet OLEDB:Database Password into the linked server setup, but haven't figured out how. I've posted the scripts I'm using along with the associated error messages below. Any help is greatly appreciated. I'll provide more details if needed, just ask. Thanks! Method #1 - Using the Jet 4.0 Provider When I try to run these statements to create the linked server: sp_dropserver 'Test', 'droplogins'; EXEC sp_addlinkedserver @server = N'Test', @provider = N'Microsoft.Jet.OLEDB.4.0', @srvproduct = N'Access DB', @datasrc = N'C:\Test.mdb' GO EXEC sp_addlinkedsrvlogin @rmtsrvname=N'Test', @useself=N'False',@locallogin=NULL, @rmtuser=N'Admin', @rmtpassword='####' GO I get this error when testing the connection: TITLE: Microsoft SQL Server Management Studio ------------------------------ "The test connection to the linked server failed." ------------------------------ ADDITIONAL INFORMATION: An exception occurred while executing a Transact-SQL statement or batch. (Microsoft.SqlServer.ConnectionInfo) ------------------------------ The OLE DB provider "Microsoft.Jet.OLEDB.4.0" for linked server "Test" reported an error. Authentication failed. Cannot initialize the data source object of OLE DB provider "Microsoft.Jet.OLEDB.4.0" for linked server "Test". OLE DB provider "Microsoft.Jet.OLEDB.4.0" for linked server "Test" returned message "Cannot start your application. The workgroup information file is missing or opened exclusively by another user.". (Microsoft SQL Server, Error: 7399) ------------------------------ Method #2 - Using the ODBC Provider... sp_dropserver 'Test', 'droplogins'; EXEC sp_addlinkedserver @server = N'Test', @provider = N'MSDASQL', @srvproduct = N'ODBC', @datasrc = N'Test:DSN' GO EXEC sp_addlinkedsrvlogin @rmtsrvname=N'Test', @useself=N'False',@locallogin=NULL, @rmtuser=N'Admin', @rmtpassword='####' GO I get this error: TITLE: Microsoft SQL Server Management Studio ------------------------------ "The test connection to the linked server failed." ------------------------------ ADDITIONAL INFORMATION: An exception occurred while executing a Transact-SQL statement or batch. (Microsoft.SqlServer.ConnectionInfo) ------------------------------ Cannot initialize the data source object of OLE DB provider "MSDASQL" for linked server "Test". OLE DB provider "MSDASQL" for linked server "Test" returned message "[Microsoft][ODBC Driver Manager] Driver's SQLSetConnectAttr failed". OLE DB provider "MSDASQL" for linked server "Test" returned message "[Microsoft][ODBC Driver Manager] Driver's SQLSetConnectAttr failed". OLE DB provider "MSDASQL" for linked server "Test" returned message "[Microsoft][ODBC Microsoft Access Driver] Cannot open database '(unknown)'. It may not be a database that your application recognizes, or the file may be corrupt.". (Microsoft SQL Server, Error: 7303)

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  • Bypass BIOS password set by faulty Toshiba firmware on Satellite A55 laptop?

    - by Brian
    How can the CMOS be cleared on the Toshiba Satellite A55-S1065? I have this 7 year old laptop that has been crippled by a glitch in its BIOS: 'A "Password =" prompt may be displayed when the computer is turned on, even though no power-on password has been set. If this happens, there is no password that will satisfy the password request. The computer will be unusable until this problem is resolved. [..] The occurrence of this problem on any particular computer is unpredictable -- it may never happen, but it could happen any time that the computer is turned on. [..] Toshiba will cover the cost of this repair under warranty until Dec 31, 2010.' -Toshiba As they stated, this machine is "unusable." The escape key does not bypass the prompt (nor does any other key), thus no operating system can be booted and no firmware updates can be installed. After doing some research, I found solutions that have been suggested for various Toshiba Satellite models afflicted by this glitch: "Make arrangements with a Toshiba Authorized Service Provider to have this problem resolved." -Toshiba (same link). Even prior to the expiration of Toshiba's support ("repair under warranty until Dec 31, 2010"), there have been reports that this solution is prohibitively expensive, labor charges accruing even when the laptop is still under warranty, and other reports that are generally discouraging: "They were unable to fix it and the guy who worked on it said he couldn’t find the jumpers on the motherboard to clear the BIOS. I paid $39 for my troubles and still have the password problem." - Steve. Since the costs of the repairs can now exceed the value of the hardware, it would seem this is a DIY solution, or a non-solution (i.e. the hardware is trash). Build a Toshiba parallel loopback by stripping and soldering the wires on a DB25 plug to connect connect these pins: 1-5-10, 2-11, 3-17, 4-12, 6-16, 7-13, 8-14, 9-15, 18-25. -CGSecurity. According to a list of supported models on pwcrack, this will likely not work for my Satellite A55-1065 (as well as many other models of similar age). -pwcrack Disconnect the laptop battery for an extended period of time. Doesn't work, laptop sat in a closet for several years without the battery connected and I forgot about the whole thing for awhile. The poor thing. Clear CMOS by setting the proper jumper setting or by removing the CMOS (RTC) battery, or by short circuiting a (hidden?) jumper that looks like a pair of solder marks -various sources for various Satellite models: Satellite A105: "you will see C88 clearly labeled right next the jack that the wireless card plugs into. There are two little solder squares (approx 1/16") at this location" -kerneltrap Satellite 1800: "Underneath the RAM there is black sticker, peel off the black sticker and you will reveal two little solder marks which are actually 'jumpers'. Very carefully hold a flat-head screwdriver touching both points and power on the unit briefly, effectively 'shorting' this circuit." -shadowfax2020 Satellite L300: "Short the B500 solder pads on the system board." -Lester Escobar Satellite A215: "Short the B500 solder pads on the system board." -fixya Clearing the CMOS could resolve the issue, but I cannot locate a jumper or a battery on this board. Nothing that looks remotely like a battery can be removed (everything is soldered). I have looked closely at the area around the memory and do not see any obvious solder pads that could be a secret jumper. Here are pictures (click for full resolution) : Where is the jumper (or solder pads) to short circuit and wipe the CMOS on this board? Possibly related questions: Remove Toshiba laptop BIOS password? Password Problem Toshiba Satellite..

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  • How can I switch an existing set of Subversion repositories to use ActiveDirectory?

    - by jpierson
    I have a set of private Subversion repositories on a Windows Server 2003 box which developers access via SVNServe over the svn:// protocol. Currently we have been using the authz and passwd files for each repository to control access however with the growing number of repositories and developers I'm considering switching to using their credentials from ActiveDirectory. We run in an all Microsoft shop and use IIS instead of Apache on all of our web servers so I would prefer to continue to use SVNServe if possible. Besides it being possible, I'm also concerned about how to migrate our repositories so that the history for the existing users map to the correct ActiveDirectory accounts. Keep in mind also that I'm not the network administrator and I'm not terrible familiar with ActiveDirectory so I'll probably have to go through some other people to get the changes made in ActiveDirectory if necessary. What are my options? UPDATE 1: It appears from the SVN documentation that by using SASL I should be able to get SVNServe to authenticate using ActiveDirectory. To clarify, the answer that I'm looking for is how to go about configuring SVNServe (if possible) to use ActiveDirectory for authentication and then how to modify an existing repository to remap existing svn users to their ActiveDirectory domain login accounts. UPDATE 2: It appears that the SASL support in SVNServe works off of a plugin model and the documentation only shows as an example. Looking at the Cyrus SASL Library it looks like a number of authentication "mechanisms" are supported but I'm not sure which one is to be used for ActiveDirectory support nor can I find any documentation about such matters. UPDATE 3: Ok, well it looks like in order to communication with ActiveDirectory I'm looking to use saslauthd instead of sasldb for the *auxprop_plugin* property. Unfortunately it appears that according to some posts (possibly outdated and inaccurate) saslauthd does not build on Windows and such endeavors are considered a work in progress. UPDATE 4: The lastest post I've found on this topic makes it sound as though the proper binaries () are available through the MIT Kerberos Library but it sounds like the author of this post on Nabble.com is still having issues getting things working. UPDATE 5: It looks like from the TortoiseSVN discussions and also this post on svn.haxx.se that even if saslgssapi.dll or whatever necessary binaries are available and configured on the Windows server that the clients will also need the same customization in order to work with these repositories. If this is true, we will only be able to get ActiveDirectory support from a windows client only if changes are made in these clients such as TortoiseSVN and CollabNet build of the client binaries to support such authentication schemes. Although thats what these posts suggest, this is contradictory from what I originally assumed from other reading in that being SASL compatible should require no changes on the client but instead only that the server be setup to handle the authentication mechanism. After reading a bit more carefully in the document about Cyrus SASL in Subversion section 5 states "1.5+ clients with Cyrus SASL support will be able to authenticate against 1.5+ servers with SASL enabled, provided at least one of the mechanisms supported by the server is also supported by the client." So clearly GSSAPI support (which I understand is required for Active Directory) must be available within the client and the server. I have to say, I'm learning way too much about the internals of how Subversion handles authentication than I ever wanted to and I juts simply want to get an answer about whether I can have Active Directory authentication support when using SVNServe on a Windows server and accessing this from Windows clients. According to the official documentation it seems that this is possible however you can see that the configuration is not trivial if even possible at all.

