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  • DirectX particle system. ConstantBuffer

    - by Liuka
    I'm new in DirectX and I'm making a 2D game. I want to use a particle system to simulate a 3D starfield, so each star has to set its own constant buffer for the vertexshader es. to set it's world matrix. So if i have 500 stars (that move every frame) i need to call 500 times VSsetconstantbuffer, and map/unmap each buffer. with 500 stars i have an average of 220 fps and that's quite good. My bottelneck is Vs/PsSetconstantbuffer. If i dont call this function i have 400 fps(obliviously nothing is display, since i dont set the position of the stars). So is there a method to speed up the render of the particle system?? Ps. I'm using intel integrate graphic (hd 2000-3000). with a nvidia (or amd) gpu will i have the same bottleneck?? If, for example, i dont call setshaderresource i have 10-20 fps more (for 500 objcets), that is not 180.Why does SetConstantBuffer take so long?? LPVOID VSdataPtr = VSmappedResource.pData; memcpy(VSdataPtr, VSdata, CszVSdata); context->Unmap(VertexBuffer, 0); result = context->Map(PixelBuffer, 0, D3D11_MAP_WRITE_DISCARD, 0, &PSmappedResource); if (FAILED(result)) { outputResult.OutputErrorMessage(TITLE, L"Cannot map the PixelBuffer", &result, OUTPUT_ERROR_FILE); return; } LPVOID PSdataPtr = PSmappedResource.pData; memcpy(PSdataPtr, PSdata, CszPSdata); context->Unmap(PixelBuffer, 0); context->VSSetConstantBuffers(0, 1, &VertexBuffer); context->PSSetConstantBuffers(0, 1, &PixelBuffer); this update and set the buffer. It's part of the render method of a sprite class that contains a the vertex buffer and the texture to apply to the quads(it's a 2d game) too. I have an array of 500 stars (sprite setup with a star texture). Every frame: clear back buffer; draw the array of stars; present the backbuffer; draw also call the function update( which calculate the position of the sprite on screen based on a "camera class") Ok, create a vertex buffer with the vertices of each quads(stars) seems to be good, since the stars don't change their "virtual" position; so.... In a particle system (where particles move) it's better to have all the object in only one vertices array, rather then an array of different sprite/object in order to update all the vertices' position with a single setbuffer call. In this case i have to use a dynamic vertex buffer with the vertices positions like this: verticesForQuad={{ XMFLOAT3((float)halfDImensions.x-1+pos.x, (float)halfDImensions.y-1+pos.y, 1.0f), XMFLOAT2(1.0f, 0.0f) }, { XMFLOAT3((float)halfDImensions.x-1+pos.x, -(float)halfDImensions.y-1+pos.y, 1.0f), XMFLOAT2(1.0f, 1.0f) }, { XMFLOAT3(-(float)halfDImensions.x-1+pos.x, (float)halfDImensions.y-1.pos.y, 1.0f), XMFLOAT2(0.0f, 0.0f) }, { XMFLOAT3(-(float)halfDImensions.x-1.pos.x, -(float)halfDImensions.y-1+pos.y, 1.0f), XMFLOAT2(0.0f, 1.0f) }, ....other quads} where halfDimensions is the halfsize in pixel of a texture and pos the virtual position of a star. than create an array of verticesForQuad and create the vertex buffer ZeroMemory(&vertexDesc, sizeof(vertexDesc)); vertexDesc.Usage = D3D11_USAGE_DEFAULT; vertexDesc.BindFlags = D3D11_BIND_VERTEX_BUFFER; vertexDesc.ByteWidth = sizeof(VertexType)* 4*numStars; ZeroMemory(&resourceData, sizeof(resourceData)); resourceData.pSysMem = verticesForQuad; result = device->CreateBuffer(&vertexDesc, &resourceData, &CvertexBuffer); and call each frame Context->IASetVertexBuffers(0, 1, &CvertexBuffer, &stride, &offset); But if i want to add and remove obj i have to recreate the buffer each time, havent i?? There is a faster way? I think i can create a vertex buffer with a max size (es. 10000 objs) and when i update it set only the 250 position (for 250 onjs for example) and pass this number as the vertexCount to the draw function (numObjs*4), or i'm worng

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  • Spotlight on Oracle Social Relationship Management. Social Enable Your Enterprise with Oracle SRM.

    - by Pat Ma
    Facebook is now the most popular site on the Internet. People are tweeting more than they send email. Because there are so many people on social media, companies and brands want to be there too. They want to be able to listen to social chatter, engage with customers on social, create great-looking Facebook pages, and roll out social-collaborative work environments within their organization. This is where Oracle Social Relationship Management (SRM) comes in. Oracle SRM is a product that allows companies to manage their presence with prospects and customers on social channels. Let's talk about two popular use cases with Oracle SRM. Easy Publishing - Companies now have an average of 178 social media accounts - with every product or geography or employee group creating their own social media channel. For example, if you work at an international hotel chain with every single hotel creating their own Facebook page for their location, that chain can have well over 1,000 social media accounts. Managing these channels is a mess - with logging in and out of every account, making sure that all accounts are on brand, and preventing rogue posts from destroying the brand. This is where Oracle SRM comes in. With Oracle Social Relationship Management, you can log into one window and post messages to all 1,000+ social channels at once. You can set up approval flows and have each account generate their own content but that content must be approved before publishing. The benefits of this are easy social media publishing, brand consistency across all channels, and protection of your brand from inappropriate posts. Monitoring and Listening - People are writing and talking about your company right now on social media. 75% of social media users have written a negative post about a brand after a poor customer service experience. Think about all the negative posts you see in your Facebook news feed about delayed flights or being on hold for 45 minutes. There is so much social chatter going on around your brand that it's almost impossible to keep up or comprehend what's going on. That's where Oracle SRM comes in. With Social Relationship Management, a company can monitor and listen to what people are saying about them on social channels. They can drill down into individual posts or get a high level view of trends and mentions. The benefits of this are comprehending what's being said about your brand and its competitors, understanding customers and their intent, and responding to negative posts before they become a PR crisis. Oracle SRM is part of Oracle Cloud. The benefits of cloud deployment for customers are faster deployments, less maintenance, and lower cost of ownership versus on-premise deployments. Oracle SRM also fits into Oracle's vision to social enable your enterprise. With Oracle SRM, social media is not just a marketing channel. Social media is also mechanism for sales, customer support, recruiting, and employee collaboration. For more information about how Oracle SRM can social enable your enterprise, please visit oracle.com/social. For more information about Oracle Cloud, please visit cloud.oracle.com.

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  • Challenges in Corporate Reporting - New Independent Research

    - by ndwyouell
    Earlier this year, Oracle and Accenture sponsored a global study on trends in financial close and reporting. We surveyed 1,123 finance professionals in large organizations in 12 countries around the world during February and March. Financial Consolidation and Reporting is the most mature aspect of Enterprise Performance Management with mainstream solutions having been around for over 30 years. But of course over this time there have been many changes and very significant increases in regulation. So just what is the current state is Financial Consolidation and Reporting in our major corporations across the world? We commissioned this independent research to find out. Highlights of the result are: •          Seeking change: Businesses recognize they need to invest in financial reporting to address the challenges they currently face. 47 percent of companies have made substantial investments over the last year to the financial close, filing, and reporting processes. •          Ineffective investments: Despite these investments, spreadsheets (72 percent) and e-mails (68 percent) are still being used daily to track and manage reporting, suggesting that new investments are falling short of expectations. •          Increased costs and uncertainty: The situation is so opaque that managers across the finance function are unable to fully understand the financial impact or cost implications of reporting, with 60 percent of respondents admitting they did not know the total cost of managing and publicizing their financial results. •          Persistent challenges: 68 percent of respondents admitted that they have inadequate visibility into reporting processes, while 84 percent of finance managers surveyed said they find it difficult to control the quality of financial data across the entire reporting process. •          Decreased effectiveness: 71 percent of finance managers feel their effectiveness is limited in some way by data-analysis–related issues, while 39 percent of C-level or VP-level respondents say their effectiveness is impaired by limited visibility. •          Missed deadlines: Due to late changes to the chart of accounts, 15 percent of global businesses have missed statutory filings, putting their companies at risk of financial penalties and potentially impacting share value. The report makes it clear that investments made to date by these large organizations around the world have been uneven across the close, reporting, and filing processes, which has led to the challenges these organizations currently face in the overall process. Regardless of whether companies are using a variety of solutions or a single solution, the report shows they continue to witness increased costs, ineffectual data management, and missed reporting, which—in extreme circumstances—can impact a company’s corporate image and share value. The good news is that businesses realize that these problems persist and 86 percent of companies are likely to make a significant investment during the next five years to address these issues. While they should invest, it is critical that they direct investments correctly to address the key issues this research identified: •          Improving data integrity •          Optimizing processes •          Integrating the extended financial close process By addressing these issues and with clear guidance on how to implement the correct business processes, infrastructure, and software solutions, finance teams will find that their reporting processes are much more effective, cost-efficient, and aligned with their performance expectations. To get a copy of the full report: http://www.oracle.com/webapps/dialogue/ns/dlgwelcome.jsp?p_ext=Y&p_dlg_id=11747758&src=7300117&Act=92 To replay a webcast discussing the findings: http://www.cfo.com/webcast.cfm?webcast=14639438&pcode=ORA061912_ORA

