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  • How to apply %% to a variable in a script with input parameter?

    - by murxx
    Hi, I asked a question yesterday about how to manage to get %% around a variable without getting the evaluation. Well, the thing is, it does not work that way in my case... I have a .bat file which gets an input parameter. Later on I want to use the value of this input parameter and put %...% around, like: call script.bat testValue script.bat: set inputPar=%1 set newValue=%%inputPar%% Now I get: echo %inputPar% testValue echo %newValue% %inputPar% But I would like to get: echo %newValue% %testValue% Is that somehow possible?

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  • Red Gate Coder interviews: Alex Davies

    - by Michael Williamson
    Alex Davies has been a software engineer at Red Gate since graduating from university, and is currently busy working on .NET Demon. We talked about tackling parallel programming with his actors framework, a scientific approach to debugging, and how JavaScript is going to affect the programming languages we use in years to come. So, if we start at the start, how did you get started in programming? When I was seven or eight, I was given a BBC Micro for Christmas. I had asked for a Game Boy, but my dad thought it would be better to give me a proper computer. For a year or so, I only played games on it, but then I found the user guide for writing programs in it. I gradually started doing more stuff on it and found it fun. I liked creating. As I went into senior school I continued to write stuff on there, trying to write games that weren’t very good. I got a real computer when I was fourteen and found ways to write BASIC on it. Visual Basic to start with, and then something more interesting than that. How did you learn to program? Was there someone helping you out? Absolutely not! I learnt out of a book, or by experimenting. I remember the first time I found a loop, I was like “Oh my God! I don’t have to write out the same line over and over and over again any more. It’s amazing!” When did you think this might be something that you actually wanted to do as a career? For a long time, I thought it wasn’t something that you would do as a career, because it was too much fun to be a career. I thought I’d do chemistry at university and some kind of career based on chemical engineering. And then I went to a careers fair at school when I was seventeen or eighteen, and it just didn’t interest me whatsoever. I thought “I could be a programmer, and there’s loads of money there, and I’m good at it, and it’s fun”, but also that I shouldn’t spoil my hobby. Now I don’t really program in my spare time any more, which is a bit of a shame, but I program all the rest of the time, so I can live with it. Do you think you learnt much about programming at university? Yes, definitely! I went into university knowing how to make computers do anything I wanted them to do. However, I didn’t have the language to talk about algorithms, so the algorithms course in my first year was massively important. Learning other language paradigms like functional programming was really good for breadth of understanding. Functional programming influences normal programming through design rather than actually using it all the time. I draw inspiration from it to write imperative programs which I think is actually becoming really fashionable now, but I’ve been doing it for ages. I did it first! There were also some courses on really odd programming languages, a bit of Prolog, a little bit of C. Having a little bit of each of those is something that I would have never done on my own, so it was important. And then there are knowledge-based courses which are about not programming itself but things that have been programmed like TCP. Those are really important for examples for how to approach things. Did you do any internships while you were at university? Yeah, I spent both of my summers at the same company. I thought I could code well before I went there. Looking back at the crap that I produced, it was only surpassed in its crappiness by all of the other code already in that company. I’m so much better at writing nice code now than I used to be back then. Was there just not a culture of looking after your code? There was, they just didn’t hire people for their abilities in that area. They hired people for raw IQ. The first indicator of it going wrong was that they didn’t have any computer scientists, which is a bit odd in a programming company. But even beyond that they didn’t have people who learnt architecture from anyone else. Most of them had started straight out of university, so never really had experience or mentors to learn from. There wasn’t the experience to draw from to teach each other. In the second half of my second internship, I was being given tasks like looking at new technologies and teaching people stuff. Interns shouldn’t be teaching people how to do their jobs! All interns are going to have little nuggets of things that you don’t know about, but they shouldn’t consistently be the ones who know the most. It’s not a good environment to learn. I was going to ask how you found working with people who were more experienced than you… When I reached Red Gate, I found some people who were more experienced programmers than me, and that was difficult. I’ve been coding since I was tiny. At university there were people who were cleverer than me, but there weren’t very many who were more experienced programmers than me. During my internship, I didn’t find anyone who I classed as being a noticeably more experienced programmer than me. So, it was a shock to the system to have valid criticisms rather than just formatting criticisms. However, Red Gate’s not so big on the actual code review, at least it wasn’t when I started. We did an entire product release and then somebody looked over all of the UI of that product which I’d written and say what they didn’t like. By that point, it was way too late and I’d disagree with them. Do you think the lack of code reviews was a bad thing? I think if there’s going to be any oversight of new people, then it should be continuous rather than chunky. For me I don’t mind too much, I could go out and get oversight if I wanted it, and in those situations I felt comfortable without it. If I was managing the new person, then maybe I’d be keener on oversight and then the right way to do it is continuously and in very, very small chunks. Have you had any significant projects you’ve worked on outside of a job? When I was a teenager I wrote all sorts of stuff. I used to write games, I derived how to do isomorphic projections myself once. I didn’t know what the word was so I couldn’t Google for it, so I worked it out myself. It was horrifically complicated. But it sort of tailed off when I started at university, and is now basically zero. If I do side-projects now, they tend to be work-related side projects like my actors framework, NAct, which I started in a down tools week. Could you explain a little more about NAct? It is a little C# framework for writing parallel code more easily. Parallel programming is difficult when you need to write to shared data. Sometimes parallel programming is easy because you don’t need to write to shared data. When you do need to access shared data, you could just have your threads pile in and do their work, but then you would screw up the data because the threads would trample on each other’s toes. You could lock, but locks are really dangerous if you’re using more than one of them. You get interactions like deadlocks, and that’s just nasty. Actors instead allows you to say this piece of data belongs to this thread of execution, and nobody else can read it. If you want to read it, then ask that thread of execution for a piece of it by sending a message, and it will send the data back by a message. And that avoids deadlocks as long as you follow some obvious rules about not making your actors sit around waiting for other actors to do something. There are lots of ways to write actors, NAct allows you to do it as if it was method calls on other objects, which means you get all the strong type-safety that C# programmers like. Do you think that this is suitable for the majority of parallel programming, or do you think it’s only suitable for specific cases? It’s suitable for most difficult parallel programming. If you’ve just got a hundred web requests which are all independent of each other, then I wouldn’t bother because it’s easier to just spin them up in separate threads and they can proceed independently of each other. But where you’ve got difficult parallel programming, where you’ve got multiple threads accessing multiple bits of data in multiple ways at different times, then actors is at least as good as all other ways, and is, I reckon, easier to think about. When you’re using actors, you presumably still have to write your code in a different way from you would otherwise using single-threaded code. You can’t use actors with any methods that have return types, because you’re not allowed to call into another actor and wait for it. If you want to get a piece of data out of another actor, then you’ve got to use tasks so that you can use “async” and “await” to await asynchronously for it. But other than that, you can still stick things in classes so it’s not too different really. Rather than having thousands of objects with mutable state, you can use component-orientated design, where there are only a few mutable classes which each have a small number of instances. Then there can be thousands of immutable objects. If you tend to do that anyway, then actors isn’t much of a jump. If I’ve already built my system without any parallelism, how hard is it to add actors to exploit all eight cores on my desktop? Usually pretty easy. If you can identify even one boundary where things look like messages and you have components where some objects live on one side and these other objects live on the other side, then you can have a granddaddy object on one side be an actor and it will parallelise as it goes across that boundary. Not too difficult. If we do get 1000-core desktop PCs, do you think actors will scale up? It’s hard. There are always in the order of twenty to fifty actors in my whole program because I tend to write each component as actors, and I tend to have one instance of each component. So this won’t scale to a thousand cores. What you can do is write data structures out of actors. I use dictionaries all over the place, and if you need a dictionary that is going to be accessed concurrently, then you could build one of those out of actors in no time. You can use queuing to marshal requests between different slices of the dictionary which are living on different threads. So it’s like a distributed hash table but all of the chunks of it are on the same machine. That means that each of these thousand processors has cached one small piece of the dictionary. I reckon it wouldn’t be too big a leap to start doing proper parallelism. Do you think it helps if actors get baked into the language, similarly to Erlang? Erlang is excellent in that it has thread-local garbage collection. C# doesn’t, so there’s a limit to how well C# actors can possibly scale because there’s a single garbage collected heap shared between all of them. When you do a global garbage collection, you’ve got to stop all of the actors, which is seriously expensive, whereas in Erlang garbage collections happen per-actor, so they’re insanely cheap. However, Erlang deviated from all the sensible language design that people have used recently and has just come up with crazy stuff. You can definitely retrofit thread-local garbage collection to .NET, and then it’s quite well-suited to support actors, even if it’s not baked into the language. Speaking of language design, do you have a favourite programming language? I’ll choose a language which I’ve never written before. I like the idea of Scala. It sounds like C#, only with some of the niggles gone. I enjoy writing static types. It means you don’t have to writing tests so much. When you say it doesn’t have some of the niggles? C# doesn’t allow the use of a property as a method group. It doesn’t have Scala case classes, or sum types, where you can do a switch statement and the compiler checks that you’ve checked all the cases, which is really useful in functional-style programming. Pattern-matching, in other words. That’s actually the major niggle. C# is pretty good, and I’m quite happy with C#. And what about going even further with the type system to remove the need for tests to something like Haskell? Or is that a step too far? I’m quite a pragmatist, I don’t think I could deal with trying to write big systems in languages with too few other users, especially when learning how to structure things. I just don’t know anyone who can teach me, and the Internet won’t teach me. That’s the main reason I wouldn’t use it. If I turned up at a company that writes big systems in Haskell, I would have no objection to that, but I wouldn’t instigate it. What about things in C#? For instance, there’s contracts in C#, so you can try to statically verify a bit more about your code. Do you think that’s useful, or just not worthwhile? I’ve not really tried it. My hunch is that it needs to be built into the language and be quite mathematical for it to work in real life, and that doesn’t seem to have ended up true for C# contracts. I don’t think anyone who’s tried them thinks they’re any good. I might be wrong. On a slightly different note, how do you like to debug code? I think I’m quite an odd debugger. I use guesswork extremely rarely, especially if something seems quite difficult to debug. I’ve been bitten spending hours and hours on guesswork and not being scientific about debugging in the past, so now I’m scientific to a fault. What I want is to see the bug happening in the debugger, to step through the bug happening. To watch the program going from a valid state to an invalid state. When there’s a bug and I can’t work out why it’s happening, I try to find some piece of evidence which places the bug in one section of the code. From that experiment, I binary chop on the possible causes of the bug. I suppose that means binary chopping on places in the code, or binary chopping on a stage through a processing cycle. Basically, I’m very stupid about how I debug. I won’t make any guesses, I won’t use any intuition, I will only identify the experiment that’s going to binary chop most effectively and repeat rather than trying to guess anything. I suppose it’s quite top-down. Is most of the time then spent in the debugger? Absolutely, if at all possible I will never debug using print statements or logs. I don’t really hold much stock in outputting logs. If there’s any bug which can be reproduced locally, I’d rather do it in the debugger than outputting logs. And with SmartAssembly error reporting, there’s not a lot that can’t be either observed in an error report and just fixed, or reproduced locally. And in those other situations, maybe I’ll use logs. But I hate using logs. You stare at the log, trying to guess what’s going on, and that’s exactly what I don’t like doing. You have to just look at it and see does this look right or wrong. We’ve covered how you get to grip with bugs. How do you get to grips with an entire codebase? I watch it in the debugger. I find little bugs and then try to fix them, and mostly do it by watching them in the debugger and gradually getting an understanding of how the code works using my process of binary chopping. I have to do a lot of reading and watching code to choose where my slicing-in-half experiment is going to be. The last time I did it was SmartAssembly. The old code was a complete mess, but at least it did things top to bottom. There wasn’t too much of some of the big abstractions where flow of control goes all over the place, into a base class and back again. Code’s really hard to understand when that happens. So I like to choose a little bug and try to fix it, and choose a bigger bug and try to fix it. Definitely learn by doing. I want to always have an aim so that I get a little achievement after every few hours of debugging. Once I’ve learnt the codebase I might be able to fix all the bugs in an hour, but I’d rather be using them as an aim while I’m learning the codebase. If I was a maintainer of a codebase, what should I do to make it as easy as possible for you to understand? Keep distinct concepts in different places. And name your stuff so that it’s obvious which concepts live there. You shouldn’t have some variable that gets set miles up the top of somewhere, and then is read miles down to choose some later behaviour. I’m talking from a very much SmartAssembly point of view because the old SmartAssembly codebase had tons and tons of these things, where it would read some property of the code and then deal with it later. Just thousands of variables in scope. Loads of things to think about. If you can keep concepts separate, then it aids me in my process of fixing bugs one at a time, because each bug is going to more or less be understandable in the one place where it is. And what about tests? Do you think they help at all? I’ve never had the opportunity to learn a codebase which has had tests, I don’t know what it’s like! What about when you’re actually developing? How useful do you find tests in finding bugs or regressions? Finding regressions, absolutely. Running bits of code that would be quite hard to run otherwise, definitely. It doesn’t happen very often that a test finds a bug in the first place. I don’t really buy nebulous promises like tests being a good way to think about the spec of the code. My thinking goes something like “This code works at the moment, great, ship it! Ah, there’s a way that this code doesn’t work. Okay, write a test, demonstrate that it doesn’t work, fix it, use the test to demonstrate that it’s now fixed, and keep the test for future regressions.” The most valuable tests are for bugs that have actually happened at some point, because bugs that have actually happened at some point, despite the fact that you think you’ve fixed them, are way more likely to appear again than new bugs are. Does that mean that when you write your code the first time, there are no tests? Often. The chance of there being a bug in a new feature is relatively unaffected by whether I’ve written a test for that new feature because I’m not good enough at writing tests to think of bugs that I would have written into the code. So not writing regression tests for all of your code hasn’t affected you too badly? There are different kinds of features. Some of them just always work, and are just not flaky, they just continue working whatever you throw at them. Maybe because the type-checker is particularly effective around them. Writing tests for those features which just tend to always work is a waste of time. And because it’s a waste of time I’ll tend to wait until a feature has demonstrated its flakiness by having bugs in it before I start trying to test it. You can get a feel for whether it’s going to be flaky code as you’re writing it. I try to write it to make it not flaky, but there are some things that are just inherently flaky. And very occasionally, I’ll think “this is going to be flaky” as I’m writing, and then maybe do a test, but not most of the time. How do you think your programming style has changed over time? I’ve got clearer about what the right way of doing things is. I used to flip-flop a lot between different ideas. Five years ago I came up with some really good ideas and some really terrible ideas. All of them seemed great when I thought of them, but they were quite diverse ideas, whereas now I have a smaller set of reliable ideas that are actually good for structuring code. So my code is probably more similar to itself than it used to be back in the day, when I was trying stuff out. I’ve got more disciplined about encapsulation, I think. There are operational things like I use actors more now than I used to, and that forces me to use immutability more than I used to. The first code that I wrote in Red Gate was the memory profiler UI, and that was an actor, I just didn’t know the name of it at the time. I don’t really use object-orientation. By object-orientation, I mean having n objects of the same type which are mutable. I want a constant number of objects that are mutable, and they should be different types. I stick stuff in dictionaries and then have one thing that owns the dictionary and puts stuff in and out of it. That’s definitely a pattern that I’ve seen recently. I think maybe I’m doing functional programming. Possibly. It’s plausible. If you had to summarise the essence of programming in a pithy sentence, how would you do it? Programming is the form of art that, without losing any of the beauty of architecture or fine art, allows you to produce things that people love and you make money from. So you think it’s an art rather than a science? It’s a little bit of engineering, a smidgeon of maths, but it’s not science. Like architecture, programming is on that boundary between art and engineering. If you want to do it really nicely, it’s mostly art. You can get away with doing architecture and programming entirely by having a good engineering mind, but you’re not going to produce anything nice. You’re not going to have joy doing it if you’re an engineering mind. Architects who are just engineering minds are not going to enjoy their job. I suppose engineering is the foundation on which you build the art. Exactly. How do you think programming is going to change over the next ten years? There will be an unfortunate shift towards dynamically-typed languages, because of JavaScript. JavaScript has an unfair advantage. JavaScript’s unfair advantage will cause more people to be exposed to dynamically-typed languages, which means other dynamically-typed languages crop up and the best features go into dynamically-typed languages. Then people conflate the good features with the fact that it’s dynamically-typed, and more investment goes into dynamically-typed languages. They end up better, so people use them. What about the idea of compiling other languages, possibly statically-typed, to JavaScript? It’s a reasonable idea. I would like to do it, but I don’t think enough people in the world are going to do it to make it pick up. The hordes of beginners are the lifeblood of a language community. They are what makes there be good tools and what makes there be vibrant community websites. And any particular thing which is the same as JavaScript only with extra stuff added to it, although it might be technically great, is not going to have the hordes of beginners. JavaScript is always to be quickest and easiest way for a beginner to start programming in the browser. And dynamically-typed languages are great for beginners. Compilers are pretty scary and beginners don’t write big code. And having your errors come up in the same place, whether they’re statically checkable errors or not, is quite nice for a beginner. If someone asked me to teach them some programming, I’d teach them JavaScript. If dynamically-typed languages are great for beginners, when do you think the benefits of static typing start to kick in? The value of having a statically typed program is in the tools that rely on the static types to produce a smooth IDE experience rather than actually telling me my compile errors. And only once you’re experienced enough a programmer that having a really smooth IDE experience makes a blind bit of difference, does static typing make a blind bit of difference. So it’s not really about size of codebase. If I go and write up a tiny program, I’m still going to get value out of writing it in C# using ReSharper because I’m experienced with C# and ReSharper enough to be able to write code five times faster if I have that help. Any other visions of the future? Nobody’s going to use actors. Because everyone’s going to be running on single-core VMs connected over network-ready protocols like JSON over HTTP. So, parallelism within one operating system is going to die. But until then, you should use actors. More Red Gater Coder interviews

