Search Results

Search found 20883 results on 836 pages for 'wont say'.

Page 441/836 | < Previous Page | 437 438 439 440 441 442 443 444 445 446 447 448  | Next Page >

  • Isn't it better to use a single try catch instead of tons of TryParsing and other error handling sometimes?

    - by Ryan Peschel
    I know people say it's bad to use exceptions for flow control and to only use exceptions for exceptional situations, but sometimes isn't it just cleaner and more elegant to wrap the entire block in a try-catch? For example, let's say I have a dialog window with a TextBox where the user can type input in to be parsed in a key-value sort of manner. This situation is not as contrived as you might think because I've inherited code that has to handle this exact situation (albeit not with farm animals). Consider this wall of code: class Animals { public int catA, catB; public float dogA, dogB; public int mouseA, mouseB, mouseC; public double cow; } class Program { static void Main(string[] args) { string input = "Sets all the farm animals CAT 3 5 DOG 21.3 5.23 MOUSE 1 0 1 COW 12.25"; string[] splitInput = input.Split(' '); string[] animals = { "CAT", "DOG", "MOUSE", "COW", "CHICKEN", "GOOSE", "HEN", "BUNNY" }; Animals animal = new Animals(); for (int i = 0; i < splitInput.Length; i++) { string token = splitInput[i]; if (animals.Contains(token)) { switch (token) { case "CAT": animal.catA = int.Parse(splitInput[i + 1]); animal.catB = int.Parse(splitInput[i + 2]); break; case "DOG": animal.dogA = float.Parse(splitInput[i + 1]); animal.dogB = float.Parse(splitInput[i + 2]); break; case "MOUSE": animal.mouseA = int.Parse(splitInput[i + 1]); animal.mouseB = int.Parse(splitInput[i + 2]); animal.mouseC = int.Parse(splitInput[i + 3]); break; case "COW": animal.cow = double.Parse(splitInput[i + 1]); break; } } } } } In actuality there are a lot more farm animals and more handling than that. A lot of things can go wrong though. The user could enter in the wrong number of parameters. The user can enter the input in an incorrect format. The user could specify numbers too large or too small for the data type to handle. All these different errors could be handled without exceptions through the use of TryParse, checking how many parameters the user tried to use for a specific animal, checking if the parameter is too large or too small for the data type (because TryParse just returns 0), but every one should result in the same thing: A MessageBox appearing telling the user that the inputted data is invalid and to fix it. My boss doesn't want different message boxes for different errors. So instead of doing all that, why not just wrap the block in a try-catch and in the catch statement just display that error message box and let the user try again? Maybe this isn't the best example but think of any other scenario where there would otherwise be tons of error handling that could be substituted for a single try-catch. Is that not the better solution?

    Read the article

  • Appropriate programming design questions.

    - by Edward
    I have a few questions on good programming design. I'm going to first describe the project I'm building so you are better equipped to help me out. I am coding a Remote Assistance Tool similar to TeamViewer, Microsoft Remote Desktop, CrossLoop. It will incorporate concepts like UDP networking (using Lidgren networking library), NAT traversal (since many computers are invisible behind routers nowadays), Mirror Drivers (using DFMirage's Mirror Driver (http://www.demoforge.com/dfmirage.htm) for realtime screen grabbing on the remote computer). That being said, this program has a concept of being a client-server architecture, but I made only one program with both the functionality of client and server. That way, when the user runs my program, they can switch between giving assistance and receiving assistance without having to download a separate client or server module. I have a Windows Form that allows the user to choose between giving assistance and receiving assistance. I have another Windows Form for a file explorer module. I have another Windows Form for a chat module. I have another Windows Form form for a registry editor module. I have another Windows Form for the live control module. So I've got a Form for each module, which raises the first question: 1. Should I process module-specific commands inside the code of the respective Windows Form? Meaning, let's say I get a command with some data that enumerates the remote user's files for a specific directory. Obviously, I would have to update this on the File Explorer Windows Form and add the entries to the ListView. Should I be processing this code inside the Windows Form though? Or should I be handling this in another class (although I have to eventually pass the data to the Form to draw, of course). Or is it like a hybrid in which I process most of the data in another class and pass the final result to the Form to draw? So I've got like 5-6 forms, one for each module. The user starts up my program, enters the remote machine's ID (not IP, ID, because we are registering with an intermediary server to enable NAT traversal), their password, and connects. Now let's suppose the connection is successful. Then the user is presented with a form with all the different modules. So he can open up a File Explorer, or he can mess with the Registry Editor, or he can choose to Chat with his buddy. So now the program is sort of idle, just waiting for the user to do something. If the user opens up Live Control, then the program will be spending most of it's time receiving packets from the remote machine and drawing them to the form to provide a 'live' view. 2. Second design question. A spin off question #1. How would I pass module-specific commands to their respective Windows Forms? What I mean is, I have a class like "NetworkHandler.cs" that checks for messages from the remote machine. NetworkHandler.cs is a static class globally accessible. So let's say I get a command that enumerates the remote user's files for a specific directory. How would I "give" that command to the File Explorer Form. I was thinking of making an OnCommandReceivedEvent inside NetworkHandler, and having each form register to that event. When the NetworkHandler received a command, it would raise the event, all forms would check it to see if it was relevant, and the appropriate form would take action. Is this an appropriate/the best solution available? 3. The networking library I'm using, Lidgren, provides two options for checking networking messages. One can either poll ReadMessage() to return null or a message, or one can use an AutoResetEvent OnMessageReceived (I'm guessing this is like an event). Which one is more appropriate?

    Read the article

  • [C#][Design] Appropriate programming design questions.

    - by Edward
    I have a few questions on good programming design. I'm going to first describe the project I'm building so you are better equipped to help me out. I am coding a Remote Assistance Tool similar to TeamViewer, Microsoft Remote Desktop, CrossLoop. It will incorporate concepts like UDP networking (using Lidgren networking library), NAT traversal (since many computers are invisible behind routers nowadays), Mirror Drivers (using DFMirage's Mirror Driver (http://www.demoforge.com/dfmirage.htm) for realtime screen grabbing on the remote computer). That being said, this program has a concept of being a client-server architecture, but I made only one program with both the functionality of client and server. That way, when the user runs my program, they can switch between giving assistance and receiving assistance without having to download a separate client or server module. I have a Windows Form that allows the user to choose between giving assistance and receiving assistance. I have another Windows Form for a file explorer module. I have another Windows Form for a chat module. I have another Windows Form form for a registry editor module. I have another Windows Form for the live control module. So I've got a Form for each module, which raises the first question: 1. Should I process module-specific commands inside the code of the respective Windows Form? Meaning, let's say I get a command with some data that enumerates the remote user's files for a specific directory. Obviously, I would have to update this on the File Explorer Windows Form and add the entries to the ListView. Should I be processing this code inside the Windows Form though? Or should I be handling this in another class (although I have to eventually pass the data to the Form to draw, of course). Or is it like a hybrid in which I process most of the data in another class and pass the final result to the Form to draw? So I've got like 5-6 forms, one for each module. The user starts up my program, enters the remote machine's ID (not IP, ID, because we are registering with an intermediary server to enable NAT traversal), their password, and connects. Now let's suppose the connection is successful. Then the user is presented with a form with all the different modules. So he can open up a File Explorer, or he can mess with the Registry Editor, or he can choose to Chat with his buddy. So now the program is sort of idle, just waiting for the user to do something. If the user opens up Live Control, then the program will be spending most of it's time receiving packets from the remote machine and drawing them to the form to provide a 'live' view. 2. Second design question. A spin off question #1. How would I pass module-specific commands to their respective Windows Forms? What I mean is, I have a class like "NetworkHandler.cs" that checks for messages from the remote machine. NetworkHandler.cs is a static class globally accessible. So let's say I get a command that enumerates the remote user's files for a specific directory. How would I "give" that command to the File Explorer Form. I was thinking of making an OnCommandReceivedEvent inside NetworkHandler, and having each form register to that event. When the NetworkHandler received a command, it would raise the event, all forms would check it to see if it was relevant, and the appropriate form would take action. Is this an appropriate/the best solution available? 3. The networking library I'm using, Lidgren, provides two options for checking networking messages. One can either poll ReadMessage() to return null or a message, or one can use an AutoResetEvent OnMessageReceived (I'm guessing this is like an event). Which one is more appropriate?

