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  • Generic Type constraint in .net

    - by Jose
    Okay I'm looking for some input, I'm pretty sure this is not currently supported in .NET 3.5 but here goes. I want to require a generic type passed into my class to have a constructor like this: new(IDictionary<string,object>) so the class would look like this public MyClass<T> where T : new(IDictionary<string,object>) { T CreateObject(IDictionary<string,object> values) { return new T(values); } } But the compiler doesn't support this, it doesn't really know what I'm asking. Some of you might ask, why do you want to do this? Well I'm working on a pet project of an ORM so I get values from the DB and then create the object and load the values. I thought it would be cleaner to allow the object just create itself with the values I give it. As far as I can tell I have two options: 1) Use reflection(which I'm trying to avoid) to grab the PropertyInfo[] array and then use that to load the values. 2) require T to support an interface like so: public interface ILoadValues { void LoadValues(IDictionary values); } and then do this public MyClass<T> where T:new(),ILoadValues { T CreateObject(IDictionary<string,object> values) { T obj = new T(); obj.LoadValues(values); return obj; } } The problem I have with the interface I guess is philosophical, I don't really want to expose a public method for people to load the values. Using the constructor the idea was that if I had an object like this namespace DataSource.Data { public class User { protected internal User(IDictionary<string,object> values) { //Initialize } } } As long as the MyClass<T> was in the same assembly the constructor would be available. I personally think that the Type constraint in my opinion should ask (Do I have access to this constructor? I do, great!) Anyways any input is welcome.

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  • markdown to HTML with customised WMD editor

    - by spirytus
    For my application I customized slightly the way WMD behaves so when user enters empty lines, these are reflected in HTML output as <br />'s. Now I came to a point when I should store it somewhere at backend and so after going thru SO posts for a while I'm not sure what is the best way to do it. I have few options and if you could point out which their pros/cons that would be much appreciated. send to server and store as markdown rather than HTML. To me obvious advantage would be keeping exactly same formatting as user originally entered. But then how can I convert it back to HTML for display to a client? It seems very troublesome to convert it on client side as even if it would be possible what would happen if JS would be disabled? If I wanted to do it on the server, then standard server side implementations of markup to HTML might be resource expensive. Would that be an issue in your opinion? Even if it wouldn't be the case then as I mentioned my WMD implementation is customised and those server side solutions wouldn't probably do the right conversion to markdown anyway and there always would be a risk that something would convert wrong. Send to server as converted HTML. Same as above.. conversion on client side would be difficult, server side same with possibility of getting it wrong. send original markdown and converted HTML and store both. No performance issues related to converting markdown to HTML on client side, nor on server side. Users would have always same markdown they originally entered and same HTML they originally saw in preview (possibly sanitized in php though). It would have to take twice that much storage space though and that is my biggest worry. I tend to lean towards 3rd solution as it seems simplest, but there is a worry of doubled storage space needed for this solution. Please bear in mind that my implementation of WMD is slightly modified and also I'm going with PHP/MySql server side implementation. So apart from 3 options I listed above, are there any other possible solutions to my problem? Did I miss anything important that would make one of the options above better then the rest? And what other pros/cons would apply to each solution I listed? Also how is it implemented on SO? I read somwhere that they using option 3, and so if its good enough for SO would be good enough for me :) but not sure if its true anyway, so how is it done? Also please forgive me, but at least for once I got to say that StackOverflow IS THE BEST DAMN RESOURCE ON THE WEB and I truly appreciate all the people trying to help others here! The site and users here are simply amazing!

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  • What is fastest way to convert bool to byte?

    - by Amir Rezaei
    What is fastest way to convert bool to byte? I want this mapping: False=0, True=1 Note: I don't want to use any if statement. Update: I don't want to use conditional statement. I don't want the CPU to halt or guess next statement. I want to optimize this code: private static string ByteArrayToHex(byte[] barray) { char[] c = new char[barray.Length * 2]; byte k; for (int i = 0; i < barray.Length; ++i) { k = ((byte)(barray[i] >> 4)); c[i * 2] = (char)(k > 9 ? k + 0x37 : k + 0x30); k = ((byte)(barray[i] & 0xF)); c[i * 2 + 1] = (char)(k > 9 ? k + 0x37 : k + 0x30); } return new string(c); } Update: The length of the array is very large, it's in terabyte order! Therefore I need to do optimization if possible. I shouldn't need to explain my self. The question is still valid. Update: I'm working on a project and looking at others code. That's why I didn't provide with the function at first place. I didn't want to spend time on explaining for people when they have opinion about the code. I shouldn’y need to provide in my question the background of my work, and a function that is not written by me. I have started to optimize it part by part. If I needed help with the whole function I would asked that in another question. That is why I asked this very simple at the beginning. Unfortunately people couldn’t keep themselves to the question. So please if you want to help answer the question. Update: For dose who want to see the point of this question. This example shows how two if statement are reduced from the code. byte A = k > 9 ; //If it was possible (k>9) == 0 || 1 c[i * 2] = A * (k + 0x30) - (A - 1) * (k + 0x30);

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  • itertools.product eliminating repeated reversed tuples

    - by genclik27
    I asked a question yesterday and thanks to Tim Peters, it is solved. The question is here; itertools.product eliminating repeated elements The new question is further version of this. This time I will generate tuples inside of tuples. Here is an example; lis = [[(1,2), (3,4)], [(5,2), (1,2)], [(2,1), (1,2)]] When I use it in itertools.product function this is what I get, ((1, 2), (5, 2), (2, 1)) ((1, 2), (5, 2), (1, 2)) ((1, 2), (1, 2), (2, 1)) ((1, 2), (1, 2), (1, 2)) ((3, 4), (5, 2), (2, 1)) ((3, 4), (5, 2), (1, 2)) ((3, 4), (1, 2), (2, 1)) ((3, 4), (1, 2), (1, 2)) I want to change it in a way that if a sequence has (a,b) inside of it, then it can not have (b,a). In this example if you look at this sequence ((3, 4), (1, 2), (2, 1)) it has (1,2) and (2,1) inside of it. So, this sequence ((3, 4), (1, 2), (2, 1)) should not be considered in the results. As I said, I asked similar question before, in that case it was not considering duplicate elements. I try to adapt it to my problem. Here is modified code. Changed parts in old version are taken in comments. def reverse_seq(seq): s = [] for i in range(len(seq)): s.append(seq[-i-1]) return tuple(s) def uprod(*seqs): def inner(i): if i == n: yield tuple(result) return for elt in sets[i] - reverse: #seen.add(elt) rvrs = reverse_seq(elt) reverse.add(rvrs) result[i] = elt for t in inner(i+1): yield t #seen.remove(elt) reverse.remove(rvrs) sets = [set(seq) for seq in seqs] n = len(sets) #seen = set() reverse = set() result = [None] * n for t in inner(0): yield t In my opinion this code should work but I am getting error for the input lis = [[(1,2), (3,4)], [(5,2), (1,2)], [(2,1), (1,2)]]. I could not understand where I am wrong. for i in uprod(*lis): print i Output is, ((1, 2), (1, 2), (1, 2)) Traceback (most recent call last): File "D:\Users\SUUSER\workspace tree\sequence_covering _array\denemeler_buraya.py", line 39, in <module> for i in uprod(*lis): File "D:\Users\SUUSER\workspace tree\sequence_covering _array\denemeler_buraya.py", line 32, in uprod for t in inner(0): File "D:\Users\SUUSER\workspace tree\sequence_covering _array\denemeler_buraya.py", line 22, in inner for t in inner(i+1): File "D:\Users\SUUSER\workspace tree\sequence_covering _array\denemeler_buraya.py", line 25, in inner reverse.remove(rvrs) KeyError: (2, 1) Thanks,