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  • broadcom 5722 NIC not installed on Ubuntu Server, although driver present

    - by Bastien
    Hello, I just installed Ubuntu Server 10.04 LTS, running kernel 2.6.32-24-server, on a brand new Dell T110 server, supposedly fully compatible with Ubuntu Server. I have two NICs: one ONBOARD, the other additional on PCI. both of them are Broadcom netXtreme 5572. on the first boot of the system, I could see both cards as eth0 and eth1 (with ifconfig) I configured eth0 as static IP (as planned), and did not configure eth1. after rebooting, one of the two NICs "disappeared": it does not appear in ifconfig at all. the one that disappeared is the ONBOARD one. I investigated a bit and found the following things: the card is SEEN, but not "installed", it appears as "UNCLAIMED" in lshw: *-network UNCLAIMED description: Ethernet controller product: NetXtreme BCM5722 Gigabit Ethernet PCI Express vendor: Broadcom Corporation physical id: 0 bus info: pci@0000:04:00.0 version: 00 width: 64 bits clock: 33MHz capabilities: pm vpd msi pciexpress cap_list configuration: latency=0 resources: memory:df9f0000-df9fffff *-network description: Ethernet interface product: NetXtreme BCM5722 Gigabit Ethernet PCI Express vendor: Broadcom Corporation physical id: 0 bus info: pci@0000:05:00.0 logical name: eth0 version: 00 serial: 00:10:18:60:23:64 size: 100MB/s capacity: 1GB/s width: 64 bits clock: 33MHz capabilities: pm vpd msi pciexpress bus_master cap_list ethernet physical tp 10bt 10bt-fd 100bt 100bt-fd 1000bt 1000bt-fd autonegotiation configuration: autonegotiation=on broadcast=yes driver=tg3 driverversion=3.102 duplex=full firmware=5722-v3.09 ip=10.129.167.25 latency=0 link=yes multicast=yes port=twisted pair speed=100MB/s resources: irq:35 memory:dfaf0000-dfafffff so I checked my dmesg and found a few strange lines, showing, there actually is a problem bringing up this card: [ 3.737506] tg3: Could not obtain valid ethernet address, aborting. [ 3.737527] tg3 0000:04:00.0: PCI INT A disabled [ 3.737535] tg3: probe of 0000:04:00.0 failed with error -22 [ 3.737553] alloc irq_desc for 17 on node -1 [ 3.737555] alloc kstat_irqs on node -1 [ 3.737560] tg3 0000:05:00.0: PCI INT A -> GSI 17 (level, low) -> IRQ 17 [ 3.737566] tg3 0000:05:00.0: setting latency timer to 64 [ 3.793529] eth0: Tigon3 [partno(BCM95722A2202G) rev a200] (PCI Express) MAC address 00:10:18:60:23:64 [ 3.793532] eth0: attached PHY is 5722/5756 (10/100/1000Base-T Ethernet) (WireSpeed[1]) [ 3.793534] eth0: RXcsums[1] LinkChgREG[0] MIirq[0] ASF[0] TSOcap[1] [ 3.793536] eth0: dma_rwctrl[76180000] dma_mask[64-bit] that actually shows that one NIC is recognized, the other is not. I researched a bit more, with lspci -v: 04:00.0 Ethernet controller: Broadcom Corporation NetXtreme BCM5722 Gigabit Ethernet PCI Express Subsystem: Broadcom Corporation NetXtreme BCM5722 Gigabit Ethernet PCI Express Flags: fast devsel, IRQ 16 Memory at df9f0000 (64-bit, non-prefetchable) [size=64K] Capabilities: [48] Power Management version 3 Capabilities: [50] Vital Product Data <?> Capabilities: [58] Vendor Specific Information <?> Capabilities: [e8] Message Signalled Interrupts: Mask- 64bit+ Queue=0/0 Enable- Capabilities: [d0] Express Endpoint, MSI 00 Capabilities: [100] Advanced Error Reporting <?> Capabilities: [13c] Virtual Channel <?> Capabilities: [160] Device Serial Number 00-00-00-fe-ff-00-00-00 Kernel modules: tg3 05:00.0 Ethernet controller: Broadcom Corporation NetXtreme BCM5722 Gigabit Ethernet PCI Express Subsystem: Broadcom Corporation NetXtreme BCM5722 Gigabit Ethernet PCI Express Flags: bus master, fast devsel, latency 0, IRQ 35 Memory at dfaf0000 (64-bit, non-prefetchable) [size=64K] Expansion ROM at <ignored> [disabled] Capabilities: [48] Power Management version 3 Capabilities: [50] Vital Product Data <?> Capabilities: [58] Vendor Specific Information <?> Capabilities: [e8] Message Signalled Interrupts: Mask- 64bit+ Queue=0/0 Enable+ Capabilities: [d0] Express Endpoint, MSI 00 Capabilities: [100] Advanced Error Reporting <?> Capabilities: [13c] Virtual Channel <?> Capabilities: [160] Device Serial Number 64-23-60-fe-ff-18-10-00 Capabilities: [16c] Power Budgeting <?> Kernel driver in use: tg3 Kernel modules: tg3 here I could see that the MAC address is 00-00-00-FE-FF-00-00-00, which, according to some forum posts on several websites, could be an issue. I've researched everything I could on the net, and found out several people having slightly comparable issues, but they usually involve different HW, and do not provide a proper explanation / solution... I would appreciate if anyone around here has some info to share ! thanks

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  • Network Restructure Method for Double-NAT network

    - by Adrian
    Due to a series of poor network design decisions (mostly) made many years ago in order to save a few bucks here and there, I have a network that is decidedly sub-optimally architected. I'm looking for suggestions to improve this less-than-pleasant situation. We're a non-profit with a Linux-based IT department and a limited budget. (Note: None of the Windows equipment we have runs does anything that talks to the Internet nor do we have any Windows admins on staff.) Key points: We have a main office and about 12 remote sites that essentially double NAT their subnets with physically-segregated switches. (No VLANing and limited ability to do so with current switches) These locations have a "DMZ" subnet that are NAT'd on an identically assigned 10.0.0/24 subnet at each site. These subnets cannot talk to DMZs at any other location because we don't route them anywhere except between server and adjacent "firewall". Some of these locations have multiple ISP connections (T1, Cable, and/or DSLs) that we manually route using IP Tools in Linux. These firewalls all run on the (10.0.0/24) network and are mostly "pro-sumer" grade firewalls (Linksys, Netgear, etc.) or ISP-provided DSL modems. Connecting these firewalls (via simple unmanaged switches) is one or more servers that must be publically-accessible. Connected to the main office's 10.0.0/24 subnet are servers for email, tele-commuter VPN, remote office VPN server, primary router to the internal 192.168/24 subnets. These have to be access from specific ISP connections based on traffic type and connection source. All our routing is done manually or with OpenVPN route statements Inter-office traffic goes through the OpenVPN service in the main 'Router' server which has it's own NAT'ing involved. Remote sites only have one server installed at each site and cannot afford multiple servers due to budget constraints. These servers are all LTSP servers several 5-20 terminals. The 192.168.2/24 and 192.168.3/24 subnets are mostly but NOT entirely on Cisco 2960 switches that can do VLAN. The remainder are DLink DGS-1248 switches that I am not sure I trust well enough to use with VLANs. There is also some remaining internal concern about VLANs since only the senior networking staff person understands how it works. All regular internet traffic goes through the CentOS 5 router server which in turns NATs the 192.168/24 subnets to the 10.0.0.0/24 subnets according to the manually-configured routing rules that we use to point outbound traffic to the proper internet connection based on '-host' routing statements. I want to simplify this and ready All Of The Things for ESXi virtualization, including these public-facing services. Is there a no- or low-cost solution that would get rid of the Double-NAT and restore a little sanity to this mess so that my future replacement doesn't hunt me down? Basic Diagram for the main office: These are my goals: Public-facing Servers with interfaces on that middle 10.0.0/24 network to be moved in to 192.168.2/24 subnet on ESXi servers. Get rid of the double NAT and get our entire network on one single subnet. My understanding is that this is something we'll need to do under IPv6 anyway, but I think this mess is standing in the way.