    Read the article

  • Using inheritance and polymorphism to solve a common game problem

    - by Barry Brown
    I have two classes; let's call them Ogre and Wizard. (All fields are public to make the example easier to type in.) public class Ogre { int weight; int height; int axeLength; } public class Wizard { int age; int IQ; int height; } In each class I can create a method called, say, battle() that will determine who will win if an Ogre meets and Ogre or a Wizard meets a Wizard. Here's an example. If an Ogre meets an Ogre, the heavier one wins. But if the weight is the same, the one with the longer axe wins. public Ogre battle(Ogre o) { if (this.height > o.height) return this; else if (this.height < o.height) return o; else if (this.axeLength > o.axeLength) return this; else if (this.axeLength < o.axeLength) return o; else return this; // default case } We can make a similar method for Wizards. But what if a Wizard meets an Ogre? We could of course make a method for that, comparing, say, just the heights. public Wizard battle(Ogre o) { if (this.height > o.height) return this; else if (this.height < o.height) return o; else return this; } And we'd make a similar one for Ogres that meet Wizard. But things get out of hand if we have to add more character types to the program. This is where I get stuck. One obvious solution is to create a Character class with the common traits. Ogre and Wizard inherit from the Character and extend it to include the other traits that define each one. public class Character { int height; public Character battle(Character c) { if (this.height > c.height) return this; else if (this.height < c.height) return c; else return this; } } Is there a better way to organize the classes? I've looked at the strategy pattern and the mediator pattern, but I'm not sure how either of them (if any) could help here. My goal is to reach some kind of common battle method, so that if an Ogre meets an Ogre it uses the Ogre-vs-Ogre battle, but if an Ogre meets a Wizard, it uses a more generic one. Further, what if the characters that meet share no common traits? How can we decide who wins a battle?

    Read the article

  • Very different font sizes across browsers

    - by Yang
    Chrome/WebKit and Firefox have different rendering engines which render fonts differently, in particular with differing dimensions. This isn't too surprising, but what's surprising is the magnitude of some of the differences. I can always tweak individual elements on a page to be more similar, but that's tedious, to say the least. I've been searching for more systematic solutions, but many resources (e.g. SO answers) simply say "use a reset package." While I'm sure this fixes a bunch of other things like padding and spacing, it doesn't seem to make any difference for font dimensions. For instance, if I take the reset package from http://html5reset.org/, I can show pretty big differences (note the layout dimensions shown in the inspectors). [The images below are actually higher res than shown/resized in this answer.] <h1 style="font-size:64px; background-color: #eee;">Article Header</h1> With Helvetica, Chrome is has the shorter height instead. <h1 style="font-size:64px; background-color: #eee; font-family: Helvetica">Article Header</h1> Using a different font, Chrome again renders a much taller font, but additionally the letter spacing goes haywire (probably due to the boldification of the font): <style> @font-face { font-family: "MyriadProRegular"; src: url("fonts/myriadpro-regular-webfont.eot"); src: local("?"), url("fonts/myriadpro-regular-webfont.woff") format("woff"), url("fonts/myriadpro-regular-webfont.ttf") format("truetype"), url("fonts/myriadpro-regular-webfont.svg#webfonteknRmz0m") format("svg"); font-weight: normal; font-style: normal; } @font-face { font-family: "MyriadProLight"; src: url("fonts/myriadpro-light-webfont.eot"); src: local("?"), url("fonts/myriadpro-light-webfont.woff") format("woff"), url("fonts/myriadpro-light-webfont.ttf") format("truetype"), url("fonts/myriadpro-light-webfont.svg#webfont2SBUkD9p") format("svg"); font-weight: normal; font-style: normal; } @font-face { font-family: "MyriadProSemibold"; src: url("fonts/myriadpro-semibold-webfont.eot"); src: local("?"), url("fonts/myriadpro-semibold-webfont.woff") format("woff"), url("fonts/myriadpro-semibold-webfont.ttf") format("truetype"), url("fonts/myriadpro-semibold-webfont.svg#webfontM3ufnW4Z") format("svg"); font-weight: normal; font-style: normal; } </style> ... <h1 style="font-size:64px; background-color: #eee; font-family: Helvetica">Article Header</h1> I've tried a few resets/normalize packages to no avail. I just wanted to confirm here that this is indeed a fact of life (even omitting the more glaring offenders like IE and mobile) and I'm not missing some super-awesome solution to this mess.

    Read the article

  • Nhibernate Migration from 1.0.2.0 to 2.1.2 and many-to-one save problems

    - by Meska
    Hi, we have an old, big asp.net application with nhibernate, which we are extending and upgrading some parts of it. NHibernate that was used was pretty old ( 1.0.2.0), so we decided to upgrade to ( 2.1.2) for the new features. HBM files are generated through custom template with MyGeneration. Everything went quite smoothly, except for one thing. Lets say we have to objects Blog and Post. Blog can have many posts, so Post will have many-to-one relationship. Due to the way that this application operates, relationship is done not through primary keys, but through Blog.Reference column. Sample mapings and .cs files: <?xml version="1.0" encoding="utf-8" ?> <id name="Id" column="Id" type="Guid"> <generator class="assigned"/> </id> <property column="Reference" type="Int32" name="Reference" not-null="true" /> <property column="Name" type="String" name="Name" length="250" /> </class> <?xml version="1.0" encoding="utf-8" ?> <id name="Id" column="Id" type="Guid"> <generator class="assigned"/> </id> <property column="Reference" type="Int32" name="Reference" not-null="true" /> <property column="Name" type="String" name="Name" length="250" /> <many-to-one name="Blog" column="BlogId" class="SampleNamespace.BlogEntity,SampleNamespace" property-ref="Reference" /> </class> And class files class BlogEntity { public Guid Id { get; set; } public int Reference { get; set; } public string Name { get; set; } } class PostEntity { public Guid Id { get; set; } public int Reference { get; set; } public string Name { get; set; } public BlogEntity Blog { get; set; } } Now lets say that i have a Blog with Id 1D270C7B-090D-47E2-8CC5-A3D145838D9C and with Reference 1 In old nhibernate such thing was possible: //this Blog already exists in database BlogEntity blog = new BlogEntity(); blog.Id = Guid.Empty; blog.Reference = 1; //Reference is unique, so we can distinguish Blog by this field blog.Name = "My blog"; //this is new Post, that we are trying to insert PostEntity post = new PostEntity(); post.Id = Guid.NewGuid(); post.Name = "New post"; post.Reference = 1234; post.Blog = blog; session.Save(post); However, in new version, i get an exception that cannot insert NULL into Post.BlogId. As i understand, in old version, for nhibernate it was enough to have Blog.Reference field, and it could retrieve entity by that field, and attach it to PostEntity, and when saving PostEntity, everything would work correctly. And as i understand, new NHibernate tries only to retrieve by Blog.Id. How to solve this? I cannot change DB design, nor can i assign an Id to BlogEntity, as objects are out of my control (they come prefilled as generic "ojbects" like this from external source)

    Read the article

  • MySQL and INT auto_increment fields

    - by PHPguy
    Hello folks, I'm developing in LAMP (Linux+Apache+MySQL+PHP) since I remember myself. But one question was bugging me for years now. I hope you can help me to find an answer and point me into the right direction. Here is my challenge: Say, we are creating a community website, where we allow our users to register. The MySQL table where we store all users would look then like this: CREATE TABLE `users` ( `uid` int(2) unsigned NOT NULL auto_increment COMMENT 'User ID', `name` varchar(20) NOT NULL, `password` varchar(32) NOT NULL COMMENT 'Password is saved as a 32-bytes hash, never in plain text', `email` varchar(64) NOT NULL, `created` int(11) unsigned NOT NULL default '0' COMMENT 'Timestamp of registration', `updated` int(11) unsigned NOT NULL default '0' COMMENT 'Timestamp of profile update, e.g. change of email', PRIMARY KEY (`uid`) ) ENGINE=MyISAM DEFAULT CHARSET=utf8; So, from this snippet you can see that we have a unique and automatically incrementing for every new user 'uid' field. As on every good and loyal community website we need to provide users with possibility to completely delete their profile if they want to cancel their participation in our community. Here comes my problem. Let's say we have 3 registered users: Alice (uid = 1), Bob (uid = 2) and Chris (uid = 3). Now Bob want to delete his profile and stop using our community. If we delete Bob's profile from the 'users' table then his missing 'uid' will create a gap which will be never filled again. In my opinion it's a huge waste of uid's. I see 3 possible solutions here: 1) Increase the capacity of the 'uid' field in our table from SMALLINT (int(2)) to, for example, BIGINT (int(8)) and ignore the fact that some of the uid's will be wasted. 2) introduce the new field 'is_deleted', which will be used to mark deleted profiles (but keep them in the table, instead of deleting them) to re-utilize their uid's for newly registered users. The table will look then like this: CREATE TABLE `users` ( `uid` int(2) unsigned NOT NULL auto_increment COMMENT 'User ID', `name` varchar(20) NOT NULL, `password` varchar(32) NOT NULL COMMENT 'Password is saved as a 32-bytes hash, never in plain text', `email` varchar(64) NOT NULL, `is_deleted` int(1) unsigned NOT NULL default '0' COMMENT 'If equal to "1" then the profile has been deleted and will be re-used for new registrations', `created` int(11) unsigned NOT NULL default '0' COMMENT 'Timestamp of registration', `updated` int(11) unsigned NOT NULL default '0' COMMENT 'Timestamp of profile update, e.g. change of email', PRIMARY KEY (`uid`) ) ENGINE=MyISAM DEFAULT CHARSET=utf8; 3) Write a script to shift all following user records once a previous record has been deleted. E.g. in our case when Bob (uid = 2) decides to remove his profile, we would replace his record with the record of Chris (uid = 3), so that uid of Chris becomes qual to 2 and mark (is_deleted = '1') the old record of Chris as vacant for the new users. In this case we keep the chronological order of uid's according to the registration time, so that the older users have lower uid's. Please, advice me now which way is the right way to handle the gaps in the auto_increment fields. This is just one example with users, but such cases occur very often in my programming experience. Thanks in advance!

