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  • PowerShell – Show a Notification Balloon

    - by BuckWoody
    In my presentations for PowerShell I sometimes want to start a process (like a backup) that will take some time. I normally pop up a notification “balloon” at the start, and then do the bulk of the work, and then pop up a balloon at the end to let me know it’s done. You can actually try out this little sample (on a test system, of course) without any other code to see what it does. Then just put the other PowerShell commands in the #Do Some Work part. Oh – throw an icon (.ico file) in a c:\temp directory or point that somewhere else. (No, this probably isn’t original. Can’t remember where I saw the original code, but I’ve modified it a bit anyway, so if you’re the original author and this looks slightly familiar, post a comment.) [void] [System.Reflection.Assembly]::LoadWithPartialName("System.Windows.Forms") $objBalloon = New-Object System.Windows.Forms.NotifyIcon $objBalloon.Icon = "C:\temp\Folder.ico" # You can use the value Info, Warning, Error $objBalloon.BalloonTipIcon = "Info" # Put what you want to say here for the Start of the process $objBalloon.BalloonTipTitle = "Begin Title" $objBalloon.BalloonTipText = "Begin Message" $objBalloon.Visible = $True $objBalloon.ShowBalloonTip(10000) # Do some work # Put what you want to say here for the completion of the process $objBalloon.BalloonTipTitle = "End Title" $objBalloon.BalloonTipText = "End Message" $objBalloon.Visible = $True $objBalloon.ShowBalloonTip(10000) Script Disclaimer, for people who need to be told this sort of thing: Never trust any script, including those that you find here, until you understand exactly what it does and how it will act on your systems. Always check the script on a test system or Virtual Machine, not a production system. Yes, there are always multiple ways to do things, and this script may not work in every situation, for everything. It’s just a script, people. All scripts on this site are performed by a professional stunt driver on a closed course. Your mileage may vary. Void where prohibited. Offer good for a limited time only. Keep out of reach of small children. Do not operate heavy machinery while using this script. If you experience blurry vision, indigestion or diarrhea during the operation of this script, see a physician immediately. Share this post: email it! | bookmark it! | digg it! | reddit! | kick it! | live it!

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  • Storing game objects with generic object information

    - by Mick
    In a simple game object class, you might have something like this: public abstract class GameObject { protected String name; // other properties protected double x, y; public GameObject(String name, double x, double y) { // etc } // setters, getters } I was thinking, since a lot of game objects (ex. generic monsters) will share the same name, movement speed, attack power, etc, it would be better to have all that information shared between all monsters of the same type. So I decided to have an abstract class "ObjectData" to hold all this shared information. So whenever I create a generic monster, I would use the same pre-created "ObjectData" for it. Now the above class becomes more like this: public abstract class GameObject { protected ObjectData data; protected double x, y; public GameObject(ObjectData data, double x, double y) { // etc } // setters, getters public String getName() { return data.getName(); } } So to tailor this specifically for a Monster (could be done in a very similar way for Npcs, etc), I would add 2 classes. Monster which extends GameObject, and MonsterData which extends ObjectData. Now I'll have something like this: public class Monster extends GameObject { public Monster(MonsterData data, double x, double y) { super(data, x, y); } } This is where my design question comes in. Since MonsterData would hold data specific to a generic monster (and would vary with what say NpcData holds), what would be the best way to access this extra information in a system like this? At the moment, since the data variable is of type ObjectData, I'll have to cast data to MonsterData whenever I use it inside the Monster class. One solution I thought of is this, but this might be bad practice: public class Monster extends GameObject { private MonsterData data; // <- this part here public Monster(MonsterData data, double x, double y) { super(data, x, y); this.data = data; // <- this part here } } I've read that for one I should generically avoid overwriting the underlying classes variables. What do you guys think of this solution? Is it bad practice? Do you have any better solutions? Is the design in general bad? How should I redesign this if it is? Thanks in advanced for any replies, and sorry about the long question. Hopefully it all makes sense!

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  • Parallel MSBuild FTW - Build faster in parallel

    - by deadlydog
    Hey everyone, I just discovered this great post yesterday that shows how to have msbuild build projects in parallel Basically all you need to do is pass the switches “/m:[NumOfCPUsToUse] /p:BuildInParallel=true” into MSBuild. Example to use 4 cores/processes (If you just pass in “/m” it will use all CPU cores): MSBuild /m:4 /p:BuildInParallel=true "C:\dev\Client.sln" Obviously this trick will only be useful on PCs with multi-core CPUs (which we should all have by now) and solutions with multiple projects; So there’s no point using it for solutions that only contain one project.  Also, testing shows that using multiple processes does not speed up Team Foundation Database deployments either in case you’re curious Also, I found that if I didn’t explicitly use “/p:BuildInParallel=true” I would get many build errors (even though the MSDN documentation says that it is true by default). The poster boasts compile time improvements up to 59%, but the performance boost you see will vary depending on the solution and its project dependencies.  I tested with building a solution at my office, and here are my results (runs are in seconds): # of Processes 1st Run 2nd Run 3rd Run Avg Performance 1 192 195 200 195.67 100% 2 155 156 156 155.67 79.56% 4 146 149 146 147.00 75.13% 8 136 136 138 136.67 69.85%   So I updated all of our build scripts to build using 2 cores (~20% speed boost), since that gives us the biggest bang for our buck on our solution without bogging down a machine, and developers may sometimes compile more than 1 solution at a time.  I’ve put the any-PC-safe batch script code at the bottom of this post. The poster also has a follow-up post showing how to add a button and keyboard shortcut to the Visual Studio IDE to have VS build in parallel as well (so you don’t have to use a build script); if you do this make sure you use the .Net 4.0 MSBuild, not the 3.5 one that he shows in the screenshot.  While this did work for me, I found it left an MSBuild.exe process always hanging around afterwards for some reason, so watch out (batch file doesn’t have this problem though).  Also, you do get build output, but it may not be the same that you’re used to, and it doesn’t say “Build succeeded” in the status bar when completed, so I chose to not make this my default Visual Studio build option, but you may still want to. Happy building! ------------------------------------------------------------------------------------- :: Calculate how many Processes to use to do the build. SET NumberOfProcessesToUseForBuild=1  SET BuildInParallel=false if %NUMBER_OF_PROCESSORS% GTR 2 (                 SET NumberOfProcessesToUseForBuild=2                 SET BuildInParallel=true ) MSBuild /maxcpucount:%NumberOfProcessesToUseForBuild% /p:BuildInParallel=%BuildInParallel% "C:\dev\Client.sln"

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  • How can a solo programmer become a good team player?