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  • Using NSpec at various architectural layers

    - by nono
    Having read the quick start at nspec.org, I realized that NSpec might be a useful tool in a scenario which was becoming a bit cumbersome with NUnit alone. I'm adding an OAuth (or, DotNetOpenAuth) to a website and quickly made a mess of writing test methods such as [Test] public void UserIsLoggedInLocallyPriorToInvokingExternalLoginAndExternalLoginSucceedsAndExternalProviderIdIsNotAlreadyAssociatedWithUserAccount() { ... } ... and I wound up with maybe a dozen permutations of this theme, for the user already being logged in locally and not locally, the external login succeeding or failing, etc. Not only were the method names unwieldy, but every test needed a setup that contained parts in common with a different set of other tests. I realized that NSpec's incremental setup capabilities would work great for this, and for a while I was trucking a long wonderfully, with code like act = () => { actionResult = controller.ExternalLoginCallback(returnUrl); }; context["The user is already logged in"] = () => { before = () => identity.Setup(x => x.IsAuthenticated).Returns(true); context["The external login succeeds"] = () => { before = () => oauth.Setup(x => x.VerifyAuthentication(It.IsAny<string>())).Returns(new AuthenticationResult(true, providerName, "provideruserid", "username", new Dictionary<string, string>())); context["External login already exists for current user"] = () => { before = () => authService.Setup(x => x.ExternalLoginExistsForUser(It.IsAny<string>(), It.IsAny<string>(), It.IsAny<string>())).Returns(true); it["Should add 'login sucessful' alert"] = () => { var alerts = (IList<Alert>)controller.TempData[TempDataKeys.AlertCollection]; alerts[0].Message.should_be_same("Login successful"); alerts[0].AlertType.should_be(AlertType.Success); }; it["Should return a redirect result"] = () => actionResult.should_cast_to<RedirectToRouteResult>(); }; context["External login already exists for another user"] = () => { before = () => authService.Setup(x => x.ExternalLoginExistsForAnyOtherUser(It.IsAny<string>(), It.IsAny<string>(), It.IsAny<string>())).Returns(true); it["Adds an error alert"] = () => { var alerts = (IList<Alert>)controller.TempData[TempDataKeys.AlertCollection]; alerts[0].Message.should_be_same("The external login you requested is already associated with a different user account"); alerts[0].AlertType.should_be(AlertType.Error); }; it["Should return a redirect result"] = () => actionResult.should_cast_to<RedirectToRouteResult>(); }; This approach seemed to work magnificently until I prepared to write test code for my ApplicationServices layer, to which I delegate viewmodel manipulation from my MVC controllers, and which coordinates the operations of the lower data repository layer: public void CreateUserAccountFromExternalLogin(RegisterExternalLoginModel model) { throw new NotImplementedException(); } public void AssociateExternalLoginWithUser(string userName, string provider, string providerUserId) { throw new NotImplementedException(); } public string GetLocalUserName(string provider, string providerUserId) { throw new NotImplementedException(); } I have no idea what in the world to name the test class, the test methods, or even if I should perhaps include the testing for this layer into the test class from my large code snippet above, so that a single feature or user action could be tested without regard to architectural layering. I can't find any tutorials or blog posts which cover more than simple examples, so I would appreciate any recommendations or pointing in the right direction. I would even welcome "your question is invalid"-type answers as long as some explanation is provided.

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  • Cannot establish maximum resolution on ASUS PB278Q

    - by dentuzhik
    I've recently bought brand new ASUS PB278Q monitor. When trying to connect to my laptop, everything works great, except that I can't get the native resolution of my monitor (2560x1440) working. The automatic is 1920x1080. My graphic card is Nvidia GeForce 320m. Here's output from lspci for it: ~$ lspci | grep VGA 02:00.0 VGA compatible controller: NVIDIA Corporation GT216M [GeForce GT 320M] (rev a2) and also xrandr: ~$ xrandr Screen 0: minimum 8 x 8, current 3286 x 1437, maximum 8192 x 8192 VGA-0 disconnected (normal left inverted right x axis y axis) LVDS-0 connected primary 1366x768+0+669 (normal left inverted right x axis y axis) 344mm x 193mm 1366x768 60.0*+ HDMI-0 connected 1920x1080+1366+0 (normal left inverted right x axis y axis) 600mm x 340mm 1920x1080 60.0*+ 59.9 50.0 30.0 25.0 24.0 60.0 50.0 1680x1050 60.0 1440x900 59.9 1280x1024 75.0 60.0 1280x960 60.0 1280x800 59.8 1280x720 60.0 59.9 50.0 1152x864 75.0 1024x768 75.0 70.1 60.0 800x600 75.0 72.2 60.3 56.2 720x576 50.0 720x480 59.9 640x480 75.0 59.9 59.9 480x576 50.0 480x480 59.9 I have proprietary drivers installed on my machine, here's the info about the monitor from nvidia-settings (Actually I don't have enough reputation to post images, so here's the text): Chip Location: Internal Signal: TDMS Connection link: Single Native resolution: 2560x1440 Refresh rate: 60.00 Hz The monitor is connected to laptop via HDMI cable, and honestly I have no idea what version it is, and what version is my HDMI output of my graphics card. I tried to find how I can figure it out on the web, but had no luck. Also my video card has only VGA and HDMI outs so I can't test neither DVI-D cable nor DisplayPort. So apparently, there's some problem over there. At least I want to know exactly what's going on. I've tried to see if it a linux-specific problem, but windows also gave me the same resolution by default. What I've already tried: Connect through VGA (stupid one, of course it gave me 1920x1080). Checked two HDMI cables (not sure if they're the same or not, as mentioned above). Played around with xrandr and adding custom modes. Didn't help. Surfed for the info a lot on the web, but couldn't get appropriate results. Actually xrandr gives me the following: ~$ cvt 2560 1440 60 # 2560x1440 59.96 Hz (CVT 3.69M9) hsync: 89.52 kHz; pclk: 312.25 MHz Modeline "2560x1440_60.00" 312.25 2560 2752 3024 3488 1440 1443 1448 1493 -hsync +vsync ~$ xrandr --newmode "2560x1440_60.00" 312.25 2560 2752 3024 3488 1440 1443 1448 1493 -hsync +vsync ~$ xrandr Screen 0: minimum 8 x 8, current 3286 x 1437, maximum 8192 x 8192 VGA-0 disconnected (normal left inverted right x axis y axis) LVDS-0 connected 1366x768+0+669 (normal left inverted right x axis y axis) 344mm x 193mm 1366x768 60.0*+ HDMI-0 connected primary 1920x1080+1366+0 (normal left inverted right x axis y axis) 600mm x 340mm 1920x1080 60.0*+ 59.9 50.0 30.0 25.0 24.0 60.0 50.0 1680x1050 60.0 1440x900 59.9 1280x1024 75.0 60.0 1280x960 60.0 1280x800 59.8 1280x720 60.0 59.9 50.0 1152x864 75.0 1024x768 75.0 70.1 60.0 800x600 75.0 72.2 60.3 56.2 720x576 50.0 720x480 59.9 640x480 75.0 59.9 59.9 480x576 50.0 480x480 59.9 2560x1440_60.00 (0x34f) 312.2MHz h: width 2560 start 2752 end 3024 total 3488 skew 0 clock 89.5KHz v: height 1440 start 1443 end 1448 total 1493 clock 60.0Hz ~$ xrandr --addmode HDMI-0 2560x1440_60.00 X Error of failed request: BadMatch (invalid parameter attributes) Major opcode of failed request: 140 (RANDR) Minor opcode of failed request: 18 (RRAddOutputMode) Serial number of failed request: 29 Current serial number in output stream: 30 What I intend to do next: Try another HDMI cable? Try HDMI to DVI-D cable? Try HDMI to DisplayPort cable? Another type of adapters? VGA to DVI-D? Buy another laptop with another graphic card. Damn. My ideas pretty much end here. Any ideas? Any explanations why it isn't working are appreciated.