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  • Which style is preferable when writing this boolean expression?

    - by Jeppe Stig Nielsen
    I know this question is to some degree a matter of taste. I admit this is not something I don't understand, it's just something I want to hear others' opinion about. I need to write a method that takes two arguments, a boolean and a string. The boolean is in a sense (which will be obvious shortly) redundant, but it is part of a specification that the method must take in both arguments, and must raise an exception with a specific message text if the boolean has the "wrong" value. The bool must be true if and only if the string is not null or empty. So here are some different styles to write (hopefully!) the same thing. Which one do you find is the most readable, and compliant with good coding practice? // option A: Use two if, repeat throw statement and duplication of message string public void SomeMethod(bool useName, string name) { if (useName && string.IsNullOrEmpty(name)) throw new SomeException("..."); if (!useName && !string.IsNullOrEmpty(name)) throw new SomeException("..."); // rest of method } // option B: Long expression but using only && and || public void SomeMethod(bool useName, string name) { if (useName && string.IsNullOrEmpty(name) || !useName && !string.IsNullOrEmpty(name)) throw new SomeException("..."); // rest of method } // option C: With == operator between booleans public void SomeMethod(bool useName, string name) { if (useName == string.IsNullOrEmpty(name)) throw new SomeException("..."); // rest of method } // option D1: With XOR operator public void SomeMethod(bool useName, string name) { if (!(useName ^ string.IsNullOrEmpty(name))) throw new SomeException("..."); // rest of method } // option D2: With XOR operator public void SomeMethod(bool useName, string name) { if (useName ^ !string.IsNullOrEmpty(name)) throw new SomeException("..."); // rest of method } Of course you're welcome to suggest other possibilities too. Message text "..." would be something like "If 'useName' is true a name must be given, and if 'useName' is false no name is allowed".

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  • Thinking about introducing PHP/MySQL into a .NET/SQL Server environment. Thoughts?

    - by abszero
    I posted this over at reddit but it didn't gain any momentum. So here is what is going on: our company was recently purchased by another web shop and I was promoted to head of development here in our office. Our office is completely .NET/SQL Server and the company who purchased us is a *nix/PHP/MySQL shop. Now several of our large clients who are on the .NET platform are up for complete rewrites (the sites are from '04 and are running on the 1.x framework.) While reviewing the proposal for one client with my superior I came across a pretty extensive module which would require several hundred man hours to complete and voiced some concern about it in relation to the quote. One of the guys from the PHP group happen to hear this and told me of a module that they (PHP Group) use in Drupal that does exactly what the proposal in front of me was describing and it only took, at most, 8 hours to completely setup / configure. My superior suggested that I take a look at Drupal and the module in question over the weekend but stressed that we should only go that route if it really made sense. So this weekend I spun up a CentOS instance in VirtualBox and started playing around with Drupal. I am still fleshing it out so don't have a solid opinion on it just yet. Anyway I have some questions / fears that I was hoping progit could help me out in! Has anyone had experience doing this and, if so, how did it turn out? I am completely ignorant to what IDE's (if any) are available to for PHP. The last time I worked with PHP it was in Notepad and that was less than intuitive. So is there are more intuitive IDE out there for PHP dev? I don't want to scare my .NET guys. Since the merger all of our new business clients that have had relatively small websites have gone on Drupal with the larger sites going on .NET. My concern is that if they see a large site go onto Drupal that they might start getting anxious and start handing out their resumes. For the foreseeable future there are no plans to liquidate the .NET platform and really we can't just from a support standpoint. What would be the best way to approach this? Any other helpful info? Thanks!

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  • Creating multiple heads in remote repository

    - by Jab
    We are looking to move our team (~10 developers) from SVN to mercurial. We are trying to figure out how to manage our workflow. In particular, we are trying to see if creating remote heads is the right solution. We currently have a very large repository with multiple, related projects. They share a lot of code, but pieces of the project are deployed by different teams (3 teams) independent of other portions of the code-base. So each team is working on concurrent large features. The way we currently handles this in SVN are branches. Team1 has a branch for Feature1, same deal for the other teams. When Team1 finishes their change, it gets merged into the trunk and deployed out. The other teams follow suite when their project is complete, merging of course. So my initial thought are using Named Branches for these situations. Team1 makes a Feature1 branch off of the default branch in Hg. Now, here is the question. Should the team PUSH that branch, in it's current/half-state to the repository. This will create a second head in the core repo. My initial reaction was "NO!" as it seems like a bad idea. Handling multiple heads on our repository just sounds awful, but there are some advantages... First, the teams want to setup Continuous Integration to build this branch during their development cycle(months long). This will only work if the CI can pull this branch from the repo. This is something we do now with SVN, copy a CI build and change the branch. Easy. Second, it makes it easier for any team member to jump onto the branch and start working. Without pushing to the core repo, they would have to receive a push from a developer on that team with the changeset information. It is also possible to lose local commits to hardware failure. The chances increase a lot if it's a branch by a single developer who has followed the "don't push until finished" approach. And lastly is just for ease of use. The developers can easily just commit and push on their branch at any time without consequence(as they do today, in their SVN branches). Is there a better way to handle this scenario that I may be missing? I just want a veteran's opinion before moving forward with the strategy. For bug fixes we like the general workflow of mecurial, anonymous branches that only consist of 1-2 commits. The simplicity is great for those cases. By the way, I've read this , great article which seems to favor Named branches.

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  • Setting default values for inherited property without using accessor in Objective-C?