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

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

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  • Ubuntu: Failure to login with multiple video adapters

    - by tsilb
    Forgive my ignorance, for I am a complete linux noob. I have a computer with three video cards and six monitors. Works great on Windows. Trying to get it to run Ubuntu as well. It loads fine when I have it configured to run on one adapter; detects both screens, runs ok. But I want to turn the other 4 monitors on and run the whole thing as one extended desktop (one session, etc). So I downloaded and installed the newest ATI driver for Linux, which seems to work, kinda. I ran this to set up the screens: aticonfig --adapter=all --initial -f Now when I boot, Ubuntu seems to turn on all the screens (3 viewports, each with two cloned displays from what I can tell). When I enter my login info OR move the mouse off the main screen, the screens freeze and the kbd/ms become unresponsive. aticonfig generated xorg.conf included below. Have tried the following: aticonfig -initial -f - works, but only detects the primary adapter and 2 screens aticccle - Tells me I have to reboot after enabling the other cards. Then goes into above described freezing state. aticonfig --adapter=all --initial -f - see above Manually editing xorg.conf file with my limited knowledge - Was able to get two adapters running, but only the second adapter initialized while the primary stopped at the Ubuntu boot screen. Was unable to see the login prompt. Froze after I logged in blindly (was able to hear the login sound). Using generic "radeon" driver instead of ATI Proprietary driver with the above init attempts Toggling xinerama Various combinations of the above Hardware: Intel Core 2 Quad q6600 8GB DDR2 (3x) ATI Radeon HD 4680 5 monitors (21W, 21W, 22W Portrait, 22W Portrait, 19")and an HDTV (26"W, HDMI) in a horizontal arrangement I know next to nothing about Linux/Ubuntu aside from basic filesystem navigation, editing text files, and accessing my local and networked Windows stores and shares. Basically this is the most advanced thing I've had to do. I installed today. Please advise how to make this configuration work. my xorg.conf: Section "ServerLayout" Identifier "Layout0" Screen 0 "aticonfig-Screen[0]-0" 0 0 Screen "aticonfig-Screen[1]-0" RightOf "aticonfig-Screen[0]-0" Screen "aticonfig-Screen[2]-0" RightOf "aticonfig-Screen[1]-0" Option "RenderAccel" "true" Option "AllowGLXWithComposite" "true" EndSection Section "Files" EndSection Section "Module" EndSection Section "ServerFlags" Option "Xinerama" "0" EndSection Section "Monitor" Identifier "aticonfig-Monitor[0]-0" Option "VendorName" "ATI Proprietary Driver" Option "ModelName" "Generic Autodetecting Monitor" Option "DPMS" "true" EndSection Section "Monitor" Identifier "aticonfig-Monitor[1]-0" Option "VendorName" "ATI Proprietary Driver" Option "ModelName" "Generic Autodetecting Monitor" Option "DPMS" "true" EndSection Section "Monitor" Identifier "aticonfig-Monitor[2]-0" Option "VendorName" "ATI Proprietary Driver" Option "ModelName" "Generic Autodetecting Monitor" Option "DPMS" "true" EndSection Section "Device" Identifier "aticonfig-Device[0]-0" Driver "fglrx" BusID "PCI:1:0:0" EndSection Section "Device" Identifier "aticonfig-Device[1]-0" Driver "fglrx" BusID "PCI:3:0:0" EndSection Section "Device" Identifier "aticonfig-Device[2]-0" Driver "fglrx" BusID "PCI:4:0:0" EndSection Section "Screen" Identifier "aticonfig-Screen[0]-0" Device "aticonfig-Device[0]-0" Monitor "aticonfig-Monitor[0]-0" DefaultDepth 24 SubSection "Display" Viewport 0 0 Depth 24 EndSubSection EndSection Section "Screen" Identifier "aticonfig-Screen[1]-0" Device "aticonfig-Device[1]-0" Monitor "aticonfig-Monitor[1]-0" DefaultDepth 24 SubSection "Display" Viewport 0 0 Depth 24 EndSubSection EndSection Section "Screen" Identifier "aticonfig-Screen[2]-0" Device "aticonfig-Device[2]-0" Monitor "aticonfig-Monitor[2]-0" DefaultDepth 24 SubSection "Display" Viewport 0 0 Depth 24 EndSubSection EndSection

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  • Need help with yum,python and php in CentOS. (I made a complete mess!)

    - by pek
    a while back I wanted to install some plugins for Trac but it required python 2.5 I tried installing it (I don't remember how) and the only thing I managed was to have two versions of python (2.4 and 2.5). Trac still uses the old version but the console uses 2.5 (python -V = Python 2.5.2). Anyway, the problem is not python, the problem is yum (which uses python). I am trying to upgrade my PHP version from 5.1.x to 5.2.x. I tried following this tutorial but when I reach the step with yum I get this error: >[root@XXX]# yum update Loading "installonlyn" plugin Setting up Update Process Setting up repositories Reading repository metadata in from local files Traceback (most recent call last): File "/usr/bin/yum", line 29, in ? yummain.main(sys.argv[1:]) File "/usr/share/yum-cli/yummain.py", line 94, in main result, resultmsgs = base.doCommands() File "/usr/share/yum-cli/cli.py", line 381, in doCommands return self.yum_cli_commands[self.basecmd].doCommand(self, self.basecmd, self.extcmds) File "/usr/share/yum-cli/yumcommands.py", line 150, in doCommand return base.updatePkgs(extcmds) File "/usr/share/yum-cli/cli.py", line 672, in updatePkgs self.doRepoSetup() File "/usr/share/yum-cli/cli.py", line 109, in doRepoSetup self.doSackSetup(thisrepo=thisrepo) File "/usr/lib/python2.4/site-packages/yum/__init__.py", line 338, in doSackSetup self.repos.populateSack(which=repos) File "/usr/lib/python2.4/site-packages/yum/repos.py", line 200, in populateSack sack.populate(repo, with, callback, cacheonly) File "/usr/lib/python2.4/site-packages/yum/yumRepo.py", line 91, in populate dobj = repo.cacheHandler.getPrimary(xml, csum) File "/usr/lib/python2.4/site-packages/yum/sqlitecache.py", line 100, in getPrimary return self._getbase(location, checksum, 'primary') File "/usr/lib/python2.4/site-packages/yum/sqlitecache.py", line 86, in _getbase (db, dbchecksum) = self.getDatabase(location, metadatatype) File "/usr/lib/python2.4/site-packages/yum/sqlitecache.py", line 82, in getDatabase db = self.makeSqliteCacheFile(filename,cachetype) File "/usr/lib/python2.4/site-packages/yum/sqlitecache.py", line 245, in makeSqliteCacheFile self.createTablesPrimary(db) File "/usr/lib/python2.4/site-packages/yum/sqlitecache.py", line 165, in createTablesPrimary cur.execute(q) File "/usr/lib/python2.4/site-packages/sqlite/main.py", line 244, in execute self.rs = self.con.db.execute(SQL) _sqlite.DatabaseError: near "release": syntax error Any help? Thank you. Update OK, so I've managed to update yum hoping it would solve my problems but now I get a slightly different version of the same error: [root@XXX]# yum -y update Loaded plugins: fastestmirror Loading mirror speeds from cached hostfile * addons: mirror.skiplink.com * base: www.gtlib.gatech.edu * epel: mirrors.tummy.com * extras: yum.singlehop.com * updates: centos-distro.cavecreek.net (process:30840): GLib-CRITICAL **: g_timer_stop: assertion `timer != NULL' failed (process:30840): GLib-CRITICAL **: g_timer_destroy: assertion `timer != NULL' failed Traceback (most recent call last): File "/usr/bin/yum", line 29, in ? yummain.user_main(sys.argv[1:], exit_code=True) File "/usr/share/yum-cli/yummain.py", line 309, in user_main errcode = main(args) File "/usr/share/yum-cli/yummain.py", line 178, in main result, resultmsgs = base.doCommands() File "/usr/share/yum-cli/cli.py", line 345, in doCommands self._getTs(needTsRemove) File "/usr/lib/python2.4/site-packages/yum/depsolve.py", line 101, in _getTs self._getTsInfo(remove_only) File "/usr/lib/python2.4/site-packages/yum/depsolve.py", line 112, in _getTsInfo pkgSack = self.pkgSack File "/usr/lib/python2.4/site-packages/yum/__init__.py", line 661, in <lambda> pkgSack = property(fget=lambda self: self._getSacks(), File "/usr/lib/python2.4/site-packages/yum/__init__.py", line 501, in _getSacks self.repos.populateSack(which=repos) File "/usr/lib/python2.4/site-packages/yum/repos.py", line 260, in populateSack sack.populate(repo, mdtype, callback, cacheonly) File "/usr/lib/python2.4/site-packages/yum/yumRepo.py", line 190, in populate dobj = repo_cache_function(xml, csum) File "/usr/lib/python2.4/site-packages/sqlitecachec.py", line 42, in getPrimary self.repoid)) TypeError: Can not create packages table: near "release": syntax error I'm guessing that this "release" thing has something to do with a repository, but I didn't find anything... I went to the sqlitecachec.py at line 42 which writes (line numbers added for convenience): 39: return self.open_database(_sqlitecache.update_primary(location, 40: checksum, 41: self.callback, 42: self.repoid)) Update 2 I think I found the problem. This post suggests that the problem is sqlite and not yum. The version of sqlite I have installed is 3.6.10 but I have no idea which version does python 2.4 uses. ld.so.config contains the following: include ld.so.conf.d/*.conf /usr/local/lib In folder /usr/local/lib I find a symbolic link named libsqlite3.so that points to libsqlite3.so.0.8.6 WHAT IS HAPPENING??????? :S