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  • Concurrent Affairs

    - by Tony Davis
    I once wrote an editorial, multi-core mania, on the conundrum of ever-increasing numbers of processor cores, but without the concurrent programming techniques to get anywhere near exploiting their performance potential. I came to the.controversial.conclusion that, while the problem loomed for all procedural languages, it was not a big issue for the vast majority of programmers. Two years later, I still think most programmers don't concern themselves overly with this issue, but I do think that's a bigger problem than I originally implied. Firstly, is the performance boost from writing code that can fully exploit all available cores worth the cost of the additional programming complexity? Right now, with quad-core processors that, at best, can make our programs four times faster, the answer is still no for many applications. But what happens in a few years, as the number of cores grows to 100 or even 1000? At this point, it becomes very hard to ignore the potential gains from exploiting concurrency. Possibly, I was optimistic to assume that, by the time we have 100-core processors, and most applications really needed to exploit them, some technology would be around to allow us to do so with relative ease. The ideal solution would be one that allows programmers to forget about the problem, in much the same way that garbage collection removed the need to worry too much about memory allocation. From all I can find on the topic, though, there is only a remote likelihood that we'll ever have a compiler that takes a program written in a single-threaded style and "auto-magically" converts it into an efficient, correct, multi-threaded program. At the same time, it seems clear that what is currently the most common solution, multi-threaded programming with shared memory, is unsustainable. As soon as a piece of state can be changed by a different thread of execution, the potential number of execution paths through your program grows exponentially with the number of threads. If you have two threads, each executing n instructions, then there are 2^n possible "interleavings" of those instructions. Of course, many of those interleavings will have identical behavior, but several won't. Not only does this make understanding how a program works an order of magnitude harder, but it will also result in irreproducible, non-deterministic, bugs. And of course, the problem will be many times worse when you have a hundred or a thousand threads. So what is the answer? All of the possible alternatives require a change in the way we write programs and, currently, seem to be plagued by performance issues. Software transactional memory (STM) applies the ideas of database transactions, and optimistic concurrency control, to memory. However, working out how to break down your program into sufficiently small transactions, so as to avoid contention issues, isn't easy. Another approach is concurrency with actors, where instead of having threads share memory, each thread runs in complete isolation, and communicates with others by passing messages. It simplifies concurrent programs but still has performance issues, if the threads need to operate on the same large piece of data. There are doubtless other possible solutions that I haven't mentioned, and I would love to know to what extent you, as a developer, are considering the problem of multi-core concurrency, what solution you currently favor, and why. Cheers, Tony.

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  • How to get the height of an image and apply that height to a div? [migrated]

    - by Mick79
    I am building a mobile web app and I'm using jquerytools slider on it. i want te slider to show (in proper ratio) across all mobile devices so width of the images is 100% and height is auto in css. However as all the elements are floated and jquerytools slider requires the position be set to absolute, the containing div (#header) doesn't stretch to fit the content. I am trying to use jquery to get the height of the height of the img and apply that height to the header.... however I am having no luck. CSS: #header{ width:100%; position:relative; z-index: 20; /* box-shadow: 0 0 10px white; */ overflow: auto; } .scrollable { position:relative; overflow:hidden; width: 100%; height: 100%; /* box-shadow: 0 0 20px purple; */ /* height:198px; */ z-index: 20; overflow: auto; } .scrollable .items { /* this cannot be too large */ width:1000%; position:absolute; clear:both; /* box-shadow: 0 0 30px green; */ } .items div { float:left; width:10%; height:100%; } /* single scrollable item */ .scrollable img { /* float:left; */ width:100%; height: auto; /* height:198px; */ } /* active item */ .scrollable .active { border:2px solid #000; position:relative; cursor:default; } HTML <div id=header><!-- root element for scrollable --> <div class="scrollable" id="scrollable"> <!-- root element for the items --> <div class="items"> <div> <img src="img/img2.jpg" /> </div> <div> <img src="img/img1.jpg" /> </div> <div> <img src="img/img3.jpg" /> </div> <div> <img src="img/img4.jpg" /> </div> <div> <img src="img/img6.jpg" /> </div> </div><!-- items --> </div><!-- scrollable --> </div><!-- header -->

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  • Webcast On-Demand: Building Java EE Apps That Scale

    - by jeckels
    With some awesome work by one of our architects, Randy Stafford, we recently completed a webcast on scaling Java EE apps efficiently. Did you miss it? No problem. We have a replay available on-demand for you. Just hit the '+' sign drop-down for access.Topics include: Domain object caching Service response caching Session state caching JSR-107 HotCache and more! Further, we had several interesting questions asked by our audience, and we thought we'd share a sampling of those here for you - just in case you had the same queries yourself. Enjoy! What is the largest Coherence deployment out there? We have seen deployments with over 500 JVMs in the Coherence cluster, and deployments with over 1000 JVMs using the Coherence jar file, in one system. On the management side there is an ecosystem of monitoring tools from Oracle and third parties with dashboards graphing values from Coherence's JMX instrumentation. For lifecycle management we have seen a lot of custom scripting over the years, but we've also integrated closely with WebLogic to leverage its management ecosystem for deploying Coherence-based applications and managing process life cycles. That integration introduces a new Java EE archive type, the Grid Archive or GAR, which embeds in an EAR and can be seen by a WAR in WebLogic. That integration also doesn't require any extra WebLogic licensing if Coherence is licensed. How is Coherence different from a NoSQL Database like MongoDB? Coherence can be considered a NoSQL technology. It pre-dates the NoSQL movement, having been first released in 2001 whereas the term "NoSQL" was coined in 2009. Coherence has a key-value data model primarily but can also be used for document data models. Coherence manages data in memory currently, though disk persistence is in a future release currently in beta testing. Where the data is managed yields a few differences from the most well-known NoSQL products: access latency is faster with Coherence, though well-known NoSQL databases can manage more data. Coherence also has features that well-known NoSQL database lack, such as grid computing, eventing, and data source integration. Finally Coherence has had 15 years of maturation and hardening from usage in mission-critical systems across a variety of industries, particularly financial services. Can I use Coherence for local caching? Yes, you get additional features beyond just a java.util.Map: you get expiration capabilities, size-limitation capabilities, eventing capabilites, etc. Are there APIs available for GoldenGate HotCache? It's mostly a black box. You configure it, and it just puts objects into your caches. However you can treat it as a glass box, and use Coherence event interceptors to enhance its behavior - and there are use cases for that. Are Coherence caches updated transactionally? Coherence provides several mechanisms for concurrency control. If a project insists on full-blown JTA / XA distributed transactions, Coherence caches can participate as resources. But nobody does that because it's a performance and scalability anti-pattern. At finer granularity, Coherence guarantees strict ordering of all operations (reads and writes) against a single cache key if the operations are done using Coherence's "EntryProcessor" feature. And Coherence has a unique feature called "partition-level transactions" which guarantees atomic writes of multiple cache entries (even in different caches) without requiring JTA / XA distributed transaction semantics.

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  • Making user input/math on data fast, unlike excel type programs

    - by proGrammar
    I'm creating a research platform solely for myself to do some research on data. Programs like excel are terribly slow for me so I'm trying to come up with another solution. Originally I used excel. A1 was the cell that contained the data and all other cells in use calculated something on A1, or on other cells, that all could be in the end traced to A1. A1 was like an element of an array, I then I incremented it to go through all my data. This was way too slow. So the only other option I found originally was to hand code in c# the calculations inside a loop. Then I simply recompiled each time I changed my math. This was terribly slow to do and I had to order everything correctly so things would update correctly (dependencies). I could have also used events, but hand coding events for each cell like calculation would also be very slow. Next I created an application to read Excel and to perfectly imitate it. Which is what I now use. Basically I write formulas onto a fraction of my data to get live results inside excel. Then my program reads excel, writes another c# program, compiles it, and runs that program which runs my excel created formulas through a lot more data a whole lot faster. The advantage being my application dependency sorts everything (or I could use events) so I don't have to (like excel does) And of course the speed. But now its not a single application anymore. Instead its 2 applications, one which only reads my formulas and writes another program. The other one being the result which only lives for a short while before I do other runs through my data with different formulas / settings. So I can't see multiple results at one time without introducing even more programs like a database or at least having the 2 applications talking to each other. My idea was to have a dll that would be written, compiled, loaded, and unloaded again and again. So a self-updating program, sort of. But apparently that's not possible without another appdomain which means data has to be marshaled to be moved between the appdomains. Which would slow things down, not for summaries, but for other stuff I need to do with all my data. I'm also forgetting to mention a huge problem with restarting an application again and again which is having to reload ALL my data into memory again and again. But its still a whole lot faster than excel. I'm really super puzzled as to what people do when they want to research data fast. I'm completely unable to have a program accept user input and having it fast. My understanding is that it would have to do things like excel which is to evaluate strings again and again. So my only option is to repeatedly compile applications. Do I have a correct understanding on computer science? I've only just began programming, and didn't think I would have to learn much to do some simple math on data. My understanding is its either compiling my user defined stuff to a program or evaluating them from a string or something stupid again and again. And my only option is to probably switch operating systems or something to be able to have a program compile and run itself without stopping (writing/compiling dll, loading dll to program, unloading, and repeating). Can someone give me some idea on how computers work? Is anything better possible? Like a running program, that can accept user input and compile it and then unload it later? I mean heck operating systems dont need to be RESTARTED with every change to user input. What is this the cave man days? Sorry, it's just so super frustrating not knowing what one can do, and can't do. If only I could understand and learn this stuff fast enough.