    Read the article

  • "C variable type sizes are machine dependent." Is it really true? signed & unsigned numbers ;

    - by claws
    Hello, I've been told that C types are machine dependent. Today I wanted to verify it. void legacyTypes() { /* character types */ char k_char = 'a'; //Signedness --> signed & unsigned signed char k_char_s = 'a'; unsigned char k_char_u = 'a'; /* integer types */ int k_int = 1; /* Same as "signed int" */ //Signedness --> signed & unsigned signed int k_int_s = -2; unsigned int k_int_u = 3; //Size --> short, _____, long, long long short int k_s_int = 4; long int k_l_int = 5; long long int k_ll_int = 6; /* real number types */ float k_float = 7; double k_double = 8; } I compiled it on a 32-Bit machine using minGW C compiler _legacyTypes: pushl %ebp movl %esp, %ebp subl $48, %esp movb $97, -1(%ebp) # char movb $97, -2(%ebp) # signed char movb $97, -3(%ebp) # unsigned char movl $1, -8(%ebp) # int movl $-2, -12(%ebp)# signed int movl $3, -16(%ebp) # unsigned int movw $4, -18(%ebp) # short int movl $5, -24(%ebp) # long int movl $6, -32(%ebp) # long long int movl $0, -28(%ebp) movl $0x40e00000, %eax movl %eax, -36(%ebp) fldl LC2 fstpl -48(%ebp) leave ret I compiled the same code on 64-Bit processor (Intel Core 2 Duo) on GCC (linux) legacyTypes: .LFB2: .cfi_startproc pushq %rbp .cfi_def_cfa_offset 16 movq %rsp, %rbp .cfi_offset 6, -16 .cfi_def_cfa_register 6 movb $97, -1(%rbp) # char movb $97, -2(%rbp) # signed char movb $97, -3(%rbp) # unsigned char movl $1, -12(%rbp) # int movl $-2, -16(%rbp)# signed int movl $3, -20(%rbp) # unsigned int movw $4, -6(%rbp) # short int movq $5, -32(%rbp) # long int movq $6, -40(%rbp) # long long int movl $0x40e00000, %eax movl %eax, -24(%rbp) movabsq $4620693217682128896, %rax movq %rax, -48(%rbp) leave ret Observations char, signed char, unsigned char, int, unsigned int, signed int, short int, unsigned short int, signed short int all occupy same no. of bytes on both 32-Bit & 64-Bit Processor. The only change is in long int & long long int both of these occupy 32-bit on 32-bit machine & 64-bit on 64-bit machine. And also the pointers, which take 32-bit on 32-bit CPU & 64-bit on 64-bit CPU. Questions: I cannot say, what the books say is wrong. But I'm missing something here. What exactly does "Variable types are machine dependent mean?" As you can see, There is no difference between instructions for unsigned & signed numbers. Then how come the range of numbers that can be addressed using both is different? I was reading http://stackoverflow.com/questions/2511246/how-to-maintain-fixed-size-of-c-variable-types-over-different-machines I didn't get the purpose of the question or their answers. What maintaining fixed size? They all are the same. I didn't understand how those answers are going to ensure the same size.

    Read the article

  • How to design a data model that deals with (real) contracts?

    - by Geoffrey
    I was looking for some advice on designing a data model for contract administration. The general life cycle of a contract is thus: Contract is created and in a "draft" state. It is viewable internally and changes may be made. Contract goes out to vendor, status is set to "pending" Contract is rejected by vendor. At this state, nothing can be done to the contract. No statuses may be added to the collection. Contract is accepted by vendor. At this state, nothing can be done to the contract. No statuses may be added to the collection. I obviously want to avoid a situation where the contract is accepted and, say, the amount is changed. Here are my classes: [EnforceNoChangesAfterDraftState] public class VendorContract { public virtual Vendor Vendor { get; set; } public virtual decimal Amount { get; set; } public virtual VendorContact VendorContact { get; set; } public virtual string CreatedBy { get; set; } public virtual DateTime CreatedOn { get; set; } public virtual FileStore Contract { get; set; } public virtual IList<VendorContractStatus> ContractStatus { get; set; } } [EnforceCorrectWorkflow] public class VendorContractStatus { public virtual VendorContract VendorContract { get; set; } public virtual FileStore ExecutedDocument { get; set; } public virtual string Status { get; set; } public virtual string Reason { get; set; } public virtual string CreatedBy { get; set; } public virtual DateTime CreatedOn { get; set; } } I've omitted the filestore class, which is basically a key/value lookup to find the document based on its guid. The VendorContractStatus is mapped as a many-to-one in Nhibernate. I then use a custom validator as described here. If anything but draft is returned in the VendorContractStatus collection, no changes are allowed. Furthermore the VendorContractStatus must follow the correct workflow (you can add a rejected after a pending, but you can't add anything else to the collection if a reject or accepted exists, etc.). All sounds alright? Well a colleague has argued that we should simply add an "IsDraft" bool property to VendorContract and not accept updates if IsDraft is false. Then we should setup a method inside of VendorContractStatus for updating the status, if something gets added after a draft, it sets the IsDraft property of VendorContract to false. I do not like this as it feels like I'm dirtying up the POCOs and adding logic that should persist in the validation area, that no rules should really exist in these classes and they shouldn't be aware of their states. Any thoughts on this and what is the better practice from a DDD perspective? From my view, if in the future we want more complex rules, my way will be more maintainable over the long run. Say we have contracts over a certain amount to be approved by a manager. I would think it would be better to have a one-to-one mapping with a VendorContractApproval class, rather than adding IsApproved properties, but that's just speculation. This might be splitting hairs, but this is the first real gritty enterprise software project we've done. Any advice would be appreciated!

    Read the article

  • More localized, efficient Lowest Common Ancestor algorithm given multiple binary trees?

    - by mstksg
    I have multiple binary trees stored as an array. In each slot is either nil (or null; pick your language) or a fixed tuple storing two numbers: the indices of the two "children". No node will have only one child -- it's either none or two. Think of each slot as a binary node that only stores pointers to its children, and no inherent value. Take this system of binary trees: 0 1 / \ / \ 2 3 4 5 / \ / \ 6 7 8 9 / \ 10 11 The associated array would be: 0 1 2 3 4 5 6 7 8 9 10 11 [ [2,3] , [4,5] , [6,7] , nil , nil , [8,9] , nil , [10,11] , nil , nil , nil , nil ] I've already written simple functions to find direct parents of nodes (simply by searching from the front until there is a node that contains the child) Furthermore, let us say that at relevant times, both all trees are anywhere between a few to a few thousand levels deep. I'd like to find a function P(m,n) to find the lowest common ancestor of m and n -- to put more formally, the LCA is defined as the "lowest", or deepest node in which have m and n as descendants (children, or children of children, etc.). If there is none, a nil would be a valid return. Some examples, given our given tree: P( 6,11) # => 2 P( 3,10) # => 0 P( 8, 6) # => nil P( 2,11) # => 2 The main method I've been able to find is one that uses an Euler trace, which turns the given tree, with a node A to be the invisible parent of 0 and 1 with a depth of -1, into: A-0-2-6-2-7-10-7-11-7-2-0-3-0-A-1-4-1-5-8-5-9-5-1-A And from that, simply find the node between your given m and n that has the lowest number; For example, to find P(6,11), look for a 6 and an 11 on the trace. The number between them that is the lowest is 2, and that's your answer. If A is in between them, return nil. -- Calculating P(6,11) -- A-0-2-6-2-7-10-7-11-7-2-0-3-0-A-1-4-1-5-8-5-9-5-1-A ^ ^ ^ | | | m lowest n Unfortunately, I do believe that finding the Euler trace of a tree that can be several thousands of levels deep is a bit machine-taxing...and because my tree is constantly being changed throughout the course of the programming, every time I wanted to find the LCA, I'd have to re-calculate the Euler trace and hold it in memory every time. Is there a more memory efficient way, given the framework I'm using? One that maybe iterates upwards? One way I could think of would be the "count" the generation/depth of both nodes, and climb the lowest node until it matched the depth of the highest, and increment both until they find someone similar. But that'd involve climbing up from level, say, 3025, back to 0, twice, to count the generation, and using a terribly inefficient climbing-up algorithm in the first place, and then re-climbing back up. Are there any other better ways?