    - by Nick
    I've been programming (obsessively) since I was 12. I am fairly knowledgeable across the spectrum of languages out there, from assembly, to C++, to Javascript, to Haskell, Lisp, and Qi. But all of my projects have been by myself. I got my degree in chemical engineering, not CS or computer engineering, but for the first time this fall I'll be working on a large programming project with other people, and I have no clue how to prepare. I've been using Windows all of my life, but this project is going to be very unix-y, so I purchased a Mac recently in the hopes of familiarizing myself with the environment. I was fortunate to participate in a hackathon with some friends this past year -- both CS majors -- and excitingly enough, we won. But I realized as I worked with them that their workflow was very different from mine. They used Git for version control. I had never used it at the time, but I've since learned all that I can about it. They also used a lot of frameworks and libraries. I had to learn what Rails was pretty much overnight for the hackathon (on the other hand, they didn't know what lexical scoping or closures were). All of our code worked well, but they didn't understand mine, and I didn't understand theirs. I hear references to things that real programmers do on a daily basis -- unit testing, code reviews, but I only have the vaguest sense of what these are. I normally don't have many bugs in my little projects, so I have never needed a bug tracking system or tests for them. And the last thing is that it takes me a long time to understand other people's code. Variable naming conventions (that vary with each new language) are difficult (__mzkwpSomRidicAbbrev), and I find the loose coupling difficult. That's not to say I don't loosely couple things -- I think I'm quite good at it for my own work, but when I download something like the Linux kernel or the Chromium source code to look at it, I spend hours trying to figure out how all of these oddly named directories and files connect. It's a programming sin to reinvent the wheel, but I often find it's just quicker to write up the functionality myself than to spend hours dissecting some library. Obviously, people who do this for a living don't have these problems, and I'll need to get to that point myself. Question: What are some steps that I can take to begin "integrating" with everyone else? Thanks!

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  • Enterprise Service Bus (ESB): Important architectural piece to a SOA or is it just vendor hype?

    Is an Enterprise Service Bus (ESB) an important architectural piece to a Service-Oriented Architecture (SOA), or is it just vendor hype in order to sell a particular product such as SOA-in-a-box? According to IBM.com, an ESB is a flexible connectivity infrastructure for integrating applications and services; it offers a flexible and manageable approach to service-oriented architecture implementation. With this being said, it is my personal belief that ESBs are an important architectural piece to any SOA. Additionally, generic design patterns have been created around the integration of web services in to ESB regardless of any vendor. ESB design patterns, according to Philip Hartman, can be classified in to the following categories: Interaction Patterns: Enable service interaction points to send and/or receive messages from the bus Mediation Patterns: Enable the altering of message exchanges Deployment Patterns: Support solution deployment into a federated infrastructure Examples of Interaction Patterns: One-Way Message Synchronous Interaction Asynchronous Interaction Asynchronous Interaction with Timeout Asynchronous Interaction with a Notification Timer One Request, Multiple Responses One Request, One of Two Possible Responses One Request, a Mandatory Response, and an Optional Response Partial Processing Multiple Application Interactions Benefits of the Mediation Pattern: Mediator promotes loose coupling by keeping objects from referring to each other explicitly, and it lets you vary their interaction independently Design an intermediary to decouple many peers Promote the many-to-many relationships between interacting peers to “full object status” Examples of Interaction Patterns: Global ESB: Services share a single namespace and all service providers are visible to every service requester across an entire network Directly Connected ESB: Global service registry that enables independent ESB installations to be visible Brokered ESB: Bridges services that are reluctant to expose requesters or providers to ESBs in other domains Federated ESB: Service consumers and providers connect to the master or to a dependent ESB to access services throughout the network References: Mediator Design Pattern. (2011). Retrieved 2011, from SourceMaking.com: http://sourcemaking.com/design_patterns/mediator Hartman, P. (2006, 24 1). ESB Patterns that "Click". Retrieved 2011, from The Art and Science of Being an IT Architect: http://artsciita.blogspot.com/2006/01/esb-patterns-that-click.html IBM. (2011). WebSphere DataPower XC10 Appliance Version 2.0. Retrieved 2011, from IBM.com: http://publib.boulder.ibm.com/infocenter/wdpxc/v2r0/index.jsp?topic=%2Fcom.ibm.websphere.help.glossary.doc%2Ftopics%2Fglossary.html Oracle. (2005). 12 Interaction Patterns. Retrieved 2011, from Oracle® BPEL Process Manager Developer's Guide: http://docs.oracle.com/cd/B31017_01/integrate.1013/b28981/interact.htm#BABHHEHD

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  • What are some reasonable stylistic limits on type inference?

    - by Jon Purdy
    C++0x adds pretty darn comprehensive type inference support. I'm sorely tempted to use it everywhere possible to avoid undue repetition, but I'm wondering if removing explicit type information all over the place is such a good idea. Consider this rather contrived example: Foo.h: #include <set> class Foo { private: static std::set<Foo*> instances; public: Foo(); ~Foo(); // What does it return? Who cares! Just forward it! static decltype(instances.begin()) begin() { return instances.begin(); } static decltype(instances.end()) end() { return instances.end(); } }; Foo.cpp: #include <Foo.h> #include <Bar.h> // The type need only be specified in one location! // But I do have to open the header to find out what it actually is. decltype(Foo::instances) Foo::instances; Foo() { // What is the type of x? auto x = Bar::get_something(); // What does do_something() return? auto y = x.do_something(*this); // Well, it's convertible to bool somehow... if (!y) throw "a constant, old school"; instances.insert(this); } ~Foo() { instances.erase(this); } Would you say this is reasonable, or is it completely ridiculous? After all, especially if you're used to developing in a dynamic language, you don't really need to care all that much about the types of things, and can trust that the compiler will catch any egregious abuses of the type system. But for those of you that rely on editor support for method signatures, you're out of luck, so using this style in a library interface is probably really bad practice. I find that writing things with all possible types implicit actually makes my code a lot easier for me to follow, because it removes nearly all of the usual clutter of C++. Your mileage may, of course, vary, and that's what I'm interested in hearing about. What are the specific advantages and disadvantages to radical use of type inference?

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  • How to become a good team player?

    - by Nick
    I've been programming (obsessively) since I was 12. I am fairly knowledgeable across the spectrum of languages out there, from assembly, to C++, to Javascript, to Haskell, Lisp, and Qi. But all of my projects have been by myself. I got my degree in chemical engineering, not CS or computer engineering, but for the first time this fall I'll be working on a large programming project with other people, and I have no clue how to prepare. I've been using Windows all of my life, but this project is going to be very unix-y, so I purchased a Mac recently in the hopes of familiarizing myself with the environment. I was fortunate to participate in a hackathon with some friends this past year -- both CS majors -- and excitingly enough, we won. But I realized as I worked with them that their workflow was very different from mine. They used Git for version control. I had never used it at the time, but I've since learned all that I can about it. They also used a lot of frameworks and libraries. I had to learn what Rails was pretty much overnight for the hackathon (on the other hand, they didn't know what lexical scoping or closures were). All of our code worked well, but they didn't understand mine, and I didn't understand theirs. I hear references to things that real programmers do on a daily basis -- unit testing, code reviews, but I only have the vaguest sense of what these are. I normally don't have many bugs in my little projects, so I have never needed a bug tracking system or tests for them. And the last thing is that it takes me a long time to understand other people's code. Variable naming conventions (that vary with each new language) are difficult (__mzkwpSomRidicAbbrev), and I find the loose coupling difficult. That's not to say I don't loosely couple things -- I think I'm quite good at it for my own work, but when I download something like the Linux kernel or the Chromium source code to look at it, I spend hours trying to figure out how all of these oddly named directories and files connect. It's a programming sin to reinvent the wheel, but I often find it's just quicker to write up the functionality myself than to spend hours dissecting some library. Obviously, people who do this for a living don't have these problems, and I'll need to get to that point myself. Question: What are some steps that I can take to begin "integrating" with everyone else? Thanks!