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  • High Resolution Timeouts

    - by user12607257
    The default resolution of application timers and timeouts is now 1 msec in Solaris 11.1, down from 10 msec in previous releases. This improves out-of-the-box performance of polling and event based applications, such as ticker applications, and even the Oracle rdbms log writer. More on that in a moment. As a simple example, the poll() system call takes a timeout argument in units of msec: System Calls poll(2) NAME poll - input/output multiplexing SYNOPSIS int poll(struct pollfd fds[], nfds_t nfds, int timeout); In Solaris 11, a call to poll(NULL,0,1) returns in 10 msec, because even though a 1 msec interval is requested, the implementation rounds to the system clock resolution of 10 msec. In Solaris 11.1, this call returns in 1 msec. In specification lawyer terms, the resolution of CLOCK_REALTIME, introduced by POSIX.1b real time extensions, is now 1 msec. The function clock_getres(CLOCK_REALTIME,&res) returns 1 msec, and any library calls whose man page explicitly mention CLOCK_REALTIME, such as nanosleep(), are subject to the new resolution. Additionally, many legacy functions that pre-date POSIX.1b and do not explicitly mention a clock domain, such as poll(), are subject to the new resolution. Here is a fairly comprehensive list: nanosleep pthread_mutex_timedlock pthread_mutex_reltimedlock_np pthread_rwlock_timedrdlock pthread_rwlock_reltimedrdlock_np pthread_rwlock_timedwrlock pthread_rwlock_reltimedwrlock_np mq_timedreceive mq_reltimedreceive_np mq_timedsend mq_reltimedsend_np sem_timedwait sem_reltimedwait_np poll select pselect _lwp_cond_timedwait _lwp_cond_reltimedwait semtimedop sigtimedwait aiowait aio_waitn aio_suspend port_get port_getn cond_timedwait cond_reltimedwait setitimer (ITIMER_REAL) misc rpc calls, misc ldap calls This change in resolution was made feasible because we made the implementation of timeouts more efficient a few years back when we re-architected the callout subsystem of Solaris. Previously, timeouts were tested and expired by the kernel's clock thread which ran 100 times per second, yielding a resolution of 10 msec. This did not scale, as timeouts could be posted by every CPU, but were expired by only a single thread. The resolution could be changed by setting hires_tick=1 in /etc/system, but this caused the clock thread to run at 1000 Hz, which made the potential scalability problem worse. Given enough CPUs posting enough timeouts, the clock thread could be a performance bottleneck. We fixed that by re-implementing the timeout as a per-CPU timer interrupt (using the cyclic subsystem, for those familiar with Solaris internals). This decoupled the clock thread frequency from timeout resolution, and allowed us to improve default timeout resolution without adding CPU overhead in the clock thread. Here are some exceptions for which the default resolution is still 10 msec. The thread scheduler's time quantum is 10 msec by default, because preemption is driven by the clock thread (plus helper threads for scalability). See for example dispadmin, priocntl, fx_dptbl, rt_dptbl, and ts_dptbl. This may be changed using hires_tick. The resolution of the clock_t data type, primarily used in DDI functions, is 10 msec. It may be changed using hires_tick. These functions are only used by developers writing kernel modules. A few functions that pre-date POSIX CLOCK_REALTIME mention _SC_CLK_TCK, CLK_TCK, "system clock", or no clock domain. These functions are still driven by the clock thread, and their resolution is 10 msec. They include alarm, pcsample, times, clock, and setitimer for ITIMER_VIRTUAL and ITIMER_PROF. Their resolution may be changed using hires_tick. Now back to the database. How does this help the Oracle log writer? Foreground processes post a redo record to the log writer, which releases them after the redo has committed. When a large number of foregrounds are waiting, the release step can slow down the log writer, so under heavy load, the foregrounds switch to a mode where they poll for completion. This scales better because every foreground can poll independently, but at the cost of waiting the minimum polling interval. That was 10 msec, but is now 1 msec in Solaris 11.1, so the foregrounds process transactions faster under load. Pretty cool.

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  • Disabling the right-click sub menu using JQuery

    - by nikolaosk
    Recently I needed to disable the right-click contextual menu in an HTML page for a very simple HTML application I was creating for a friend.This is going to be a short post where I will demonstrate how to disable the right-click contextual menu.I will use the very popular JQuery Library. Please download the library (minified version) from http://jquery.com/downloadPlease find here all my posts regarding JQuery.In this hands-on example I will be using Expression Web 4.0.This application is not a free application. You can use any HTML editor you like.You can use Visual Studio 2012 Express edition. You can download it here. I am going to create a very simple HTML 5 page with some text and an image. The HTML markup for the page follows. <!DOCTYPE html><html lang="en">  <head>    <title>HTML 5, CSS3 and JQuery</title>        <meta http-equiv="Content-Type" content="text/html;charset=utf-8" >    <link rel="stylesheet" type="text/css" href="style.css">     <script type="text/javascript" src="jquery-1.8.2.min.js">        </script><script type="text/javascript"> (function ($) { $(document).bind('contextmenu', function () { return false;}); })(jQuery); </script>       </head>  <body>      <div id="header">      <h1>Learn cutting edge technologies</h1>      <h2>HTML 5, JQuery, CSS3</h2>    </div>      <figure>  <img src="html5.png" alt="HTML 5"></figure>        <div id="main">          <h2>HTML 5</h2>                        <article>          <p>            HTML5 is the latest version of HTML and XHTML. The HTML standard defines a single language that can be written in HTML and XML. It attempts to solve issues found in previous iterations of HTML and addresses the needs of Web Applications, an area previously not adequately covered by HTML.          </p>          </article>      </div>             </body>  </html> This is the JQuery code, I use (function ($) { $(document).bind('contextmenu', function () { return false;}); })(jQuery); I simply disable/cancel the contextmenu event.When I load the simple page on the browser and I right-click the context menu does not appear.Hope it helps!!!

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  • My Favorite Free Windows Phone Twitter App

    - by Tim Murphy
    Windows Phone 7 has been out for about two years now.  In that time I have switched back and forth with different free Twitter apps.  Mostly the has been because someone has mentioned one or another that they like.  I figured I would give a quick run down of what I felt were the pros and cons of each.  These are only the ones that I have used and your mileage may vary.  So here we go. WP7 Built-In Twitter Functionality While it is great that Microsoft put this functionality in, it is extremely limited in usefulness.  Some apps leverage it to allow you to share pictures or information they contain.  In all though, I don’t use it unless it is the quickest way to get something out. Official Twitter App The official Twitter app isn’t a very big step up from the phone functionality.  It gives you a better timeline view and better attachment handling, but it makes you bounce to a browser page to see images that are linked to a tweet. TweetCaster This was my main Twitter app for quite a while.  It is the only one with InstaPaper integration so that you can save you a tweet and review it later.  My main problem is that it crashes too much when it can’t find a connection.  It also only previews yfrog and twitpic images and only once you go to the detail of a tweet.  Other than that it is a solid Twitter client. moTweets This is my current favorite. It has nice image display in your timeline which I have not seen on any of the other apps.  There are two modes that you can use with this app.  The first is standard to most Twitter apps that allows you to navigate to a tweet and do the usual operations.  The second is what they call Quick Buttons.  In this case you do not see the content of the tweet but go straight to the let’s get something done stage.  It is an interesting take.  I do miss the Instapaper integration and it has a tendency to show a blank timeline list once in a while after you view detail entry.  If you scroll the list it restore your timeline, but you lose you place and are put to the first entry. Seesmic I am not very fond of this app.  The first thing is that it makes you pick a “Space” when you enter the app.  This is really “which account do you want to see”.  On top of that it does not show who retweeted an entry in your timeline and then only tells you how many people RT the post when you look at the detail.  There is a Speak feature that will read you a single tweet, but you have to navigate to the tweet and then to a menu to make it work.  We will have to see if this gets better with the features in Windows Phone 8.  Other than that it is another basic feature app.  Summary In the end I am sticking with moTweets.  I would appreciate it if they added the Instapaper capability and fixed the one bug.  If they did that I would be really happy with the product. del.icio.us Tags: Twitter,Windows Phone 7,WP7,TweetCaster,moTweets,Seesmic

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  • I thought the new AUTO_SAMPLE_SIZE in Oracle Database 11g looked at all the rows in a table so why do I see a very small sample size on some tables?