    Read the article

  • Defend PHP; convince me it isn't horrible

    - by Jason L
    I made a tongue-in-cheek comment in another question thread calling PHP a terrible language and it got down-voted like crazy. Apparently there are lots of people here who love PHP. So I'm genuinely curious. What am I missing? What makes PHP a good language? Here are my reasons for disliking it: PHP has inconsistent naming of built-in and library functions. Predictable naming patterns are important in any design. PHP has inconsistent parameter ordering of built-in functions, eg array_map vs. array_filter which is annoying in the simple cases and raises all sorts of unexpected behaviour or worse. The PHP developers constantly deprecate built-in functions and lower-level functionality. A good example is when they deprecated pass-by-reference for functions. This created a nightmare for anyone doing, say, function callbacks. A lack of consideration in redesign. The above deprecation eliminated the ability to, in many cases, provide default keyword values for functions. They fixed this in PHP 5, but they deprecated the pass-by-reference in PHP 4! Poor execution of name spaces (formerly no name spaces at all). Now that name spaces exist, what do we use as the dereference character? Backslash! The character used universally for escaping, even in PHP! Overly-broad implicit type conversion leads to bugs. I have no problem with implicit conversions of, say, float to integer or back again. But PHP (last I checked) will happily attempt to magically convert an array to an integer. Poor recursion performance. Recursion is a fundamentally important tool for writing in any language; it can make complex algorithms far simpler. Poor support is inexcusable. Functions are case insensitive. I have no idea what they were thinking on this one. A programming language is a way to specify behavior to both a computer and a reader of the code without ambiguity. Case insensitivity introduces much ambiguity. PHP encourages (practically requires) a coupling of processing with presentation. Yes, you can write PHP that doesn't do so, but it's actually easier to write code in the incorrect (from a sound design perspective) manner. PHP performance is abysmal without caching. Does anyone sell a commercial caching product for PHP? Oh, look, the designers of PHP do. Worst of all, PHP convinces people that designing web applications is easy. And it does indeed make much of the effort involved much easier. But the fact is, designing a web application that is both secure and efficient is a very difficult task. By convincing so many to take up programming, PHP has taught an entire subgroup of programmers bad habits and bad design. It's given them access to capabilities that they lack the understanding to use safely. This has led to PHP's reputation as being insecure. (However, I will readily admit that PHP is no more or less secure than any other web programming language.) What is it that I'm missing about PHP? I'm seeing an organically-grown, poorly-managed mess of a language that's spawning poor programmers. So convince me otherwise!

    Read the article

  • C++ templated factory constructor/de-serialization

    - by KRao
    Hi, I was looking at the boost serialization library, and the intrusive way to provide support for serialization is to define a member function with signature (simplifying): class ToBeSerialized { public: //Define this to support serialization //Notice not virtual function! template<class Archive> void serialize(Archive & ar) {.....} }; Moreover, one way to support serilization of derived class trough base pointers is to use a macro of the type: //No mention to the base class(es) from which Derived_class inherits BOOST_CLASS_EXPORT_GUID(Derived_class, "derived_class") where Derived_class is some class which is inheriting from a base class, say Base_class. Thanks to this macro, it is possible to serialize classes of type Derived_class through pointers to Base_class correctly. The question is: I am used in C++ to write abstract factories implemented through a map from std::string to (pointer to) functions which return objects of the desired type (and eveything is fine thanks to covariant types). Hover I fail to see how I could use the above non-virtual serialize template member function to properly de-serialize (i.e. construct) an object without knowing its type (but assuming that the type information has been stored by the serializer, say in a string). What I would like to do (keeping the same nomenclature as above) is something like the following: XmlArchive xmlArchive; //A type or archive xmlArchive.open("C:/ser.txt"); //Contains type information for the serialized class Base_class* basePtr = Factory<Base_class>::create("derived_class",xmlArchive); with the function on the righ-hand side creating an object on the heap of type Derived_class (via default constructor, this is the part I know how to solve) and calling the serialize function of xmlArchive (here I am stuck!), i.e. do something like: Base_class* Factory<Base_class>::create("derived_class",xmlArchive) { Base_class* basePtr = new Base_class; //OK, doable, usual map string to pointer to function static_cast<Derived_class*>( basePtr )->serialize( xmlArchive ); //De-serialization, how????? return basePtr; } I am sure this can be done (boost serialize does it but its code is impenetrable! :P), but I fail to figure out how. The key problem is that the serialize function is a template function. So I cannot have a pointer to a generic templated function. As the point in writing the templated serialize function is to make the code generic (i.e. not having to re-write the serialize function for different Archivers), it does not make sense then to have to register all the derived classes for all possible archive types, like: MY_CLASS_REGISTER(Derived_class, XmlArchive); MY_CLASS_REGISTER(Derived_class, TxtArchive); ... In fact in my code I relies on overloading to get the correct behaviour: void serialize( XmlArchive& archive, Derived_class& derived ); void serialize( TxtArchive& archive, Derived_class& derived ); ... The key point to keep in mind is that the archive type is always known, i.e. I am never using runtime polymorphism for the archive class...(again I am using overloading on the archive type). Any suggestion to help me out? Thank you very much in advance! Cheers

    Read the article

  • How can I determine/use $(this) in js callback script

    - by Rabbott
    I am using Rails and jQuery, making an ajax call initiated by clicking a link. I setup my application.js file to look like the one proposed here and it works great. The problem I'm having is how can I use $(this) in my say.. update.js.erb file to represent the link I clicked? I don't want to have to assign an ID to every one, then recompile that id in the callback script.. EDIT To give a simple example of something similar to what I'm trying to do (and much easier to explain): If a user clicks on a link, that deletes that element from a list, the controller would handle the callback, and the callback (which is in question here) would delete the element I clicked on, so in the callback delete.js.erb would just say $(this).fadeOut(); This is why I want to use $(this) so that I dont have to assign an ID to every element (which would be the end of the world, just more verbose markup) application.js jQuery.ajaxSetup({ 'beforeSend': function(xhr) {xhr.setRequestHeader("Accept", "text/javascript,application/javascript,text/html")} }) function _ajax_request(url, data, callback, type, method) { if (jQuery.isFunction(data)) { callback = data; data = {}; } return jQuery.ajax({ type: method, url: url, data: data, success: callback, dataType: type }); } jQuery.extend({ put: function(url, data, callback, type) { return _ajax_request(url, data, callback, type, 'PUT'); }, delete_: function(url, data, callback, type) { return _ajax_request(url, data, callback, type, 'DELETE'); } }); jQuery.fn.submitWithAjax = function() { this.unbind('submit', false); this.submit(function() { $.post(this.action, $(this).serialize(), null, "script"); return false; }) return this; }; // Send data via get if <acronym title="JavaScript">JS</acronym> enabled jQuery.fn.getWithAjax = function() { this.unbind('click', false); this.click(function() { $.get($(this).attr("href"), $(this).serialize(), null, "script"); return false; }) return this; }; // Send data via Post if <acronym title="JavaScript">JS</acronym> enabled jQuery.fn.postWithAjax = function() { this.unbind('click', false); this.click(function() { $.post($(this).attr("href"), $(this).serialize(), null, "script"); return false; }) return this; }; jQuery.fn.putWithAjax = function() { this.unbind('click', false); this.click(function() { $.put($(this).attr("href"), $(this).serialize(), null, "script"); return false; }) return this; }; jQuery.fn.deleteWithAjax = function() { this.removeAttr('onclick'); this.unbind('click', false); this.click(function() { $.delete_($(this).attr("href"), $(this).serialize(), null, "script"); return false; }) return this; }; // This will "ajaxify" the links function ajaxLinks(){ $('.ajaxForm').submitWithAjax(); $('a.get').getWithAjax(); $('a.post').postWithAjax(); $('a.put').putWithAjax(); $('a.delete').deleteWithAjax(); } show.html.erb <%= link_to 'Link Title', article_path(a, :sentiment => Article::Sentiment['Neutral']), :class => 'put' %> The combination of the two things will call update.js.erb in rails, the code in that file is used as the callback of the ajax ($.put in this case) update.js.erb // user feedback $("#notice").html('<%= flash[:notice] %>'); // update the background color $(this OR e.target).attr("color", "red");