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  • How are Reads Distributed in a Workload

    - by Bill Graziano
    People have uploaded nearly one millions rows of trace data to TraceTune.  That’s enough data to start to look at the results in aggregate.  The first thing I want to look at is logical reads.  This is the easiest metric to identify and fix. When you upload a trace, I rank each statement based on the total number of logical reads.  I also calculate each statement’s percentage of the total logical reads.  I do the same thing for CPU, duration and logical writes.  When you view a statement you can see all the details like this: This single statement consumed 61.4% of the total logical reads on the system while we were tracing it.  I also wanted to see the distribution of reads across statements.  That graph looks like this: On average, the highest ranked statement consumed just under 50% of the reads on the system.  When I tune a system, I’m usually starting in one of two modes: this “piece” is slow or the whole system is slow.  If a given piece (screen, report, query, etc.) is slow you can usually find the specific statements behind it and tune it.  You can make that individual piece faster but you may not affect the whole system. When you’re trying to speed up an entire server you need to identity those queries that are using the most disk resources in aggregate.  Fixing those will make them faster and it will leave more disk throughput for the rest of the queries. Here are some of the things I’ve learned querying this data: The highest ranked query averages just under 50% of the total reads on the system. The top 3 ranked queries average 73% of the total reads on the system. The top 10 ranked queries average 91% of the total reads on the system. Remember these are averages across all the traces that have been uploaded.  And I’m guessing that people mainly upload traces where there are performance problems so your mileage may vary. I also learned that slow queries aren’t the problem.  Before I wrote ClearTrace I used to identify queries by filtering on high logical reads using Profiler.  That picked out individual queries but those rarely ran often enough to put a large load on the system. If you look at the execution count by rank you’d see that the highest ranked queries also have the highest execution counts.  The graph would look very similar to the one above but flatter.  These queries don’t look that bad individually but run so often that they hog the disk capacity. The take away from all this is that you really should be tuning the top 10 queries if you want to make your system faster.  Tuning individually slow queries will help those specific queries but won’t have much impact on the system as a whole.

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  • Unable to ping inside or outside network with default gateway 0.0.0.0

    - by agentroadkill
    I've been around here before and I could usually piece together everything to more or less get myself up and running, but this time I'm truly stumped. I'm trying to connect my new 14.04 install to a network, and I'm forced to be behind my college's router. Now I've tested the vary cable that is right now plugged into my Ubuntu box on a Windows, Mac OS X, and even my friend's Ubuntu 14.04 box, and they all connect no problem. I've been trying to track this down for about two days, but every time I get close to it, the bug jumps to some other piece of my connection. Anyway, as it sits ifconfig -a gives: eth2 Lninkencap:Ethernet HWaddr:00:1f:bc:08:31:1d inet addr:10.32.51.51 Bcast:10.32.51.155 Mask: 255.255.255.0 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 RX bytes:0 TX bytes:0 as well as the local loopback, but I'm assuming that is not an issue here. sudo dhclient -v eth2 returns: Listening on LPF/<hardware address of my integrated NIC, above> Sending on <same> Sending on Socket/fallback DHCPREQUEST of 10.32.51.51 on eth2 to 255.255.255.255 port 67 (xid=0x6f4a66ba) <two more lines of same> DHCPDISCOVER on eth2 to 255.255.255.255 port 67 interval 3 (xid=0x156f9fb4) <many more of above with varying intervals> No DHCPOFFERS received. Trying recorded lease 10.32.51.51 RTNETLINK answers: File exists bound: renewal in <large number> seconds If I then try ping 8.8.8.8, I get: connect: Network is unreachable /etc/resolv.conf only contains the two lines telling you not to edit it, while /etc/network/interfaces only has the loopback interface block in it. I've tried commenting out the "option rfc3442" line in /etc/dhcp/dhclient.conf which seemed to fix this issue for many people, as well as adding the line send vendor-class-indentifier "MSFT5.0" to dhclient.conf as well to tell the router I'm a windows box, in case they don't like Linux. Finally, route -n reveals: Destination Gateway Genmask Flags Metric Ref Use Iface 10.32.51.0 0.0.0.0 255.255.255.0 U 0 0 0 eth2 I would like to apologize in advance for the doubtless butchered text alignment, but I'm obviously typing this all by hand, reading from the terminal as I type commands. I'm hoping this is an interesting problem, and not something I blithely stumbled past in my (apparent) over-confidence. TIA! Quick addendum before posting: The activity light on the ethernet port are lit and one blinks during boot, but they rarely (and seemingly randomly) do so afterwards (both are dark) even while running dhclient in the foreground. When I had the Ubuntu box tethered to my MacBook earlier, I got what looked like a normal power/uplink blinking pattern, but was unable to ping one from the other.

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  • 2D camera perspective projection from 3D coordinates -- HOW?

    - by Jack
    I am developing a camera for a 2D game with a top-down view that has depth. It's almost a 3D camera. Basically, every object has a Z even though it is in 2D, and similarly to parallax layers their position, scale and rotation speed vary based on their Z. I guess this would be a perspective projection. But I am having trouble converting the objects' 3D coordinates into the 2D space of the screen so that everything has correct perspective and scale. I never learned matrices though I did dig the topic a bit today. I tried without using matrices thanks to this article but every attempt gave awkward results. I'm using ActionScript 3 and Flash 11+ (Starling), where the screen coordinates work like this: Left-handed coordinates system illustration I can explain further what I did if you want to help me sort out what's wrong, or you can directly tell me how you would do it properly. In case you prefer the former, read on. These are images showing the formulas I used: upload.wikimedia.org/math/1/c/8/1c89722619b756d05adb4ea38ee6f62b.png upload.wikimedia.org/math/d/4/0/d4069770c68cb8f1aa4b5cfc57e81bc3.png (Sorry new users can't post images, but both are from the wikipedia article linked above, section "Perspective projection". That's where you'll find what all variables mean, too) The long formula is greatly simplified because I believe a normal top-down 2D camera has no X/Y/Z rotation values (correct ?). Then it becomes d = a - c. Still, I can't get it to work. Maybe you could explain what numbers I should put in a(xyz), c(xyz), theta(xyz), and particularly, e(xyz) ? I don't quite get how e is different than c in my case. c.z is also an issue to me. If the Z of the camera's target object is 0, should the camera's Z be something like -600 ? ( = focal length of 600) Whatever I do, it's wrong. I only got it to work when I used arbitrary calculations that "looked" right, like most cameras with parallax layers seem to do, but that's fake! ;) If I want objects to travel between Z layers I might as well do it right. :) Thanks a lot for your help!