    - by Ben Stock
    I always see people debating whether or not to use a property's setter in the -init method. I don't know enough about the Objective-C language yet to have an opinion one way or the other. With that said, lately I've been sticking to ivars exclusively. It seems cleaner in a way. I don't know. I digress. Anyway, here's my problem … Say we have a class called Dude with an interface that looks like this: @interface Dude : NSObject { @private NSUInteger _numberOfGirlfriends; } @property (nonatomic, assign) NSUInteger numberOfGirlfriends; @end And an implementation that looks like this: @implementation Dude - (instancetype)init { self = [super init]; if (self) { _numberOfGirlfriends = 0; } } @end Now let's say I want to extend Dude. My subclass will be called Playa. And since a playa should have mad girlfriends, when Playa gets initialized, I don't want him to start with 0; I want him to have 10. Here's Playa.m: @implementation Playa - (instancetype)init { self = [super init]; if (self) { // Attempting to set the ivar directly will result in the compiler saying, // "Instance variable `_numberOfGirlfriends` is private." // _numberOfGirlfriends = 10; <- Can't do this. // Thus, the only way to set it is with the mutator: self.numberOfGirlfriends = 10; // Or: [self setNumberOfGirlfriends:10]; } } @end So what's a Objective-C newcomer to do? Well, I mean, there's only one thing I can do, and that's set the property. Unless there's something I'm missing. Any ideas, suggestions, tips, or tricks? Sidenote: The other thing that bugs me about setting the ivar directly — and what a lot of ivar-proponents say is a "plus" — is that there are no KVC notifications. A lot of times, I want the KVC magic to happen. 50% of my setters end in [self setNeedsDisplay:YES], so without the notification, my UI doesn't update unless I remember to manually add -setNeedsDisplay. That's a bad example, but the point stands. I utilize KVC all over the place, so without notifications, things can act wonky. Anyway, any info is much appreciated. Thanks!

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  • C++ Class Access Specifier Verbosity

    - by PolyTex
    A "traditional" C++ class (just some random declarations) might resemble the following: class Foo { public: Foo(); explicit Foo(const std::string&); ~Foo(); enum FooState { Idle, Busy, Unknown }; FooState GetState() const; bool GetBar() const; void SetBaz(int); private: struct FooPartialImpl; void HelperFunction1(); void HelperFunction2(); void HelperFunction3(); FooPartialImpl* m_impl; // smart ptr FooState m_state; bool m_bar; int m_baz; }; I always found this type of access level specification ugly and difficult to follow if the original programmer didn't organize his "access regions" neatly. Taking a look at the same snippet in a Java/C# style, we get: class Foo { public: Foo(); public: explicit Foo(const std::string&); public: ~Foo(); public: enum FooState { Idle, Busy, Unknown }; public: FooState GetState() const; public: bool GetBar() const; public: void SetBaz(int); private: struct FooPartialImpl; private: void HelperFunction1(); private: void HelperFunction2(); private: void HelperFunction3(); private: FooPartialImpl* m_impl; // smart ptr private: FooState m_state; private: bool m_bar; private: int m_baz; }; In my opinion, this is much easier to read in a header because the access specifier is right next to the target, and not a bunch of lines away. I found this especially true when working with header-only template code that wasn't separated into the usual "*.hpp/*.inl" pair. In that scenario, the size of the function implementations overpowered this small but important information. My question is simple and stems from the fact that I've never seen anyone else actively do this in their C++ code. Assuming that I don't have a "Class View" capable IDE, are there any obvious drawbacks to using this level of verbosity? Any other style recommendations are welcome!

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  • Model Binding with Parent/Child Relationship

    - by user296297
    I'm sure this has been answered before, but I've spent the last three hours looking for an acceptable solution and have been unable to find anything, so I apologize for what I'm sure is a repeat. I have two domain objects, Player and Position. Player's have a Position. My domain objects are POCOs tied to my database with NHibernate. I have an Add action that takes a Player, so I'm using the built in model binding. On my view I have a drop down list that lets a user select the Position for the Player. The value of the drop down list is the Id of the position. Everything gets populated correctly except that my Position object fails validation (ModelState.IsValid) because at the point of model binding it only has an Id and none of it's other required attributes. What is the preferred solution for solving this with ASP.NET MVC 2? Solutions I've tried... Fetch the Position from the database based on the Id before ModelState.IsValid is called in the Add action of my controller. I can't get the model to run the validation again, so ModelState.IsValid always returns false. Create a custom ModelBinder that inherits from the default binder and fetch the Position from the database after the base binder is called. The ModelBinder seems to be doing the validation so if I use anything from the default binder I'm hosed. Which means I have to completely roll my own binder and grab every value from the form...this seems really wrong and inefficient for such a common use-case. Solutions I think might work, I just can't figure out how to do... Turn off the validation for the Position class when used in Player. Write a custom ModelBinder leverages the default binder for most of the property binding, but lets me get the Position from the database BEFORE the default binder runs validation. So, how do the rest of you solve this? Thanks, Dan P.S. In my opinion having a PositionId on Player just for this case is not a good solution. There has to be solvable in a more elegant fashion.

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  • Hidden divs for "lazy javascript" loading? Possible security/other issues?

    - by xyld
    I'm curious about people's opinion's and thoughts about this situation. The reason I'd like to lazy load javascript is because of performance. Loading javascript at the end of the body reduces the browser blocking and ends up with much faster page loads. But there is some automation I'm using to generate the html (django specifically). This automation has the convenience of allowing forms to be built with "Widgets" that output content it needs to render the entire widget (extra javascript, css, ...). The problem is that the widget wants to output javascript immediately into the middle of the document, but I want to ensure all javascript loads at the end of the body. When the following widget is added to a form, you can see it renders some <script>...</script> tags: class AutoCompleteTagInput(forms.TextInput): class Media: css = { 'all': ('css/jquery.autocomplete.css', ) } js = ( 'js/jquery.bgiframe.js', 'js/jquery.ajaxQueue.js', 'js/jquery.autocomplete.js', ) def render(self, name, value, attrs=None): output = super(AutoCompleteTagInput, self).render(name, value, attrs) page_tags = Tag.objects.usage_for_model(DataSet) tag_list = simplejson.dumps([tag.name for tag in page_tags], ensure_ascii=False) return mark_safe(u'''<script type="text/javascript"> jQuery("#id_%s").autocomplete(%s, { width: 150, max: 10, highlight: false, scroll: true, scrollHeight: 100, matchContains: true, autoFill: true }); </script>''' % (name, tag_list,)) + output What I'm proposing is that if someone uses a <div class=".lazy-js">...</div> with some css (.lazy-js { display: none; }) and some javascript (jQuery('.lazy-js').each(function(index) { eval(jQuery(this).text()); }), you can effectively force all javascript to load at the end of page load: class AutoCompleteTagInput(forms.TextInput): class Media: css = { 'all': ('css/jquery.autocomplete.css', ) } js = ( 'js/jquery.bgiframe.js', 'js/jquery.ajaxQueue.js', 'js/jquery.autocomplete.js', ) def render(self, name, value, attrs=None): output = super(AutoCompleteTagInput, self).render(name, value, attrs) page_tags = Tag.objects.usage_for_model(DataSet) tag_list = simplejson.dumps([tag.name for tag in page_tags], ensure_ascii=False) return mark_safe(u'''<div class="lazy-js"> jQuery("#id_%s").autocomplete(%s, { width: 150, max: 10, highlight: false, scroll: true, scrollHeight: 100, matchContains: true, autoFill: true }); </div>''' % (name, tag_list,)) + output Nevermind all the details of my specific implementation (the specific media involved), I'm looking for a consensus on whether the method of using lazy-loaded javascript through hidden a hidden tags can pose issues whether security or other related? One of the most convenient parts about this is that it follows the DRY principle rather well IMO because you don't need to hack up a specific lazy-load for each instance in the page. It just "works". UPDATE: I'm not sure if django has the ability to queue things (via fancy template inheritance or something?) to be output just before the end of the </body>?