    - by Maria Colgan
    I recently got asked this question and thought it was worth a quick blog post to explain in a little more detail what is going on with the new AUTO_SAMPLE_SIZE in Oracle Database 11g and what you should expect to see in the dictionary views. Let’s take the SH.CUSTOMERS table as an example.  There are 55,500 rows in the SH.CUSTOMERS tables. If we gather statistics on the SH.CUSTOMERS using the new AUTO_SAMPLE_SIZE but without collecting histogram we can check what sample size was used by looking in the USER_TABLES and USER_TAB_COL_STATISTICS dictionary views. The sample sized shown in the USER_TABLES is 55,500 rows or the entire table as expected. In USER_TAB_COL_STATISTICS most columns show 55,500 rows as the sample size except for four columns (CUST_SRC_ID, CUST_EFF_TO, CUST_MARTIAL_STATUS, CUST_INCOME_LEVEL ). The CUST_SRC_ID and CUST_EFF_TO columns have no sample size listed because there are only NULL values in these columns and the statistics gathering procedure skips NULL values. The CUST_MARTIAL_STATUS (38,072) and the CUST_INCOME_LEVEL (55,459) columns show less than 55,500 rows as their sample size because of the presence of NULL values in these columns. In the SH.CUSTOMERS table 17,428 rows have a NULL as the value for CUST_MARTIAL_STATUS column (17428+38072 = 55500), while 41 rows have a NULL values for the CUST_INCOME_LEVEL column (41+55459 = 55500). So we can confirm that the new AUTO_SAMPLE_SIZE algorithm will use all non-NULL values when gathering basic table and column level statistics. Now we have clear understanding of what sample size to expect lets include histogram creation as part of the statistics gathering. Again we can look in the USER_TABLES and USER_TAB_COL_STATISTICS dictionary views to find the sample size used. The sample size seen in USER_TABLES is 55,500 rows but if we look at the column statistics we see that it is same as in previous case except  for columns  CUST_POSTAL_CODE and  CUST_CITY_ID. You will also notice that these columns now have histograms created on them. The sample size shown for these columns is not the sample size used to gather the basic column statistics. AUTO_SAMPLE_SIZE still uses all the rows in the table - the NULL rows to gather the basic column statistics (55,500 rows in this case). The size shown is the sample size used to create the histogram on the column. When we create a histogram we try to build it on a sample that has approximately 5,500 non-null values for the column.  Typically all of the histograms required for a table are built from the same sample. In our example the histograms created on CUST_POSTAL_CODE and the CUST_CITY_ID were built on a single sample of ~5,500 (5,450 rows) as these columns contained only non-null values. However, if one or more of the columns that requires a histogram has null values then the sample size maybe increased in order to achieve a sample of 5,500 non-null values for those columns. n addition, if the difference between the number of nulls in the columns varies greatly, we may create multiple samples, one for the columns that have a low number of null values and one for the columns with a high number of null values.  This scheme enables us to get close to 5,500 non-null values for each column. +Maria Colgan

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  • Xenserver 5.6 SR_BACKEND_FAILURE_47 no such volume group, but it is there

    - by Juan Carlos
    I've looked everywhere (Google, here, a bunch of other sites), and while I have found people with similar problems, I couldn't find a single one with a solution to this. Last night our xenserver 5.6 box corrupted the /var/xapi/state.db, and I couldn't fix the xml, no matter what I did. After a good hour fiddling with the file, I figured it would be faster to just reinstall. The server had one 2tb hard drive running Xen and its VMs, and since Xen's install said it would erase the hard drive it was installed on, I plugged a new harddrive and installed Xen on it, without selecting any hard drives for storage. I Figured I could make it happen after install, using the partition on the old harddrive with all my VMs on it. After instalation finished and the system booted I did: #fdisk -l found the old partition at /dev/sda3 #ll /dev/disk/by-id found the partition at /dev/disk/by-id/scsi-3600188b04c02f100181ab3a48417e490-part3 #xe host-list uuid ( RO) : a019d93e-4d84-4a4b-91e3-23572b5bd8a4 name-label ( RW): xenserver-scribfourteen name-description ( RW): Default install of XenServer #pvscan PV /dev/sda3 VG VG_XenStorage-405a2ece-d10e-d6c5-ede2-e1ad2c29c68d lvm2 [1.81 TB / 204.85 GB free] Total: 1 [1.81 TB] / in use: 1 [1.81 TB] / in no VG: 0 [0 ] #vgscan Reading all physical volumes. This may take a while... Found volume group "VG_XenStorage-405a2ece-d10e-d6c5-ede2-e1ad2c29c68d" using metadata type lvm2 # pvdisplay --- Physical volume --- PV Name /dev/sda3 VG Name VG_XenStorage-405a2ece-d10e-d6c5-ede2-e1ad2c29c68d PV Size 1.81 TB / not usable 6.97 MB Allocatable yes PE Size (KByte) 4096 Total PE 474747 Free PE 52441 Allocated PE 422306 PV UUID U03Gt9-WtHi-8Nnu-QB2Q-c7BV-CO9A-cFpYWW # xe sr-introduce name-label="VMs" type=lvm uuid=U03Gt9-WtHi-8Nnu-QB2Q-c7BV-CO9A-cFpYWW name-description="VMs Local HD Storage" content-type=user shared=false device-config=:device=/dev/disk/by-id/scsi-3600188b04c02f100181ab3a483f9f0ae-part3 U03Gt9-WtHi-8Nnu-QB2Q-c7BV-CO9A-cFpYWW # xe pbd-create host-uuid=a019d93e-4d84-4a4b-91e3-23572b5bd8a4 sr-uuid=U03Gt9-WtHi-8Nnu-QB2Q-c7BV-CO9A-cFpYWW device-config:device=/dev/disk/by-id/scsi-3600188b04c02f100181ab3a483f9f0ae-part3 adf92b7f-ad40-828f-0728-caf94d2a0ba1 # xe pbd-plug uuid=adf92b7f-ad40-828f-0728-caf94d2a0ba1 Error code: SR_BACKEND_FAILURE_47 Error parameters: , The SR is not available [opterr=no such volume group: VG_XenStorage-U03Gt9-WtHi-8Nnu-QB2Q-c7BV-CO9A-cFpYWW] At this point I did a # vgrename VG_XenStorage-405a2ece-d10e-d6c5-ede2-e1ad2c29c68d VG_XenStorage-U03Gt9-WtHi-8Nnu-QB2Q-c7BV-CO9A-cFpYWW cause the VG name was different, but pdb-plug still gives me the same error. So, now I'm kinda lost about what to do, I'm not used to Xen and most sites I've been finding are really unhelpful. I hope someone can guide me in the right way to fix this. I cant lose those VMs (got backups, but from inside the guests, not the VMs themselves).

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  • SQL Server Licensing in a VMware vSphere Cluster

    - by Helvick
    If I have SQL Server 2008 instances running in virtual machines on a VMware vSphere cluster with vMotion\DRS enabled so that the VM's can (potentially) run on any one of the physical servers in the cluster what precisely are the license requirements? For example assume that I have 4 physical ESX Hosts with dual physical CPU's and 3 separate single vCPU Virtual Machines running SQL Server 2008 running in that cluster. How many SQL Standard Processor licenses would I need? Is it 3 (one per VM) or 12 (one per VM on each physical host) or something else? How many SQL Enterprise Processor licenses would I need? Is it 3 (one per VM) or 8 (one for each physical CPU in the cluster) or, again, something else? The range in the list prices for these options goes from $17k to $200k so getting it right is quite important. Bonus question: If I choose the Server+CAL licensing model do I need to buy multiple Server instance licenses for each of the ESX hosts (so 12 copies of the SQL Server Standard server license so that there are enough licenses on each host to run all VM's) or again can I just license the VM and what difference would using Enterprise per server licensing make? Edited to Add Having spent some time reading the SQL 2008 Licensing Guide (63 Pages! Includes Maps!*) I've come across this: • Under the Server/CAL model, you may run unlimited instances of SQL Server 2008 Enterprise within the server farm, and move those instances freely, as long as those instances are not running on more servers than the number of licenses assigned to the server farm. • Under the Per Processor model, you effectively count the greatest number of physical processors that may support running instances of SQL Server 2008 Enterprise at any one time across the server farm and assign that number of Processor licenses And earlier: ..For SQL Server, these rule changes apply to SQL Server 2008 Enterprise only. By my reading this means that for my 3 VM's I only need 3 SQL 2008 Enterprise Processor Licenses or one copy of Server Enterprise + CALs for the cluster. By implication it means that I have to license all processors if I choose SQL 2008 Standard Processor licensing or that I have to buy a copy of SQL Server 2008 Standard for each ESX host if I choose to use CALs. *There is a map to demonstrate that a Server Farm cannot extend across an area broader than 3 timezones unless it's in the European Free Trade Area, I wasn't expecting that when I started reading it.

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  • SQL Server Licensing in a VMware vSphere Cluster

    - by Helvick
    If I have SQL Server 2008 instances running in virtual machines on a VMware vSphere cluster with vMotion\DRS enabled so that the VM's can (potentially) run on any one of the physical servers in the cluster what precisely are the license requirements? For example assume that I have 4 physical ESX Hosts with dual physical CPU's and 3 separate single vCPU Virtual Machines running SQL Server 2008 running in that cluster. How many SQL Standard Processor licenses would I need? Is it 3 (one per VM) or 12 (one per VM on each physical host) or something else? How many SQL Enterprise Processor licenses would I need? Is it 3 (one per VM) or 8 (one for each physical CPU in the cluster) or, again, something else? The range in the list prices for these options goes from $17k to $200k so getting it right is quite important. Bonus question: If I choose the Server+CAL licensing model do I need to buy multiple Server instance licenses for each of the ESX hosts (so 12 copies of the SQL Server Standard server license so that there are enough licenses on each host to run all VM's) or again can I just license the VM and what difference would using Enterprise per server licensing make? Edited to Add Having spent some time reading the SQL 2008 Licensing Guide (63 Pages! Includes Maps!*) I've come across this: • Under the Server/CAL model, you may run unlimited instances of SQL Server 2008 Enterprise within the server farm, and move those instances freely, as long as those instances are not running on more servers than the number of licenses assigned to the server farm. • Under the Per Processor model, you effectively count the greatest number of physical processors that may support running instances of SQL Server 2008 Enterprise at any one time across the server farm and assign that number of Processor licenses And earlier: ..For SQL Server, these rule changes apply to SQL Server 2008 Enterprise only. By my reading this means that for my 3 VM's I only need 3 SQL 2008 Enterprise Processor Licenses or one copy of Server Enterprise + CALs for the cluster. By implication it means that I have to license all processors if I choose SQL 2008 Standard Processor licensing or that I have to buy a copy of SQL Server 2008 Standard for each ESX host if I choose to use CALs. *There is a map to demonstrate that a Server Farm cannot extend across an area broader than 3 timezones unless it's in the European Free Trade Area, I wasn't expecting that when I started reading it.