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  • Download files from a SharePoint site using the RSSBus SSIS Components

    - by dataintegration
    In this article we will show how to use a stored procedure included in the RSSBus SSIS Components for SharePoint to download files from SharePoint. While the article uses the RSSBus SSIS Components for SharePoint, the same process will work for any of our SSIS Components. Step 1: Open Visual Studio and create a new Integration Services Project. Step 2: Add a new Data Flow Task to the Control Flow screen and open the Data Flow Task. Step 3: Add an RSSBus SharePoint Source to the Data Flow Task. Step 4: In the RSSBus SharePoint Source, add a new Connection Manager, and add your credentials for the SharePoint site. Step 5: Now from the Table or View dropdown, choose the name of the Document Library that you are going to back up and close the wizard. Step 6: Add a Script Component to the Data Flow Task and drag an output arrow from the 'RSSBus SharePoint Source' to it. Step 7: Open the Script Component, go to edit the Input Columns, and choose all the columns. Step 8: This will open a new Visual Studio instance, with a project in it. In this project add a reference to the RSSBus.SSIS2008.SharePoint assembly available in the RSSBus SSIS Components for SharePoint installation directory. Step 9: In the 'ScriptMain' class, add the System.Data.RSSBus.SharePoint namespace and go to the 'Input0_ProcessInputRow' method (this method's name may vary depending on the input name in the Script Component). Step 10: In the 'Input0_ProcessInputRow' method, you can add code to use the DownloadDocument stored procedure. Below we show the sample code: String connString = "Offline=False;Password=PASSWORD;User=USER;URL=SHAREPOINT-SITE"; String downloadDir = "C:\\Documents\\"; SharePointConnection conn = new SharePointConnection(connString); SharePointCommand comm = new SharePointCommand("DownloadDocument", conn); comm.CommandType = CommandType.StoredProcedure; comm.Parameters.Clear(); String file = downloadDir+Row.LinkFilenameNoMenu.ToString(); comm.Parameters.Add(new SharePointParameter("@File", file)); String list = Row.ServerUrl.ToString().Split('/')[1].ToString(); comm.Parameters.Add(new SharePointParameter("@Library", list)); String remoteFile = Row.LinkFilenameNoMenu.ToString(); comm.Parameters.Add(new SharePointParameter("@RemoteFile", remoteFile)); comm.ExecuteNonQuery(); After saving your changes to the Script Component, you can execute the project and find the downloaded files in the download directory. SSIS Sample Project To help you with getting started using the SharePoint Data Provider within SQL Server SSIS, download the fully functional sample package. You will also need the SharePoint SSIS Connector to make the connection. You can download a free trial here. Note: Before running the demo, you will need to change your connection details in both the 'Script Component' code and the 'Connection Manager'.

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  • Support ARMv7 instruction set in Windows Embedded Compact applications

    - by Valter Minute
    On of the most interesting new features of Windows Embedded Compact 7 is support for the ARMv5, ARMv6 and ARMv7 instruction sets instead of the ARMv4 “generic” support provided by the previous releases. This means that code build for Windows Embedded Compact 7 can leverage features (like the FPU unit for ARMv6 and v7) and instructions of the recent ARM cores and improve their performances. Those improvements are noticeable in graphics, floating point calculation and data processing. The ARMv7 instruction set is supported by the latest Cortex-A8, A9 and A15 processor families. Those processor are currently used in tablets, smartphones, in-car navigation systems and provide a great amount of processing power and a low amount of electric power making them very interesting for portable device but also for any kind of device that requires a rich user interface, processing power, connectivity and has to keep its power consumption low. The bad news is that the compiler provided with Visual Studio 2008 does not provide support for ARMv7, building native applications using just the ARMv4 instruction set. Porting a Visual Studio “Smart Device” native C/C++ project to Platform Builder is not easy and you’ll lack many of the features that the VS2008 application development environment provides. You’ll also need access to the BSP and OSDesign configuration for your device to be able to build and debug your application inside Platform Builder and this may prevent independent software vendors from using the new compiler to improve their applications performances. Adeneo Embedded now provides a whitepaper and a Visual Studio plug-in that allows usage of the new ARMv7 enabled compiler to build applications inside Visual Studio 2008. I worked on the whitepaper and the tools, with the help of my colleagues and now the results can be downloaded from Adeneo Embedded’s website: http://www.adeneo-embedded.com/OS-Technologies/Windows-Embedded (Click on the “WEC7 ARMv7 Whitepaper tab to access the download links, free registration required) A very basic benchmark showed a very good performance improvement in integer and floating-point operations. Obviously your mileage may vary and we can’t promise the same amount of improvement on any application, but with a small effort on your side (even smaller if you use the plug-in) you can try on your own application. ARMv7 support is provided using Platform Builder’s compiler and VS2008 application debugger is not able to debut ARMv7 code, so you may need to put in place some workaround like keeping ARMv4 code for debugging etc.

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  • Travelling at Magenic

    - by Chris G. Williams
    I occasionally get asked if we travel "a lot" at Magenic. Sometimes the question comes from job candidates. Other times it's clients, recruiters or friends. To give a simple yes or no answer would be a disservice to the person asking the question. So here is my standard answer:It depends.(That was the short version.  Here's the long version...)We do have some guys that are more "national" in focus, and they can travel a fair amount. They also receive a little extra in compensation for doing so. It's a balancing act, and not necessarily a one-size-fits-all situation. Not everyone is well suited to constant travel. Some folks enjoy it and some folks hate it.With our local guys, our general policy is to TRY and keep them close to home whenever possible, but sometimes the needs of the client will dictate otherwise. (As Spock would say... the needs of the many outweigh the needs of the few, or the one.)In most cases though, we really do try to avoid sending our guys on extended travel gigs (i.e. every week for 6 months) when a simple kickoff trip and occasional visit will do. This depends on the nature of the gig, of course. Some types of work lend themselves to this model better than others. Additionally, this can and does vary by office. If one office is having trouble staffing a gig (not enough available bodies) and another office has a few too many folks on the bench, well... you can connect the dots. But again, we try to keep that to a minimum.Lastly, we all have our own thresholds for what we consider "a lot" of travel. There are two parts to this threshold:Half of it is whatever you're accustomed to already. The other half is being honest with yourself about how much you [like/hate] dealing with airports, car rentals, taxis, hotels, disruptions to your workout schedule, time away from friends/family, etc.Knowing a bit about yourself will definitely help you decide how much travel is too much for you.