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  • Where do you put your unit test?

    - by soulmerge
    I have found several conventions to housekeeping unit tests in a project and I'm not sure which approach would be suitable for our next PHP project. I am trying to find the best convention to encourage easy development and accessibility of the tests when reviewing the source code. I would be very interested in your experience/opinion regarding each: One folder for productive code, another for unit tests: This separates unit tests from the logic files of the project. This separation of concerns is as much a nuisance as it is an advantage: Someone looking into the source code of the project will - so I suppose - either browse the implementation or the unit tests (or more commonly: the implementation only). The advantage of unit tests being another viewpoint to your classes is lost - those two viewpoints are just too far apart IMO. Annotated test methods: Any modern unit testing framework I know allows developers to create dedicated test methods, annotating them (@test) and embedding them in the project code. The big drawback I see here is that the project files get cluttered. Even if these methods are separated using a comment header (like UNIT TESTS below this line) it just bloats the class unnecessarily. Test files within the same folders as the implementation files: Our file naming convention dictates that PHP files containing classes (one class per file) should end with .class.php. I could imagine that putting unit tests regarding a class file into another one ending on .test.php would render the tests much more present to other developers without tainting the class. Although it bloats the project folders, instead of the implementation files, this is my favorite so far, but I have my doubts: I would think others have come up with this already, and discarded this option for some reason (i.e. I have not seen a java project with the files Foo.java and FooTest.java within the same folder.) Maybe it's because java developers make heavier use of IDEs that allow them easier access to the tests, whereas in PHP no big editors have emerged (like eclipse for java) - many devs I know use vim/emacs or similar editors with little support for PHP development per se. What is your experience with any of these unit test placements? Do you have another convention I haven't listed here? Or am I just overrating unit test accessibility to reviewers?

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  • What the best approach to iterate and "store" files over a directory in C (Linux) ?

    - by Andrei Ciobanu
    I have written a function that checks if to files are duplicates or not. This function signature is: int check_dup_memmap(char *f1_name, char *f2_name) It returns: (-1) - If something went wrong; (0) - If the two files are similar; (+1) - If the two files are different; The next step is to write a function that iterates through all the files in a certain directory,apply the previous function, and gives a report on every existing duplicates. Initially I've thought to write a function that generates a file with all the filenames in a certain directory and then, read that file again and gain and compare every two files. Here is that version of the function, that gets all the filenames in a certain directory. void *build_dir_tree(char *dirname, FILE *f) { DIR *cdir = NULL; struct dirent *ent = NULL; struct stat buf; if(f == NULL){ fprintf(stderr, "NULL file submitted. [build_dir_tree].\n"); exit(-1); } if(dirname == NULL){ fprintf(stderr, "NULL dirname submitted. [build_dir_tree].\n"); exit(-1); } if((cdir = opendir(dirname)) == NULL){ char emsg[MFILE_LEN]; sprintf(emsg, "Cannot open dir: %s [build_dir_tree]\t",dirname); perror(emsg); } chdir(dirname); while ((ent = readdir(cdir)) != NULL) { lstat(ent->d_name, &buf); if (S_ISDIR(buf.st_mode)) { if (strcmp(".", ent->d_name) == 0 || strcmp("..", ent->d_name) == 0) { continue; } build_dir_tree(ent->d_name, f); } else{ fprintf(f, "/%s/%s\n",util_get_cwd(),ent->d_name); } } chdir(".."); closedir(cdir); } Still I consider this approach a little inefficient, as I have to parse the file again and again. In your opinion what are other approaches should I follow: Write a datastructure and hold the files instead of writing them in the file ? I think for a directory with a lot of files, the memory will become very fragmented. Hold all the filenames in auto-expanding array, so that I can easy access every file by their index, because they will in a contiguous memory location. Map this file in memory using mmap() ? But mmap may fail, as the file gets to big. Any opinions on this. I want to choose the most efficient path, and access as few resources as possible. This is the requirement of the program... EDIT: Is there a way to get the numbers of files in a certain directory, without iterating through it ?

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  • Security review of an authenticated Diffie Hellman variant

    - by mtraut
    EDIT I'm still hoping for some advice on this, i tried to clarify my intentions... When i came upon device pairing in my mobile communication framework i studied a lot of papers on this topic and and also got some input from previous questions here. But, i didn't find a ready to implement protocol solution - so i invented a derivate and as i'm no crypto geek i'm not sure about the security caveats of the final solution: The main questions are Is SHA256 sufficient as a commit function? Is the addition of the shared secret as an authentication info in the commit string safe? What is the overall security of the 1024 bit group DH I assume at most 2^-24 bit probability of succesful MITM attack (because of 24 bit challenge). Is this plausible? What may be the most promising attack (besides ripping the device out off my numb, cold hands) This is the algorithm sketch For first time pairing, a solution proposed in "Key agreement in peer-to-peer wireless networks" (DH-SC) is implemented. I based it on a commitment derived from: A fix "UUID" for the communicating entity/role (128 bit, sent at protocol start, before commitment) The public DH key (192 bit private key, based on the 1024 bit Oakley group) A 24 bit random challenge Commit is computed using SHA256 c = sha256( UUID || DH pub || Chall) Both parties exchange this commitment, open and transfer the plain content of the above values. The 24 bit random is displayed to the user for manual authentication DH session key (128 bytes, see above) is computed When the user opts for persistent pairing, the session key is stored with the remote UUID as a shared secret Next time devices connect, commit is computed by additionally hashing the previous DH session key before the random challenge. For sure it is not transfered when opening. c = sha256( UUID || DH pub || DH sess || Chall) Now the user is not bothered authenticating when the local party can derive the same commitment using his own, stored previous DH session key. After succesful connection the new DH session key becomes the new shared secret. As this does not exactly fit the protocols i found so far (and as such their security proofs), i'd be very interested to get an opinion from some more crypto enabled guys here. BTW. i did read about the "EKE" protocol, but i'm not sure what the extra security level is.