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  • Apple Airport Express, Extreme and Time Capsules, BT Home Hub, Wireless Extenders confusion

    - by Jamie Hartnoll
    I post quite frequently in Stack Overflow, but use Superuser less frequently. Mainly as I don't change hardware often and rarely have software issues! I live in a small stone cottage, and have an office in a separate building across a yard. I have a BT Homehub which is located in the cottage and a series of Ethernet cables running across the yard to the office. This is fine for my wired stuff. My main office computers are PCs running Windows 7 Ultimate, and one on Win7 Home, all working fine. I also have an old laptop on Win XP which works fine wirelessly in the house for those evenings in front of the TV catching up on a bit of work. I also have an iPhone and an iPad. Recently, I have been trying to get WiFi in the office so I can use Adobe Shadow (or whatever it now is!) to improve mobile web development efficiency using my iPhone and iPad, so I bought this: http://www.ebuyer.com/393462-zyxel-wre2205-500mbps-powerline-wireless-n300-range-extender-wre2205-gb0101f Thinking that would be lovely just plugged into the socket by the door in the office, extending the perimeter of the WiFi from my Homehub. I can't get it to work properly! If I plug a laptop into its ethernet port I can get it to connect to the Homehub and give me a kinda of wired, wireless extender. If, however, I plug the ethernet port into my home hub, it then seems to extend the network, but only my iOs devices work, and all my wired stuff stops working, and seems to create an infinite loop where windows connects to my homehob, and then rather to the internet, it then connects back to the extender thing. Anyway... in the meantime, I took a fatal trip to the Apple Store, where I purchased an Airport Express... solely for the purpose of hooking my iOs devices up as wireless music players in the house. I knew it had WiFi, but didn't want to use that part as an extender, I didn't think it would work on a Homehub anyway. It doesn't work on a Homehub! I now have a new wireless network in the house, which, when anything connects to it cannot connect to the Internet, so it works ONLY as a wireless music player. I then borrowed some Powerline Adaptors from someone and realised that this whole thing was getting totally out of control! It seems all the technology is out there but it's so complicated to get the right series of devices. To further add to the confusion, I wouldn't mind a network hard drive. I bought one that broke and lost everything, so now we're on to looking at the Apple Time Capsules. So my question is... IF... I buy an Apple Time Capsule, can I: Hook that up to my Homehub, leaving the homehub connected to the Internet so my Hub phones still work, then disable wireless on the homehub Link up my Airport Express to the Time Capsule PROPERLY so it will connect to the Internet Do the above with an Apple TV box should I buy one in future Use the Time Capsule as a network hard drive to store video and music that can be viewed/listened to via my iOS devices/Apple TV/Aiport Express anywhere even with my main PC off (this currently stores all this data) Hope that the IOS devices like the WiFi from the TimeCapsule better than the Homehub and work without extension, or buy another Airport Express to get WiFI in the office. Or... should I buy an Airport Extreme and use a USB hard drive for the network drive?

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  • Cannot turn on "Network Discovery and File Sharing" when Windows Firewall is enabled

    - by Cheeso
    I have a problem similar to this one. Windows Firewall prevents File and Printer sharing from working and Why does File and Printer Sharing keep turning off in Windows 7? I cannot turn on Network Discovery. This is Windows 7 Home Premium, x64. It's a Dell XPS 1340 and Windows came installed from the OEM. This used to work. Now it doesn't. I don't know what has changed. In windows Explorer, the UI looks like this: When I click the yellow panel that says "Click to change...", the panel disappears, then immediately reappears, with exactly the same text. If I go through the control panel "Network and Sharing Center" thing, the UI looks like this: If I tick the box to "turn on network discovery", the "Save Changes" button becomes enabled. If I then click that button, the dialog box just closes, with no message or confirmation. Re-opening the same dialog box shows that Network Discovery has not been turned on. If I turn off Windows Firewall, I can then turn on Network Discovery via either method. The machine is connected to a wireless home network, via a router. The network is marked as "Home Network" in the Network and Sharing Center, which I think corresponds to the "Private" profile in Windows Firewall Advanced Settings app. (Confirm?) The PC is not part of a domain, and has never been part of a domain. The machine is not bridging any networks. There is a regular 100baseT connector but I have the network adapter for that disabled in Windows. Something else that seems odd. Within Windows Firewall Advanced Settings, there are no predefined rules available. If I click the "New Rule...." Action on the action pane, the "Predefined" option is greyed out. like this: In order to attempt to allow the network discovery protocols through on the private network, I hand-coded a bunch of rules, intending to allow the necessary UPnP and WDP protocols supporting network discovery. I copied them from a working Windows 7 Ultimate PC, running on the same network. This did not work. Even with the hand-coded rules, I still cannot turn on Network Discovery. I looked on the interwebs, and the only solution that appears to work is a re-install of Windows. Seriously? If I try netsh advfirewall firewall set rule group="Network Discovery" new enable=Yes ...it says "No rules match the specified criteria" EDIT: by the way, these services are running. DNS Client Function Discovery Resource Publication SSDP Discovery UPnP Device Host in any case, since it works with no firewall, I would assume all necessary services are present and running. The issue is a firewall thing, but I don't know how to diagnose further, or fix it. Q1: Is there a way to definitively insure the correct holes are punched through the Windows Firewall to allow Network Discovery to function? Q2: Should I expect the "predefined" firewall rules to be greyed out? Q3: Why did this change?

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  • Kernel Mode Rootkit

    - by Pajarito
    On the other 3 computers in my family, I believe that we have a kernel-mode rootkit for windows. It appears that the same rootkit is on all of them. We think. We changed all the important passwords from my computer, running linux right now. On all of the infected computers is Symantic Endpoint Protection, because it's free from the university where my mom and dad work. In my opinion symantec is a piece of crap, seeing as it didn't even manager to delete the tracking cookies it found when I tried it on my own computer. The Computers and their set-ups: Computer A: Vista Business; symantec antivirus. runs it as admin, no password. IE8. no other security software other than what comes with windows. IE8 security settings the default Computer B: XP Home Premium; symantec antivirus. runs as normal user, no password, admin account with weak password, spybot, uses IE8 with default settings, sometimes Firefox Computer C: XP Home Premium; symantec antivirus. runs as normal user, no password, admin account with weak password, uses IE8 with default settings, no other security programs except what came with windows This is what's happening. Cut and pasted from my dad's forum post. -- When I scanned my laptop (Dell XPS M1330 with Windows Vista Small Business), Symantec Endpoint Protection hangs for a while, perhaps 10 seconds or so, on some of the following files 9129837.exe, hide_evr2.sys, VirusRemoval.vbs, NewVirusRemoval.vbs, dll.dll, alsmt.ext, and _epnt.sys. It does this if a run a scan that I set up to run on a new thumbnail drive and it does this even if the thumbnail is not plugged in. It doesn't seem to do this if I scan only the C: drive. I've check for problems with symantec endpoint protection and also with Microsoft Security Essentials and Malwarebytes Anti-Malware. They found nothing and I can't find anything by searching for hidden files. Next I tried microsoft's rootkitrevealer. It (rootkitrevealer) finds 279660 (or so) discrepancies and the interface is so glitchy after that I can't really figure out what is going on. The screen is squirrely. The rootkitrevealer pulls up many files in the folder \programdata\applicationdata and there are numberous appended \applicationdata on the end of that as well. -- As you can see, what we did was install MSE and MBAM and scan with both of them. Nothing but a tracking cookie. Then I took over and ran rootkitrevealer.exe from MicroSoft from a flash drive. It found a bunch of discrepancies, but only about 20 or so where security related, the rest being files that you just couldn't see from Windows Explorer. I couldn't see whether of not the files list above, the ones that the scan was hanging on, where in the list. The other thing is, I have no idea what to do about the things the scan comes up with. Then we checked the other computers and they do the same thing when you scan with Symantec. The people at the university seen to think that dad might not have a virus, but 2 of the computers slowed down noticably AND IE8 started acting all funny. None of my family is very computer oriented, and 2 of the possible causes for the rootkit are: -My dad bought a new flash drive, which shipped with a data security executable on it -My dad has to download lots of articles for his work Those are the only things that stand out, but it could have been anything. We are currently backing up our data, and I'll post again after trying IceSword 1.22. I just looked at my dad's forum topic, and someone recommended GMER. I'll try that too.

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  • CPU Temperature sensor wrong?

    - by Matias Nino
    Everest Ultimate is suddenly telling me that the CPU temperature (and core temps) for my E6850 Core 2 Duo is 72 degrees Celsius. When I stress-test the machine, the temp goes up to 91 degrees and the CPU actually throttles. System remains stable though. For over a year now, my CPU has run very cool (40's) with a large commercial copper heatsink/fan that I bought separately. To top it off, I removed the cover of the box and felt the cpu heatsink and it wasn't even warm. Is there such a thing as a CPU temp sensor showing the wrong readings? Any tips would help. UPDATE #1 Temp is also just as high in BIOS. So that leads me to believe it's a CPU seating issue (even though I used thermal paste to seat it two years ago when I built the machine) UPDATE #2 Well. I removed the heatsink and cleaned off the original thermal paste (which was somewhat crusty). I polished the surface, re-applied some new paste, and reseated the heat sink. After powering it up, there was no noticeable change in the temp - ideling at 74. Ran the stress test and it went up to 94 degrees before being 100% throttled. I let it sit at 94 degrees for 20 minutes straight and the computer didn't even flinch. I then immediately shut it off and opened the case and felt around. The heatsink was completely cold to the touch. Even the copper rods were cold. The area near contact with the CPU was slightly warm but not hot to touch. Then I ran REALTEMP, which is supposedly more accurate and it told me the CPU was at 104 degrees. (LOL) At this point, I'm thinking no doubt the cpu's sensor is wrong. Sidenote: the BIOS has the latest version so no option to flash there. Reverting hasn't been known to help from what I've read. What pisses me off is the false temps force the CPU to artificially throttle from 3GHz down to 2GHz and my CPU fan is cranking at full force all the time. Should I call intel and tell them to send me another E6850? SOLUTION UPDATE I switched the processor out with another one and got the same obscene temperatures with the new processor followed by a heatsink that was cool to touch. My suspicion in the heatsink was suddenly renewed. I swapped it out with the stock heatsink/fan and lo and behold the temperatures returned to the normal 35C-50C. Even though the thermal paste was visibly flattened out every time I removed it, it looks like the heatsink was still not pressing hard enough on the CPU to effectively conduct the heat. The heatsink is a Masscool 8Wa741, which screws into a standard position on a mount on the back of the MOBO. Only thing I can surmise after 2 years of use was that, over time, the heatsink pressure on the CPU gave way until the heat began to be ineffectively conducted. Lessons learned: Intel CPU's can run SUPER HOT (upwards of 95C) and still be stable. Heatsink's need to be VERY firmly pressed against the CPU to conduct heat.