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  • Setting MTU on Exalogic

    - by csoto
    For many reasons, a system administrator may want to change the MTU settings of a server. But in a system like Exalogic which contains lots of interconnected nodes and other various components, it's important to understand how this applies to the different networks. For example, when bringing up bonding of InfiniBand an error like the following may be thrown: Bringing up interface bond1: SIOCSIFMTU: Invalid argument Both scripts ifcfg-ib0 and ifcfg-ib1 (from the /etc/sysconfig/network-scripts/ direectory) have MTU set to 65500, which is a valid MTU value only if all IPoIB slaves operate in connected mode and are configured with the same value, so the line below must be added to both network scripts and then restart the network: CONNECTED_MODE=yes By the way, an error of the form “SIOCSIFMTU: Invalid argument” indicates that the requested MTU was rejected by the kernel. Typically this would be due to it exceeding the maximum value supported by the interface hardware. In that case you must either reduce the MTU to a value that is supported or obtain more capable hardware. This problem has been seen when trying to modify the MTU using the ifconfig command, like the output of the example below: [root@elxxcnxx ~]# ifconfig ib1 mtu 65520 SIOCSIFMTU: Invalid argument It's important to insist that in most cases the nodes must be rebooted after the MTU size has been changed. Although in some circumstances it may work without a reboot, it is not how it is typically documented. Now, in order to achieve a reduced memory consumption and improve performance for network traffic received on IPoIB related interfaces, it is recommend to reduce the MTU value in interface configuration files for IPoIB related bonds from 65520 to 64000. The change needs to be made to interface configuration files under the /etc/sysconfig/network-scripts directory and applies to the interface configuration files for bonds over IPoIB related slave devices, for example /etc/sysconfig/network-scripts/ifcfg-bond1. However, keep in mind that the numeric portion of the interface filenames that corresponding to IPoIB interfaces is expected to vary across compute nodes and vServers and so cannot be relied upon to identify which interface files are for bonds are over IPoIB rather than EoIB related slave interfaces. To fix these MTU values to the recommended settings, there are very useful instructions and a script on the MOS Note 1624434.1, and it's applicable physical and virtual configurations of Exalogic. Regarding the recommended MTU value for EoIB related interfaces, its maximum appropriate value is 1500. If for some reason a vServer has been created with a higher value (set on the /etc/sysconfig/network-scripts/ifcfg-bond0 file), then it must be fixed. An error like the following could be thrown under this circumstance: [root@vServer ~]# service network restart ... Bringing up interface bond0:  SIOCSIFMTU: Invalid argument Also an error like the one below can be seen on the /var/log/messages file of the vServer: kernel: T5074835532 [mlx4_vnic] eth1:vnic_change_mtu:360: failed: new_mtu 64000 2026 The MOS Note 1611657.1 is very useful for this purpose.

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  • MVC Can the model know ANYTHING about the view?

    - by AwDogsGo2Heaven
    I'm working on a game, and without getting into any details I am using MVC "patterns", "rules" or whatever you want to call it to make the game. The view includes everything needed to draw things on the screen, the controller passes input to the model. And the Model contains game logic. Here's my problem, the game is being made for mobile devices that vary in screen sizes. I feel my model needs to know the view size so it can appropriately adjust for it. I've thought about it for a while I could put all the adjustments in the view, but it just seems inefficient to translate the model positioning to the view's needed positioning every time. If the model knows the size it only needed to adjust itself once. So my question is can I pass the view size to the model without 'breaking' MVC? I feel personally they are still decoupled this way because a size is just a number, I could still change the view any time as long as it has a size. But I wanted to get a community response on this because I haven't seen many discussions of MVC being used in a game. (And to be clear I don't want an answer of why I shouldn't use it in a game, but do I break MVC by letting it know the view's size) EDIT - More details on the game's needs and why I wanted to pass the view. Some objects positions need to be set relative to the edge of the screen (such as UI elements) so that they appear relatively in the same place. Sprite sizes are not stretched if the window size is wider such as an IPhone 5 screen. They will just be placed relatively to the screen edge. .If I gave it to the view to handle this, I will need a flag saying that this element is positioned say x number of pixels from TOP BOTTOM RIGHT LEFT. Then allow the view to draw it with that information. Its acceptable, I just wanted to know if there was a better way while still being MVC because it seems this way will be repeating a logic over and over, where as if I knew the view size in the model, I could convert all the relative positions into absolute positions in one run, but with this I have to convert on every update to the screen.

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  • Using dd-wrt Dynamic DNS client with CloudFlare

    - by Roman
    I'm trying to configure Dynamic DNS client on my router with dd-wrt (v24-sp2) firmware so it would dynamically change IP address in one of the DNS records. Unfortunately I encountered a problem… Here is an example request from their ddclient configuration: https://www.cloudflare.com/api.html?a=DIUP&u=<my_login>&tkn=<my_token>&ip=<my_ip>&hosts=<my_record> It works if I use it in browser, but in dd-wrt I get this output: Tue Jan 24 00:36:47 2012: INADYN: Started 'INADYN Advanced version 1.96-ADV' - dynamic DNS updater. Tue Jan 24 00:36:47 2012: I:INADYN: IP address for alias '<my_record>' needs update to '<my_ip>' Tue Jan 24 00:36:48 2012: W:INADYN: Error validating DYNDNS svr answer. Check usr,pass,hostname! (HTTP/1.1 303 See Other Server: cloudflare-nginx Date: Mon, 23 Jan 2012 14:36:48 GMT Content-Type: text/plain Connection: close Expires: Sun, 25 Jan 1981 05:00:00 GMT Cache-Control: no-store, no-cache, must-revalidate, post-check=0, pre-check=0 Pragma: no-cache Location: https://www.cloudflare.com/api.html?a=DIUP&u=<my_login>&tkn=<my_token>&ip=<my_ip>&hosts=<my_record> Vary: Accept-Encoding Set-Cookie: __cfduid=<id>; expires=Mon, 23-Dec-2019 23:50:00 GMT; path=/; domain=.cloudflare.com Set-Cookie: __cfduid=<id>; expires=Mon, 23-Dec-2019 23:50:00 GMT; path=/; domain=.www.cloudflare.com You must include an `a' paramiter, with a value of DIUP|wl|chl|nul|ban|comm_news|devmode|sec_lvl|ipv46|ob|cache_lvl|fpurge_ts|async|pre_purge|minify|stats|direct|zone_check|zone_ips|zone_errors|zone_agg|zone_search|zone_time|zone_grab|app|rec_se URL from "Location" works perfectly and parameter "a" is included. What's the problem?