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  • Webservice for uploading data: security considerations

    - by Philip Daubmeier
    Hi everyone! Im not sure about what authentification method I should use for my webservice. I've searched on SO, and found nothing that helped me. Preliminary Im building an application that uploads data from a local database to a server (running my webservice), where all records are merged and stored in a central database. I am currently binary serializing a DataTable, that holds a small fragment of the local database, where all uninteresting stuff is already filtered out. The byte[] (serialized DataTable), together with the userid and a hash of the users password is then uploaded to the webservice via SOAP. The application together with the webservice already work exactly like intended. The Problem The issue I am thinking about is now: What is if someone just sniffs the network traffic, 'steals' the users id and password hash to send his own SOAP message with modified data that corrupts my database? Options The approaches to solving that problem, I already thought of, are: Using ssl + certificates for establishing the connection: I dont really want to use ssl, I would prefer a simpler solution. After all, every information that is transfered to the webservice can be seen on the website later on. What I want to say is: there is no secret/financial/business-critical information, that has to be hidden. I think ssl would be sort of an overkill for that task. Encrypting the byte[]: I think that would be a performance killer, considering that the goal of the excercise was simply to authenticate the user. Hashing the users password together with the data: I kind of like the idea: Creating a checksum from the data, concatenating that checksum with the password-hash and hashing this whole thing again. That would assure the data was sent from this specific user, and the data wasnt modified. The actual question So, what do you think is the best approach in terms of meeting the following requirements? Rather simple solution (As it doesnt have to be super secure; no secret/business-critical information transfered) Easily implementable retrospectively (Dont want to write it all again :) ) Doesnt impact to much on performance What do you think of my prefered solution, the last one in the list above? Is there any alternative solution I didnt mention, that would fit better? You dont have to answer every question in detail. Just push me in the right direction. I very much appreciate every well-grounded opinion. Thanks in advance!

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  • Which work process in my company should I Improve first?

    - by shoren
    I've just started to work in a new place, and I see several things they do that I find really terrible, and I want to know if they are indeed so wrong, or I am just too strict. Please let me know if my criticism is in place, and your opinion on which problem is the worst and should be fixed first. The developement is all in Java. 1) Not using svnignore. This means svn stat can't be used, and developers forget to add files and break the build. 2) Generated files go to same folders as committed files. Can't use simple maven clean, have to find them one by one. Maven clean doesn't remove all of them. 3) Not fixing IDE analyze warnings. Analyze code returns about 5,000 warning, of many different kinds. 4) Not following conventions: spring beans names sometimes start with uppercase and sometimes not, ant properties sometimes with underline and sometimes with dots delimiter, etc. 5) Incremental build takes 6 minutes, even when nothing is changed. 6) Developers only use remote debug, and don't know how to run the Tomcat server internally from the IDE. 7) Developers always restart the server after every compilation, instead of dynamically reloading the class and saving the server's state. It takes them at least 10 minutes to start checking any change in the code. 8) Developers only compile from command line. When there are compilation errors, they manually open the file and go the the problematic line. 9) A complete mess in project dependencies. Over 200 open sources are depended on, and no one knows what is indeed needed and why. They do know that not all dependencies are necessary. 10) Mixing Maven and Ant in a way that disables the benefits of both. In one case, even dependency checks are not done by Maven. 11) Not using generics properly. 12) Developers don't use Subversion integration with IDE (Eclipse, Intellij Idea). What do you think? Where should I start? Is any of the things I mentioned not really a problem?

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  • organizing unit test

    - by soulmerge
    I have found several conventions to housekeeping unit tests in a project and I'm not sure which approach would be suitable for our next PHP project. I am trying to find the best convention to encourage easy development and accessibility of the tests when reviewing the source code. I would be very interested in your experience/opinion regarding each: One folder for productive code, another for unit tests: This separates unit tests from the logic files of the project. This separation of concerns is as much a nuisance as it is an advantage: Someone looking into the source code of the project will - so I suppose - either browse the implementation or the unit tests (or more commonly: the implementation only). The advantage of unit tests being another viewpoint to your classes is lost - those two viewpoints are just too far apart IMO. Annotated test methods: Any modern unit testing framework I know allows developers to create dedicated test methods, annotating them (@test) and embedding them in the project code. The big drawback I see here is that the project files get cluttered. Even if these methods are separated using a comment header (like UNIT TESTS below this line) it just bloats the class unnecessarily. Test files within the same folders as the implementation files: Our file naming convention dictates that PHP files containing classes (one class per file) should end with .class.php. I could imagine that putting unit tests regarding a class file into another one ending on .test.php would render the tests much more present to other developers without tainting the class. Although it bloats the project folders, instead of the implementation files, this is my favorite so far, but I have my doubts: I would think others have come up with this already, and discarded this option for some reason (i.e. I have not seen a java project with the files Foo.java and FooTest.java within the same folder.) Maybe it's because java developers make heavier use of IDEs that allow them easier access to the tests, whereas in PHP no big editors have emerged (like eclipse for java) - many devs I know use vim/emacs or similar editors with little support for PHP development per se. What is your experience with any of these unit test placements? Do you have another convention I haven't listed here? Or am I just overrating unit test accessibility to reviewing developers?