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  • Ubuntu 10.04: OpenVZ Kernel and pure-ftpd issues on HOST (no guest setup yet)

    - by Seidr
    After compiling and installing the OpenVZ flavour of kernel under Ubuntu 10.04, I am unable to browse to certain directories when connecting to the pure-ftpd server. The clients are dropping into PASSIVE mode, which is fine. This behaviour was happening before the change of kernel, however now when I browse to certain directories the connection just gets dropped. This only happens with a few directories under one login (web in specific), where as with another login it happens as soon as I connect. I've got the nf_conntrack_ftp kernel module installed (required to keep track of passive FTP connections as I understand, and an alias of the ip_conntrack_ftp module), however this has provided no alleviation of my problem. This module was actually required upon initial setup of my OS to get passive FTP working correctly, however when I compiled the OpenVZ kernel a lot of these modules were missing (iptables, conntrack etc). I recompiled the kernel with the missing modules, but to no effect. I've turned verbosity for the pure-ftpd server up, and still no clues have been spotted in either syslog or the transfer log. Neither did an strace provide any clues (that I could discern anyway) - although one strange thing is both in the output to the client and in the strace I notice that it does infact probe the directory and return the number of matches - it just fails after that. One more thing to mention is that if I FTP using the same credentials locally, everything works fine. This suggests that it is in fact an issue with either the conntrack_ftp module not functioning as expected, or a deeper networking issue. The Kernel was compiled and installed following the instructions at https://help.ubuntu.com/community/OpenVZ - bar the changes to the Kernel configuration (such as add iptables as a module). Below is an example of the log sent to the data (under FileZilla). Status: Resolving address of xxxx.co.uk Status: Connecting to 78.46.xxx.xxx:21... Status: Connection established, waiting for welcome message... Response: 220---------- Welcome to Pure-FTPd [privsep] [TLS] ---------- Response: 220-You are user number 4 of 10 allowed. Response: 220-Local time is now 08:52. Server port: 21. Response: 220-This is a private system - No anonymous login Response: 220-IPv6 connections are also welcome on this server. Response: 220 You will be disconnected after 15 minutes of inactivity. Command: USER xxx Response: 331 User xxx OK. Password required Command: PASS ******** Response: 230-User xxx has group access to: client1 sshusers Response: 230 OK. Current restricted directory is / Command: OPTS UTF8 ON Response: 200 OK, UTF-8 enabled Status: Connected Status: Retrieving directory listing... Command: PWD Response: 257 "/" is your current location Status: Directory listing successful Status: Retrieving directory listing... Command: CWD /web Response: 250 OK. Current directory is /web Command: TYPE I Response: 200 TYPE is now 8-bit binary Command: PORT 10,0,2,30,14,143 Response: 500 I won't open a connection to 10.0.2.30 (only to 188.220.xxx.xxx) Command: PASV Response: 227 Entering Passive Mode (78,46,79,147,234,110) Command: MLSD Response: 150 Accepted data connection Response: 226-ASCII Response: 226-Options: -a -l Response: 226 57 matches total Error: Could not read from transfer socket: ECONNRESET - Connection reset by peer Error: Failed to retrieve directory listing Any suggestions please? I'm willing to try anything!

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  • All my sites are 403 but the server is running. Errors on startup

    - by Craig
    We gave access to a contractor to install a firewall and somehow while he was doing it he fracked something up. Everything went off-line about 24 hours ago and we are effectively out of business until I solve this and the person who messed up the thing is not returning calls. I found a few errors. First, I'm not a server guy - I can look at log files and normally everything runs fine. All 'services' are running according to 1and1 server monitoring and mail is being delivered just fine. The whole thing was off-line until I (probably stupidly) updated the kernel from 6.2 to 6.3 this morning and I got everything back except the http access. All the domains (~200 of them) are returning a 403 error and nothing is recorded in the access log. On every restart I see this error in the messages log file: init: Failed to spawn ttyS0 main process: unable to execute: No such file or directory and a little later these: kernel: WARNING: at kernel/sched.c:5914 thread_return+0x232/0x79d() (Not tainted) kernel: Hardware name: X9SCL/X9SCM kernel: Modules linked in: xt_iprange iptable_filter ip_tables ip6t_REJECT nf_conntrack_ipv6 nf_defrag_ipv6 xt_state nf_conntrack ip6table_filter ip6_tables ipv6 ext4 jbd2 serio_raw i2c_i801 i2c_core sg iTCO_wdt iTCO_vendor_support e1000e ext3 jbd mbcache raid1 sd_mod crc_t10dif ahci dm_mirror dm_region_hash dm_log dm_mod [last unloaded: scsi_wait_scan] kernel: Pid: 367, comm: md3_raid1 Not tainted 2.6.32-220.2.1.el6.x86_64 #1 kernel: Call Trace: kernel: [<ffffffff81069997>] ? warn_slowpath_common+0x87/0xc0 kernel: [<ffffffff810699ea>] ? warn_slowpath_null+0x1a/0x20 kernel: [<ffffffff814eccc5>] ? thread_return+0x232/0x79d kernel: [<ffffffff8126a4d9>] ? cpumask_next_and+0x29/0x50 kernel: [<ffffffff813e9c05>] ? md_super_wait+0x55/0x90 kernel: [<ffffffff81090a10>] ? autoremove_wake_function+0x0/0x40 kernel: [<ffffffff813ebf46>] ? md_update_sb+0x206/0x3f0 kernel: [<ffffffff813ee922>] ? md_check_recovery+0x3f2/0x6d0 kernel: [<ffffffffa005b129>] ? raid1d+0x49/0x1050 [raid1] kernel: [<ffffffff814ed985>] ? schedule_timeout+0x215/0x2e0 kernel: [<ffffffff814ef447>] ? _spin_unlock_irqrestore+0x17/0x20 kernel: [<ffffffff813eb336>] ? md_thread+0x116/0x150 kernel: [<ffffffff81090a10>] ? autoremove_wake_function+0x0/0x40 kernel: [<ffffffff813eb220>] ? md_thread+0x0/0x150 kernel: [<ffffffff810906a6>] ? kthread+0x96/0xa0 kernel: [<ffffffff8100c14a>] ? child_rip+0xa/0x20 kernel: [<ffffffff81090610>] ? kthread+0x0/0xa0 kernel: [<ffffffff8100c140>] ? child_rip+0x0/0x20 And something is wrong with the Named/BIND resulting in the same error for all domains: zone DOMAINEXAMPLE.com/IN: loading from master file DOMAINEXAMPLE.com failed: file not found zone DOMAINEXAMPLE.com/IN: not loaded due to errors. _default/DOMAINEXAMPLE.com/IN: file not found I'm pretty sure this is not enough information to solve the problem, but I'm willing to engage someone who can work this out for me. Any help would be greatly appreciated.

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  • Can't re-mount existing RAID10 on Ubuntu

    - by Zoran
    I saw similar questions, but didn't find what solution to my problem. After power-cut, one of RAID10 (4 disks were) appears to be malfunctioning. I make tha array active one, but can not mount it. Always the same error: mount: you must specify the filesystem type So, here is what I have when type mdadm --detail /dev/md0 /dev/md0: Version : 00.90.03 Creation Time : Tue Sep 1 11:00:40 2009 Raid Level : raid10 Array Size : 1465148928 (1397.27 GiB 1500.31 GB) Used Dev Size : 732574464 (698.64 GiB 750.16 GB) Raid Devices : 4 Total Devices : 3 Preferred Minor : 0 Persistence : Superblock is persistent Update Time : Mon Jun 11 09:54:27 2012 State : clean, degraded Active Devices : 3 Working Devices : 3 Failed Devices : 0 Spare Devices : 0 Layout : near=2, far=1 Chunk Size : 64K UUID : 1a02e789:c34377a1:2e29483d:f114274d Events : 0.166 Number Major Minor RaidDevice State 0 8 16 0 active sync /dev/sdb 1 0 0 1 removed 2 8 48 2 active sync /dev/sdd 3 8 64 3 active sync /dev/sde At the /etc/mdadm/mdadm.conf I have by default, scan all partitions (/proc/partitions) for MD superblocks. alternatively, specify devices to scan, using wildcards if desired. DEVICE partitions auto-create devices with Debian standard permissions CREATE owner=root group=disk mode=0660 auto=yes automatically tag new arrays as belonging to the local system HOMEHOST <system> instruct the monitoring daemon where to send mail alerts MAILADDR root definitions of existing MD arrays ARRAY /dev/md0 level=raid10 num-devices=4 UUID=1a02e789:c34377a1:2e29483d:f114274d ARRAY /dev/md1 level=raid1 num-devices=2 UUID=9b592be7:c6a2052f:2e29483d:f114274d This file was auto-generated... So, my question is, how can I mount md0 array (md1 has been mounted without problem) in order to preserve existing data? One more thing, fdisk -l command gives the following result: Disk /dev/sdb: 750.1 GB, 750156374016 bytes 255 heads, 63 sectors/track, 91201 cylinders Units = cylinders of 16065 * 512 = 8225280 bytes Disk identifier: 0x660a6799 Device Boot Start End Blocks Id System /dev/sdb1 * 1 88217 708603021 83 Linux /dev/sdb2 88218 91201 23968980 5 Extended /dev/sdb5 88218 91201 23968948+ 82 Linux swap / Solaris Disk /dev/sdc: 750.1 GB, 750156374016 bytes 255 heads, 63 sectors/track, 91201 cylinders Units = cylinders of 16065 * 512 = 8225280 bytes Disk identifier: 0x0008f8ae Device Boot Start End Blocks Id System /dev/sdc1 1 88217 708603021 83 Linux /dev/sdc2 88218 91201 23968980 5 Extended /dev/sdc5 88218 91201 23968948+ 82 Linux swap / Solaris Disk /dev/sdd: 750.1 GB, 750156374016 bytes 255 heads, 63 sectors/track, 91201 cylinders Units = cylinders of 16065 * 512 = 8225280 bytes Disk identifier: 0x4be1abdb Device Boot Start End Blocks Id System Disk /dev/sde: 750.1 GB, 750156374016 bytes 255 heads, 63 sectors/track, 91201 cylinders Units = cylinders of 16065 * 512 = 8225280 bytes Disk identifier: 0xa4d5632e Device Boot Start End Blocks Id System Disk /dev/sdf: 750.1 GB, 750156374016 bytes 255 heads, 63 sectors/track, 91201 cylinders Units = cylinders of 16065 * 512 = 8225280 bytes Disk identifier: 0xdacb141c Device Boot Start End Blocks Id System Disk /dev/sdg: 750.1 GB, 750156374016 bytes 255 heads, 63 sectors/track, 91201 cylinders Units = cylinders of 16065 * 512 = 8225280 bytes Disk identifier: 0xdacb141c Device Boot Start End Blocks Id System Disk /dev/md1: 750.1 GB, 750156251136 bytes 2 heads, 4 sectors/track, 183143616 cylinders Units = cylinders of 8 * 512 = 4096 bytes Disk identifier: 0xdacb141c Device Boot Start End Blocks Id System Warning: ignoring extra data in partition table 5 Warning: ignoring extra data in partition table 5 Warning: ignoring extra data in partition table 5 Warning: invalid flag 0x7b6e of partition table 5 will be corrected by w(rite) Disk /dev/md0: 1500.3 GB, 1500312502272 bytes 255 heads, 63 sectors/track, 182402 cylinders Units = cylinders of 16065 * 512 = 8225280 bytes Disk identifier: 0x660a6799 Device Boot Start End Blocks Id System /dev/md0p1 * 1 88217 708603021 83 Linux /dev/md0p2 88218 91201 23968980 5 Extended /dev/md0p5 ? 121767 155317 269488144 20 Unknown And one more thing. When using mdadm --examine command, here ise result: mdadm -v --examine --scan /dev/sdb /dev/sdc /dev/sdd /dev/sde /dev/sdf /dev/sd ARRAY /dev/md1 level=raid1 num-devices=2 UUID=9b592be7:c6a2052f:2e29483d:f114274d devices=/dev/sdf ARRAY /dev/md0 level=raid10 num-devices=4 UUID=1a02e789:c34377a1:2e29483d:f114274d devices=/dev/sdb,/dev/sdc,/dev/sdd,/dev/sde md0 has 3 devices which are active. Can someone instruct me how to solve this issue? If it is possible, I would like not to removing faulty HDD. Please advise

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  • Open-source and client-side IRC-Client

    - by user125197
    I'm administrating a homepage with an IRC-Webclient. At the moment, it uses a Java-client, modified to be SSL-compatible and compatible with whatever design I want to use. The usual PJIRC thing, with users wishes and concerns implemented (I'll give you SSL-PJIRC if you want - no problem. The login-script will be improved for IE6-support this weekend, means write the object tag if an old IE is found, that's it. Rest runs perfectly well). But still, the better user experience would be a client that only requires the user to enable JavaScript. So, before I go into rewriting and customizing a complete Java-IRC-solution, I'd like to ask some other people - after researching for half a year of time. The requirements are: a) Free hosting, no cgi-bin is acceptable. A lot of hosters also have CGI supported, but it doesn't allow IRC-access. So, seems you are limited to x10-hosting for CGI:IRC. b) Solution must be open source (not free as in free beer, but free as in free speech - I want to be able to read every line of the code). c) Hosting cannot be limited to Windows-hosting. Free operation systems (anything with Linux, BSD, whatever - I think you get what I mean) must be possible. d) No server-side technology. I'm limited to everything that runs client-side only. (Well, a Java-Applet does...). d) Solution must support SSL. e) The side must be able to move. Means, you register to a new free hoster, load your things up via FileZilla and the thing simply runs. No server-concerning implementations needed. f) Old browsers must be supported. From all my research, there are two ways a) Java-Applet b) Use HTML5-Websockets (it is impossible to use the JavaScript-library I found, as it depends on ActiveX - means, Windows hosting). b) means "you are going to enter very unstable content". I worked with HTML5 for month, and, well ... Though, HTML5 also is not suitable for older browsers (old browser support is an absolutely necessary requirement). PHP by the way is a server side solution ... so no line of PHP. My idea now was a Java Applet, and a fallback which again is embedded and proprietary, but only requieres JavaScript (wsirc or Mibbit are great here, whilst I prefer wsirc ... and zooming fonts that are plainly to small, but worse, cannot read the code). So my question is - with free hosting, not installation on server, plain upload - do you see and open source, complelety client-side, ssl-compatible way to use or write a client? This wouldn't be my first programming project, I'm not afraid of the code. If there was a way and I had to write it, even from scratch, I'd do. From all I know, there sadly is none. But maybe some experienced admin has an idea? Best Wishes, yetanotheruser Sry my English isn't perfect. It's enough for programming at least.