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  • How and where to implement basic authentication in Kibana 3

    - by Jabb
    I have put my elasticsearch server behind a Apache reverse proxy that provides basic authentication. Authenticating to Apache directly from the browser works fine. However, when I use Kibana 3 to access the server, I receive authentication errors. Obviously because no auth headers are sent along with Kibana's Ajax calls. I added the below to elastic-angular-client.js in the Kibana vendor directory to implement authentication quick and dirty. But for some reason it does not work. $http.defaults.headers.common.Authorization = 'Basic ' + Base64Encode('user:Password'); What is the best approach and place to implement basic authentication in Kibana? /*! elastic.js - v1.1.1 - 2013-05-24 * https://github.com/fullscale/elastic.js * Copyright (c) 2013 FullScale Labs, LLC; Licensed MIT */ /*jshint browser:true */ /*global angular:true */ 'use strict'; /* Angular.js service wrapping the elastic.js API. This module can simply be injected into your angular controllers. */ angular.module('elasticjs.service', []) .factory('ejsResource', ['$http', function ($http) { return function (config) { var // use existing ejs object if it exists ejs = window.ejs || {}, /* results are returned as a promise */ promiseThen = function (httpPromise, successcb, errorcb) { return httpPromise.then(function (response) { (successcb || angular.noop)(response.data); return response.data; }, function (response) { (errorcb || angular.noop)(response.data); return response.data; }); }; // check if we have a config object // if not, we have the server url so // we convert it to a config object if (config !== Object(config)) { config = {server: config}; } // set url to empty string if it was not specified if (config.server == null) { config.server = ''; } /* implement the elastic.js client interface for angular */ ejs.client = { server: function (s) { if (s == null) { return config.server; } config.server = s; return this; }, post: function (path, data, successcb, errorcb) { $http.defaults.headers.common.Authorization = 'Basic ' + Base64Encode('user:Password'); console.log($http.defaults.headers); path = config.server + path; var reqConfig = {url: path, data: data, method: 'POST'}; return promiseThen($http(angular.extend(reqConfig, config)), successcb, errorcb); }, get: function (path, data, successcb, errorcb) { $http.defaults.headers.common.Authorization = 'Basic ' + Base64Encode('user:Password'); path = config.server + path; // no body on get request, data will be request params var reqConfig = {url: path, params: data, method: 'GET'}; return promiseThen($http(angular.extend(reqConfig, config)), successcb, errorcb); }, put: function (path, data, successcb, errorcb) { $http.defaults.headers.common.Authorization = 'Basic ' + Base64Encode('user:Password'); path = config.server + path; var reqConfig = {url: path, data: data, method: 'PUT'}; return promiseThen($http(angular.extend(reqConfig, config)), successcb, errorcb); }, del: function (path, data, successcb, errorcb) { $http.defaults.headers.common.Authorization = 'Basic ' + Base64Encode('user:Password'); path = config.server + path; var reqConfig = {url: path, data: data, method: 'DELETE'}; return promiseThen($http(angular.extend(reqConfig, config)), successcb, errorcb); }, head: function (path, data, successcb, errorcb) { $http.defaults.headers.common.Authorization = 'Basic ' + Base64Encode('user:Password'); path = config.server + path; // no body on HEAD request, data will be request params var reqConfig = {url: path, params: data, method: 'HEAD'}; return $http(angular.extend(reqConfig, config)) .then(function (response) { (successcb || angular.noop)(response.headers()); return response.headers(); }, function (response) { (errorcb || angular.noop)(undefined); return undefined; }); } }; return ejs; }; }]); UPDATE 1: I implemented Matts suggestion. However, the server returns a weird response. It seems that the authorization header is not working. Could it have to do with the fact, that I am running Kibana on port 81 and elasticsearch on 8181? OPTIONS /solar_vendor/_search HTTP/1.1 Host: 46.252.46.173:8181 User-Agent: Mozilla/5.0 (Windows NT 6.1; WOW64; rv:25.0) Gecko/20100101 Firefox/25.0 Accept: text/html,application/xhtml+xml,application/xml;q=0.9,*/*;q=0.8 Accept-Language: de-de,de;q=0.8,en-us;q=0.5,en;q=0.3 Accept-Encoding: gzip, deflate Origin: http://46.252.46.173:81 Access-Control-Request-Method: POST Access-Control-Request-Headers: authorization,content-type Connection: keep-alive Pragma: no-cache Cache-Control: no-cache This is the response HTTP/1.1 401 Authorization Required Date: Fri, 08 Nov 2013 23:47:02 GMT WWW-Authenticate: Basic realm="Username/Password" Vary: Accept-Encoding Content-Encoding: gzip Content-Length: 346 Connection: close Content-Type: text/html; charset=iso-8859-1 UPDATE 2: Updated all instances with the modified headers in these Kibana files root@localhost:/var/www/kibana# grep -r 'ejsResource(' . ./src/app/controllers/dash.js: $scope.ejs = ejsResource({server: config.elasticsearch, headers: {'Access-Control-Request-Headers': 'Accept, Origin, Authorization', 'Authorization': 'Basic XXXXXXXXXXXXXXXXXXXXXXXXXXXXX=='}}); ./src/app/services/querySrv.js: var ejs = ejsResource({server: config.elasticsearch, headers: {'Access-Control-Request-Headers': 'Accept, Origin, Authorization', 'Authorization': 'Basic XXXXXXXXXXXXXXXXXXXXXXXXXXXXX=='}}); ./src/app/services/filterSrv.js: var ejs = ejsResource({server: config.elasticsearch, headers: {'Access-Control-Request-Headers': 'Accept, Origin, Authorization', 'Authorization': 'Basic XXXXXXXXXXXXXXXXXXXXXXXXXXXXX=='}}); ./src/app/services/dashboard.js: var ejs = ejsResource({server: config.elasticsearch, headers: {'Access-Control-Request-Headers': 'Accept, Origin, Authorization', 'Authorization': 'Basic XXXXXXXXXXXXXXXXXXXXXXXXXXXXX=='}}); And modified my vhost conf for the reverse proxy like this <VirtualHost *:8181> ProxyRequests Off ProxyPass / http://127.0.0.1:9200/ ProxyPassReverse / https://127.0.0.1:9200/ <Location /> Order deny,allow Allow from all AuthType Basic AuthName “Username/Password” AuthUserFile /var/www/cake2.2.4/.htpasswd Require valid-user Header always set Access-Control-Allow-Methods "GET, POST, DELETE, OPTIONS, PUT" Header always set Access-Control-Allow-Headers "Content-Type, X-Requested-With, X-HTTP-Method-Override, Origin, Accept, Authorization" Header always set Access-Control-Allow-Credentials "true" Header always set Cache-Control "max-age=0" Header always set Access-Control-Allow-Origin * </Location> ErrorLog ${APACHE_LOG_DIR}/error.log </VirtualHost> Apache sends back the new response headers but the request header still seems to be wrong somewhere. Authentication just doesn't work. Request Headers OPTIONS /solar_vendor/_search HTTP/1.1 Host: 46.252.26.173:8181 User-Agent: Mozilla/5.0 (Windows NT 6.1; WOW64; rv:25.0) Gecko/20100101 Firefox/25.0 Accept: text/html,application/xhtml+xml,application/xml;q=0.9,*/*;q=0.8 Accept-Language: de-de,de;q=0.8,en-us;q=0.5,en;q=0.3 Accept-Encoding: gzip, deflate Origin: http://46.252.26.173:81 Access-Control-Request-Method: POST Access-Control-Request-Headers: authorization,content-type Connection: keep-alive Pragma: no-cache Cache-Control: no-cache Response Headers HTTP/1.1 401 Authorization Required Date: Sat, 09 Nov 2013 08:48:48 GMT Access-Control-Allow-Methods: GET, POST, DELETE, OPTIONS, PUT Access-Control-Allow-Headers: Content-Type, X-Requested-With, X-HTTP-Method-Override, Origin, Accept, Authorization Access-Control-Allow-Credentials: true Cache-Control: max-age=0 Access-Control-Allow-Origin: * WWW-Authenticate: Basic realm="Username/Password" Vary: Accept-Encoding Content-Encoding: gzip Content-Length: 346 Connection: close Content-Type: text/html; charset=iso-8859-1 SOLUTION: After doing some more research, I found out that this is definitely a configuration issue with regard to CORS. There are quite a few posts available regarding that topic but it appears that in order to solve my problem, it would be necessary to to make some very granular configurations on apache and also make sure that the right stuff is sent from the browser. So I reconsidered the strategy and found a much simpler solution. Just modify the vhost reverse proxy config to move the elastisearch server AND kibana on the same http port. This also adds even better security to Kibana. This is what I did: <VirtualHost *:8181> ProxyRequests Off ProxyPass /bigdatadesk/ http://127.0.0.1:81/bigdatadesk/src/ ProxyPassReverse /bigdatadesk/ http://127.0.0.1:81/bigdatadesk/src/ ProxyPass / http://127.0.0.1:9200/ ProxyPassReverse / https://127.0.0.1:9200/ <Location /> Order deny,allow Allow from all AuthType Basic AuthName “Username/Password” AuthUserFile /var/www/.htpasswd Require valid-user </Location> ErrorLog ${APACHE_LOG_DIR}/error.log </VirtualHost>