    Read the article

  • C# ambiguity in Func + extension methods + lambdas

    - by Hobbes
    I've been trying to make my way through this article: http://blogs.msdn.com/wesdyer/archive/2008/01/11/the-marvels-of-monads.aspx ... And something on page 1 made me uncomfortable. In particular, I was trying to wrap my head around the Compose<() function, and I wrote an example for myself. Consider the following two Func's: Func<double, double> addTenth = x => x + 0.10; Func<double, string> toPercentString = x => (x * 100.0).ToString() + "%"; No problem! It's easy to understand what these two do. Now, following the example from the article, you can write a generic extension method to compose these functions, like so: public static class ExtensionMethods { public static Func<TInput, TLastOutput> Compose<TInput, TFirstOutput, TLastOutput>( this Func<TFirstOutput, TLastOutput> toPercentString, Func<TInput, TFirstOutput> addTenth) { return input => toPercentString(addTenth(input)); } } Fine. So now you can say: string x = toPercentString.Compose<double, double, string>(addTenth)(0.4); And you get the string "50%" So far, so good. But there's something ambiguous here. Let's say you write another extension method, so now you have two functions: public static class ExtensionMethods { public static Func<TInput, TLastOutput> Compose<TInput, TFirstOutput, TLastOutput>( this Func<TFirstOutput, TLastOutput> toPercentString, Func<TInput, TFirstOutput> addTenth) { return input => toPercentString(addTenth(input)); } public static Func<double, string> Compose<TInput, TFirstOutput, TLastOutput>(this Func<double, string> toPercentString, Func<double, double> addTenth) { return input => toPercentString(addTenth(input + 99999)); } } Herein is the ambiguity. Don't these two function have overlapping signatures? Yes. Does this even compile? Yes. Which one get's called? The second one (which clearly gives you the "wrong" result) gets called. If you comment out either function, it still compiles, but you get different results. It seems like nitpicking, but there's something that deeply offends my sensibilities here, and I can't put my finger on it. Does it have to do with extension methods? Does it have to do with lambdas? Or does it have to do with how Func< allows you to parameterize the return type? I'm not sure. I'm guessing that this is all addressed somewhere in the spec, but I don't even know what to Google to find this. Help!

    Read the article

  • HTML 5 <video> tag vs Flash video. What are the pros and cons?

    - by Vilx-
    Seems like the new <video> tag is all the hype these days, especially since Firefox now supports it. News of this are popping up in blogs all over the place, and everyone seems to be excited. But what about? As much as I searched I could not find anything that would make it better than the good old Flash video. In fact, I see only problems with it: It will still be some time before all the browsers start supporting it, and much more time before most people upgrade; Flash is available already and everyone has it; You can couple Flash with whatever fancy UI you want for controlling the playback. I gather that the tag will be controllable as well (via JavaScript probably), but will it be able to go fullscreen? The only two pros for a <video> tag that I can see are: It is more "semantic" - which probably holds no importance to a whole lot of people, including me; It is not dependent on a single commercial 3rd party entity (Adobe) - which I also don't see as a compelling reason to switch, because free players and video converters are already available, and Adobe is not hindering the whole process in any way (it's not in their interests even). So... what's the big deal? Added: OK, so there is one more Pro... maybe. Support for mobile devices. Hard to say though. A number of thoughts race through my head about the subject: How many mobile devices are actually able to decode video at a decent speed anyway, Flash or otherwise? How long until mainstream mobile devices get the <video> support? Even if it is available through updates, how many people actually do that? How many people watch videos on web pages on their mobile phones at all? As for the semantics part - I understand that search engines might be able to detect videos better now, but... what will they do with them anyway? OK, so they know that there is a video in the page. And? They can't index a video! I'd like some more arguments here. Added: Just thought of another Cons. This opens up a whole new area of cross-browser incompatibility. HTML and CSS is quite messy already in this aspect. Flash at least is the same everywhere. But it's enough for at least one major browser vendor to decide against the <video> tag (can anyone say "Internet Explorer"?) and we have a nice new area of hell to explore. Added: A Pro just came in. More competition = more innovation. That's true. Giving Adobe more competition will probably force them to improve Flash in areas it has been lacking so far. Linux seems to be a weak spot for it, cited by many.

    Read the article

  • What's my best approach on this simple hierarchy Java Problem?

    - by Nazgulled
    First, I'm sorry for the question title but I can't think of a better one to describe my problem. Feel free to change it :) Let's say I have this abstract class Box which implements a couple of constructors, methods and whatever on some private variables. Then I have a couple of sub classes like BoxA and BoxB. Both of these implement extra things. Now I have another abstract class Shape and a few sub classes like Square and Circle. For both BoxA and BoxB I need to have a list of Shape objects but I need to make sure that only Square objects go into BoxA's list and only Circle objects go into BoxB's list. For that list (on each box), I need to have a get() and set() method and also a addShape() and removeShape() methods. Another important thing to know is that for each box created, either BoxA or BoxB, each respectively Shape list is exactly the same. Let's say I create a list of Square's named ls and two BoxA objects named boxA1 and boxA2. No matter what, both boxA1 and boxA2 must have the same ls list. This is my idea: public abstract class Box { // private instance variables public Box() { // constructor stuff } // public instance methods } public class BoxA extends Box { // private instance variables private static List<Shape> list; public BoxA() { // constructor stuff } // public instance methods public static List<Square> getList() { List<Square> aux = new ArrayList<Square>(); for(Square s : list.values()) { aux.add(s.clone()); // I know what I'm doing with this clone, don't worry about it } return aux; } public static void setList(List<Square> newList) { list = new ArrayList<Square>(newList); } public static void addShape(Square s) { list.add(s); } public static void removeShape(Square s) { list.remove(list.indexOf(s)); } } As the list needs to be the same for that type of object, I declared as static and all methods that work with that list are also static. Now, for BoxB the class would be almost the same regarding the list stuff. I would only replace Square by Triangle and the problem was solved. So, for each BoxA object created, the list would be only one and the same for each BoxB object created, but a different type of list of course. So, what's my problem you ask? Well, I don't like the code... The getList(), setList(), addShape() and removeShape() methods are basically repeated for BoxA and BoxB, only the type of the objects that the list will hold is different. I can't think of way to do it in the super class Box instead. Doing it statically too, using Shape instead of Square and Triangle, wouldn't work because the list would be only one and I need it to be only one but for each sub class of Box. How could I do this differently and better? P.S: I could not describe my real example because I don't know the correct words in English for the stuff I'm doing, so I just used a box and shapes example, but it's basically the same.

    Read the article

  • How to make MySQL utilize available system resources, or find "the real problem"?