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  • Issue in setting up VPN connection (IKEv1) using android (ICS vpn client) with Strongswan 4.5.0 server

    - by Kushagra Bhatnagar
    I am facing issues in setting up VPN connection(IKEv1) using android (ICS vpn client) and Strongswan 4.5.0 server. Below is the set up: Strongswan server is running on ubuntu linux machine which is connected to some wifi hotspot. Using the steps in this guide link, I generated CA, server and client certificate. Once certificates are generated, following (clientCert.p12 and caCert.pem) are sent to mobile via mail and installed on android device. Below are the ip addresses assigned to various interfaces Linux server wlan0 interface ip where server is running: 192.168.43.212, android device eth0 interface ip address: 192.168.43.62; Android device is also attached with the same wifi hotspot. On the Android device, I uses IPsec Xauth RSA option for setting up VPN authentication configuration. I am using the following ipsec.conf configuration: # basic configuration config setup plutodebug=all # crlcheckinterval=600 # strictcrlpolicy=yes # cachecrls=yes nat_traversal=yes # charonstart=yes plutostart=yes # Add connections here. # Sample VPN connections conn ios1 keyexchange=ikev1 authby=xauthrsasig xauth=server left=%defaultroute leftsubnet=0.0.0.0/0 leftfirewall=yes leftcert=serverCert.pem right=192.168.43.62 rightsubnet=10.0.0.0/24 rightsourceip=10.0.0.2 rightcert=clientCert.pem pfs=no auto=add      With the above configurations when I enable VPN on android device, VPN connection is not successful and it gets timed out in Authentication phase. I ran wireshark on both the android device and strongswan server, from the tcpdump below are the observations. Initially Identity Protection (Main mode) exchanges happens between device and server and all are successful. After all successful Identity Protection (Main mode) exchanges server is sending Transaction (Config mode) to device. In reply android device is sending Informational message instead of Transaction (Config mode) message. Further server is keep on sending Transaction (Config mode) message and device is again sending Identity Protection (Main mode) messages. Finally timeout happens and connection fails. I also capture Strongswan server logs and below are the snippets from the server logs which also verifies the same(described above). Apr 27 21:09:40 Linux pluto[12105]: | **parse ISAKMP Message: Apr 27 21:09:40 Linux pluto[12105]: | initiator cookie: Apr 27 21:09:40 Linux pluto[12105]: | 06 fd 61 b8 86 82 df ed Apr 27 21:09:40 Linux pluto[12105]: | responder cookie: Apr 27 21:09:40 Linux pluto[12105]: | 73 7a af 76 74 f0 39 8b Apr 27 21:09:40 Linux pluto[12105]: | next payload type: ISAKMP_NEXT_HASH Apr 27 21:09:40 Linux pluto[12105]: | ISAKMP version: ISAKMP Version 1.0 Apr 27 21:09:40 Linux pluto[12105]: | exchange type: ISAKMP_XCHG_INFO Apr 27 21:09:40 Linux pluto[12105]: | flags: ISAKMP_FLAG_ENCRYPTION Apr 27 21:09:40 Linux pluto[12105]: | message ID: a2 80 ad 82 Apr 27 21:09:40 Linux pluto[12105]: | length: 92 Apr 27 21:09:40 Linux pluto[12105]: | ICOOKIE: 06 fd 61 b8 86 82 df ed Apr 27 21:09:40 Linux pluto[12105]: | RCOOKIE: 73 7a af 76 74 f0 39 8b Apr 27 21:09:40 Linux pluto[12105]: | peer: c0 a8 2b 3e Apr 27 21:09:40 Linux pluto[12105]: | state hash entry 25 Apr 27 21:09:40 Linux pluto[12105]: | state object not found Apr 27 21:09:40 Linux pluto[12105]: packet from 192.168.43.62:500: Informational Exchange is for an unknown (expired?) SA Apr 27 21:09:40 Linux pluto[12105]: | next event EVENT_RETRANSMIT in 10 seconds for #9 Can anyone please provide update on this issue. Why the VPN connection gets timed out and why the ISAKMP exchanges are not proper between Android and strongswan server.

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  • Postfix sasl login failing no mechanism found

    - by Nat45928
    following the link here: http://flurdy.com/docs/postfix/ with posfix, courier, MySql, and sasl gave me a web server that has imap functionality working fine but when i go to log into the server to send a message using the same user id and password for connecting the the imap server it rejects my login to the smtp server. If i do not specify a login for the outgoing mail server then it will send the message just fine. the error in postfix's log is: Jul 6 17:26:10 Sj-Linux postfix/smtpd[19139]: connect from unknown[10.0.0.50] Jul 6 17:26:10 Sj-Linux postfix/smtpd[19139]: warning: SASL authentication failure: unable to canonify user and get auxprops Jul 6 17:26:10 Sj-Linux postfix/smtpd[19139]: warning: unknown[10.0.0.50]: SASL DIGEST-MD5 authentication failed: no mechanism available Jul 6 17:26:10 Sj-Linux postfix/smtpd[19139]: warning: unknown[10.0.0.50]: SASL LOGIN authentication failed: no mechanism available Ive checked all usernames and passwords for mysql. what could be going wrong? edit: here is some other information: installed libraires for postfix, courier and sasl: aptitude install postfix postfix-mysql aptitude install libsasl2-modules libsasl2-modules-sql libgsasl7 libauthen-sasl-cyrus-perl sasl2-bin libpam-mysql aptitude install courier-base courier-authdaemon courier-authlib-mysql courier-imap courier-imap-ssl courier-ssl and here is my /etc/postfix/main.cf myorigin = domain.com smtpd_banner = $myhostname ESMTP $mail_name biff = no # appending .domain is the MUA's job. append_dot_mydomain = no # Uncomment the next line to generate "delayed mail" warnings #delay_warning_time = 4h readme_directory = no # TLS parameters smtpd_tls_cert_file=/etc/ssl/certs/ssl-cert-snakeoil.pem smtpd_tls_key_file=/etc/ssl/private/ssl-cert-snakeoil.key smtpd_use_tls=yes smtpd_tls_session_cache_database = btree:${data_directory}/smtpd_scache smtp_tls_session_cache_database = btree:${data_directory}/smtp_scache # See /usr/share/doc/postfix/TLS_README.gz in the postfix-doc package for # information on enabling SSL in the smtp client. #myhostname = my hostname alias_maps = hash:/etc/aliases alias_database = hash:/etc/aliases myorigin = /etc/mailname local_recipient_maps = mydestination = relayhost = mynetworks = 127.0.0.0/8 [::ffff:127.0.0.0]/104 [::1]/128 mailbox_size_limit = 0 recipient_delimiter = + inet_interfaces = all mynetworks_style = host # how long if undelivered before sending warning update to sender delay_warning_time = 4h # will it be a permanent error or temporary unknown_local_recipient_reject_code = 450 # how long to keep message on queue before return as failed. # some have 3 days, I have 16 days as I am backup server for some people # whom go on holiday with their server switched off. maximal_queue_lifetime = 7d # max and min time in seconds between retries if connection failed minimal_backoff_time = 1000s maximal_backoff_time = 8000s # how long to wait when servers connect before receiving rest of data smtp_helo_timeout = 60s # how many address can be used in one message. # effective stopper to mass spammers, accidental copy in whole address list # but may restrict intentional mail shots. # but may restrict intentional mail shots. smtpd_recipient_limit = 16 # how many error before back off. smtpd_soft_error_limit = 3 # how many max errors before blocking it. smtpd_hard_error_limit = 12 # Requirements for the HELO statement smtpd_helo_restrictions = permit_mynetworks, permit # Requirements for the sender details smtpd_sender_restrictions = permit_sasl_authenticated, permit_mynetworks, warn_if_reject reject_non_fqdn_sender, reject_unknown_sender_domain, reject_unauth_pipelining, permit # Requirements for the connecting server smtpd_client_restrictions = reject_rbl_client sbl.spamhaus.org, reject_rbl_client blackholes.easynet.nl, reject_rbl_client dnsbl.njabl.org # Requirement for the recipient address smtpd_recipient_restrictions = reject_unauth_pipelining, permit_mynetworks, permit_sasl_authenticated, reject_non_fqdn_recipient, reject_unknown_recipient_domain, reject_unauth_destination, permit smtpd_data_restrictions = reject_unauth_pipelining # require proper helo at connections smtpd_helo_required = yes # waste spammers time before rejecting them smtpd_delay_reject = yes disable_vrfy_command = yes # not sure of the difference of the next two # but they are needed for local aliasing alias_maps = hash:/etc/postfix/aliases alias_database = hash:/etc/postfix/aliases # this specifies where the virtual mailbox folders will be located virtual_mailbox_base = /var/spool/mail/virtual # this is for the mailbox location for each user virtual_mailbox_maps = mysql:/etc/postfix/mysql_mailbox.cf # and this is for aliases virtual_alias_maps = mysql:/etc/postfix/mysql_alias.cf # and this is for domain lookups virtual_mailbox_domains = mysql:/etc/postfix/mysql_domains.cf # this is how to connect to the domains (all virtual, but the option is there) # not used yet # transport_maps = mysql:/etc/postfix/mysql_transport.cf virtual_uid_maps = static:5000 virtual_gid_maps = static:5000 # SASL smtpd_sasl_auth_enable = yes # If your potential clients use Outlook Express or other older clients # this needs to be set to yes broken_sasl_auth_clients = yes smtpd_sasl_security_options = noanonymous smtpd_sasl_local_domain =