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  • Dynamic swappable Data Access Layer

    - by Andy
    I'm writing a data driven WPF client. The client will typically pull data from a WCF service, which queries a SQL db, but I'd like the option to pull the data directly from SQL or other arbitrary data sources. I've come up with this design and would like to hear your opinion on whether it is the best design. First, we have some data object we'd like to extract from SQL. // The Data Object with a single property public class Customer { private string m_Name = string.Empty; public string Name { get { return m_Name; } set { m_Name = value;} } } Then I plan on using an interface which all data access layers should implement. Suppose one could also use an abstract class. Thoughts? // The interface with a single method interface ICustomerFacade { List<Customer> GetAll(); } One can create a SQL implementation. // Sql Implementation public class SqlCustomrFacade : ICustomerFacade { public List<Customer> GetAll() { // Query SQL db and return something useful // ... return new List<Customer>(); } } We can also create a WCF implementation. The problem with WCF is is that it doesn't use the same data object. It creates its own local version, so we would have to copy the details over somehow. I suppose one could use reflection to copy the values of similar fields across. Thoughts? // Wcf Implementation public class WcfCustomrFacade : ICustomerFacade { public List<Customer> GetAll() { // Get date from the Wcf Service (not defined here) List<WcfService.Customer> wcfCustomers = wcfService.GetAllCustomers(); // The list we're going to return List<Customer> customers = new List<Customer>(); // This is horrible foreach(WcfService.Customer wcfCustomer in wcfCustomers) { Customer customer = new Customer(); customer.Name = wcfCustomer.Name; customers.Add(customer); } return customers; } } I also plan on using a factory to decide which facade to use. // Factory pattern public class FacadeFactory() { public static ICustomerFacade CreateCustomerFacade() { // Determine the facade to use if (ConfigurationManager.AppSettings["DAL"] == "Sql") return new SqlCustomrFacade(); else return new WcfCustomrFacade(); } } This is how the DAL would typically be used. // Test application public class MyApp { public static void Main() { ICustomerFacade cf = FacadeFactory.CreateCustomerFacade(); cf.GetAll(); } } I appreciate your thoughts and time.

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  • Advice needed: stay with Java team or move to C++ team?

    - by user68759
    Some background - I have been programming in Java as a professional for the last few years. This is mainly using Java SE. I have also touched bits and pieces of other various Java technologies and have some basic knowledge about them. I consider my self as an intermediate Java programmer. I like Java very much. I think it is only going to get bigger. Recently, my manager asked my opinion on whether I would like to be transferred to another team within the company that is developing a product in C++. This is mainly because my current Java team simply didn't make enough money due to poor sales and the economic downturn. Now, I have never had any experience with C++ nor have I ever coded a single line of code in C++. I have always wanted to learn it and now is my chance. But I really want to make sure I get benefit out of it in the future, in the sense that I will have the skills that will still be on-demand in the future. So, what do you experts think? Is C++ still the language to learn these days to secure yourself for the future? What will I learn more in C++ but not in Java? And are they worthy to learn considering the current and possible future demands in IT industry? (Apart from the obvious more control over memory management and something along that line.) What is a good excuse to refuse the offer in order to stay with the Java team? I don't want to blatantly refuse it because you can never predict the future and I could possibly come back to my manager in the future and ask him to transfer me to the C++ team. How do I say it nicely that I am taking the offer but I would like to still be involved with Java one way or another, such as when there is a new Java project I would like to be considered. I have to admit that I am kind of 50-50 at the moment. I want to learn C++ for the sake of improving my skills and also helping my company to reduce the fund required for the Java team. But it is also hard for me to leave Java because I know Java is going to get bigger, so I am afraid of getting behind when I start concentrating on C++. I could, of course, decide to just join the C++ team, and then spend my free time reading about Java to keep in touch with it, but I thought I would ask anyway in case some people can point out the strong points of either over the other given the current and possibly future circumstances.

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  • Is there a way to apply a CSS class from within a style?

    - by zashu
    I'm trying to be more modular in my CSS style sheets and was wondering if there is some feature like an include or apply that allows the author to apply a set of styles dynamically. Since I am having a hard time wording the question, perhaps an example will make more sense. Let's say, for example, I have the following CSS: .red {color:#e00b0b} #footer a {font-size:0.8em} h2 {font-size:1.4em; font-weight:bold;} In my page, let's say that I want both the footer links and h2 elements to use the special red color (there may be other locations I would like to use it as well). Ideally, I would like to do something like the following: .red {color:#e00b0b} #footer a {font-size:0.8em; apply-class:".red";} h2 {font-size:1.4em; font-weight:bold; apply-class:".red";} To me, this feels "modular" in a way because I can make modifications to the .red class without having to worry so much about where it is used, and other locations can use the styles in that class without worrying about, specifically, what they are. I understand that I have the following options and have included why, in my fairly inexperienced opinion, they are less-than-perfect: Add the color property to every element I want to be that color. Not ideal because, if I change the color, I have to update every rule to match the new color. Add the red class to every element I want to be red. Not ideal because it means that my HTML is dictating presentation. Create an additional rule that selects every element I want to be red and apply the color property to that. Not ideal because it is harder to find all of the rules that style a specific element, making maintenance more of a challenge Maybe I'm just being an ass and the following options are the only options and I should stick with them. I'm wondering, however, if the "ideal" (well, my ideal) method exists and, if so, what is the proper syntax? If it doesn't exist, option 3 above seems like my best bet. However, I would like to get confirmation.

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  • How can i merge arrays like a gearwheel

    - by JuKe
    i have 3 arrays like this $a = array( 0 => 'a1', 1 => 'a2', 2 => 'a3' ); $b = array( 0 => 'b1', 1 => 'b2', 2 => 'b3' ); $c = array( 0 => 'c1', 1 => 'c2', 2 => 'c3' ); and i like to have somethink like this: $r = array( 0 => 'a1', 1 => 'b1', 2 => 'c1', 3 => 'a2', 4 => 'b2', 5 => 'c2', 6 => 'a3', .... ... ); how can i do this, with the opinion of more than 3 arrays EDIT: i have tried this: $a = array( 0 => 'a1', 1 => 'a2', 2 => 'a3', 3 => 'a4' ); $b = array( 0 => 'b1', 1 => 'b2', 2 => 'b3' ); $c = array( 0 => 'c1', 1 => 'c2', 2 => 'c3', 3 => 'c3', 4 => 'c3', 5 => 'c3' ); $l['a'] = count($a); $l['b'] = count($b); $l['c'] = count($c); arsort($l); $largest = key($l); $result = array(); foreach ($$largest as $key => $value) { $result[] = $a[$key]; if(array_key_exists($key, $b)) $result[] = $b[$key]; if(array_key_exists($key, $c)) $result[] = $c[$key]; } print_r($result); this work.. but the code isn't nice. so anyone has a better solution?

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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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  • How Hacker Can Access VPS CentOS 6 content?