    - by anonymous coward
    This is a MySQL 5.0.26 server, running on SuSE Enterprise 10. This may be a Serverfault question. The web user interface that uses these particular queries (below) is showing sometimes 30+, even up to 120+ seconds at the worst, to generate the pages involved. On development, when the queries are run alone, they take up to 20 seconds on the first run (with no query cache enabled) but anywhere from 2 to 7 seconds after that - I assume because the tables and indexes involved have been placed into ram. From what I can tell, the longest load times are caused by Read/Update Locking. These are MyISAM tables. So it looks like a long update comes in, followed by a couple 7 second queries, and they're just adding up. And I'm fine with that explanation. What I'm not fine with is that MySQL doesn't appear to be utilizing the hardware it's on, and while the bottleneck seems to be the database, I can't understand why. I would say "throw more hardware at it", but we did and it doesn't appear to have changed the situation. Viewing a 'top' during the slowest times never shows much cpu or memory utilization by mysqld, as if the server is having no trouble at all - but then, why are the queries taking so long? How can I make MySQL use the crap out of this hardware, or find out what I'm doing wrong? Extra Details: On the "Memory Health" tab in the MySQL Administrator (for Windows), the Key Buffer is less than 1/8th used - so all the indexes should be in RAM. I can provide a screen shot of any graphs that might help. So desperate to fix this issue. Suffice it to say, there is legacy code "generating" these queries, and they're pretty much stuck the way they are. I have tried every combination of Indexes on the tables involved, but any suggestions are welcome. Here's the current Create Table statement from development (the 'experimental' key I have added, seems to help a little, for the example query only): CREATE TABLE `registration_task` ( `id` varchar(36) NOT NULL default '', `date_entered` datetime NOT NULL default '0000-00-00 00:00:00', `date_modified` datetime NOT NULL default '0000-00-00 00:00:00', `assigned_user_id` varchar(36) default NULL, `modified_user_id` varchar(36) default NULL, `created_by` varchar(36) default NULL, `name` varchar(80) NOT NULL default '', `status` varchar(255) default NULL, `date_due` date default NULL, `time_due` time default NULL, `date_start` date default NULL, `time_start` time default NULL, `parent_id` varchar(36) NOT NULL default '', `priority` varchar(255) NOT NULL default '9', `description` text, `order_number` int(11) default '1', `task_number` int(11) default NULL, `depends_on_id` varchar(36) default NULL, `milestone_flag` varchar(255) default NULL, `estimated_effort` int(11) default NULL, `actual_effort` int(11) default NULL, `utilization` int(11) default '100', `percent_complete` int(11) default '0', `deleted` tinyint(1) NOT NULL default '0', `wf_task_id` varchar(36) default '0', `reg_field` varchar(8) default '', `date_offset` int(11) default '0', `date_source` varchar(10) default '', `date_completed` date default '0000-00-00', `completed_id` varchar(36) default NULL, `original_name` varchar(80) default NULL, PRIMARY KEY (`id`), KEY `idx_reg_task_p` (`deleted`,`parent_id`), KEY `By_Assignee` (`assigned_user_id`,`deleted`), KEY `status_assignee` (`status`,`deleted`), KEY `experimental` (`deleted`,`status`,`assigned_user_id`,`parent_id`,`date_due`) ) ENGINE=MyISAM DEFAULT CHARSET=latin1 And one of the ridiculous queries in question: SELECT users.user_name assigned_user_name, registration.FIELD001 parent_name, registration_task.status status, registration_task.date_modified date_modified, registration_task.date_due date_due, registration.FIELD240 assigned_wf, if(LENGTH(registration_task.description)>0,1,0) has_description, registration_task.* FROM registration_task LEFT JOIN users ON registration_task.assigned_user_id=users.id LEFT JOIN registration ON registration_task.parent_id=registration.id where (registration_task.status != 'Completed' AND registration.FIELD001 LIKE '%' AND registration_task.name LIKE '%' AND registration.FIELD060 LIKE 'GN001472%') AND registration_task.deleted=0 ORDER BY date_due asc LIMIT 0,20; my.cnf - '[mysqld]' section. [mysqld] port = 3306 socket = /var/lib/mysql/mysql.sock skip-locking key_buffer = 384M max_allowed_packet = 100M table_cache = 2048 sort_buffer_size = 2M net_buffer_length = 100M read_buffer_size = 2M read_rnd_buffer_size = 160M myisam_sort_buffer_size = 128M query_cache_size = 16M query_cache_limit = 1M EXPLAIN above query, without additional index: +----+-------------+-------------------+--------+--------------------------------+----------------+---------+------------------------------------------------+---------+-----------------------------+ | id | select_type | table | type | possible_keys | key | key_len | ref | rows | Extra | +----+-------------+-------------------+--------+--------------------------------+----------------+---------+------------------------------------------------+---------+-----------------------------+ | 1 | SIMPLE | registration_task | ref | idx_reg_task_p,status_assignee | idx_reg_task_p | 1 | const | 1067354 | Using where; Using filesort | | 1 | SIMPLE | registration | eq_ref | PRIMARY,gbl | PRIMARY | 8 | sugarcrm401.registration_task.parent_id | 1 | Using where | | 1 | SIMPLE | users | ref | PRIMARY | PRIMARY | 38 | sugarcrm401.registration_task.assigned_user_id | 1 | | +----+-------------+-------------------+--------+--------------------------------+----------------+---------+------------------------------------------------+---------+-----------------------------+ EXPLAIN above query, with 'experimental' index: +----+-------------+-------------------+--------+-----------------------------------------------------------+------------------+---------+------------------------------------------------+--------+-----------------------------+ | id | select_type | table | type | possible_keys | key | key_len | ref | rows | Extra | +----+-------------+-------------------+--------+-----------------------------------------------------------+------------------+---------+------------------------------------------------+--------+-----------------------------+ | 1 | SIMPLE | registration_task | range | idx_reg_task_p,status_assignee,NewIndex1,tcg_experimental | tcg_experimental | 259 | NULL | 103345 | Using where; Using filesort | | 1 | SIMPLE | registration | eq_ref | PRIMARY,gbl | PRIMARY | 8 | sugarcrm401.registration_task.parent_id | 1 | Using where | | 1 | SIMPLE | users | ref | PRIMARY | PRIMARY | 38 | sugarcrm401.registration_task.assigned_user_id | 1 | | +----+-------------+-------------------+--------+-----------------------------------------------------------+------------------+---------+------------------------------------------------+--------+-----------------------------+

    Read the article

  • I never really understood: what is Application Binary Interface (ABI)?

    - by claws
    I never clearly understood what is an ABI. I'm sorry for such a lengthy question. I just want to clearly understand things. Please don't point me to wiki article, If could understand it, I wouldn't be here posting such a lengthy post. This is my mindset about different interfaces: TV remote is an interface between user and TV. It is an existing entity but useless (doesn't provide any functionality) by itself. All the functionality for each of those buttons on the remote is implemented in the Television set. Interface: It is a "existing entity" layer between the functionality and consumer of that functionality. An, interface by itself is doesn't do anything. It just invokes the functionality lying behind. Now depending on who the user is there are different type of interfaces. Command Line Interface(CLI) commands are the existing entities, consumer is the user and functionality lies behind. functionality: my software functionality which solves some purpose to which we are describing this interface. existing entities: commands consumer: user Graphical User Interface(GUI) window,buttons etc.. are the existing entities, again consumer is the user and functionality lies behind. functionality: my software functionality which solves some purpose to which we are describing this interface. existing entities: window,buttons etc.. consumer: user Application Programming Interface(API) functions or to be more correct, interfaces (in interfaced based programming) are the existing entities, consumer here is another program not a user. and again functionality lies behind this layer. functionality: my software functionality which solves some purpose to which we are describing this interface. existing entities: functions, Interfaces(array of functions). consumer: another program/application. Application Binary Interface (ABI) Here is my problem starts. functionality: ??? existing entities: ??? consumer: ??? I've wrote few softwares in different languages and provided different kind of interfaces (CLI, GUI, API) but I'm not sure, if I ever, provided any ABI. http://en.wikipedia.org/wiki/Application_binary_interface says: ABIs cover details such as data type, size, and alignment; the calling convention, which controls how functions' arguments are passed and return values retrieved; the system call numbers and how an application should make system calls to the operating system; Other ABIs standardize details such as the C++ name mangling,[2] . exception propagation,[3] and calling convention between compilers on the same platform, but do not require cross-platform compatibility. Who needs these details? Please don't say, OS. I know assembly programming. I know how linking & loading works. I know what exactly happens inside. Where did C++ name mangling come in between? I thought we are talking at the binary level. Where did languages come in between? anyway, I've downloaded the [PDF] System V Application Binary Interface Edition 4.1 (1997-03-18) to see what exactly it contains. Well, most of it didn't make any sense. Why does it contain 2 chapters (4th & 5th) which describe the ELF file format.Infact, these are the only 2 significant chapters that specification. Rest of all the chapters "Processor Specific". Anyway, I thought that it is completely different topic. Please don't say that ELF file format specs are the ABI. It doesn't qualify to be Interface according to the definition. I know, since we are talking at such low level it must be very specific. But I'm not sure how is it "Instruction Set Architecture(ISA)" specific? Where can I find MS Window's ABI? So, these are the major queries that are bugging me.