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  • pfSense Firewall or Linsys/Cisco router for small offices

    - by Tim Meers
    I'm about to start switching some networks around for multiple small offices. Each office has about 10 to 15 users and 10 to 15 computers. Each office has a spread of generic routers and access points. The routers vary from being used as routers, to just being an access point for wireless. Nothing formal has really ever beem implemented for each of the 10 offices. What I'm wanting is to set up a pfSense box for each office to configure things like: traffic shaping (for VoIP QOS) URL Filtering DHCP static routing multiple VLANs I'll then use some of the existing hardware for wireless. Maybe even integrate the wireless right into the firewall depending on the office layout. So my question, would this be better to do a full blown firewall box, or but a new business class or high end consumer class Linksys router to do the URL filtering, QOS and DHPC? Each option could allow for remote access and VPN for remote maintnance and each would only cost a nominal about of money for something decent, i.e. under $250.

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  • USB Hardware vs. Software Write Lock

    - by TreyK
    I'm in the market for a USB flash drive, and remember this cool feature a tiny 32MB flash drive of mine had: a write lock switch. This seemed like it would be an amazing feature to have as a shield against any nastiness happening to the drive on an unfamiliar computer. However, very few drives on the market offer this feature. Instead, it seems that forms of software protection are the more prominent method. This software protection causes me a bit of uneasiness, as it seems like this software wouldn't be nearly as bulletproof as a physical switch. Also, levels of protection seem to vary from product to product. Being able to protect certain folders from reading and/or writing would be nice, but is the security trade-off worth it? Just how effective can this software protection be? Wouldn't a simple format be able to clean any drive with software protection? My drive must also be compatible with Windows XP, Vista, and 7, as well as Linux and Mac. What would be the best way forward for getting a well-sized (~8GB) flash drive with a strong write protection implementation, for little or no more than a regular drive? Thanks.

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  • Is there a way to tell what the download speed is from a site/server

    - by Memor-X
    i'm looking into which ISP i should go with at the place i'm moving into, one ISP which i have been told good things about has data limits (which when breached will drop your speed to dial-up speed) but multiple memberships which, apart from the cheapest membership, have the same data limits (the cheapest has a 10GB data limit) in their fine print, they say that each different membership has different port speeds, one particular part jumps out at me These speeds are the NBN (National Broadband Network) port speed and not the actual Internet data speed which will vary based on numerous factors including destination you are reaching, your network equipment, network congestion etc. i plan to use the net to download DLC and patch updates for games (particular the insanely large update for the Wii U) and games from Steam (if i find any good one other than this one JRPG) and downloading development resources from free sites like Deposit Files and Mediafire since one membership with a 1000GB data limit is $145 with the port speed being 12Mbps/1Mbps (cheapest) while another with the same limit is $190 with the port speed of 100Mbps/40Mbps (expensive) i am wondering how i can tell what the speed coming from site is since i don't want to be wasting money on speed that makes no difference (unlike memory which i rather have to spare) NOTE: the speeds are for a fiber optic network which where my new place is can only connect via fixed wireless which i may not be able to get with this ISP but if i can get this network then good NOTE 2: most of the resources i get from Deposit Files are always about 200 MB or less, if a resource pack is greater then it's split into multiple archives (like .7z.part) while Mediafire i have to see one bigger than 150MB NOTE 3: one update patch for a PS3 game is close to 4 GB (Disgaea 4) which i need to get access to the DLC and on the weekend i downloaded 5 GB for the Final Fantasy XIV Open Beta for the PS3 which took almost 5 hours

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  • Apache Reverse Proxy not working inside a VirtualHost running a Mono Web Application

    - by Arwen
    I have a mono web application running with this virtual host below. It is running on Apache 2.2.20 / Ubuntu 11.10. I tried to add a reverse proxy inside this virtualhost so I can make asynchronous or AJAX type calls back to this same domain. My asynchronous requests would have problems in many browsers calling services that are on another domain (cross domain requests problem). I am wanting to do reverse proxy calls to this other service using http://www.whatever.com/monkey/. So, I added the directive and top directive to try to make this work. It is weird though...nothing I do seems to have any effect. I can put the exact same markup in my default website virtualhost file and it works great. What is the deal? Are some of these Mono directives causing problems? <VirtualHost *:80> ServerName www.whatever.com ServerAlias whatever.com *.whatever.com ServerAdmin [email protected] DocumentRoot /home/myuser/web/whatever ProxyRequests off <Proxy *> Order allow,deny Allow from all </Proxy> <Location /monkey/> ProxyPass http://www.google.com/ ProxyPassReverse http://www.google.com/ </Location> MonoServerPath www.whatever.com "/usr/bin/mod-mono-server2" MonoSetEnv www.whatever.com MONO_IOMAP=all MonoApplications www.whatever.com "/:/home/myuser/web/whatever" <Location "/"> Allow from all Order allow,deny MonoSetServerAlias www.whatever.com SetHandler mono SetOutputFilter DEFLATE SetEnvIfNoCase Request_URI "\.(?:gif|jpe?g|png)$" no-gzip dont-vary </Location> <IfModule mod_deflate.c> AddOutputFilterByType DEFLATE text/html text/plain text/xml text/javascript </IfModule> </VirtualHost>

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  • How to generate customized sudoers files in puppet depending on the environment they're deployed to?