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  • Since upgrading to Solaris 11, my ARC size has consistently targeted 119MB, despite having 30GB RAM. What? Why?

    - by growse
    I ran a NAS/SAN box on Solaris 11 Express before Solaris 11 was released. The box is an HP X1600 with an attached D2700. In all, 12x 1TB 7200 SATA disks, 12x 300GB 10k SAS disks in separate zpools. Total RAM is 30GB. Services provided are CIFS, NFS and iSCSI. All was well, and I had a ZFS memory usage graph looking like this: A fairly healthy Arc size of around 23GB - making use of the available memory for caching. However, I then upgraded to Solaris 11 when that came out. Now, my graph looks like this: Partial output of arc_summary.pl is: System Memory: Physical RAM: 30701 MB Free Memory : 26719 MB LotsFree: 479 MB ZFS Tunables (/etc/system): ARC Size: Current Size: 915 MB (arcsize) Target Size (Adaptive): 119 MB (c) Min Size (Hard Limit): 64 MB (zfs_arc_min) Max Size (Hard Limit): 29677 MB (zfs_arc_max) It's targetting 119MB while sitting at 915MB. It's got 30GB to play with. Why? Did they change something? Edit To clarify, arc_summary.pl is Ben Rockwood's, and the relevent lines generating the above stats are: my $mru_size = ${Kstat}->{zfs}->{0}->{arcstats}->{p}; my $target_size = ${Kstat}->{zfs}->{0}->{arcstats}->{c}; my $arc_min_size = ${Kstat}->{zfs}->{0}->{arcstats}->{c_min}; my $arc_max_size = ${Kstat}->{zfs}->{0}->{arcstats}->{c_max}; my $arc_size = ${Kstat}->{zfs}->{0}->{arcstats}->{size}; The Kstat entries are there, I'm just getting odd values out of them. Edit 2 I've just re-measured the arc size with arc_summary.pl - I've verified these numbers with kstat: System Memory: Physical RAM: 30701 MB Free Memory : 26697 MB LotsFree: 479 MB ZFS Tunables (/etc/system): ARC Size: Current Size: 744 MB (arcsize) Target Size (Adaptive): 119 MB (c) Min Size (Hard Limit): 64 MB (zfs_arc_min) Max Size (Hard Limit): 29677 MB (zfs_arc_max) The thing that strikes me is that the Target Size is 119MB. Looking at the graph, it's targeted the exact same value (124.91M according to cacti, 119M according to arc_summary.pl - think the difference is just 1024/1000 rounding issues) ever since Solaris 11 was installed. It looks like the kernel's making zero effort to shift the target size to anything different. The current size is fluctuating as the needs of the system (large) fight with the target size, and it appears equilibrium is between 700 and 1000MB. So the question is now a little more pointed - why is Solaris 11 hard setting my ARC target size to 119MB, and how do I change it? Should I raise the min size to see what happens? I've stuck the output of kstat -n arcstats over at http://pastebin.com/WHPimhfg Edit 3 Ok, weirdness now. I know flibflob mentioned that there was a patch to fix this. I haven't applied this patch yet (still sorting out internal support issues) and I've not applied any other software updates. Last thursday, the box crashed. As in, completely stopped responding to everything. When I rebooted it, it came back up fine, but here's what my graph now looks like. It seems to have fixed the problem. This is proper la la land stuff now. I've literally no idea what's going on. :(

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  • Yahoo is sending our server's transactional email to the Spam folder, even though we have set up SPF and DKIM

    - by Derrick Miller
    Yahoo Mail is sending our server's transactional emails to the Spam folder, even though we have taken quite a few anti-spam steps. By contrast, Gmail allows the messages through to the inbox just fine. Here are the things which are in place: SPF is set up for the domain holsteinplaza.com. Yahoo reports spf=pass in the message headers. DKIM is set up for the domain holsteinplaza.com. Yahoo reports dkim=pass in the message headers. We have a proper reverse DNS entry for the sending mail server. Name - IP matches IP - Name. Neither Domainkeys nor SenderID are set up. From what I can tell, DKIM is the way of the future, and there is not much to be gained from adding Domainkeys or SenderID. Following are the headers. Any ideas what more I should do to get Yahoo to stop flagging the emails as spam? From Holstein Plaza Auctions Sat Jun 25 18:30:08 2011 X-Apparently-To: [email protected] via 98.138.90.132; Sat, 25 Jun 2011 18:30:11 -0700 Return-Path: <[email protected]> X-YahooFilteredBulk: 70.32.113.42 Received-SPF: pass (domain of holsteinplaza.com designates 70.32.113.42 as permitted sender) X-YMailISG: i_vaA_QWLDuLOmXhDjUv3aBKJl5Un6EiP6Yk2m4yn3jeEuYK MkhpqIt9zDUbHARCwXrhl9pqjTANurGVca7gytSs.mryWVQcbWBx.DaItWRb VcyrIzwMzXKCSeu06H2a.cJ7HG5vJLJaKmHUUI_1ttXKn_Aegiu5yHvFX83R Lpth0witO9zfaKvOMaJV3LAxpIpFOydwvq1cqjZ8nURxQbxM3Cl.QW7MxxrC 09qLVn_D_xSdU94QdU22IsVmlaRHv.uU5dnIazu.KSkhKpYykDoZA2SH0SY4 JmTZj3LP8N926xXVDzYQ5K6QvKuJL5g0d9pYZx3KC59sgIu5oHlJ3Q15RdKb f3OJw0PR6oIyJ2yStVr8vfbDgOfj3qig03.Tw6g6MMNpv1G7Cuol4oJeUaYP xELxX6dHgBgCSuWMcbsrxbK4BIXcS2qhpMqYQ4Isk.XXyA8uvmFXyvgc1ds5 8jo0rW.Wsw.55Z.KTPaQ0gHXj0T3OGppYMELSJv1iuhPyyAnZpmq01CU0Qd5 CcRgdyW3HaqhmpXqJCS0Clo16zXA4HmAjR0tgIQrHRLc3D9N02AOzvmDgCb1 vCh0p00QeKVq8UNkcShPRxZFKi9khtkLhPBlXEKkhJ76zyDmHUxTY.dQHVVD 8D2hx7BxbqI9DINI8x5oR5Q8hYkZqHYQsmGNkaU77O2BnsEv5WxMEmzrBJ4Z h8zGCidgYPiZycZfnfaBp0Xb4tya2WMTN45W02JFcO1qq_UMJ9xPeqZhPEj. j9YvBAC8324GGF.c8eWcNB2VB34QHgTcVUl3.c0XUCuncls9Cyg4L7AoIdCi HvAklSzDDu9nW6732VEipV9FJ_JkDupDNQU2hfiPG.3OeF8GwTnVYnEn0EiZ aO0NCnZhXuLDcN3K7ml3846yRdASvzPFs9s4aJkzR0FkhVvptiMBEOdRkKdG wHWmvWpK4GTZpW4yU7CnKpW2MiWWn1MP0h_CCZFKs5.3mfmfPjPVIABN_RuU Q8ex5hdKnKlQiqK56LzcPRnYmNtrwdsUX9CYn9d6cPpXR_Bi5jrNJMNzdFvq lGO0CBT4QPe2V45U8PtpMitttuDA1cCvmyBPFswxNlL0jyX0a_W.vl0YW5.d HhDItpHhDxKRUscM28IR.exetq4QCzyM X-Originating-IP: [70.32.113.42] Authentication-Results: mta1267.mail.ac4.yahoo.com from=holsteinplaza.com; domainkeys=neutral (no sig); from=holsteinplaza.com; dkim=pass (ok) Received: from 127.0.0.1 (EHLO predator.axis80.com) (70.32.113.42) by mta1267.mail.ac4.yahoo.com with SMTP; Sat, 25 Jun 2011 18:30:11 -0700 Received: (qmail 1440 invoked by uid 48); 25 Jun 2011 21:30:09 -0400 To: [email protected] Subject: this is a test X-PHPMAILER-DKIM: phpmailer.worxware.com DKIM-Signature: v=1; a=rsa-sha1; q=dns/txt; l=203; s=auction; t=1309051808; c=relaxed/simple; h=From:To:Subject; d=holsteinplaza.com; [email protected]; z=From:=20Holstein=20Plaza=20Auctions=20<[email protected]> |To:[email protected] |Subject:=20this=20is=20a=20test; bh=B3Tw5AQb1va627KEoazuFEBZ0fg=; b=oQ5uFq+oekPTGhszyIritjuuIAi3qPNyeitu+aWMhdx3oC6O2j5hJsDFpK0sS5fms7QdnBkBcEzT0iekEvn9EfAdCkGZ2KrtEC0yv7QKQcrjXxy07GJpj9nq0LYbgOuPdw8mGvKxlRZ+jFBX0DRJm0xXFLkr+MEaILw7adHTCCM= Date: Sat, 25 Jun 2011 21:30:08 -0400 From: Holstein Plaza Auctions <[email protected]> Reply-to: Holstein Plaza Auctions <[email protected]> Message-ID: <[email protected]> X-Priority: 3 X-Mailer: PHPMailer 5.1 (phpmailer.sourceforge.net) MIME-Version: 1.0 Content-Transfer-Encoding: 8bit Content-Type: text/plain; charset="iso-8859-1" Content-Length: 195