    Read the article

  • show hidden div tag from another page

    - by neueweblernen
    I'm trying to link to an all-inclusive FAQ page from various pages. The answers are contained in tags, nested within a line item of an unordered list housed by categories. The FAQ page has the following categories: Practical Nurse Exam Online Renewal Practice Hours etc. Under Practical Nurse Exam, there are sub categories, subjects, with questions below in tags that expand onClick. (e.g. Examination Day, Exam Results, etc.) Let's say I'm on a different page called Registration and there's a link to the FAQs for Exam Results. I'm able to link to the page and included the hashtag on the anchor or Exam Results, but it does not expand the subcategory. I've read this thread but it didn't work for me. Please help! The code is below: <script type="text/javascript"> function toggle(Info,pic) { var CState = document.getElementById(Info); CState.style.display = (CState.style.display != 'block') ? 'block' : 'none'; } window.onload = function() { var hash = window.location.hash; // would be "#div1" or something if(hash != "") { var id = hash.substr(1); // get rid of # document.getElementById(id).style.display = 'block'; } } </script> <style type="text/css"> .FAQ { cursor:hand; cursor:pointer; } .FAA { display:none; padding-left:20px; text-indent:-20px; } #FAQlist li { list-style-type: none; } #FAQlist ul { margin-left:0px; } headingOne{ font-family:Arial, Helvetica, sans-serif; color:#66BBFF; font-size:20px; font-weight:bold;} </style> Here's the body (part of it anyway) <headingOne class="FAQ" onClick="toggle('CPNRE', this)">PRACTICAL NURSE EXAM</headingOne> <div class="FAA" id="CPNRE"> <h3><a name="applying">Applying to write the CPNRE</a></h3> <ul id="FAQlist" style="width:450px;"> <li class="FAQ"> <p onclick="toggle('faq1',this)"> <strong>Q: How much does it cost to write the exam?</strong></p> <div class="FAA" id="faq1"> <b>A.</b> In 2013, the cost for the first exam writing is $600.00 which includes the interim license fee. See <a href="https://www.clpnbc.org/What-is-an-LPN/Becoming-an-LPN/Canadian-Practical-Nurse-Registration-Examination/Fees-and-Deadlines.aspx"> fee schedule</a>.</div> <hr /> </li> and here's the body of the other page that contains the link and the same script syntax as the all-inclusive FAQ page. This is just a test, that's not exactly what it will say: <a onclick="toggle('CPNRE', this)" href="file:///S|/Designs/Web stuff/FAQ all inclusive.html#applying"> click here</a>

    Read the article

  • If I use a facade class with generic methods to access the JPA API, how should I provide additional processing for specific types?

    - by Shaun
    Let's say I'm making a fairly simple web application using JAVA EE specs (I've heard this is possible). In this app, I only have about 10 domain/data objects, and these are represented by JPA Entities. Architecturally, I would consider the JPA API to perform the role of a DAO. Of course, I don't want to use the EntityManager directly in my UI (JSF) and I need to manage transactions, so I delegate these tasks to the so-called service layer. More specifically, I would like to be able to handle these tasks in a single DataService class (often also called CrudService) with generic methods. See this article by Adam Bien for an example interface: http://www.adam-bien.com/roller/abien/entry/generic_crud_service_aka_dao My project differs from that article in that I can't use EJBs, so my service classes are essentially just named beans and I handle transactions manually. Regardless, what I want is a single interface for simple CRUD operations on my data objects because having a different class for each data type would lead to a lot of duplicate and/or unnecessary code. Ideally, my views would be able to use a method such as public <T> List<T> findAll(Class<T> type) { ... } to retrieve data. Using JSF, it might look something like this: <h:dataTable value="#{dataService.findAll(data.class)}" var="d"> ... </h:dataTable> Similarly, after validating forms, my controller could submit the data with a method such as: public <T> void add(T entity) { ... } Granted, you'd probably actually want to return something useful to the caller. In any case, this works well if your data can be treated as homogenous in this manner. Alas, it breaks down when you need to perform additional processing on certain objects before passing them on to JPA. For example, let's say I'm dealing with Books and Authors which have a many-to-many relationship. Each Book has a set of IDs referring to its authors, and each Author has a set of IDs referring to their books. Normally, JPA can manage this kind of relationship for you, but in some cases it can't (for example, the google app engine JPA provider doesn't support this). Thus, when I persist a new book for example, I may need to update the corresponding author entities. My question, then, is if there's an elegant way to handle this or if I should reconsider the sanity of my whole design. Here's a couple ways I see of dealing with it: The instanceof operator. I could use this to target certain classes when special processing is needed. Perhaps maintainability suffers and it isn't beautiful code, but if there's only 10 or so domain objects it can't be all that bad... could it? Make a different service for each entity type (ie, BookService and AuthorService). All services would inherit from a generic DataService base class and override methods if special processing is needed. At this point, you could probably also just call them DAOs instead. As always, I appreciate the help. Let me know if any clarifications are needed, as I left out many smaller details.

    Read the article

  • Setting up a VPN connection to Amazon VPC - routing

    - by Keeno
    I am having some real issues setting up a VPN between out office and AWS VPC. The "tunnels" appear to be up, however I don't know if they are configured correctly. The device I am using is a Netgear VPN Firewall - FVS336GV2 If you see in the attached config downloaded from VPC (#3 Tunnel Interface Configuration), it gives me some "inside" addresses for the tunnel. When setting up the IPsec tunnels do I use the inside tunnel IP's (e.g. 169.254.254.2/30) or do I use my internal network subnet (10.1.1.0/24) I have tried both, when I tried the local network (10.1.1.x) the tracert stops at the router. When I tried with the "inside" ips, the tracert to the amazon VPC (10.0.0.x) goes out over the internet. this all leads me to the next question, for this router, how do I set up stage #4, the static next hop? What are these seemingly random "inside" addresses and where did amazon generate them from? 169.254.254.x seems odd? With a device like this, is the VPN behind the firewall? I have tweaked any IP addresses below so that they are not "real". I am fully aware, this is probably badly worded. Please if there is any further info/screenshots that will help, let me know. Amazon Web Services Virtual Private Cloud IPSec Tunnel #1 ================================================================================ #1: Internet Key Exchange Configuration Configure the IKE SA as follows - Authentication Method : Pre-Shared Key - Pre-Shared Key : --- - Authentication Algorithm : sha1 - Encryption Algorithm : aes-128-cbc - Lifetime : 28800 seconds - Phase 1 Negotiation Mode : main - Perfect Forward Secrecy : Diffie-Hellman Group 2 #2: IPSec Configuration Configure the IPSec SA as follows: - Protocol : esp - Authentication Algorithm : hmac-sha1-96 - Encryption Algorithm : aes-128-cbc - Lifetime : 3600 seconds - Mode : tunnel - Perfect Forward Secrecy : Diffie-Hellman Group 2 IPSec Dead Peer Detection (DPD) will be enabled on the AWS Endpoint. We recommend configuring DPD on your endpoint as follows: - DPD Interval : 10 - DPD Retries : 3 IPSec ESP (Encapsulating Security Payload) inserts additional headers to transmit packets. These headers require additional space, which reduces the amount of space available to transmit application data. To limit the impact of this behavior, we recommend the following configuration on your Customer Gateway: - TCP MSS Adjustment : 1387 bytes - Clear Don't Fragment Bit : enabled - Fragmentation : Before encryption #3: Tunnel Interface Configuration Your Customer Gateway must be configured with a tunnel interface that is associated with the IPSec tunnel. All traffic transmitted to the tunnel interface is encrypted and transmitted to the Virtual Private Gateway. The Customer Gateway and Virtual Private Gateway each have two addresses that relate to this IPSec tunnel. Each contains an outside address, upon which encrypted traffic is exchanged. Each also contain an inside address associated with the tunnel interface. The Customer Gateway outside IP address was provided when the Customer Gateway was created. Changing the IP address requires the creation of a new Customer Gateway. The Customer Gateway inside IP address should be configured on your tunnel interface. Outside IP Addresses: - Customer Gateway : 217.33.22.33 - Virtual Private Gateway : 87.222.33.42 Inside IP Addresses - Customer Gateway : 169.254.254.2/30 - Virtual Private Gateway : 169.254.254.1/30 Configure your tunnel to fragment at the optimal size: - Tunnel interface MTU : 1436 bytes #4: Static Routing Configuration: To route traffic between your internal network and your VPC, you will need a static route added to your router. Static Route Configuration Options: - Next hop : 169.254.254.1 You should add static routes towards your internal network on the VGW. The VGW will then send traffic towards your internal network over the tunnels. IPSec Tunnel #2 ================================================================================ #1: Internet Key Exchange Configuration Configure the IKE SA as follows - Authentication Method : Pre-Shared Key - Pre-Shared Key : --- - Authentication Algorithm : sha1 - Encryption Algorithm : aes-128-cbc - Lifetime : 28800 seconds - Phase 1 Negotiation Mode : main - Perfect Forward Secrecy : Diffie-Hellman Group 2 #2: IPSec Configuration Configure the IPSec SA as follows: - Protocol : esp - Authentication Algorithm : hmac-sha1-96 - Encryption Algorithm : aes-128-cbc - Lifetime : 3600 seconds - Mode : tunnel - Perfect Forward Secrecy : Diffie-Hellman Group 2 IPSec Dead Peer Detection (DPD) will be enabled on the AWS Endpoint. We recommend configuring DPD on your endpoint as follows: - DPD Interval : 10 - DPD Retries : 3 IPSec ESP (Encapsulating Security Payload) inserts additional headers to transmit packets. These headers require additional space, which reduces the amount of space available to transmit application data. To limit the impact of this behavior, we recommend the following configuration on your Customer Gateway: - TCP MSS Adjustment : 1387 bytes - Clear Don't Fragment Bit : enabled - Fragmentation : Before encryption #3: Tunnel Interface Configuration Outside IP Addresses: - Customer Gateway : 217.33.22.33 - Virtual Private Gateway : 87.222.33.46 Inside IP Addresses - Customer Gateway : 169.254.254.6/30 - Virtual Private Gateway : 169.254.254.5/30 Configure your tunnel to fragment at the optimal size: - Tunnel interface MTU : 1436 bytes #4: Static Routing Configuration: Static Route Configuration Options: - Next hop : 169.254.254.5 You should add static routes towards your internal network on the VGW. The VGW will then send traffic towards your internal network over the tunnels. EDIT #1 After writing this post, I continued to fiddle and something started to work, just not very reliably. The local IPs to use when setting up the tunnels where indeed my network subnets. Which further confuses me over what these "inside" IP addresses are for. The problem is, results are not consistent what so ever. I can "sometimes" ping, I can "sometimes" RDP using the VPN. Sometimes, Tunnel 1 or Tunnel 2 can be up or down. When I came back into work today, Tunnel 1 was down, so I deleted it and re-created it from scratch. Now I cant ping anything, but Amazon AND the router are telling me tunnel 1/2 are fine. I guess the router/vpn hardware I have just isnt up to the job..... EDIT #2 Now Tunnel 1 is up, Tunnel 2 is down (I didn't change any settings) and I can ping/rdp again. EDIT #3 Screenshot of route table that the router has built up. Current state (tunnel 1 still up and going string, 2 is still down and wont re-connect)