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  • Where to download replacement "Explorerframe.DLL" Files for x64 Windows 7 Pro?

    - by Ben Franchuk
    After posting this question, I did some research to reveal what the problem likely was, and found what I need to fix. Following this is the original post, then my updated question. A few months ago I ended up requiring to change my computer's SID to fix a problem it had been having- Instead of fixing the problem, though, it screwed up my at-the-time current install of windows, to the point of me needing to do a fresh install. As I am in possession of an OEM copy of Windows 7 Pro 64 Bit, I successfully reinstalled over the dead copy with that (all the files that were on the computer previous to this windows install were put in a Windows.old folder). Everything installed and worked absolutely fine, except for one thing. The problem I am experiencing is that, in some Windows Explorer windows, the explorer pane doesn't show. Instead, it simply shows a white area where the pane would show. This makes some software not usable, I recently realized; Software such as Cubase, which use just the explorer pane to select file save locations, cannot save at all as the pane itself is... not operational. Below is a screenshot of this problem as it occurs in cubase; ...and again as it shows in UTorrent in the save location selector window. The highlighted area is where the sidebar would NORMALLY be. Pardon my scribbling over some of the things in the window- I would personally rather the internet did not get a glimpse of my files. I have yet to find a common reason why the pane works in some applications when they pull explorer, and others not. I have yet to see it go away, and the software affected by it has been affected since I reinstalled my copy of windows. Initially, I was able to live with it as I can type out save locations in the file name bar to navigate, but with software like Cubase, I do not have this option. Reinstalling windows again is NOT an option. Here's the updated question. After posting this question originally, I did some research on the problem in question, and it turns out that this is extremely easily fixable via replacing the file "ExplorerFrame.DLL" which is located in the System32 and SystemWOW64 Folders, in the windows folder, on the C:\ drive. As I quite frequently customize my computer, this is a normal thing for me to do and I know exactly how to safely and properly replace this file. The only problem is that I cannot for the life of me find a download of this file that actually works with my computer. I tried a couple from some different sites but they all caused explorer to not restart (I was given an error when starting the application from Task Manager) and was forced to revert to the broken .DLL files. Since there are 2 separate "ExplorerFrame.DLL" files; one for 64 bit and the other for 32 bit, I am assuming that I will need to download 2 separate versions to replace the corrupted ones. Where can I acquire these files? I am currently running Windows 7 Professional x64 Bit.

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  • iptables 1.4 and passive FTP on custom port

    - by Cracky
    after the upgrade from debian squeeze to wheezy I've got a problem with passive FTP connection. I could narrow it to be iptables related, as I could connect via FTP w/o problems after adding my IP to the iptables ACCEPT rule. Before the upgrade I was able just to do modprobe nf_conntract_ftp ports=21332 and adding iptables -A THRU -p tcp --dport 21332 -m state --state NEW,ESTABLISHED,RELATED -j ACCEPT now..it doesn't help anymore. The INPUT rule is being triggered as I can see in the counter, but the directory listing is the last thing it does. Setting up a passive-port range is the last thing I want to do, I dislike open ports. I also tried the trick with helper mod by adding following rule before the actual rule for 21332 iptables -A THRU -p tcp -i eth0 --dport 21332 -m state --state NEW -m helper --helper ftp-21332 -j ACCEPT but it doesn't help and is even not being triggered according to counter. The rule in the next line (w/o helper) is being triggered.. here some info: # iptables --version iptables v1.4.14 # lsmod |grep nf_ nf_nat_ftp 12460 0 nf_nat 18242 1 nf_nat_ftp nf_conntrack_ftp 12605 1 nf_nat_ftp nf_conntrack_ipv4 14078 32 nf_nat nf_defrag_ipv4 12483 1 nf_conntrack_ipv4 nf_conntrack 52720 7 xt_state,nf_conntrack_ipv4,xt_conntrack,nf_conntrack_ftp,nf_nat,nf_nat_ftp,xt_helper # uname -a Linux loki 3.2.0-4-amd64 #1 SMP Debian 3.2.46-1 x86_64 GNU/Linux # iptables-save # Generated by iptables-save v1.4.14 on Sun Jun 30 03:54:28 2013 *filter :INPUT ACCEPT [0:0] :FORWARD ACCEPT [0:0] :OUTPUT ACCEPT [0:0] :BLACKLIST - [0:0] :LOGDROP - [0:0] :SPAM - [0:0] :THRU - [0:0] :WEB - [0:0] :fail2ban-dovecot-pop3imap - [0:0] :fail2ban-pureftpd - [0:0] :fail2ban-ssh - [0:0] -A INPUT -p tcp -m multiport --dports 110,995,143,993 -j fail2ban-dovecot-pop3imap -A INPUT -p tcp -m multiport --dports 21,21332 -j fail2ban-pureftpd -A INPUT -p tcp -m multiport --dports 22 -j fail2ban-ssh -A INPUT -p tcp -m multiport --dports 110,995,143,993 -j fail2ban-dovecot-pop3imap -A INPUT -i lo -j ACCEPT -A INPUT -i eth0 -p tcp -m tcp --tcp-flags FIN,SYN,RST,PSH,ACK,URG NONE -j DROP -A INPUT -i eth0 -p tcp -m tcp --tcp-flags FIN,SYN FIN,SYN -j DROP -A INPUT -i eth0 -p tcp -m tcp --tcp-flags SYN,RST SYN,RST -j DROP -A INPUT -i eth0 -p tcp -m tcp --tcp-flags FIN,RST FIN,RST -j DROP -A INPUT -i eth0 -p tcp -m tcp --tcp-flags FIN,ACK FIN -j DROP -A INPUT -i eth0 -p tcp -m tcp --tcp-flags ACK,URG URG -j DROP -A INPUT -m conntrack --ctstate RELATED,ESTABLISHED -j ACCEPT -A INPUT -j BLACKLIST -A INPUT -j THRU -A INPUT -j LOGDROP -A OUTPUT -j ACCEPT -A OUTPUT -s 93.223.38.223/32 -j ACCEPT -A BLACKLIST -s 38.113.165.0/24 -j LOGDROP -A BLACKLIST -s 202.177.216.0/24 -j LOGDROP -A BLACKLIST -s 130.117.190.0/24 -j LOGDROP -A BLACKLIST -s 117.79.92.0/24 -j LOGDROP -A BLACKLIST -s 72.47.228.0/24 -j LOGDROP -A BLACKLIST -s 195.200.70.0/24 -j LOGDROP -A BLACKLIST -s 195.200.71.0/24 -j LOGDROP -A LOGDROP -m limit --limit 5/sec -j LOG --log-prefix drop_packet_ --log-level 7 -A LOGDROP -p tcp -m tcp --dport 25 -m limit --limit 2/sec -j LOG --log-prefix spam_blacklist --log-level 7 -A LOGDROP -p tcp -m tcp --dport 80 -m limit --limit 2/sec -j LOG --log-prefix web_blacklist --log-level 7 -A LOGDROP -p tcp -m tcp --dport 22 -m limit --limit 2/sec -j LOG --log-prefix ssh_blacklist --log-level 7 -A LOGDROP -j REJECT --reject-with icmp-host-prohibited -A THRU -p icmp -m limit --limit 1/sec -m icmp --icmp-type 8 -j ACCEPT -A THRU -i eth0 -p tcp -m tcp --dport 25 -j ACCEPT -A THRU -i eth0 -p udp -m udp --dport 53 -j ACCEPT -A THRU -i eth0 -p tcp -m tcp --dport 80 -j ACCEPT -A THRU -i eth0 -p tcp -m tcp --dport 110 -j ACCEPT -A THRU -i eth0 -p tcp -m tcp --dport 143 -j ACCEPT -A THRU -i eth0 -p tcp -m tcp --dport 465 -j ACCEPT -A THRU -i eth0 -p tcp -m tcp --dport 585 -j ACCEPT -A THRU -i eth0 -p tcp -m tcp --dport 993 -j ACCEPT -A THRU -i eth0 -p tcp -m tcp --dport 995 -j ACCEPT -A THRU -i eth0 -p tcp -m tcp --dport 2008 -m state --state NEW,RELATED,ESTABLISHED -j ACCEPT -A THRU -i eth0 -p tcp -m tcp --dport 10011 -m state --state NEW,RELATED,ESTABLISHED -j ACCEPT -A THRU -i eth0 -p tcp -m tcp --dport 21332 -m state --state NEW,RELATED,ESTABLISHED -j ACCEPT -A THRU -i eth0 -p tcp -m tcp --dport 30033 -m state --state NEW,RELATED,ESTABLISHED -j ACCEPT -A fail2ban-dovecot-pop3imap -j RETURN -A fail2ban-dovecot-pop3imap -j RETURN -A fail2ban-pureftpd -j RETURN -A fail2ban-pureftpd -j RETURN -A fail2ban-ssh -j RETURN -A fail2ban-ssh -j RETURN COMMIT # Completed on Sun Jun 30 03:54:28 2013 So, as I said, I have no problems with connecting when adding my IP to go through..but that's not a solution as noone except me can connect anymore~ If someone got an idea what the problem is, please help me! Thanks Cracky

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  • Hosed Windows 7 permissons

    - by Anthony
    Here is the most interesting thing I've noticed since the problems started: If I go into a control panel/system module (in this case the Resource Monitor) that has a "Check Online" type option, Firefox (my default browser) opens right up without a problem. But if I just start Firefox from any shortcuts (start menu, desktop, etc), the Firefox process starts up (and the start menu icon starts glowing) only to end without notice a few seconds later. Possibly related: If I start up in Safe-Mode (w/o Networking, but haven't tried with yet), I can start up FF or Chrome just fine, but if I attempt to open Chrome normally, I get a permissions error. Opera and Safari seem to be okay (mostly). Safari crashes when I try to download any files. All of the above leads me to believe that some (but clearly not all) core files have messed up permissions. Or rather, that I no longer have permission. System still does, based on Firefox opening without fail when the system initiates it. I've run MS Forefront once in normal mode, Malwarebytes twice in normal mode and once in safe-mode. One trojan found and deleted, but the problem persists. Two other things worth mentioning: I accidentally duplicated my library... I thought I'd try to add the "Internet" folder to my start menu, next to music and downloads. The first advanced thing I tried was "create new library". I clearly misunderstood what this means. I thought it was a way to add virtual folders to the library (which I thought, in turn, would allow me to choose it as a link on the start menu), but instead it recreated my already existing user folder, AppData and all. I didn't notice this until today. Then I tried setting permissions for my User folder to full control, recursively... Confused but not giving up,I thought I could maybe create a shortcut to the NetHood folder manually, but instead got hit with an access denied error. So I tried to change the permission levels for all sub-folders to my user folder so that I had full control. I got several access denied errors along the way. At this point I gave up, went out, ended up caught in the rain and stuck on a friend's couch and showing up late for work the next day. Thanks for nothing, Microsoft. When I finally got home today (20 hours later), I noticed that Firefox was acting really strange. I tried opening Chrome to see if the problem was client side or server side, and instead got the above-mentioned "you don't have permission to open this program" alert. And I think that's the whole story. Oh, I also did a system restore, but not chose a point from this morning (an auto update), and it worked but the problem wasn't fixed. And then all the earlier restore points were gone. So the questions are: a) is there a way to set the admin and user privs back to "default"? b) would this, in anyone's expert opinion, fix the problems I'm having? c) how come being logged in as an admin isn't the same as being logged in with admin privs? It seems that half the time I have to do run as admin for fairy standard things because i'm being treated as me-theuser and not me-theadmin. Thanks for reading.