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  • Apache2 name based virtual host always redirect 301

    - by Francesco
    I've got a server (runnging Debian Squeeze) with Apache 2.2, there are 4 site running there. I'm using namebased virtulhosts because I've got a single IP. Initial configuration has been made with Webmin and probably something has been messed up.. firstdomain.com is my default domain and is working correctly, seconddomain.com is another site that is working. Now I want to add lastdomain.tk as a new site, so I've made this config file: root@webamp:/etc/apache2# cat sites-available/lastdomain.tk.conf <VirtualHost *:80> DocumentRoot /home/server/Condivisione/RAID/lastdomain.tk ServerName www.alazanes.tk ServerAlias alazanes.tk </VirtualHost> I've added it to enabled-sites and restarted apache. The problem is that if I go to lastdomain.tk (or www.lastdomain.tk) I'm redirected to firstdomain.com with a 301 redirect. Both lastdomain.tk and www.lastdomain.tk are A DNS records pointing to my IP address. Strange thing is that if a change DocumentRoot of lastdomain.tk to DocumentRoot /home/server/Condivisione/RAID/Sito_SecondDomain I correctly see seconddomain.com content without being redirected (lastdomain.tk is showed on address bar) These are the other configurations I'm using. root@webamp:/root# source /etc/apache2/envvars ; /usr/sbin/apache2 -S VirtualHost configuration: wildcard NameVirtualHosts and _default_ servers: *:443 webamp.firstdomain.com (/etc/apache2/sites-enabled/ssl.bbteam:1) *:80 is a NameVirtualHost default server firstdomain.com (/etc/apache2/sites-enabled/000-default:7) port 80 namevhost firstdomain.com (/etc/apache2/sites-enabled/000-default:7) port 80 namevhost www.lastdomain.tk (/etc/apache2/sites-enabled/lastdomain.tk.conf:1) ## other domains ## port 80 namevhost seconddomain.com (/etc/apache2/sites-enabled/seconddomain.com.conf:1) Syntax OK Content of default config file is root@webamp:/etc/apache2# cat sites-available/default <VirtualHost *:80> ServerAdmin [email protected] ServerName firstdomain.com ServerAlias www.firstdomain.com direct.firstdomain.com DocumentRoot /home/server/Condivisione/RAID/Sito_Web_Apache_su_80 ErrorLog /var/log/apache2/error.log LogLevel warn CustomLog /var/log/apache2/access.log combined </VirtualHost> content of second domain config file is root@webamp:/etc/apache2# cat sites-available/seconddomain.com.conf <VirtualHost *:80> DocumentRoot /home/server/Condivisione/RAID/Sito_SecondDomain ServerName seconddomain.com ServerAlias www.seconddomain.com direct.seconddomain.com #redirect 301 / http://www.seconddomain.com/ <Directory "/home/server/Condivisione/RAID/Sito_SecondDomain"> allow from all Options +Indexes </Directory> </VirtualHost> Probably a file permission problem? root@webamp:/root# ls -lh /home/server/Condivisione/RAID/ total 7.1M drwxrwxr-x 15 www-data server 4.0K Jun 5 13:29 Sito_SecondDomain drwxrwxrwx 23 server server 4.0K Jun 7 16:22 Sito_Web_Apache_su_80 drwxrwxr-x 17 www-data server 4.0K Jun 8 09:56 alazanes.tk Do someone have an idea of what is happening? Thanks, Francesco

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  • XRDP: window manager not starting

    - by niboshi
    I have setup my Ubuntu server so that I can connect and login to XRDP from Windows remote desktop. My problem is that after logging in, no window-manager is started. It only displays a single gnome-terminal with no border and gray meshed background. It seems that /usr/sbin/xrdp-sesman itself is running (from observation of ps and /var/run/xrdp/xrdp-sesman.pid). I put debugging line like touch /home/myname/aaaaa into ~/startwm.sh or /etc/xrdp/startwm.sh, but the file aaaaa did not generated after logging in, so these scripts have not been executed. (Both of them have chmod +x permission.) Am I missing some configuration file, or is there any way of further inspection? Any help is appreciated. Thanks. Contents of /etc/xrdp/sesman.ini [Globals] ListenAddress=127.0.0.1 ListenPort=3350 EnableUserWindowManager=0 # or 1 UserWindowManager=startwm.sh DefaultWindowManager=startwm.sh # or commented-out [Security] AllowRootLogin=1 MaxLoginRetry=4 TerminalServerUsers=tsusers TerminalServerAdmins=tsadmins [Sessions] MaxSessions=10 KillDisconnected=0 IdleTimeLimit=0 DisconnectedTimeLimit=0 [Logging] LogFile=/var/log/xrdp-sesman.log LogLevel=DEBUG EnableSyslog=0 SyslogLevel=DEBUG [X11rdp] param1=-bs param2=-ac param3=-nolisten param4=tcp [Xvnc] param1=-bs param2=-ac param3=-nolisten param4=tcp Contents of /var/log/xrdp-sesman.log after logging in: [20120402-21:29:34] [CORE ] starting sesman with pid 11064 [20120402-21:29:34] [INFO ] listening... [20120402-21:29:39] [INFO ] scp thread on sck 7 started successfully [20120402-21:29:39] [INFO ] granted TS access to user myname [20120402-21:29:39] [INFO ] starting Xvnc session... [20120402-21:29:40] [INFO ] starting xrdp-sessvc - xpid=11074 - wmpid=11073 [20120402-21:29:49] [INFO ] session 11072 - user myname- terminated Process tree Below is a part of ps aufx output during xrdp session: xrdp 12344 0.0 0.4 22856 8732 ? Sl Apr02 0:01 /usr/sbin/xrdp root 12346 0.0 0.0 15672 2000 ? S Apr02 0:00 /usr/sbin/xrdp-sesman root 24346 0.0 0.0 3780 872 ? S 00:00 0:00 \_ /usr/sbin/xrdp-sessvc 24348 24347 myname 24347 0.4 0.6 76468 13700 ? Sl 00:00 0:14 \_ gnome-terminal myname 24362 0.0 0.0 2220 716 ? S 00:00 0:00 | \_ gnome-pty-helper myname 24363 0.0 0.2 6912 5268 pts/13 Ss 00:00 0:00 | \_ bash myname 27902 0.0 0.0 2824 1096 pts/13 R+ 00:53 0:00 | \_ ps aufx myname 24348 0.0 0.9 24984 19216 ? S 00:00 0:01 \_ Xvnc :18 -geometry 1920x1080 -depth 24 -rfbauth /home/myname/.vnc/sesman_myname_passwd -bs -ac -nolisten tcp root 24349 0.0 0.0 16596 1304 ? Sl 00:00 0:00 \_ xrdp-chansrv Environment Ubuntu 11.10 Oneiric xrdp version: 0.5.0~20100303cvs-6ubuntu2

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  • Simple Backup Strategy for Amazon EC2 instances / volumes?

    - by minerj
    You have entered Introductory Backups for Amazon EC2 EBS-backed Windows Images 010... I have been browsing my brains out to find a simple backup strategy for our single windows 2008 server running SharePoint Services. This is an EBS-backed image of one server with one data volume. I don’t need anything exotic. I only need a “daily” backup (losing a day’s worth of data is not catastrophic). We have created and saved an EBS backed AMI image (Windows 2008) we are comfortable using. We started off making backups by simply creating a new EBS AMI image. This is really simple, but the running server is put offline during the first 10 – 15 minutes of creating the image – not ideal. The standard way of creating backups would seem to be creating snapshots of volumes attached to a running instance. Again it’s pretty simple and the server remains usable during the snapshot generation. The apparent Catch-22 is that you can’t simply launch a new instance directly from a snapshot. I know how to bundle a running instance to S3 storage and then register the AMI from the S3 bucket. This does allow me to capture a backup of a running instance and, if the running instance is lost, register the AMI from the S3 bucket and launch the new AMI to recover the instance, but this seems really convoluted and it seems ridiculous to have to juggle back and forth between the AWS Console and the S3 Organizer plug-in for Firefox to get this accomplished. (Please don't mention the command line approach, this is an 010 level course). From playing around with EBS-backed images, the following approach appears to work for me (all done within the AWS Console): 1.For your backups, simply snapshot the system volume (/dev/sda1) as needed. 2.If you lose your running instance, do the following: a.Create a new volume from your last snapshot backup b.Launch another instance of your starting AMI (must be EBS-backed) c.Stop this instance. d.Detach the existing system volume from the new stopped instance and discard. e.Attach the newly created volume as system volume (/dev/sda1) to the stopped instance. f.Re-start the new instance. I have tested this out a couple of times and it seems to work for me. Question: Is there anything wrong with this approach?