    - by gozu
    the sysadmins are present in the sudoers files of all environments, but other sudoers are not. Different environments all have slightly different sudoers. Most of the time, 90% of users are the same, and 10% vary so we cannot have only one sudoers file for everything. Right now, we are using puppet with 10 different files with names like sudoers.production1, sudoers.production2, sudoers.production3, sudoers.testing1, sudoers.staging1 and so forth. Puppet then picks the file to deploy based on the server's $domain (ex: dbserver.staging1.acme.com) or $hardwaremodel. It works fine but it's a nightmare to maintain so many files. I'd like to autogenerate sudoers files based on the server's domain and have only one big file with all the sudoers permissions for all users and all environments. Something that looks like: User_Alias ADMINS = abe, bob, carol, dave case $domain { "staging1.acme.com" { #add dev1,dev2,tester1,tester2 to sudoers file } "testing2.acme.com" { #add tester1, tester3, tester4 to sudoers file } What's the best way to go about this? Suggestions for alternatives are welcome. I'd appreciate any tips. Update 1: For security reasons, we'd rather not concatenate a bunch of files from a folder located on a puppet client in case someone puts a file in there (maliciously or not) and either breaks the combined file or inserts something in it. Most importantly, for usability, we'd like to keep the number of sudoers related files (fragment or complete) on puppet server to either 3 (prod/stage/test) or preferably 1 file. this file would (somehow) generate sudoers files on the puppet server and send one customized file to each puppet client. The purpose of this would be only searching for a username in a single file and removing it quicker than doing it on 11 files. When adding a user to a bunch of environments, it won't be as quick, but only one file would need to be opened and looked at, greatly reducing the chances of an omission. our Sudo version is 1.6.9p8 so we can't use /sudoers.d folder, only a sudoers file.

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  • bind9 "error sending response: host unreachable"

    - by wolfgangsz
    of course), I have a number of DNS servers, all running bind9 (9.5.1, to be specific) under fedora. 4 of them are slaves, fed by a common master for our public DNS. These are all located on the public gateways of our various offices. One of them has tons of messages in its log files similar to these: Jul 21 17:26:18 gateway named[3487]: client 10.171.3.8#52500: view internal: error sending response: host unreachable I wonder where that comes from. The firewall is open on port 53 between the two machines (10.171.3.8 is an internal DNS server located on a Windows Domain Controller). The internal domains do NOT list the gateway as a name server (so there should not be any attempts of replicating the domains), and the gateway does not handle any internal DNS. The clients in these messages vary between the two domain controllers on the internal network and a third internal name server (running bind9 on debian in a different segment of the network). Any pointers are highly welcome. In response to the first reply: The issue with this really is that tcpdump doesn't show any problems. Here is an extract from "tcpdump -i any port 53" 09:13:38.283308 IP valine.aminocom.com.61815 ns-pri.ripe.net.domain: 14075 PTR? 166.225.58.95.in-addr.arpa. (44) 09:13:42.007410 IP gateway-eng.aminocom.com.37047 alanine.aminocom.com.domain: 35410+ PTR? 12.3.172.10.in-addr.arpa. (42) At the same time, the DNS log shows: Jul 22 09:13:38 gateway named[3487]: client 10.171.3.6#61300: view internal: error sending response: host unreachable Jul 22 09:13:40 gateway named[3487]: client 10.172.3.12#56230: view internal: error sending response: host unreachable Jul 22 09:13:40 gateway named[3487]: client 10.171.3.8#55221: view internal: error sending response: host unreachable Jul 22 09:13:49 gateway named[3487]: client 10.171.3.8#51342: view internal: error sending response: host unreachable So clearly at 09:13:40 there were two unsuccessful attempts to connect to internal machines (10.172.3.12 and 10.171.3.8, both are DNS servers), but nothing in the tcpdump output.

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  • Windows Terminal Server: occasional memory violation for applications

    - by syneticon-dj
    On a virtualized (ESXi 4.1) Windows Server 2008 SP2 32-bit machine which is used as a terminal server, I occasionally (approximately 1-3 event log entries a day) see applications fail with an 0xc0000005 error - apparently a memory access violation. The problem seems quite random and only badly reproducable - applications may run for hours, fail with 0xc0000005 and restart quite fine or just throw the access violation at startup and start flawlessly at the second attempt. The names of executables, modules and offset addresses vary, although a single executable tends to fail with same modules and the same memory offset addresses (like "OUTLOOK.EXE" repeatedly failing on module "olmapi32.dll" with the offset "0x00044b7a") - even across multiple user's logons and with several days passing without a single failure inbetween. The offset addresses seem to change across reboots, however. Only selective executables seem affected by the problem, although I may simply not be seeing a sufficient number of application runs from the other ones. I first suspected a possible problem with the physical machine's RAM, but ruled this out as a rather unlikely cause - the memory comes with ECC and I've already moved the virtual machine across several times, without any perceptable change. I've seen that DEP was enabled in "OptOut" mode on this machine: C:\Users\administrator>wmic OS Get DataExecutionPrevention_SupportPolicy DataExecutionPrevention_SupportPolicy 3 and tried changing the policy to OptIn via startup options: bcdedit.exe /set {current} nx OptIn but have yet to see any effect - I also would expect Outlook 12 or Adobe Reader 9 (both affected applications) to play well with DEP. Any other ideas why the apps may be failing?

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  • How can a CentOS 6 guest running in VirtualBox be configured as a LAMP server that can be accessed from the Windows host?

    - by jtt89
    I was able to conect Centos6 on Virtual Box to Windows (I can ping in both directions) with Host-only Adapter (for connection between the two) and NAT Adapter (to enable Linux on VB to connect to the Internet). I want to set up httpd, mysql and vsftpd servers and in the end easily connect to httpd from Windows based browser and ftp server with a Windows based client as well. I would also want to have access through SSH. I have a general idea of the steps that are involved, but there is also a configuration that I am not sure about at this point. Lets say I follow these steps: yum install httpd yum install php php-pear php-mysql yum install mysql-server mysql_secure_installation yum install vsftpd yum install mod_ssl Technically I have everything installed, but what would be the next steps that I need to take (from the networking point of view, so to speak) to get it all working)? I know I need to configure, at least Apache, and ftp server, but I am not sure how is it gonna work; like where am I gonna be uloading the sites (I know this can vary), how am I gonna know what address to use in a browser if I wanna go to a website x, y, z on that installation etc. This sounds like I need to do some kind of DNS setup and I am kind of stuck at this point. If somebody could give me a general outline of what are the things that need to be done that would be great (I was looking at a lot of websites and I know about etc/sysconfig/network, httpd.config - not too much about it on Apache's site, hostname, hostname -f etc; but it is kind of hard to piece it all together at this point). I am gonna be looking at the books also, but they not always reflect the setup that I have too (VirtualBox). Thank you.

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