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  • A failed disk (Pay for professional service or SpinRite?)(new edit)

    - by huggie
    EDIT: After much negotiating and begging and seeing through promotion smoke screen, thanks to the nice representative who took my case, I now know that the engineer has already fixed my NTFS partition (I guess it might be a bad block in the partition table?). She told me that the problem was considered minor, and I should be able to boot normally and just copy stuff out. Whew..I'm glad I didn't agree to the NTD $16,000 deal. New question (should this be in a new thread?): is it safer to use the linux "dd" command or is it better to boot normally into Windows XP and just copy stuff out? EDIT2: Thanks to all the help. I give the best answer to Console as it's most directed related to my question. But many suggestion are helpful and informational. ---- ORIGINAL POST BELOW --- Hi, in my previous post (You don't need to read but it's at http://superuser.com/questions/48838/windows-xp-a-disk-read-error-occurred), I said that my hard disk was not booting and is showing "a disk read error occurred". I took it to a recovery professional. A representative responded today told me that the NTFS partitions have a "NTFS partition system crash". I have no idea what that means. The engineer handling my drive will not be available for contact till tomorrow. Now the company charges me NTD (New Taiwan Dollar) $16,000 to recover lost data, that's kind of a lot considering that my graduate student monthly stipend is currently NTD $32,000 (max. allowed by regulation, may be lower, may change depend on funding). Now I'm weighting in between the options. Option A: let the professional recovers it with the half of my monthly stipend. If file/directories I designated are not recovered I don't pay a penny. (other than the initial examination fee of NTD $1000 which I've already paid.) Option B: let me try SpinRite, if failed, back to Option A. I spoke to the representative at the company they recommended me not to handle it on my own (yeah of course that's what they all want to say, right?), and at the price tag the disk error is probably relatively minor and data recoverable. But the representative really did not have detailed information of the disk failure so I didn't take her recommendation readily. Though one thing I heed was that she said that what they would do is to duplicate the disk before attempting discovery, so there would be no data loss (Is this true? can't duplicating invoke further data loss?). That sounds very good to me. Or maybe a third option: Option C: Negotiate with them to pay them to duplicate the disk hopefully for a much smaller price tag. Let me try SpinRite, if failed, back to Option A. This is a difficult decision. Ultimately I want my data back, but if a cheaper way is available to achieve the same thing... Can operating with SpinRite also corrupt data in someway? I've no idea what happened to my drive. I'll attempt to contact the engineer and hope to get it clarified and make an edit here.

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  • Hard drive after PCB swap strange stuff

    - by ramyy
    I’ve done a PCB swap to my HDD. The HDD model is: WD6400AAKS-00A7B2. The original PCB PN matches the new one (first three letter groups), though the cache mismatches (16MB original, 8MB new). The Hardware store that made the swap told me it was hard to do the swap, they have done firmware adaptation. I can see that this firmware version does not match the original, (01.03B01 original, 05.04E05 new). Still I can see that the serial number and model of the drive is correct, the hard drive appeared normal in the BIOS, all the partitions show and everything appears normal. I have encountered three things though, I have left the drive non operated for 2-3 weeks after the swap to avoid corrupting the data or anything else the new PCB might cause, until I buy a new drive and backup the data. I got a drive, and when I powered the old drive manually (I have a laptop, I use a normal desktop power supply and a USB/SATA connector), I heard the motor start and I could hear ticking as if the motor’s somehow struggling to start, and then the motor sound starts again then the ticking, and so on.. I tried powering again it happened again. The third time it started normally and I could see everything normally. I took the chance and copied all the data over to the new drive. When I was done, I powered off the drive (after more than 25 hours of continuous operation), tried to power it up again and it did so normally, and so are the times I powered it up later; but I got very suspicious now. What could be the problem here? And what happened new, it used to power normally after the swap directly? The second thing that happened is that I found size differences with some files; some include movies, songs, (.iso) files for games, and programs. I could find the size is the same, but size on disk is a little more on the new drive for these files. . I’ve tried some of those files (with size differences) they worked fine. They are not too much but still make you suspicious of the integrity of the data copied; one cannot try if all files are working for about (580 GB) worth of data. I tried copying these files on the same partition they exist of the old drive; they are the same in size as when copied to the new drive (allocation unit size not the issue). I took an image of a partition (sector by sector including empty ones) and when I explore it, these file sizes are equal to the original (old drive); I copy them anywhere else their size on disk, increases, i.e becomes equal to the ones I copy from the old drive itself anywhere. Why the size difference and can one trust the integrity of the data?? The third thing is that when I connect my new external USB HDD, the partitions of the old HDD unmount and then mount again. Connected are: (USB mouse + Old HDD) then external HDD. Why that happens?? Considering the following: I compared the SMART reports from after the swap directly and after the copying, no error readings or reallocated sectors where reported. Here they are: http://www.image-share.com/ijpg-1939-219.html I later ran both WD data life guard tests and they came out passed. I’m worried for this drive since I must be sure the data is fine and safe on the new one, and I will consider it backup for the new one, since you can’t trust anything anymore. I hope you can forgive me for the length of the post, but couldn’t ignore any of the details, this hard drive contains very important data to me and I have to deal with the situation with great care.

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  • ActiveX component can't create Object Error? Check 64 bit Status

    - by Rick Strahl
    If you're running on IIS 7 and a 64 bit operating system you might run into the following error using ASP classic or ASP.NET with COM interop. In classic ASP applications the error will show up as: ActiveX component can't create object   (Error 429) (actually without error handling the error just shows up as 500 error page) In my case the code that's been giving me problems has been a FoxPro COM object I'd been using to serve banner ads to some of my pages. The code basically looks up banners from a database table and displays them at random. The ASP classic code that uses it looks like this: <% Set banner = Server.CreateObject("wwBanner.aspBanner") banner.BannerFile = "wwsitebanners" Response.Write(banner.GetBanner(-1)) %> Originally this code had no specific error checking as above so the ASP pages just failed with 500 error pages from the Web server. To find out what the problem is this code is more useful at least for debugging: <% ON ERROR RESUME NEXT Set banner = Server.CreateObject("wwBanner.aspBanner") Response.Write(err.Number & " - " & err.Description) banner.BannerFile = "wwsitebanners" Response.Write(banner.GetBanner(-1)) %> which results in: 429 - ActiveX component can't create object which at least gives you a slight clue. In ASP.NET invoking the same COM object with code like this: <% dynamic banner = wwUtils.CreateComInstance("wwBanner.aspBanner") as dynamic; banner.cBANNERFILE = "wwsitebanners"; Response.Write(banner.getBanner(-1)); %> results in: Retrieving the COM class factory for component with CLSID {B5DCBB81-D5F5-11D2-B85E-00600889F23B} failed due to the following error: 80040154 Class not registered (Exception from HRESULT: 0x80040154 (REGDB_E_CLASSNOTREG)). The class is in fact registered though and the COM server loads fine from a command prompt or other COM client. This error can be caused by a COM server that doesn't load. It looks like a COM registration error. There are a number of traditional reasons why this error can crop up of course. The server isn't registered (run regserver32 to register a DLL server or /regserver on an EXE server) Access permissions aren't set on the COM server (Web account has to be able to read the DLL ie. Network service) The COM server fails to load during initialization ie. failing during startup One thing I always do to check for COM errors fire up the server in a COM client outside of IIS and ensure that it works there first - it's almost always easier to debug a server outside of the Web environment. In my case I tried the server in Visual FoxPro on the server with: loBanners = CREATEOBJECT("wwBanner.aspBanner") loBanners.cBannerFile = "wwsitebanners" ? loBanners.GetBanner(-1) and it worked just fine. If you don't have a full dev environment on the server you can also use VBScript do the same thing and run the .vbs file from the command prompt: Set banner = Server.CreateObject("wwBanner.aspBanner") banner.BannerFile = "wwsitebanners" MsgBox(banner.getBanner(-1)) Since this both works it tells me the server is registered and working properly. This leaves startup failures or permissions as the problem. I double checked permissions for the Application Pool and the permissions of the folder where the DLL lives and both are properly set to allow access by the Application Pool impersonated user. Just to be sure I assigned an Admin user to the Application Pool but still no go. So now what? 64 bit Servers Ahoy A couple of weeks back I had set up a few of my Application pools to 64 bit mode. My server is Server 2008 64 bit and by default Application Pools run 64 bit. Originally when I installed the server I set up most of my Application Pools to 32 bit mainly for backwards compatibility. But as more of my code migrates to 64 bit OS's I figured it'd be a good idea to see how well code runs under 64 bit code. The transition has been mostly painless. Until today when I noticed the problem with the code above when scrolling to my IIS logs and noticing a lot of 500 errors on many of my ASP classic pages. The code in question in most of these pages deals with this single simple COM object. It took a while to figure out that the problem is caused by the Application Pool running in 64 bit mode. The issue is that 32 bit COM objects (ie. my old Visual FoxPro COM component) cannot be loaded in a 64 bit Application Pool. The ASP pages using this COM component broke on the day I switched my main Application Pool into 64 bit mode but I didn't find the problem until I searched my logs for errors by pure chance. To fix this is easy enough once you know what the problem is by switching the Application Pool to Enable 32-bit Applications: Once this is done the COM objects started working correctly again. 64 bit ASP and ASP.NET with DCOM Servers This is kind of off topic, but incidentally it's possible to load 32 bit DCOM (out of process) servers from ASP.NET and ASP classic even if those applications run in 64 bit application pools. In fact, in West Wind Web Connection I use this capability to run a 64 bit ASP.NET handler that talks to a 32 bit FoxPro COM server which allows West Wind Web Connection to run in native 64 bit mode without custom configuration (which is actually quite useful). It's probably not a common usage scenario but it's good to know that you can actually access 32 bit COM objects this way from ASP.NET. For West Wind Web Connection this works out well as the DCOM interface only makes one non-chatty call to the backend server that handles all the rest of the request processing. Application Pool Isolation is your Friend For me the recent incident of failure in the classic ASP pages has just been another reminder to be very careful with moving applications to 64 bit operation. There are many little traps when switching to 64 bit that are very difficult to track and test for. I described one issue I had a couple of months ago where one of the default ASP.NET filters was loading the wrong version (32bit instead of 64bit) which was extremely difficult to track down and was caused by a very sneaky configuration switch error (basically 3 different entries for the same ISAPI filter all with different bitness settings). It took me almost a full day to track this down). Recently I've been taken to isolate individual applications into separate Application Pools rather than my past practice of combining many apps into shared AppPools. This is a good practice assuming you have enough memory to make this work. Application Pool isolate provides more modularity and allows me to selectively move applications to 64 bit. The error above came about precisely because I moved one of my most populous app pools to 64 bit and forgot about the minimal COM object use in some of my old pages. It's easy to forget. To 64bit or Not Is it worth it to move to 64 bit? Currently I'd say -not really. In my - admittedly limited - testing I don't see any significant performance increases. In fact 64 bit apps just seem to consume considerably more memory (30-50% more in my pools on average) and performance is minimally improved (less than 5% at the very best) in the load testing I've performed on a couple of sites in both modes. The only real incentive for 64 bit would be applications that require huge data spaces that exceed the 32 bit 4 gigabyte memory limit. However I have a hard time imagining an application that needs 4 gigs of memory in a single Application Pool :-). Curious to hear other opinions on benefits of 64 bit operation. © Rick Strahl, West Wind Technologies, 2005-2011Posted in COM   ASP.NET  FoxPro  

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  • Mulit-tenant ASP.NET MVC – Controllers

    - by zowens
    Part I – Introduction Part II – Foundation   The time has come to talk about controllers in a multi-tenant ASP.NET MVC architecture. This is actually the most critical design decision you will make when dealing with multi-tenancy with MVC. In my design, I took into account the design goals I mentioned in the introduction about inversion of control and what a tenant is to my design. Be aware that this is only one way to achieve multi-tenant controllers.   The Premise MvcEx (which is a sample written by Rob Ashton) utilizes dynamic controllers. Essentially a controller is “dynamic” in that multiple action results can be placed in different “controllers” with the same name. This approach is a bit too complicated for my design. I wanted to stick with plain old inheritance when dealing with controllers. The basic premise of my controller design is that my main host defines a set of universal controllers. It is the responsibility of the tenant to decide if the tenant would like to utilize these core controllers. This can be done either by straight usage of the controller or inheritance for extension of the functionality defined by the controller. The controller is resolved by a StructureMap container that is attached to the tenant, as discussed in Part II.   Controller Resolution I have been thinking about two different ways to resolve controllers with StructureMap. One way is to use named instances. This is a really easy way to simply pull the controller right out of the container without a lot of fuss. I ultimately chose not to use this approach. The reason for this decision is to ensure that the controllers are named properly. If a controller has a different named instance that the controller type, then the resolution has a significant disconnect and there are no guarantees. The final approach, the one utilized by the sample, is to simply pull all controller types and correlate the type with a controller name. This has a bit of a application start performance disadvantage, but is significantly more approachable for maintainability. For example, if I wanted to go back and add a “ControllerName” attribute, I would just have to change the ControllerFactory to suit my needs.   The Code The container factory that I have built is actually pretty simple. That’s really all we need. The most significant method is the GetControllersFor method. This method makes the model from the Container and determines all the concrete types for IController.  The thing you might notice is that this doesn’t depend on tenants, but rather containers. You could easily use this controller factory for an application that doesn’t utilize multi-tenancy. public class ContainerControllerFactory : IControllerFactory { private readonly ThreadSafeDictionary<IContainer, IDictionary<string, Type>> typeCache; public ContainerControllerFactory(IContainerResolver resolver) { Ensure.Argument.NotNull(resolver, "resolver"); this.ContainerResolver = resolver; this.typeCache = new ThreadSafeDictionary<IContainer, IDictionary<string, Type>>(); } public IContainerResolver ContainerResolver { get; private set; } public virtual IController CreateController(RequestContext requestContext, string controllerName) { var controllerType = this.GetControllerType(requestContext, controllerName); if (controllerType == null) return null; var controller = this.ContainerResolver.Resolve(requestContext).GetInstance(controllerType) as IController; // ensure the action invoker is a ContainerControllerActionInvoker if (controller != null && controller is Controller && !((controller as Controller).ActionInvoker is ContainerControllerActionInvoker)) (controller as Controller).ActionInvoker = new ContainerControllerActionInvoker(this.ContainerResolver); return controller; } public void ReleaseController(IController controller) { if (controller != null && controller is IDisposable) ((IDisposable)controller).Dispose(); } internal static IEnumerable<Type> GetControllersFor(IContainer container) { Ensure.Argument.NotNull(container); return container.Model.InstancesOf<IController>().Select(x => x.ConcreteType).Distinct(); } protected virtual Type GetControllerType(RequestContext requestContext, string controllerName) { Ensure.Argument.NotNull(requestContext, "requestContext"); Ensure.Argument.NotNullOrEmpty(controllerName, "controllerName"); var container = this.ContainerResolver.Resolve(requestContext); var typeDictionary = this.typeCache.GetOrAdd(container, () => GetControllersFor(container).ToDictionary(x => ControllerFriendlyName(x.Name))); Type found = null; if (typeDictionary.TryGetValue(ControllerFriendlyName(controllerName), out found)) return found; return null; } private static string ControllerFriendlyName(string value) { return (value ?? string.Empty).ToLowerInvariant().Without("controller"); } } One thing to note about my implementation is that we do not use namespaces that can be utilized in the default ASP.NET MVC controller factory. This is something that I don’t use and have no desire to implement and test. The reason I am not using namespaces in this situation is because each tenant has its own namespaces and the routing would not make sense in this case.   Because we are using IoC, dependencies are automatically injected into the constructor. For example, a tenant container could implement it’s own IRepository and a controller could be defined in the “main” project. The IRepository from the tenant would be injected into the main project’s controller. This is quite a useful feature.   Again, the source code is on GitHub here.   Up Next Up next is the view resolution. This is a complicated issue, so be prepared. I hope that you have found this series useful. If you have any questions about my implementation so far, send me an email or DM me on Twitter. I have had a lot of great conversations about multi-tenancy so far and I greatly appreciate the feedback!