    - by user2118559
    Just want to understand. Please, correct mistakes and write advices Hacker can access to VPS: 1. Through (using) console terminal, for example, using PuTTY. To access, hacker need to know port number, username and password. Port number hacker can know scanning open ports and try to login. The only way to login as I understand need to know username and password. To block (make more difficult) port scanning, need to use iptables configure /etc/sysconfig/iptables. I followed this https://www.digitalocean.com/community/articles/how-to-setup-a-basic-ip-tables-configuration-on-centos-6 tutorial and got *nat :PREROUTING ACCEPT [87:4524] :POSTROUTING ACCEPT [77:4713] :OUTPUT ACCEPT [77:4713] COMMIT *mangle :PREROUTING ACCEPT [2358:200388] :INPUT ACCEPT [2358:200388] :FORWARD ACCEPT [0:0] :OUTPUT ACCEPT [2638:477779] :POSTROUTING ACCEPT [2638:477779] COMMIT *filter :INPUT DROP [1:40] :FORWARD ACCEPT [0:0] :OUTPUT ACCEPT [339:56132] -A INPUT -m state --state RELATED,ESTABLISHED -j ACCEPT -A INPUT -p tcp -m tcp --tcp-flags FIN,SYN,RST,PSH,ACK,URG NONE -j DROP -A INPUT -p tcp -m tcp ! --tcp-flags FIN,SYN,RST,ACK SYN -m state --state NEW -j DROP -A INPUT -p tcp -m tcp --tcp-flags FIN,SYN,RST,PSH,ACK,URG FIN,SYN,RST,PSH,ACK,URG -j DROP -A INPUT -i lo -j ACCEPT -A INPUT -p tcp -m tcp --dport 80 -j ACCEPT -A INPUT -p tcp -m tcp --dport 110 -j ACCEPT -A INPUT -p tcp -m tcp --dport 22 -j ACCEPT -A INPUT -s 11.111.11.111/32 -p tcp -m tcp --dport 22 -j ACCEPT -A INPUT -p tcp -m tcp --dport 21 -j ACCEPT -A INPUT -s 11.111.11.111/32 -p tcp -m tcp --dport 21 -j ACCEPT COMMIT Regarding ports that need to be opened. If does not use ssl, then seems must leave open port 80 for website. Then for ssh (default 22) and for ftp (default 21). And set ip address, from which can connect. So if hacker uses other ip address, he can not access even knowing username and password? Regarding emails not sure. If I send email, using Gmail (Send mail as: (Use Gmail to send from your other email addresses)), then port 25 not necessary. For incoming emails at dynadot.com I use Email Forwarding. Does it mean that emails “does not arrive to VPS” (before arriving to VPS, emails are forwarded, for example to Gmail)? If emails does not arrive to VPS, then seems port 110 also not necessary. If use only ssl, must open port 443 and close port 80. Do not understand regarding port 3306 In PuTTY with /bin/netstat -lnp see Proto Recv-Q Send-Q Local Address Foreign Address State PID/Program name tcp 0 0 0.0.0.0:3306 0.0.0.0:* LISTEN 992/mysqld As understand it is for mysql. But does not remember that I have opened such port (may be when installed mysql, the port is opened automatically?). Mysql is installed on the same server, where all other content. Need to understand regarding port 3306 2. Also hacker may be able access console terminal through VPS hosting provider Control Panel (serial console emergency access). As understand only using console terminal (PuTTY, etc.) can make “global” changes (changes that can not modify with ftp). 3. Hacker can access to my VPS exploiting some hole in my php code and uploading, for example, Trojan. Unfortunately, faced situation that VPS was hacked. As understand it was because I used ZPanel. On VPS ( \etc\zpanel\panel\bin) ) found one php file, that was identified as Trojan by some virus scanners (at virustotal.com). Experimented with the file on local computer (wamp). And appears that hacker can see all content of VPS, rename, delete, upload etc. From my opinion, if in PuTTY use command like chattr +i /etc/php.ini then hacker could not be able to modify php.ini. Is there any other way to get into VPS?

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  • Ubuntu server loses exactly 5 minutes once in a while