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  • Forwarding rsyslog to syslog-ng, with FQDN and facility separation

    - by Joshua Miller
    I'm attempting to configure my rsyslog clients to forward messages to my syslog-ng log repository systems. Forwarding messages works "out of the box", but my clients are logging short names, not FQDNs. As a result the messages on the syslog repo use short names as well, which is a problem because one can't determine which system the message originated from easily. My clients get their names through DHCP / DNS. I've tried a number of solutions trying to get this working, but without success. I'm using rsyslog 4.6.2 and syslog-ng 3.2.5. I've tried setting $PreserveFQDN on as the first directive in /etc/rsyslog.conf (and restarting rsyslog of course). It seems to have no effect. hostname --fqdn on the client returns the proper FQDN, so the problem isn't whether the system can actually figure out its own FQDN. $LocalHostName <fqdn> looked promising, but this directive isn't available in my version of rsyslog (Available since 4.7.4+, 5.7.3+, 6.1.3+). Upgrading isn't an option at the moment. Configuring the syslog-ng server to populate names based on reverse lookups via DNS isn't an option. There are complexities with reverse DNS and the public cloud. Specifying for the forwarder to use a custom template seems like a viable option at first glance. I can specify the following, which causes local logging to begin using the FQDN on the syslog-ng repo. $template MyTemplate, "%timestamp% <FQDN> %syslogtag%%msg%" $ActionForwardDefaultTemplate MyTemplate However, when I put this in place syslog-ng seems to be unable to categorize messages by facility or priority. Messages come in as FQDN, but everything is put in to user.log. When I don't use the custom template, messages are properly categorized under facility and priority, but with the short name. So, in summary, if I manually trick rsyslog into including the FQDN, priority and facility becomes lost details to syslog-ng. How can I get rsyslog to do FQDN logging which works properly going to a syslog-ng repository? rsyslog client config: $ModLoad imuxsock.so # provides support for local system logging (e.g. via logger command) $ModLoad imklog.so # provides kernel logging support (previously done by rklogd) $ActionFileDefaultTemplate RSYSLOG_TraditionalFileFormat *.info;mail.none;authpriv.none;cron.none /var/log/messages authpriv.* /var/log/secure mail.* -/var/log/maillog cron.* /var/log/cron *.emerg * uucp,news.crit /var/log/spooler local7.* /var/log/boot.log $WorkDirectory /var/spool/rsyslog # where to place spool files $ActionQueueFileName fwdRule1 # unique name prefix for spool files $ActionQueueMaxDiskSpace 1g # 1gb space limit (use as much as possible) $ActionQueueSaveOnShutdown on # save messages to disk on shutdown $ActionQueueType LinkedList # run asynchronously $ActionResumeRetryCount -1 # infinite retries if host is down *.* @syslog-ng1.example.com *.* @syslog-ng2.example.com syslog-ng configuration (abridged for brevity): options { flush_lines (0); time_reopen (10); log_fifo_size (1000); long_hostnames (off); use_dns (no); use_fqdn (yes); create_dirs (no); keep_hostname (yes); }; source src { unix-stream("/dev/log"); internal(); udp(ip(0.0.0.0) port(514)); }; destination per_host_destination { file( "/var/log/syslog-ng/devices/$HOST/$FACILITY.log" owner("root") group("root") perm(0644) dir_owner(root) dir_group(root) dir_perm(0775) create_dirs(yes)); }; log { source(src); destination(per_facility_destination); };

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  • Tomcat and ASP site under IIS6 with SSL

    - by Rafe
    I've been working on migrating our companies' website from it's original server to a new one and am having two different but possibly related problems. The box this is sitting on is a Windows 2003 server x64 running IIS 6. The Tomcat version is 5.5.x as it was the version the original deployment was built on. There are two other sites on the server one in plain HTML, another in PHP and the one I am trying to migrate is a combination of Java and ASP (the introductory/sign in pages being Java as well as many reports used for our clients and the administration pages being in ASP) First of all I can only access the site if I enter the ip followed by :8080 (xxx.xxx.xxx.xxx:8080). The original setup had an index.html file in the root of the site with a bit of javascript in the header that pointed the site to 'www.mysite.com/app/public' but if I try going directly to the site without the 8080 I get a 'page not found error' and the javascript redirector causes the same problem because it doesn't add the 8080 into the URL even though on the original site the 8080 wasn't present so I don't understand why it would need it now. The js redirect looks like this: <script language="JavaScript"> <!-- location.href = "/app/public/" location.replace("/app/public/"); //--> </script> When setting the site up I used the command line to unbind IIS from all of the ip's on the system (there are 12 ip's on this box) because I was led to believe Tomcat wanted to use localhost which wasn't accessible. I'm not sure if this was the right thing to do but I'm throwing it in for the sake of completeness. And actually, at this point trying to go to localhost from the server itself throws up a 'could not connect to localhost' error. If I go to localhost:8080 I get the tomcat welcome page. If I do localhost:8080/app/public I get the intro page to our website. So I'm not sure what I'm even looking at in this case, that is what the proper behavior should be. The second part of the problem is that if I do go to either the ip or localhost such as above (localhost:8080/app/public) and click on our login link it is supposed to transfer me to our login page yet instead I receive a 'could not connect' error and the url has changed to localhost:8443/app/secure. From my research I see that port 8443 is Tomcats SSL port and the server.xml alludes to it as follows: <Connector port="8080" maxHttpHeaderSize="8192" maxThreads="150" minSpareThreads="25" maxSpareThreads="75" enableLookups="false" redirectPort="8443" acceptCount="100" connectionTimeout="20000" disableUploadTimeout="true" /> I have an SSL certificate assigned to the site via IIS and was under the impression that by default Tomcat allowed IIS to handle secure connections but apparently something is munged because it's not working. There is another section in the server.xml that reads like this: <Connector port="8009" enableLookups="false" redirectPort="443" protocol="AJP/1.3" /> Which I'm not sure what it is for although port 443 is the SSL port that IIS uses so I'm confused as to what this is supposed to be doing. Another question I have is when does the isap_redirector actually come into play? How does it know when to try and serve pages through Tomcat and when not to? I've hunted around the 'net for an answer and have yet to find a clear dialogue on the subject. Anyone have any pointers as to where to look for a solution to all of this?