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  • Compiling examples for consuming the REST Endpoints for WCF Service using Agatha

    - by REA_ANDREW
    I recently made two contributions to the Agatha Project by Davy Brion over on Google Code, and one of the things I wanted to follow up with was a post showing examples and some, seemingly required tid bits.  The contributions which I made where: To support StructureMap To include REST (JSON and XML) support for the service contract The examples which I have made, I want to format them so they fit in with the current format of examples over on Agatha and hopefully create and submit a third patch which will include these examples to help others who wish to use these additions. Whilst building these examples for both XML and JSON I have learnt a couple of things which I feel are not really well documented, but are extremely good practice and once known make perfect sense.  I have chosen a real basic e-commerce context for my example Requests and Responses, and have also made use of the excellent tool AutoMapper, again on Google Code. Setting the scene I have followed the Pipes and Filters Pattern with the IQueryable interface on my Repository and exposed the following methods to query Products: IQueryable<Product> GetProducts(); IQueryable<Product> ByCategoryName(this IQueryable<Product> products, string categoryName) Product ByProductCode(this IQueryable<Product> products, String productCode) I have an interface for the IProductRepository but for the concrete implementation I have simply created a protected getter which populates a private List<Product> with 100 test products with random data.  Another good reason for following an interface based approach is that it will demonstrate usage of my first contribution which is the StructureMap support.  Finally the two Domain Objects I have made are Product and Category as shown below: public class Product { public String ProductCode { get; set; } public String Name { get; set; } public Decimal Price { get; set; } public Decimal Rrp { get; set; } public Category Category { get; set; } }   public class Category { public String Name { get; set; } }   Requirements for the REST Support One of the things which you will notice with Agatha is that you do not have to decorate your Request and Response objects with the WCF Service Model Attributes like DataContract, DataMember etc… Unfortunately from what I have seen, these are required if you want the same types to work with your REST endpoint.  I have not tried but I assume the same result can be achieved by simply decorating the same classes with the Serializable Attribute.  Without this the operation will fail. Another surprising thing I have found is that it did not work until I used the following Attribute parameters: Name Namespace e.g. [DataContract(Name = "GetProductsRequest", Namespace = "AgathaRestExample.Service.Requests")] public class GetProductsRequest : Request { }   Although I was surprised by this, things kind of explained themselves when I got round to figuring out the exact construct required for both the XML and the REST.  One of the things which you already know and are then reminded of is that each of your Requests and Responses ultimately inherit from an abstract base class respectively. This information needs to be represented in a way native to the format being used.  I have seen this in XML but I have not seen the format which is required for the JSON. JSON Consumer Example I have used JQuery to create the example and I simply want to make two requests to the server which as you will know with Agatha are transmitted inside an array to reduce the service calls.  I have also used a tool called json2 which is again over at Google Code simply to convert my JSON expression into its string format for transmission.  You will notice that I specify the type of Request I am using and the relevant Namespace it belongs to.  Also notice that the second request has a parameter so each of these two object are representing an abstract Request and the parameters of the object describe it. <script type="text/javascript"> var bodyContent = $.ajax({ url: "http://localhost:50348/service.svc/json/processjsonrequests", global: false, contentType: "application/json; charset=utf-8", type: "POST", processData: true, data: JSON.stringify([ { __type: "GetProductsRequest:AgathaRestExample.Service.Requests" }, { __type: "GetProductsByCategoryRequest:AgathaRestExample.Service.Requests", CategoryName: "Category1" } ]), dataType: "json", success: function(msg) { alert(msg); } }).responseText; </script>   XML Consumer Example For the XML Consumer example I have chosen to use a simple Console Application and make a WebRequest to the service using the XML as a request.  I have made a crude static method which simply reads from an XML File, replaces some value with a parameter and returns the formatted XML.  I say crude but it simply shows how XML Templates for each type of Request could be made and then have a wrapper utility in whatever language you use to combine the requests which are required.  The following XML is the same Request array as shown above but simply in the XML Format. <?xml version="1.0" encoding="utf-8" ?> <ArrayOfRequest xmlns="http://schemas.datacontract.org/2004/07/Agatha.Common" xmlns:i="http://www.w3.org/2001/XMLSchema-instance"> <Request i:type="a:GetProductsRequest" xmlns:a="AgathaRestExample.Service.Requests"/> <Request i:type="a:GetProductsByCategoryRequest" xmlns:a="AgathaRestExample.Service.Requests"> <a:CategoryName>{CategoryName}</a:CategoryName> </Request> </ArrayOfRequest>   It is funny because I remember submitting a question to StackOverflow asking whether there was a REST Client Generation tool similar to what Microsoft used for their RestStarterKit but which could be applied to existing services which have REST endpoints attached.  I could not find any but this is now definitely something which I am going to build, as I think it is extremely useful to have but also it should not be too difficult based on the information I now know about the above.  Finally I thought that the Strategy Pattern would lend itself really well to this type of thing so it can accommodate for different languages. I think that is about it, I have included the code for the example Console app which I made below incase anyone wants to have a mooch at the code.  As I said above I want to reformat these to fit in with the current examples over on the Agatha project, but also now thinking about it, make a Documentation Web method…{brain ticking} :-) Cheers for now and here is the final bit of code: static void Main(string[] args) { var request = WebRequest.Create("http://localhost:50348/service.svc/xml/processxmlrequests"); request.Method = "POST"; request.ContentType = "text/xml"; using(var writer = new StreamWriter(request.GetRequestStream())) { writer.WriteLine(GetExampleRequestsString("Category1")); } var response = request.GetResponse(); using(var reader = new StreamReader(response.GetResponseStream())) { Console.WriteLine(reader.ReadToEnd()); } Console.ReadLine(); } static string GetExampleRequestsString(string categoryName) { var data = File.ReadAllText(Path.Combine(Path.GetDirectoryName(Assembly.GetExecutingAssembly().Location), "ExampleRequests.xml")); data = data.Replace("{CategoryName}", categoryName); return data; } }

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  • MvcExtensions – Bootstrapping

    - by kazimanzurrashid
    When you create a new ASP.NET MVC application you will find that the global.asax contains the following lines: namespace MvcApplication1 { // Note: For instructions on enabling IIS6 or IIS7 classic mode, // visit http://go.microsoft.com/?LinkId=9394801 public class MvcApplication : System.Web.HttpApplication { public static void RegisterRoutes(RouteCollection routes) { routes.IgnoreRoute("{resource}.axd/{*pathInfo}"); routes.MapRoute( "Default", // Route name "{controller}/{action}/{id}", // URL with parameters new { controller = "Home", action = "Index", id = UrlParameter.Optional } // Parameter defaults ); } protected void Application_Start() { AreaRegistration.RegisterAllAreas(); RegisterRoutes(RouteTable.Routes); } } } As the application grows, there are quite a lot of plumbing code gets into the global.asax which quickly becomes a design smell. Lets take a quick look at the code of one of the open source project that I recently visited: public static void RegisterRoutes(RouteCollection routes) { routes.IgnoreRoute("{resource}.axd/{*pathInfo}"); routes.MapRoute("Default","{controller}/{action}/{id}", new { controller = "Home", action = "Index", id = "" }); } protected override void OnApplicationStarted() { Error += OnError; EndRequest += OnEndRequest; var settings = new SparkSettings() .AddNamespace("System") .AddNamespace("System.Collections.Generic") .AddNamespace("System.Web.Mvc") .AddNamespace("System.Web.Mvc.Html") .AddNamespace("MvcContrib.FluentHtml") .AddNamespace("********") .AddNamespace("********.Web") .SetPageBaseType("ApplicationViewPage") .SetAutomaticEncoding(true); #if DEBUG settings.SetDebug(true); #endif var viewFactory = new SparkViewFactory(settings); ViewEngines.Engines.Add(viewFactory); #if !DEBUG PrecompileViews(viewFactory); #endif RegisterAllControllersIn("********.Web"); log4net.Config.XmlConfigurator.Configure(); RegisterRoutes(RouteTable.Routes); Factory.Load(new Components.WebDependencies()); ModelBinders.Binders.DefaultBinder = new Binders.GenericBinderResolver(Factory.TryGet<IModelBinder>); ValidatorConfiguration.Initialize("********"); HtmlValidationExtensions.Initialize(ValidatorConfiguration.Rules); } private void OnEndRequest(object sender, System.EventArgs e) { if (((HttpApplication)sender).Context.Handler is MvcHandler) { CreateKernel().Get<ISessionSource>().Close(); } } private void OnError(object sender, System.EventArgs e) { CreateKernel().Get<ISessionSource>().Close(); } protected override IKernel CreateKernel() { return Factory.Kernel; } private static void PrecompileViews(SparkViewFactory viewFactory) { var batch = new SparkBatchDescriptor(); batch.For<HomeController>().For<ManageController>(); viewFactory.Precompile(batch); } As you can see there are quite a few of things going on in the above code, Registering the ViewEngine, Compiling the Views, Registering the Routes/Controllers/Model Binders, Settings up Logger, Validations and as you can imagine the more it becomes complex the more things will get added in the application start. One of the goal of the MVCExtensions is to reduce the above design smell. Instead of writing all the plumbing code in the application start, it contains BootstrapperTask to register individual services. Out of the box, it contains BootstrapperTask to register Controllers, Controller Factory, Action Invoker, Action Filters, Model Binders, Model Metadata/Validation Providers, ValueProvideraFactory, ViewEngines etc and it is intelligent enough to automatically detect the above types and register into the ASP.NET MVC Framework. Other than the built-in tasks you can create your own custom task which will be automatically executed when the application starts. When the BootstrapperTasks are in action you will find the global.asax pretty much clean like the following: public class MvcApplication : UnityMvcApplication { public void ErrorLog_Filtering(object sender, ExceptionFilterEventArgs e) { Check.Argument.IsNotNull(e, "e"); HttpException exception = e.Exception.GetBaseException() as HttpException; if ((exception != null) && (exception.GetHttpCode() == (int)HttpStatusCode.NotFound)) { e.Dismiss(); } } } The above code is taken from my another open source project Shrinkr, as you can see the global.asax is longer cluttered with any plumbing code. One special thing you have noticed that it is inherited from the UnityMvcApplication rather than regular HttpApplication. There are separate version of this class for each IoC Container like NinjectMvcApplication, StructureMapMvcApplication etc. Other than executing the built-in tasks, the Shrinkr also has few custom tasks which gets executed when the application starts. For example, when the application starts, we want to ensure that the default users (which is specified in the web.config) are created. The following is the custom task that is used to create those default users: public class CreateDefaultUsers : BootstrapperTask { protected override TaskContinuation ExecuteCore(IServiceLocator serviceLocator) { IUserRepository userRepository = serviceLocator.GetInstance<IUserRepository>(); IUnitOfWork unitOfWork = serviceLocator.GetInstance<IUnitOfWork>(); IEnumerable<User> users = serviceLocator.GetInstance<Settings>().DefaultUsers; bool shouldCommit = false; foreach (User user in users) { if (userRepository.GetByName(user.Name) == null) { user.AllowApiAccess(ApiSetting.InfiniteLimit); userRepository.Add(user); shouldCommit = true; } } if (shouldCommit) { unitOfWork.Commit(); } return TaskContinuation.Continue; } } There are several other Tasks in the Shrinkr that we are also using which you will find in that project. To create a custom bootstrapping task you have create a new class which either implements the IBootstrapperTask interface or inherits from the abstract BootstrapperTask class, I would recommend to start with the BootstrapperTask as it already has the required code that you have to write in case if you choose the IBootstrapperTask interface. As you can see in the above code we are overriding the ExecuteCore to create the default users, the MVCExtensions is responsible for populating the  ServiceLocator prior calling this method and in this method we are using the service locator to get the dependencies that are required to create the users (I will cover the custom dependencies registration in the next post). Once the users are created, we are returning a special enum, TaskContinuation as the return value, the TaskContinuation can have three values Continue (default), Skip and Break. The reason behind of having this enum is, in some  special cases you might want to skip the next task in the chain or break the complete chain depending upon the currently running task, in those cases you will use the other two values instead of the Continue. The last thing I want to cover in the bootstrapping task is the Order. By default all the built-in tasks as well as newly created task order is set to the DefaultOrder(a static property), in some special cases you might want to execute it before/after all the other tasks, in those cases you will assign the Order in the Task constructor. For Example, in Shrinkr, we want to run few background services when the all the tasks are executed, so we assigned the order as DefaultOrder + 1. Here is the code of that Task: public class ConfigureBackgroundServices : BootstrapperTask { private IEnumerable<IBackgroundService> backgroundServices; public ConfigureBackgroundServices() { Order = DefaultOrder + 1; } protected override TaskContinuation ExecuteCore(IServiceLocator serviceLocator) { backgroundServices = serviceLocator.GetAllInstances<IBackgroundService>().ToList(); backgroundServices.Each(service => service.Start()); return TaskContinuation.Continue; } protected override void DisposeCore() { backgroundServices.Each(service => service.Stop()); } } That’s it for today, in the next post I will cover the custom service registration, so stay tuned.