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  • Prevent Windows 7 User Accounts from accessing files in other User Accounts

    - by Mantis
    I'm trying to set up another User Account on my Windows 7 Professional laptop for use by another person. I do not want that person to have access to any of the files in my User Account on the same machine. This machine has a single hard disk formatted with NTFS. User accounts data is stored in the default location, C:\Users. I use the computer with a Standard Account (not an Administrator). Let's call my user account "User A." I have given the new user a Standard Account. Let's call the new user's account "User B." To be clear, I want User B to have the ability to log in to her account, to use the computer, but to be unable to access any of the files in the User A account on the same machine. Currently, User B cannot use Windows Explorer to navigate to the location C:\Users\User A. However, by simply using Windows Search, User B can easily find and open documents saved in C:\Users\User A\Documents. After opening a document, that document's full path appears in "Recent Places" in Windows Explorer, and the document appears as a file that can be opened using the "Recent" feature in Word 2010. This is not the desired behavior. User B should not have the ability to see any documents using Windows Search or anything else. I have attempted to set permissions using the following procedure. Using an Administrator account, navigate to C:\Users and right-click on the "User A" folder. Select "Properties." In the "User A Properties" window that appears, click the "Security" tab. Click the "Edit..." button to change permissions. IN the "Permissions for User B" window that appears, under "Group or User Names," select User B. Under "Permissions for User B", check the box under the "Deny" column for the "Full Control" row. Ensure that the "Deny" box is automatically checked for all the other rows, and then click "OK." The system should then begin working. The process could take several minutes. When I followed this procedure, I received several "Access Denied" errors, suggesting that the system was unable to set the permissions as I had directed. I think this might be one of the reasons why User B is still able to access files in User A's account folders. Is there any other way I could accomplish my goal here? Thank you.

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  • Double Layer DVD+R burning problem - I/O Error

    - by Mehper C. Palavuzlar
    I have a modern PC (Quad Core CPU, 4 GB RAM, Win7 Home Premium 64-bit) but I have a problem with burning .dvd images to Double Layer (8.5 GB) DVDs. I wasted too many DVD+R DL discs but to no avail. Here is a short explanation of what I did: I'm using ImgBurn v2.5.0.0 (latest version). I'm trying to burn an image file (.dvd) which is together with the related .iso file in the same folder. In ImgBurn, I select the file with .dvd extension, and set writing speed to 2.4x. Burning process starts normally, but around 7% of the process, it gives a I/O Write Error, which is as follows: I wasted 3 discs (Magic, Made in Taiwan, DVD+R DL, 8.5 GB) trying the same thing. My DVD writer is LG GH22NP20 with IDE connection type. I updated its firmware from 1.04 to 2.00 but no success in burning again. Then my cousin brought his LG (an older model) which, he claims, was successful in writing DL discs with the same brand (Magic). I plugged off my LG and plugged the older one in, and tried to burn the image again. It also gave an I/O Error even without standing till 7%. I tried another burning program (CloneCD), but failed again. Then I bought other brands (TDK and VERBATIM) and tried to burn the image. Burning process started successfully, but around 14% (for Verbatim) and 25% (for TDK) failed again. Here is a screeny from ImgBurn: I've burned lots of 4.7 GB DVD+Rs and DVD-Rs successfully, even without a single error, with this LG DVD writer, but this case is very bothering for me. What should I do? Should I buy a new DVD writer other than LG? Could this be related to Windows or my hardware configuration? Thanks for your help. Edit: My burner works on my cousin's machine. So the problem must be related to my system. What could be the reason? Latest news: I borrowed an external USB DVD writer from a friend, which is PHILIPS SPD3000CC (an old model). Guess what! It's burning DVD+R DLs successfully! How come an internal DVD writer of a brand new computer system cannot burn DL DVDs? Now I'm considering buying a new internal DVD writer with not IDE, but SATA connection...

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  • networking through openstack installed on vm

    - by Mandar Katdare
    I am trying to set up a test installation of Openstack on a Ubuntu 12.04 VM running on a ESXi server. So far I have been able to launch the VMs on the ESXi, however am unable to assign IP addresses to them. As the VM with the Openstack installation has a single public IP, I wish to assign IPs to the VMs create through Openstack so that they can directly interact with the public network itself without having a separate private network. So I feel that bridging would not be the correct option here. But am unable to find the correct documents to go ahead with such an install. My ifconfig looks as follows: eth0 Link encap:Ethernet HWaddr 00:0c:29:6f:8a:d7 inet addr:192.168.4.167 Bcast:192.168.4.255 Mask:255.255.255.0 inet6 addr: fe80::20c:29ff:fe6f:8ad7/64 Scope:Link UP BROADCAST RUNNING MULTICAST MTU:1500 Metric:1 RX packets:391640 errors:33 dropped:98 overruns:0 frame:0 TX packets:545044 errors:0 dropped:0 overruns:0 carrier:0 collisions:0 txqueuelen:1000 RX bytes:40303931 (40.3 MB) TX bytes:763127348 (763.1 MB) Interrupt:18 Base address:0x2000 lo Link encap:Local Loopback inet addr:127.0.0.1 Mask:255.0.0.0 inet6 addr: ::1/128 Scope:Host UP LOOPBACK RUNNING MTU:16436 Metric:1 RX packets:146127 errors:0 dropped:0 overruns:0 frame:0 TX packets:146127 errors:0 dropped:0 overruns:0 carrier:0 collisions:0 txqueuelen:0 RX bytes:799815763 (799.8 MB) TX bytes:799815763 (799.8 MB) virbr0 Link encap:Ethernet HWaddr 8a:80:33:32:63:a0 UP BROADCAST MULTICAST MTU:1500 Metric:1 RX packets:0 errors:0 dropped:0 overruns:0 frame:0 TX packets:0 errors:0 dropped:0 overruns:0 carrier:0 collisions:0 txqueuelen:0 RX bytes:0 (0.0 B) TX bytes:0 (0.0 B) The eth0 is the adapter that I intend to use for all communication. My nova.conf looks as follows: --dhcpbridge_flagfile=/etc/nova/nova.conf --dhcpbridge=/usr/bin/nova-dhcpbridge --logdir=/var/log/nova --state_path=/var/lib/nova --lock_path=/var/lock/nova --allow_admin_api=true --use_deprecated_auth=false --auth_strategy=keystone --scheduler_driver=nova.scheduler.simple.SimpleScheduler --s3_host=192.168.4.167 --ec2_host=192.168.4.167 --rabbit_host=192.168.4.167 --cc_host=192.168.4.167 --nova_url=http://192.168.4.167:8774/v1.1/ --routing_source_ip=192.168.4.167 --glance_api_servers=192.168.4.167:9292 --image_service=nova.image.glance.GlanceImageService --iscsi_ip_prefix=192.168.4 --sql_connection=mysql://novadbadmin:[email protected]/nova --ec2_url=http://192.168.4.167:8773/services/Cloud --keystone_ec2_url=http://192.168.4.167:5000/v2.0/ec2tokens --api_paste_config=/etc/nova/api-paste.ini --libvirt_type=kvm --libvirt_use_virtio_for_bridges=true --start_guests_on_host_boot=true --resume_guests_state_on_host_boot=true --vnc_enabled=true --vncproxy_url=http://192.168.4.167:6080 --vnc_console_proxy_url=http://192.168.4.167:6080 # network specific settings --network_manager=nova.network.manager.FlatDHCPManager --public_interface=eth0 --vmwareapi_host_ip=192.168.4.254 --vmwareapi_host_username=**** --vmwareapi_host_password=**** --vmwareapi_wsdl_loc=http://127.0.0.1:8080/wsdl/vim25/vimService.wsdl --fixed_range=192.168.4.190/24 --floating_range=192.168.4.190/24 --network_size=32 --flat_network_dhcp_start=192.168.4.190 --flat_injected=False --force_dhcp_release --iscsi_helper=tgtadm --connection_type=vmwareapi --root_helper=sudo nova-rootwrap --verbose --libvirt_use_virtio_for_bridges --ec2_private_dns_show --novnc_enabled=true --novncproxy_base_url=http://192.168.4.167:6080/vnc_auto.html --vncserver_proxyclient_address=192.168.4.167 --vncserver_listen=192.168.4.167 192.168.4.167 is my VM with the Openstack installation and 192.168.4.254 is my ESXi server on which the VM runs. Can anyone advice me about how to proceed? Thanks, Mandar

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  • What is the difference between Anycast and GeoDNS / GeoIP wrt HA?