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  • Reconstructing the disk order in RAID 6 with 7 disks

    - by rkotulla
    a little background to this question first: I am running a RAID-6 within a QNAP TS869L external RAID/NAS system. I started with 5 disks of 3 TB each back in the day, and later added another 2 disks of 3TB to the RAID. The QNAP internals handled the growing and re-syncing etc, and everything seemd to be perfectly fine. About 2 weeks ago, I had one of the disks (disk #5, disk #2 has gone bad in the mean time) fail, and somehow (I have no idea why), also disks 1 and 2 got kicked out of the array. I replaced disk #5, but the RAID didn't start working again. After some calls to QNAP technical support, they re-created the array (using mdadm --create --force --assume-clean ...), but the resulting array couldn't find a filesystem, and I was kindly referred to contact a data recovery company that I can't afford. After some digging through old log files, resetting the disk to factory default, etc, I found a few errors that were made during this re-create - I wish I still had some of the original metadata, but unfortunately i don't (I definitely learned that lesson). I'm currently at the point where I know the correct chunk-size (64K), metadata-version (1.0; factory default was 0.9, but from what I read 0.9 doesn't handle disks over 2 TB, mine are 3 TB), and I now find the ext4 filesystem that should be on the disks. Only variable left to determine is the right disk order! I started using the description found in answer #4 of "Recover RAID 5 data after created new array instead of re-using" but am a little confused on what the order should be for a proper RAID-6. RAID-5 is pretty well documented in a number of places, but RAID-6 much less so. Also, does the layout, i.e. distribution of parity and data chunks across the disks, change after the growing of the array from 5 to 7 disks, or does the re-sync re-organize them in such a way a native 7-disk RAID-6 would have been? Thanks some more mdadm output that might be helpful: mdadm version: [~] # mdadm --version mdadm - v2.6.3 - 20th August 2007 mdadm details from one of the disks in the array: [~] # mdadm --examine /dev/sda3 /dev/sda3: Magic : a92b4efc Version : 1.0 Feature Map : 0x0 Array UUID : 1c1614a5:e3be2fbb:4af01271:947fe3aa Name : 0 Creation Time : Tue Jun 10 10:27:58 2014 Raid Level : raid6 Raid Devices : 7 Used Dev Size : 5857395112 (2793.02 GiB 2998.99 GB) Array Size : 29286975360 (13965.12 GiB 14994.93 GB) Used Size : 5857395072 (2793.02 GiB 2998.99 GB) Super Offset : 5857395368 sectors State : clean Device UUID : 7c572d8f:20c12727:7e88c888:c2c357af Update Time : Tue Jun 10 13:01:06 2014 Checksum : d275c82d - correct Events : 7036 Chunk Size : 64K Array Slot : 0 (0, 1, failed, 3, failed, 5, 6) Array State : Uu_u_uu 2 failed mdadm details for the array in the current disk-order (based on my best guess reconstructed from old log-files) [~] # mdadm --detail /dev/md0 /dev/md0: Version : 01.00.03 Creation Time : Tue Jun 10 10:27:58 2014 Raid Level : raid6 Array Size : 14643487680 (13965.12 GiB 14994.93 GB) Used Dev Size : 2928697536 (2793.02 GiB 2998.99 GB) Raid Devices : 7 Total Devices : 5 Preferred Minor : 0 Persistence : Superblock is persistent Update Time : Tue Jun 10 13:01:06 2014 State : clean, degraded Active Devices : 5 Working Devices : 5 Failed Devices : 0 Spare Devices : 0 Chunk Size : 64K Name : 0 UUID : 1c1614a5:e3be2fbb:4af01271:947fe3aa Events : 7036 Number Major Minor RaidDevice State 0 8 3 0 active sync /dev/sda3 1 8 19 1 active sync /dev/sdb3 2 0 0 2 removed 3 8 51 3 active sync /dev/sdd3 4 0 0 4 removed 5 8 99 5 active sync /dev/sdg3 6 8 83 6 active sync /dev/sdf3 output from /proc/mdstat (md8, md9, and md13 are internally used RAIDs holding swap, etc; the one I'm after is md0) [~] # more /proc/mdstat Personalities : [linear] [raid0] [raid1] [raid10] [raid6] [raid5] [raid4] [multipath] md0 : active raid6 sdf3[6] sdg3[5] sdd3[3] sdb3[1] sda3[0] 14643487680 blocks super 1.0 level 6, 64k chunk, algorithm 2 [7/5] [UU_U_UU] md8 : active raid1 sdg2[2](S) sdf2[3](S) sdd2[4](S) sdc2[5](S) sdb2[6](S) sda2[1] sde2[0] 530048 blocks [2/2] [UU] md13 : active raid1 sdg4[3] sdf4[4] sde4[5] sdd4[6] sdc4[2] sdb4[1] sda4[0] 458880 blocks [8/7] [UUUUUUU_] bitmap: 21/57 pages [84KB], 4KB chunk md9 : active raid1 sdg1[6] sdf1[5] sde1[4] sdd1[3] sdc1[2] sda1[0] sdb1[1] 530048 blocks [8/7] [UUUUUUU_] bitmap: 37/65 pages [148KB], 4KB chunk unused devices: <none>

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  • How to stop windows resizing when the monitor display channel is turned off / switched to different source

    - by Heartspeace
    Have a new 6870 ati radeon adapter with its drivers set to 1080p 60hz resolution hooked up to a 2008 47" high end Samsung HDMI based TV. However, when the tv is turned to a different hdmi input -(when I come back into windows) somehow Windows decides to resize all the open apps to a lower resolution - including some of the side docked hidden pop-outs. When it resizes those though - it just sticked the pop-outs in the middle of the screen and all the resized windows from the open applications in the top left corner - all of them stacked on top of each other and resized to the smaller resolution. The things that seem to be ok after returning are the icons on the desktop, the taskbar, and the sidebar. Anyone have any knowledge of 1) how this happens 2) why it happens 3) how to stop it from resizing the applications and some of the docked pop-outs (they are not really resized after returning - they are just stuck in the middle of the screen approximately where they would be if the right or bottom sidebar should be if the screen was resized to that lower resolution). My hypothesis is that upon losing HDMI signal - that Windows is told by something (driver, or windows itself) that the resolution to be without a signal being present (noting that HDMI signals and handshakes are two way on HDMI devices. If it loses the signal or the tv is switched to another device - then the display adapter must figure that out and tell Windows or figures it out and designs randomly to change the display size). Any and all help is most appreciated. I asked AMD/ATI - but they said they don't know why or how this is happening. I was hoping that maybe this is THE place that the super users truly go to - those that develop display adapter drivers, or that dive deeply into these areas of windows. If there is better sites or just competing sites - please advise - noting I have already written AMD/ATI. HP Response / Additions 4/7/2011 It is really nice to get your reply Shinrai. (BTW is it proper etiquette on these forums to have a discussion?) Yet 'only one issue' - I am using a single display in this case - so Windows doesn't move application windows to another desktop. Windows (or something) decides to shrink the desktop it currently has and resize all windows to the maximum size of the desktop. As such I would be glad if Windows would just keep the current size of the one desktop that is in operation. I also know that this does NOT happen on monitors connected with DVI. There I have had one and two monitors setup and it doesn't resize those screens at all when disconnecting monitors, turning them off, whatever... they stay solid - everything in place - to such an extent that if you forgot the other monitor is off - you will have troubles finding some windows without using one of the control app utilities. So if I could even get the HDMI handling by Windows (or the display driver) ( 1] which is doing this anyway the display driver or Windows - and 2] where is that other resolution size (1024x768) coming from - its not the smallest and its not the largest?) to be having like DVI - Life would be golden (for this aspect anyway). ** found others with same problem in this thread: http://hardforum.com/showthread.php?t=1507324 Thanks, HP

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  • Specifying a Postfix Instance to send outbound email