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  • Validation in Silverlight

    - by Timmy Kokke
    Getting started with the basics Validation in Silverlight can get very complex pretty easy. The DataGrid control is the only control that does data validation automatically, but often you want to validate your own entry form. Values a user may enter in this form can be restricted by the customer and have to fit an exact fit to a list of requirements or you just want to prevent problems when saving the data to the database. Showing a message to the user when a value is entered is pretty straight forward as I’ll show you in the following example.     This (default) Silverlight textbox is data-bound to a simple data class. It has to be bound in “Two-way” mode to be sure the source value is updated when the target value changes. The INotifyPropertyChanged interface must be implemented by the data class to get the notification system to work. When the property changes a simple check is performed and when it doesn’t match some criteria an ValidationException is thrown. The ValidatesOnExceptions binding attribute is set to True to tell the textbox it should handle the thrown ValidationException. Let’s have a look at some code now. The xaml should contain something like below. The most important part is inside the binding. In this case the Text property is bound to the “Name” property in TwoWay mode. It is also told to validate on exceptions. This property is false by default.   <StackPanel Orientation="Horizontal"> <TextBox Width="150" x:Name="Name" Text="{Binding Path=Name, Mode=TwoWay, ValidatesOnExceptions=True}"/> <TextBlock Text="Name"/> </StackPanel>   The data class in this first example is a very simplified person class with only one property: string Name. The INotifyPropertyChanged interface is implemented and the PropertyChanged event is fired when the Name property changes. When the property changes a check is performed to see if the new string is null or empty. If this is the case a ValidationException is thrown explaining that the entered value is invalid.   public class PersonData:INotifyPropertyChanged { private string _name; public string Name { get { return _name; } set { if (_name != value) { if(string.IsNullOrEmpty(value)) throw new ValidationException("Name is required"); _name = value; if (PropertyChanged != null) PropertyChanged(this, new PropertyChangedEventArgs("Name")); } } } public event PropertyChangedEventHandler PropertyChanged=delegate { }; } The last thing that has to be done is letting binding an instance of the PersonData class to the DataContext of the control. This is done in the code behind file. public partial class Demo1 : UserControl { public Demo1() { InitializeComponent(); this.DataContext = new PersonData() {Name = "Johnny Walker"}; } }   Error Summary In many cases you would have more than one entry control. A summary of errors would be nice in such case. With a few changes to the xaml an error summary, like below, can be added.           First, add a namespace to the xaml so the control can be used. Add the following line to the header of the .xaml file. xmlns:Controls="clr-namespace:System.Windows.Controls;assembly=System.Windows.Controls.Data.Input"   Next, add the control to the layout. To get the result as in the image showed earlier, add the control right above the StackPanel from the first example. It’s got a small margin to separate it from the textbox a little.   <Controls:ValidationSummary Margin="8"/>   The ValidationSummary control has to be notified that an ValidationException occurred. This can be done with a small change to the xaml too. Add the NotifyOnValidationError to the binding expression. By default this value is set to false, so nothing would be notified. Set the property to true to get it to work.   <TextBox Width="150" x:Name="Name" Text="{Binding Name, Mode=TwoWay, ValidatesOnExceptions=True, NotifyOnValidationError=True}"/>   Data annotation Validating data in the setter is one option, but not my personal favorite. It’s the easiest way if you have a single required value you want to check, but often you want to validate more. Besides, I don’t consider it best practice to write logic in setters. The way used by frameworks like WCF Ria Services is the use of attributes on the properties. Instead of throwing exceptions you have to call the static method ValidateProperty on the Validator class. This call stays always the same for a particular property, not even when you change the attributes on the property. To mark a property “Required” you can use the RequiredAttribute. This is what the Name property is going to look like:   [Required] public string Name { get { return _name; } set { if (_name != value) { Validator.ValidateProperty(value, new ValidationContext(this, null, null){ MemberName = "Name" }); _name = value; if (PropertyChanged != null) PropertyChanged(this, new PropertyChangedEventArgs("Name")); } } }   The ValidateProperty method takes the new value for the property and an instance of ValidationContext. The properties passed to the constructor of the ValidationContextclass are very straight forward. This part is the same every time. The only thing that changes is the MemberName property of the ValidationContext. Property has to hold the name of the property you want to validate. It’s the same value you provide the PropertyChangedEventArgs with. The System.ComponentModel.DataAnnotation contains eight different validation attributes including a base class to create your own. They are: RequiredAttribute Specifies that a value must be provided. RangeAttribute The provide value must fall in the specified range. RegularExpressionAttribute Validates is the value matches the regular expression. StringLengthAttribute Checks if the number of characters in a string falls between a minimum and maximum amount. CustomValidationAttribute Use a custom method to validate the value. DataTypeAttribute Specify a data type using an enum or a custom data type. EnumDataTypeAttribute Makes sure the value is found in a enum. ValidationAttribute A base class for custom validation attributes All of these will ensure that an validation exception is thrown, except the DataTypeAttribute. This attribute is used to provide some additional information about the property. You can use this information in your own code.   [Required] [Range(0,125,ErrorMessage = "Value is not a valid age")] public int Age {   It’s no problem to stack different validation attributes together. For example, when an Age is required and must fall in the range from 0 to 125:   [Required, StringLength(255,MinimumLength = 3)] public string Name {   Or in one row like this, for a required Name with at least 3 characters and a maximum of 255:   Delayed validation Having properties marked as required can be very useful. The only downside to the technique described earlier is that you have to change the value in order to get it validated. What if you start out with empty an empty entry form? All fields are empty and thus won’t be validated. With this small trick you can validate at the moment the user click the submit button.   <TextBox Width="150" x:Name="NameField" Text="{Binding Name, Mode=TwoWay, ValidatesOnExceptions=True, NotifyOnValidationError=True, UpdateSourceTrigger=Explicit}"/>   By default, when a TwoWay bound control looses focus the value is updated. When you added validation like I’ve shown you earlier, the value is validated. To overcome this, you have to tell the binding update explicitly by setting the UpdateSourceTrigger binding property to Explicit:   private void SubmitButtonClick(object sender, RoutedEventArgs e) { NameField.GetBindingExpression(TextBox.TextProperty).UpdateSource(); }   This way, the binding is in two direction but the source is only updated, thus validated, when you tell it to. In the code behind you have to call the UpdateSource method on the binding expression, which you can get from the TextBox.   Conclusion Data validation is something you’ll probably want on almost every entry form. I always thought it was hard to do, but it wasn’t. If you can throw an exception you can do validation. If you want to know anything more in depth about something I talked about in this article let me know. I might write an entire post to that.

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  • Twitter traffic might not be what it seems

    - by Piet
    Are you using bit.ly stats to measure interest in the links you post on twitter? I’ve been hearing for a while about people claiming to get the majority of their traffic originating from twitter these days. Now, I’ve been playing with the twitter ruby gem recently, doing various experiments which I’ll not go into detail here because they could be regarded as spamming… if I’d conduct them on a large scale, that is. It’s scary to see people actually engaging with @replies crafted with some regular expressions and eliza-like trickery on status updates found using the twitter api. I’m wondering how Twitter is going to contain the coming spam-flood. When posting links I used bit.ly as url shortener, since this one seems to be the de-facto standard on twitter. A nice thing about bit.ly is that it shows some basic stats about the redirects it performs for your shortened links. To my surprise, most links posted almost immediately resulted in several visitors. Now, seeing that I was posting the links together with some information concerning what the link is about, I concluded that the people who were actually clicking the links should be very targeted visitors. This felt a bit like free adwords, and I suddenly started to understand why everyone was raving about getting traffic from twitter. How wrong I was! (and I think several 1000 online marketers with me) On the destination site I used a traffic logging solution that works by including a little javascript snippet in your pages. It seemed that somehow all visitors disappeared after the bit.ly redirect and before getting to the site, because I was hardly seeing any visitors there. So I started investigating what was happening: by looking at the logfiles of the destination site, and by making my own ’shortened’ urls by doing redirects using a very short domain name I own. This way, I could check the apache access_log before the redirects. Most user agents turned out to be bots without a doubt. Here’s an excerpt of user-agents awk’ed from apache’s access_log for a time period of about one hour, right after posting some links: AideRSS 2.0 (postrank.com) Java/1.6.0_13 Java/1.6.0_14 libwww-perl/5.816 MLBot (www.metadatalabs.com/mlbot) Mozilla/4.0 (compatible;MSIE 5.01; Windows -NT 5.0 - real-url.org) Mozilla/5.0 (compatible; Twitturls; +http://twitturls.com) Mozilla/5.0 (compatible; Viralheat Bot/1.0; +http://www.viralheat.com/) Mozilla/5.0 (Danger hiptop 4.6; U; rv:1.7.12) Gecko/20050920 Mozilla/5.0 (X11; U; Linux i686; en-us; rv:1.9.0.2) Gecko/2008092313 Ubuntu/9.04 (jaunty) Firefox/3.5 OpenCalaisSemanticProxy PycURL/7.18.2 PycURL/7.19.3 Python-urllib/1.17 Twingly Recon twitmatic Twitturly / v0.6 Wget/1.10.2 (Red Hat modified) Wget/1.11.1 (Red Hat modified) Of the few user-agents that seem ‘real’ at first, half are originating from an ip-address used by Amazon EC2. And I doubt people are setting op proxies on there. Oh yeah, Googlebot (the real deal, from a legit google owned address) is sucking up posted links like fresh oysters. I guess google is trying to make sure in advance to never be beaten by twitter in the ‘realtime search’ department. Actually, I think it’d be almost stupid NOT to post any new pages/posts/websites on Twitter, it must be one of the fastest ways to get a Googlebot visit. Same experiment with a real, established twitter account Now, because I was posting the url’s either as ’status’ messages or directed @people, on a test-account with hardly any (human) followers, I checked again using the twitter accounts from a commercial site I’m involved with. These accounts all have between 500 and 1000 targeted (I think) followers. I checked the destination access_logs and also added ‘my’ redirect after the bit.ly redirect: same results, although seemingly a bit higher real visitor/bot ratio. Btw: one of these account was ‘punished’ with a 1 week lock recently because the same (1 one!) status update was sent that was sent right before using another account. They got an email explaining the lock because the account didn’t act according to their TOS. I can’t find anything in their TOS about it, can you? I don’t think Twitter is on the right track punishing a legit account, knowing the trickery I had been doing with it’s api went totally unpunished. I might be wrong though, I often am. On the other hand: this commercial site reported targeted traffic and actual signups from visitors coming from Twitter. The ones that are really real visitors are also very targeted. I’m just not sure if the amount of work involved could hold up against an adwords campaign. Reposting the same link over and over again helps On thing I noticed: It helps to keep on reposting the same links with regular intervals. I guess most people only look at their first page when checking out recent posts of the ones they’re following, or don’t look too far back when performing a search. Now, this probably isn’t according to the twitter TOS. Actually, it might be spamming but no-one is obligated to follow anyone else of course. This way, I was getting more real visitors and less bots. To my surprise (when my programmer’s hat is on) there were still repeated visits from the same bots coming from the same ip-addresses. Did they expect to find something else when visiting for a 2nd or 3rd time? (actually,this gave me an idea: you can’t change a link once it’s posted, but you can change where it redirects to) Most bots were smart enough not to follow the same link again though. Are you successful in getting real visitors from Twitter? Are you only relying on bit.ly to provide traffic stats?

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