    - by Riyad
    Based on the Wikipedia description of Anycast, it includes both the distribution of a domain-name-to-many-IP-mapping across many DNS servers as well as replying to clients with the most geographically close (or fastest) server. In the context of a globally distributed, highly available site like google.com (or any CDN service with many global edge locations) this sounds like the two key features one would need. DNS services like Amazon's Route53, EasyDNS and DNSMadeEasy all advertise themselves as Anycast-enabled networks. Therefore my assumption is that each of these DNS services transparently offer me those two killer features: multi-IP-to-domain mapping AND routing clients to the closest node. However, each of these services seem to separate out these two functionalities, referring to the 2nd one (routing clients to closest node) as "GeoDNS", "GeoIP" or "Global Traffic Director" and charge extra for the service. If a core tenant of an Anycast-capable system is to already do this, why is this functionality being earmarked as this extra feature? What is this "GeoDNS" feature doing that a standard Anycast DNS service won't do (according to the definition of Anycast from Wikipedia -- I understand what is being advertised, just not why it isn't implied already). I get extra-confused when a DNS service like Route53 that doesn't support this nebulous "GeoDNS" feature lists functionality like: Fast – Using a global anycast network of DNS servers around the world, Route 53 is designed to automatically route your users to the optimal location depending on network conditions. As a result, the service offers low query latency for your end users, as well as low update latency for your DNS record management needs. ... which sounds exactly like what GeoDNS is intended to do, but geographically directing clients is something they explicitly don't support it yet. Ultimately I am looking for the two following features from a DNS provider: Map multiple IP addresses to a single domain name (like google.com, amazon.com, etc. does) Utilize a DNS service that will respond to client requests for that domain with the IP address of the nearest server to the requestee. As mentioned, it seems like this is all part of an "Anycast" DNS service (all of which these services are), but the features and marketing I see from them suggest otherwise, making me think I need to learn a bit more about how DNS works before making a deployment choice. Thanks in advance for any clarifications.

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  • Getting macro keys from a razer blackwidow to work on linux

    - by Journeyman Geek
    I picked up a razer blackwidow ultimate that has additional keys meant for macros that are set using a tool that's installed on windows. I'm assuming that these arn't some fancypants joojoo keys and should emit scancodes like any other keys. Firstly is there a standard way to check these scancodes in linux? Secondly how do i set these keys to do things in command line and x based linux setups? My current linux install is xubuntu 10.10, but i'll be switching to kubuntu once i have a few things fixed up. Ideally the answer should be generic and system-wide Things i have tried so far: showkeys from the built in kbd package (in a seperate vt) - macro keys not detected xev - macro keys not detected lsusb and evdev output this ahk script's output suggests the M keys are not outputting standard scancodes Things i need to try snoopy pro + reverse engineering (oh dear) Wireshark - preliminary futzing around seems to indicate no scancodes emitted when what i seem to think is the keyboard is monitored and keys pressed. Might indicate additional keys are a seperate device or need to be initialised somehow. Need to cross reference that with lsusb output from linux, in 3 scenarios - standalone, passed through to a windows VM without the drivers installed, and the same with. LSUSB only detects one device on a standalone linux install It might be useful to check if the mice use the same razer synapse driver , since that means some variation of razercfg might work (not detected. only seems to work for mice) Things i have Have worked out: In a windows system with the driver, the keyboard is seen as a keyboard and a pointing device. And said pointing device uses, in addition to your bog standard mouse drivers.. a driver for something called a razer synapse. Mouse driver seen in linux under evdev and lsusb as well Single Device under OS X apparently, though i have yet to try lsusb equivilent on that Keyboard goes into pulsing backlight mode in OS X upon initialisation with the driver. This should probably indicate that there's some initialisation sequence sent to the keyboard on activation. They are, in fact, fancypants joojoo keys. Extending this question a little I have access to a windows system so if i need to use any tools on that to help answer the question, its fine. I can also try it on systems with and without the config utility. The expected end result is still to make those keys usable on linux however. I also realise this is a very specific family of hardware. I would be willing to test anything that makes sense on a linux system if i have detailed instructions - this should open up the question to people who have linux skills, but no access to this keyboard The minimum end result i require I need these keys detected, and usable in any fashion on any of the current graphical mainstream ubuntu varients

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  • Linux boot on a raid1 software raid ?

    - by azera
    Hello I am trying to convert my single disk boot to a raid1 boot So far here is what i have: I sucessfully create the raid 1 as degraded with the new drive alone, I copied all the data on it I can mount that raid 1, see its files etc I already have a raid5 that is working on the same box (although not booting on it) I have installed grub on both drive When grub boot, it loads the kernel alright, but during the kernel boot it fails to load the "root block device" The kernel tells me : 1 - detected that root device is an md device 2 - determining root devices 3 - mounting root 4 - mounting /dev/md125 on /newroot failed: input/output error. Please enter another root device: ... At this point, if I enter /dev/sda3 (my "old" root device that isn't converted to raid yet) everything boots fine without the root. The /dev/md125 device is indeed created but it seems to be created after the error happens, as in it creates it after loading the device, when mdadm is loaded. Somehow it looks like it can't/doesn't load the raid array before it needs to mount it, and I don't know how I can solve that. My config files (taken from the system once it boots with sda3 as root device): $ cat /etc/mdadm.conf ARRAY /dev/md/md0-r5 metadata=0.90 UUID=1a118934:c831bdb3:64188b84:66721085 ARRAY /dev/md125 metadata=0.90 UUID=48ec4190:a80d4dde:64188b84:66721085 $ cat /proc/mdstat Personalities : [raid1] [raid6] [raid5] [raid4] [raid0] [raid10] md125 : active raid1 sdc3[1] 477853312 blocks [2/1] [_U] md127 : active raid5 sdd[0] sdf[3] sdb[2] sde[1] 4395415488 blocks level 5, 64k chunk, algorithm 2 [4/4] [UUUU] unused devices: <none> $ cat /boot/grub/menu.lst default 0 timeout 8 splashimage=(hd0,0)/boot/grub/splash.xpm.gz title Gentoo Linux 2.6.31-r10 root (hd0,0) #kernel /boot/kernel-genkernel-x86_64-2.6.31-gentoo-r10 root=/dev/ram0 real_root=/dev/sda3 kernel /boot/kernel-genkernel-x86_64-2.6.31-gentoo-r10 root=/dev/md125 md=125,/dev/sdc3,/dev/sda3 initrd /boot/initramfs-genkernel-x86_64-2.6.31-gentoo-r10 # blkid /dev/sda1: UUID="89fee223-b845-4e0a-8a0b-e6cf695d5bcf" TYPE="ext2" /dev/sda2: UUID="a72296a8-d7d4-447f-a34b-ee920fd1a767" TYPE="swap" /dev/sda3: UUID="97eb0a6a-c385-4a9d-bf74-c0bab1fa4dc1" TYPE="ext3" /dev/sdb: UUID="1a118934-c831-bdb3-6418-8b8466721085" TYPE="linux_raid_member" /dev/sdc1: UUID="d36537fd-19a0-b8a3-6418-8b8466721085" TYPE="linux_raid_member" /dev/sdd: UUID="1a118934-c831-bdb3-6418-8b8466721085" TYPE="linux_raid_member" /dev/sde: UUID="1a118934-c831-bdb3-6418-8b8466721085" TYPE="linux_raid_member" /dev/md127: UUID="13a41589-4cf1-4c04-91ca-37484182c783" TYPE="ext4" /dev/sdf: UUID="1a118934-c831-bdb3-6418-8b8466721085" TYPE="linux_raid_member" /dev/sdc2: UUID="a1916397-1b48-45d7-9f98-73aa521e882f" TYPE="swap" /dev/sdc3: UUID="48ec4190-a80d-4dde-6418-8b8466721085" TYPE="linux_raid_member" /dev/md125: UUID="c947ed64-1d4d-4d1d-b4d2-24669fff916e" SEC_TYPE="ext2" TYPE="ext3" # mdadm -E mdadm: No devices to examine # fdisk -l Disk /dev/sda: 500.1 GB, 500107862016 bytes 255 heads, 63 sectors/track, 60801 cylinders Units = cylinders of 16065 * 512 = 8225280 bytes Disk identifier: 0xe975e9fc Device Boot Start End Blocks Id System /dev/sda1 1 5 40131 83 Linux /dev/sda2 6 1311 10490445 82 Linux swap / Solaris /dev/sda3 1312 60801 477853425 83 Linux Disk /dev/sdc: 500.1 GB, 500107862016 bytes 255 heads, 63 sectors/track, 60801 cylinders Units = cylinders of 16065 * 512 = 8225280 bytes Disk identifier: 0xe975e9fc Device Boot Start End Blocks Id System /dev/sdc1 1 5 40131 83 Linux /dev/sdc2 6 1311 10490445 82 Linux swap / Solaris /dev/sdc3 1312 60801 477853425 83 Linux Disk /dev/md125: 489.3 GB, 489321791488 bytes 2 heads, 4 sectors/track, 119463328 cylinders Units = cylinders of 8 * 512 = 4096 bytes Disk identifier: 0x00000000 Disk /dev/md125 doesn't contain a valid partition table

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