    - by Harold Smith
    I noticed that my server, an Ubuntu server 12.04, was losing time. I figured the hardware clock was off or maybe dying due to a faulty CMOS battery. I installed NTP to ensure the drift would be corrected, but to no avail. During a day it would lose 20 minutes or so. To debug, I created a small cron job to check against a remote servers time, which I knew to be correct. The script calculates the difference in seconds between local and remote time. The result was interesting. It seems to be losing exactly 5 minutes several times during the day. Look at this log (difference from remote server noted in seconds): Tue Oct 23 03:30:02 CEST 2012: 284 Tue Oct 23 03:35:02 CEST 2012: 284 Tue Oct 23 03:40:01 CEST 2012: 285 Tue Oct 23 03:45:02 CEST 2012: 285 Tue Oct 23 03:50:02 CEST 2012: 285 Tue Oct 23 03:55:02 CEST 2012: 284 Tue Oct 23 04:00:02 CEST 2012: 284 Tue Oct 23 04:05:01 CEST 2012: 285 Tue Oct 23 04:10:01 CEST 2012: 285 Tue Oct 23 04:15:02 CEST 2012: 585 Tue Oct 23 04:20:02 CEST 2012: 584 Tue Oct 23 04:25:02 CEST 2012: 584 Tue Oct 23 04:30:02 CEST 2012: 584 Tue Oct 23 04:35:01 CEST 2012: 585 Tue Oct 23 04:40:01 CEST 2012: 585 Tue Oct 23 04:45:02 CEST 2012: 585 Tue Oct 23 04:50:02 CEST 2012: 584 Tue Oct 23 04:55:02 CEST 2012: 584 Tue Oct 23 05:00:02 CEST 2012: 584 Tue Oct 23 05:05:01 CEST 2012: 585 Tue Oct 23 05:10:01 CEST 2012: 585 Tue Oct 23 05:15:02 CEST 2012: 585 Tue Oct 23 05:20:02 CEST 2012: 584 Tue Oct 23 05:25:02 CEST 2012: 584 Tue Oct 23 05:30:02 CEST 2012: 584 Tue Oct 23 05:35:01 CEST 2012: 585 Tue Oct 23 05:40:01 CEST 2012: 585 Tue Oct 23 05:45:02 CEST 2012: 584 Tue Oct 23 05:50:02 CEST 2012: 584 Tue Oct 23 05:55:02 CEST 2012: 584 Tue Oct 23 06:00:02 CEST 2012: 584 Tue Oct 23 06:05:03 CEST 2012: 584 Tue Oct 23 06:10:02 CEST 2012: 584 Tue Oct 23 06:15:01 CEST 2012: 585 Tue Oct 23 06:20:02 CEST 2012: 584 Tue Oct 23 06:25:02 CEST 2012: 584 Tue Oct 23 06:30:02 CEST 2012: 584 Tue Oct 23 06:35:02 CEST 2012: 584 Tue Oct 23 06:40:02 CEST 2012: 584 Tue Oct 23 06:45:01 CEST 2012: 585 Tue Oct 23 06:50:02 CEST 2012: 584 Tue Oct 23 06:55:01 CEST 2012: 585 Tue Oct 23 07:00:02 CEST 2012: 584 Tue Oct 23 07:05:02 CEST 2012: 584 Tue Oct 23 07:10:02 CEST 2012: 584 Tue Oct 23 07:15:02 CEST 2012: 584 Tue Oct 23 07:20:02 CEST 2012: 584 Tue Oct 23 07:25:02 CEST 2012: 584 Tue Oct 23 07:30:01 CEST 2012: 585 Tue Oct 23 07:35:02 CEST 2012: 584 Tue Oct 23 07:40:02 CEST 2012: 584 Tue Oct 23 07:45:02 CEST 2012: 584 Tue Oct 23 07:50:02 CEST 2012: 584 Tue Oct 23 07:55:02 CEST 2012: 584 Tue Oct 23 08:00:01 CEST 2012: 585 Tue Oct 23 08:05:02 CEST 2012: 584 Tue Oct 23 08:10:02 CEST 2012: 584 Tue Oct 23 08:15:02 CEST 2012: 584 Tue Oct 23 08:20:02 CEST 2012: 584 Tue Oct 23 08:25:02 CEST 2012: 584 Tue Oct 23 08:30:01 CEST 2012: 585 Tue Oct 23 08:35:02 CEST 2012: 584 Tue Oct 23 08:40:02 CEST 2012: 584 Tue Oct 23 08:45:02 CEST 2012: 584 Tue Oct 23 08:50:02 CEST 2012: 584 Tue Oct 23 08:55:02 CEST 2012: 584 Tue Oct 23 09:00:02 CEST 2012: 584 Tue Oct 23 09:05:03 CEST 2012: 584 Tue Oct 23 09:10:02 CEST 2012: 584 Tue Oct 23 09:15:02 CEST 2012: 584 Tue Oct 23 09:20:02 CEST 2012: 584 Tue Oct 23 09:25:02 CEST 2012: 584 Tue Oct 23 09:30:01 CEST 2012: 584 Tue Oct 23 09:35:02 CEST 2012: 584 Tue Oct 23 09:40:02 CEST 2012: 584 Tue Oct 23 09:45:02 CEST 2012: 584 Tue Oct 23 09:50:02 CEST 2012: 584 Tue Oct 23 09:55:02 CEST 2012: 584 Tue Oct 23 10:00:01 CEST 2012: 584 Tue Oct 23 10:05:02 CEST 2012: 584 Tue Oct 23 10:10:07 CEST 2012: 584 Tue Oct 23 10:15:02 CEST 2012: 584 Tue Oct 23 10:20:02 CEST 2012: 884 Tue Oct 23 10:25:02 CEST 2012: 884 Tue Oct 23 10:30:02 CEST 2012: 883 Tue Oct 23 10:35:01 CEST 2012: 884 Tue Oct 23 10:40:02 CEST 2012: 884 Tue Oct 23 10:45:02 CEST 2012: 884 Tue Oct 23 10:50:02 CEST 2012: 884 Tue Oct 23 10:55:02 CEST 2012: 1184 Tue Oct 23 11:00:02 CEST 2012: 1183 Tue Oct 23 11:05:01 CEST 2012: 1184 Tue Oct 23 11:10:02 CEST 2012: 1184 Tue Oct 23 11:15:02 CEST 2012: 1184 Tue Oct 23 11:20:02 CEST 2012: 1184 This does not seem to be faulty CMOS battery in my opinion. But what do you think?

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  • Clarification On Write-Caching Policy, Its Underlying Options And How It Applies To Hard Drives And Solid-State Drives

    - by Boris_yo
    In last week after doing more research on subject matter, I have been wondering about what I have been neglecting all those years to understand write-caching policy, always leaving it on default setting. Write-caching policy improves writing performance and consists of write-back caching and write-cache buffer flushing. This is how I understand all the above, but correct me if I erred somewhere: Write-through cache / Write-through caching itself is not a part of write caching policy per se and it's when data is written to both cache and storage device so if Windows will need that data later again, it is retrieved from cache and not from storage device which means only improved read performance as there is no need for waiting for storage device to read required data again. Since data is still written to storage device, write performance isn't improved and represents no risk of data loss or corruption in case of power failure or system crash while only data in cache gets lost. This option seems to be enabled by default and is recommended for removable devices with no need to use function of "Safely Remove Hardware" on user's part. Write-back caching is similar to above but without writing data to storage device, periodically releasing data from cache and writing to storage device when it is idle. In my opinion this option improves both read and write performance but represents risk if power failure or system crash occurs with the outcome of not only losing data eventually to be written to storage device, but causing file inconsistencies or corrupted file system. Write-back caching cannot be enabled together with write-through caching and it is not recommended to be enabled if no backup power supply is availabe. Write-cache buffer flushing I reckon is similar to write-back caching but enables immediate release and writing of data from cache to storage device right before power outage occurs but I don't know if it applies also to occasional system crash. This option seem to be complementary to write-back cache reducing or potentially eliminating risk of data loss or corruption of file system. I have questions about relevance of last 2 options to today's modern SSDs in order to get best performance and with less wear on SSDs: I know that traditional hard drives come with onboard cache (I wonder what type of cache that is), but do SSDs also come with cache? Assuming they do, is this cache faster than their NAND flash and system RAM and worth taking the risk of utilizing it by enabling write-back cache? I read somewhere that generally storage device's cache is faster than RAM, but I want to be sure. Additionally I read that write-caching should be enabled since current data that is to be written later to NAND flash is kept for a while in cache and provided there is data that gets modified a lot before finally being written, holding of this data and its periodic release reduces its write times to SSD thereby reducing its wearing. Now regarding to write-cache buffer flushing, I heard that SSD controllers are so fast by themselves that enabling this option is not required, because they manage flushing. However, once again, I don't know if SSDs have their own onboard cache and whether or not it is faster than their NAND flash and system RAM because if it is, keeping this option enabled would make sense. Recently I have posted question about issue with my Intel 330 SSD 120GB which was main reason to do deeper research having suspicion of write-caching policy being the culprit of SSD's freezing issue assuming data being released is what causes freezes. Currently I have write-cache enabled and write-cache buffer flushing disabled because I believe SSD controller's management of write-cache flushing and Windows write-cache buffer flushing are conflicting with each other: Since I want to troubleshoot in small steps to finally determine the source of issue, I have decided to start with write-caching policy and the move to drivers, switching to AHCI later on and finally disabling DIPM (device initiated power management) through registry modification thanks to @TomWijsman

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