    - by Catherine Jefferson
    I have a CentOS 6.5 server running Postfix 2.6x (the default distribution) with five public IPv4 IPs bound to it. Each IP has DNS and rDNS set separately. Each uses a different hostname at a different domain. I have five Postfix instances, one bound to each IP, like this example: 192.168.34.104 red.example.com /etc/postfix 192.168.36.48 green.example.net /etc/postfix-green 192.168.36.49 pink.example.org /etc/postfix-pink 192.168.36.50 orange.example.info /etc/postfix-orange 192.168.36.51 blue.example.us /etc/postfix-blue I've tested each IP by telneting to port 25. Postfix answers and banners properly with the correct hostname. Email is received on all of these instances with no problems and is routed to the correct place. This setup, minus the final instance, has existed for a couple of years and works. I never bothered to set up outbound email to go through any but the main instance, however; there was no need. Now I need to send email from blue.example.us that actually leaves from that interface and IP, such that the Received headers show blue.example.us as the sending mailhost, so that SPF and DKIM validate, etc etc. The email that will be sent from blue.example.com is a feedback loop sent by a single shell account on the server (account5), an account that is dedicated to sending this email. The account receives the feedback loop emails from servers on other networks, saves the bodies of those emails, and then generates a new outbound email header, appends the saved body, and sends the email. It's sending by piping each email to sendmail -oi -t. We're doing it this way to mask the identities of the initial servers. The procmail script that processes these emails works correctly. However, I cannot configure this account to send email through the proper Postfix instance/IP/interface. The exact same account and script sends email through the main Postfix instance /etc/postfix without any issues. When I change MAIL_CONFIG to point to /etc/postfix-blue in either .bash_profile or the Procmail script that handles this email, though, I get this error: sendmail: fatal: User account5(###) is not allowed to submit mail I've read the manuals on Postfix.org, searched Google, and tried the suggestions in three previous answers here on ServerFault.com: Postfix - specify interface to deliver outbound mail on Postfix user is not allowed to submit mail Postfix rejects php mails I have been careful to stop and restart Postfix after each configuration change, and tested the results. Nothing has worked. The main postfix instance happily accepts outbound email from account5. The postfix-blue instance continues to reject email from account5 with the sendmail error above. As tempting as it is to blame machine hostility, I know that I must be missing something or doing something wrong. Does anybody have any suggestions as to what it might be? Please feel free to ask for further information about my setup if you need it. =-=-=-=-=-=-=-=-=-= At the request of the responder, here are main.cf and master.cf for a) the main postfix instance ("red.example.com") and b) the FBL instance ("blue.example.us") [NOTE: All parameters not specified below were left at the default Postfix 2.6 settings] MAIN: master.cf smtp inet n - n - - smtpd main.cf myhostname = red.example.com mydomain = example.com inet_interfaces = $myhostname, localhost inet_protocols = all lmtp_host_lookup = native smtp_host_lookup = native ignore_mx_lookup_error = yes mydestination = $myhostname, localhost.$mydomain, localhost local_recipient_maps = mynetworks = 192.168.34.104/32 relay_domains = example.com, example.info, example.net, example.org, example.us relayhost = [192.168.34.102] # Separate physical server, main mailserver. relay_recipient_maps = hash:/etc/postfix/relay_recipients alias_maps = hash:/etc/aliases alias_database = hash:/etc/aliases smtpd_banner = $myhostname ESMTP $mail_name multi_instance_wrapper = ${command_directory}/postmulti -p -- multi_instance_enable = yes multi_instance_directories = /etc/postfix-green /etc/postfix-pink /etc/postfix-orange /etc/postfix-blue FBL: master.cf 184.173.119.103:25 inet n - n - - smtpd main.cf myhostname = blue.example.us mydomain = blue.example.us <= Deliberately set to subdomain only. myorigin = $mydomain inet_interfaces = $myhostname lmtp_host_lookup = native smtp_host_lookup = native ignore_mx_lookup_error = yes mydestination = $myhostname local_recipient_maps = unix:passwd.byname $alias_maps $virtual_alias_maps mynetworks = 192.168.36.51/32, 192.168.35.20/31 <= Second IP is backup MX servers relay_domains = $mydestination recipient_canonical_maps = hash:/etc/postfix-blue/canonical virtual_alias_maps = hash:/etc/postfix-fbl/virtual alias_maps = hash:/etc/aliases, hash:/etc/postfix-blue/canonical alias_maps = hash:/etc/aliases, hash:/etc/postfix-blue/canonical mailbox_command = /usr/bin/procmail -a "$EXTENSION" DEFAULT=$HOME/Mail/ MAILDIR=$HOME/Mail smtpd_banner = $myhostname ESMTP $mail_name authorized_submit_users = multi_instance_name = postfix-blue multi_instance_enable = yes

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  • WNDR3700 Router + Cisco SG200-08 + LACP + Dual Uplink

    - by kobaltz
    Background I have a storage server that has several virtual machine images stored on them. I would store them locally, but I have limited space on my desktop (using SSD storage). I would like to increase the bandwidth between the desktop and the storage server by using two NICs on each computer. My original configuration allowed about 55MBps between the desktop and storage server. This storage server also has several TBs of documents, pictures, movies, vms, and ISO/programs. The storage server has 8 1.5TB hard drives in a RAID 10 configuration with a hardware RAID controller. The benchmarks on the RAID 10 are about 300MBps. Configuration In short, I am trying to bridge my switch and router. The switch is a small 8 port Cisco smart switch that supports 802.3ad LACP. I have two computers plugged into the switch, each with 2 Intel Gigabit NICs. The first computer is a Windows 7 machine that has the Intel ANS software installed. I have LACP configured with the computer and now show 3 NICs (2 Physical + 1 TEAM Virtual @ 2Gbps). It looks like this computer is configured correctly. I trunked the two ports that this computer is plugged into with the switch's web interface. The second computer is a homebrew storage box running debian. I also have the bonding enabled on this machine and the switch configured with LACP. Without having the WNDR3700 router in the picture yet, I am able to communicate between the Windows 7 machine and the debian box since they both have static IP addresses. With LACP enabled on both machines I am getting about 106-108MBps speeds. Issue I plug in a network cable from the switch into the router and enable DHCP on the desktop. I saw no need to have a static address on the desktop. My transfer rates are still from 106MBps-108MBps. While this is still a boost, I am trying to figure out how to get about 140-180MBps. I am thinking that I need to increase the bandwidth from the router to the switch. My switch allows 4 groups for port trunking. I plugged in a second network cable from the router to the switch. My question is, what is the proper way to fix this issue. Should I port trunk the two ports that are going from the switch to the router? Keep in mind that the router is a WNDR3700 and is unsure whether or not it supports LACP. I do have OpenWRT installed on the router, but it still wasn't clear in any documentation that I found if it supported 802.3ad LACP standards. I am also wondering if there needs to be anything changed within the Cisco settings. [Edit] - Corrected some numbers, wasn't really paying attention. It looks like the speeds though at least two NICs are bonded with LACP is still reaching the max bandwidth of one port. Is there a way to configure the switch so that I can increase this bandwidth? Also, on the storage server, I had a couple of extra NICs laying around and threw them on there as well. Another EDIT and More Findings I happened to look at the traffic of each individual NIC and think that I see the problem. I tested with a simple transfer for a 4GB file. I noticed that only one of the NICs was taking the load of the traffic. I then copied the file back to the Storage Server and noticed that the other NIC was sending out the traffic. I have 802.3ad LACP enabled on the two NICs and I see that it gets enabled dynamically on the switch's interface. Should I be using Static Link Aggregation?

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  • KVM with one host IP and a different subnet for machines

    - by Jguy
    I've already setup a KVM host with proper IP configurations, but my host had me create DHCP and use that to assign the IP's to the machines. I want to see if there's an easier way to do it (or better). Upon my first setting out on this, I didn't find anything that pointed me in the right direction. I'm coming off a fresh install of Debian 6.0 x64, so I have nothing installed. I've logged in, queried for the below information and changed the password from my host set one. I have a Debian 6.0 x64 system with the following initial network configuration (substituted 255 in place of my real first octave): # tail /etc/network/interfaces auto eth0 iface eth0 inet static address 255.9.24.80 broadcast 255.9.24.95 netmask 255.255.255.224 gateway 255.9.24.65 # default route to access subnet up route add -net 255.9.24.64 netmask 255.255.255.224 gw 255.9.24.65 eth0 I have a /29 subnet that I want the virtual machines to use from my host: IP: 255.46.187.152 /29 Mask: 255.255.255.248 Broadcast: 255.46.187.159 Usable IP addresses: 255.46.187.153 to 255.46.187.158 I like the interface of Cloudmin, so I want to try and use that if I can to administrate my guests. So, my questions: How do I set this up on the host system the best so that I can use the additional Subnet IP's on the guests and have them accessible from the internet? I also need to host a DNS server, which means one of these VM's has to have two IP's assigned to it and accessable from the outside world. How can I do that using Cloudmin? I had a question about this here: Multiple IP addresses assigned to one KVM VM But I just reformatted the entire server and am trying to figure out a better way of doing this. Machine information: # ip route show 255.9.24.64/27 via 255.9.24.65 dev eth0 255.9.24.64/27 dev eth0 proto kernel scope link src 255.9.24.80 default via 255.9.24.65 dev eth0 brctl is empty # ip addr list 1: lo: <LOOPBACK,UP,LOWER_UP> mtu 16436 qdisc noqueue state UNKNOWN link/loopback 00:00:00:00:00:00 brd 00:00:00:00:00:00 inet 127.0.0.1/8 scope host lo inet6 ::1/128 scope host valid_lft forever preferred_lft forever 2: eth0: <BROADCAST,MULTICAST,UP,LOWER_UP> mtu 1500 qdisc pfifo_fast state UP qlen 1000 link/ether c8:60:00:54:b5:d8 brd ff:ff:ff:ff:ff:ff inet 255.9.24.80/27 brd 255.9.24.95 scope global eth0 inet6 fe80::ca60:ff:fe54:b5d8/64 scope link valid_lft forever preferred_lft forever Thank you for any help you can provide me. EDIT: I've installed kvm and cloudmin: aptitude install qemu-kvm libvirt-bin wget http://cloudmin.virtualmin.com/gpl/scripts/cloudmin-kvm-debian-install.sh ./cloudmin-kvm-debian-install.sh Rebooted and now my network configuration looks like this: # device: eth0 iface eth0 inet manual # default route to access subnet iface br0 inet static address 255.9.24.80 netmask 255.255.255.224 broadcast 255.9.24.95 network 255.9.24.64 bridge_ports eth0 gateway 255.9.24.65 I setup in Cloudmin the Start IP as 255.46.187.153 and End IP as 255.46.187.158. The CIDR is 29 and the gateway is 255.46.187.152. I've installed a guest with ubuntuserver 12.04 x64, which was able to get and retrieve internet resources during installation, but now cannot reach anything nor can it be reached from anything. Its network configuration is: iface eth0 inet static address 255.46.187.153 netmask 255.255.255.224 broadcast 255.46.187.159 gateway 255.46.187.152 dns-nameservers <host provided nameservers> And is not able to ping google.com through DNS or direct IP, I can't ping the VM from the outside or the host. any ideas now?

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  • How does the Cloud compare to Colocation? And development too

    - by David
    Currently I/we run a SaaS web application where each subscriber has their own physical instance of the application in addition to their own database. The setup has each web application instance deployed on two different IIS boxes both for load-balancing and redundancy (the machines have their Windows Update install times 12 hours apart, for example). Databases are mirrored on two different SQL Server 2012 machines with AlwaysOn for uptime. I don't make use of SQL Server clustering (as it doesn't provide storage-level failover: we don't have a shared storage box). Because it's a Windows setup it means there are two Domain Controllers (we cheat: they're both Mac Minis, 17W each, which keeps our colo power costs low). Finally there's also an Exchange server (Mailbox, Hub Transport and Client Access). One of the SQL Servers also doubles-up as an Exchange Hub Transport. Running costs are about $700 a month for our quarter-rack colocation (which includes power and peering/transfer), then there's about $150 a month for SPLA licensing, so $850 a month in total. Then there's the hard-to-quantify cost of administration, but I reckon I spend a couple of hours a week checking-in on the servers: reviewing event logs, etc. I keep getting bombarded by ads and manufactured news stories about how great "the cloud" is. Back in 2008 when the cloud was taking off I was reading up about the proper "cloud" services like Google AppEngine, where you write in Python against Google's API and that's how they scale your application across servers and also use their database provider for scaling storage. Simple enough to understand. Then came along Amazon, and I understand how Amazon Storage works, but I'm not sure how Amazon Compute works: web application pages don't take much CPU time to compute, how do you even quantify usage anyway? Finally, RackSpace gets in the act and now I'm really confused. RackSpace advertise "Cloud" SQL Server 2012 available for about "$0.70 per hour", going by how they advertise it I thought the "hour" meant the sum of CPU time, IO blocking time, maybe time spent transferring data, so for a low-intensity application that works out pretty cheap then? Nope. I went on to a Sales Chat window and spoke to one of their advisors. They told me the $0.70/hour was actually for every hour the SQL Server is running... but who wants a SQL Server for only a few hours? You're going to need it available 24 hours a day for months on end. $0.70 * 24 * 31 works out at $520 a month, which is rediculously expensive for SQL Server. An SPLA license for SQL Server is only $50 a month or so. That $520 a month does not include "fanatical support", and you also need to stack on top the costs of the host Windows server instance too. From what I can tell, Rackspace's "Cloud" products seem like like an cynical rebranding of an overpriced VPS service, but priced by the hour. I have the same confusion about Windows Azure which uses similar terms to describe the products available, but I think that's because Azure offers both traditional shared webhosting in addition to their own APIs you can target for scalable applications.

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  • Connecting PC to TV via HDMI/DVI: Windows XP doesn't allow the appropriate screen resolution

    - by Jørgen
    I have a computer that is connected to the living room TV (a Panasonic) via HDMI. There is no other monitor connected. My problem is that the computer, which is running Windows XP, does not allow me to set the proper resolution for the TV. Both the graphics adapter and the TV should support the 1280x720 resolution, but it cannot be selected - the only available options are 1280x600 and 800x600, both in the "native" Windows dialog box and the custom Intel graphics options dialog box. Do anyone have a suggestion for a solution for this? Things I've thought of: Setting the resolution directly in the registry (where?) Installing some "custom" monitor driver (the TV manufacturer does not appear to provide any, currently the "generic" one is used) Details on the setup: Connection: DVI output on the computer via a passive DVI-HDMI adapter to the HDMI input on the TV, audio is run on a separate link, the TV is able to combine video and audio without any problem, the problem is there regardless of whether or not the audio is connected. The connection is several meters long through some walls, for this reason using a VGA cable instead is not an option. Note that the report explicitly says that the TV supports 1280x720. Still, I am not allowed to select it in Graphics Options, only 1280x600 and 800x600 is available. For 800x600, there's a lot of black around the edges; for 1280x600, the screen is "zoomed" so the edges of the monitor image (like the taskbar) is not visible. Other: The computer is running Windows XP. More recent versions of Windows are not an option (I have no licence). Linux is probably not an option (some of the video streaming sites I plan to use do not support it, I think) I wrote the rest of the details below. Thanks for any help!! TV: Panasonic TX-L32X10Y, European version; a 720p 32" quite "regular" LCD TV. Allowed resolutions according to manual: Signal name: 640x480 @60HZ Horizontal frequency: 31.47 kHz Vertical frequency: 60Hz Signal name: 750/720) /60p Horizontal frequency: 45.00 kHz Vertical frequency: 60Hz Signal name: 1,125 (1,080) / 60p Horizontal frequency: 67.50 kHz Vertical frequency: 60Hz (this is exactly how the manual presents it. PC via D-SUB (VGA cable) and "regular" HDMI have more alternatives.) Messing with the "zoom" settings on the TV does not affect the available resolution options on the computer. Computer: The following is a printout from one of the graphics adapter option pages. I think it covers most of it. The computer is a Dell. INTEL(R) EXTREME GRAPHICS 2 REPORT Report Date: 04/17/2011 Report Time[hr:mm:ss]: 20:18:02 Driver Version: 6.14.10.4396 Operating System: Windows XP* Professional, Service Pack 3 (5.1.2600) Default Language: English DirectX* Version: 9.0 Physical Memory: 1021 MB Minimum Graphics Memory: 1 MB Maximum Graphics Memory: 96 MB Graphics Memory in Use: 6 MB Processor: x86 Processor Speed: 2593 MHZ Vendor ID: 8086 Device ID: 2572 Device Revision: 02 * Accelerator Information * Accelerator in Use: Intel(R) 82865G Graphics Controller Video BIOS: 2972 Current Graphics Mode: 1280 by 600 True Color (60 Hz) * Devices Connected to the Graphics Accelerator * Active Digital Displays: 1 * Digital Display * Monitor Name: Plug and Play Monitor Display Type: Digital Gamma Value: 2.20 DDC2 Protocol: Supported Maximum Image Size: Horizontal: Not Available Vertical: Not Available Monitor Supported Modes: 1280 by 720 (50 Hz) 1280 by 720 (60 Hz) Display Power Management Support: Standby Mode: Not Supported Suspend Mode: Not Supported Active Off Mode: Not Supported (disclaimer: this question was also asked at the Wikipedia Reference Desk some time ago and might show up in a Google search. I got no useful answers there.)

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