    Read the article

  • Installing nGinX Reverse Proxy on CentOS 5

    - by heavymark
    I'm trying to install nGinX as a reverse proxy on CentOS 5 with apache. The instructions to do this are here: http://wiki.mediatemple.net/w/(dv):Configure_nginx_as_reverse_proxy_web_server Note- in the instructions, for the url to get nginx I'm using the following: http://nginx.org/download/nginx-1.0.10.tar.gz Now here is my problem. After installing the required packages and running .configure I get the following: checking for OS + Linux 2.6.18-028stab094.3 x86_64 checking for C compiler ... found + using GNU C compiler + gcc version: 4.1.2 20080704 (Red Hat 4.1.2-51) checking for gcc -pipe switch ... found checking for gcc builtin atomic operations ... found checking for C99 variadic macros ... found checking for gcc variadic macros ... found checking for unistd.h ... found checking for inttypes.h ... found checking for limits.h ... found checking for sys/filio.h ... not found checking for sys/param.h ... found checking for sys/mount.h ... found checking for sys/statvfs.h ... found checking for crypt.h ... found checking for Linux specific features checking for epoll ... found checking for sendfile() ... found checking for sendfile64() ... found checking for sys/prctl.h ... found checking for prctl(PR_SET_DUMPABLE) ... found checking for sched_setaffinity() ... found checking for crypt_r() ... found checking for sys/vfs.h ... found checking for nobody group ... found checking for poll() ... found checking for /dev/poll ... not found checking for kqueue ... not found checking for crypt() ... not found checking for crypt() in libcrypt ... found checking for F_READAHEAD ... not found checking for posix_fadvise() ... found checking for O_DIRECT ... found checking for F_NOCACHE ... not found checking for directio() ... not found checking for statfs() ... found checking for statvfs() ... found checking for dlopen() ... not found checking for dlopen() in libdl ... found checking for sched_yield() ... found checking for SO_SETFIB ... not found checking for SO_ACCEPTFILTER ... not found checking for TCP_DEFER_ACCEPT ... found checking for accept4() ... not found checking for int size ... 4 bytes checking for long size ... 8 bytes checking for long long size ... 8 bytes checking for void * size ... 8 bytes checking for uint64_t ... found checking for sig_atomic_t ... found checking for sig_atomic_t size ... 4 bytes checking for socklen_t ... found checking for in_addr_t ... found checking for in_port_t ... found checking for rlim_t ... found checking for uintptr_t ... uintptr_t found checking for system endianess ... little endianess checking for size_t size ... 8 bytes checking for off_t size ... 8 bytes checking for time_t size ... 8 bytes checking for setproctitle() ... not found checking for pread() ... found checking for pwrite() ... found checking for sys_nerr ... found checking for localtime_r() ... found checking for posix_memalign() ... found checking for memalign() ... found checking for mmap(MAP_ANON|MAP_SHARED) ... found checking for mmap("/dev/zero", MAP_SHARED) ... found checking for System V shared memory ... found checking for POSIX semaphores ... not found checking for POSIX semaphores in libpthread ... found checking for struct msghdr.msg_control ... found checking for ioctl(FIONBIO) ... found checking for struct tm.tm_gmtoff ... found checking for struct dirent.d_namlen ... not found checking for struct dirent.d_type ... found checking for PCRE library ... found checking for system md library ... not found checking for system md5 library ... not found checking for OpenSSL md5 crypto library ... found checking for sha1 in system md library ... not found checking for OpenSSL sha1 crypto library ... found checking for zlib library ... found creating objs/Makefile Configuration summary + using system PCRE library + OpenSSL library is not used + md5: using system crypto library + sha1: using system crypto library + using system zlib library nginx path prefix: "/usr/local/nginx" nginx binary file: "/usr/local/nginx/sbin/nginx" nginx configuration prefix: "/usr/local/nginx/conf" nginx configuration file: "/usr/local/nginx/conf/nginx.conf" nginx pid file: "/usr/local/nginx/logs/nginx.pid" nginx error log file: "/usr/local/nginx/logs/error.log" nginx http access log file: "/usr/local/nginx/logs/access.log" nginx http client request body temporary files: "client_body_temp" nginx http proxy temporary files: "proxy_temp" nginx http fastcgi temporary files: "fastcgi_temp" nginx http uwsgi temporary files: "uwsgi_temp" nginx http scgi temporary files: "scgi_temp" It says if you get errors to stop and make sure packages are installed. I didn't get errors but as you can see I got several "not founds". Are those considered errors? If so how do I resolve that. And as noted in the link, I cannot install through yum, because it wont work with plesk then. Thanks!

    Read the article

  • SSH is not working .. Password promt is not coming

    - by Sumanth Lingappa
    I am not able to SSH into my ubuntu server since yesterday. I am not using any keyless or public key method.. Its simple SSH with username and password everytime.. However I can do a VNC session running on my ubuntu server.. But I am afraid that if the vnc session goes out, I wont be having any way to login to the server.. My ssh-vvv output is as below.. sumanth@sumanth:~$ ssh -vvv user@serverIP OpenSSH_6.6.1, OpenSSL 1.0.1f 6 Jan 2014 debug1: Reading configuration data /etc/ssh/ssh_config debug1: /etc/ssh/ssh_config line 19: Applying options for * debug2: ssh_connect: needpriv 0 debug1: Connecting to 172.16.2.156 [172.16.2.156] port 22. debug1: Connection established. debug1: identity file /home/sumanth/.ssh/id_rsa type -1 debug1: identity file /home/sumanth/.ssh/id_rsa-cert type -1 debug1: identity file /home/sumanth/.ssh/id_dsa type -1 debug1: identity file /home/sumanth/.ssh/id_dsa-cert type -1 debug1: identity file /home/sumanth/.ssh/id_ecdsa type -1 debug1: identity file /home/sumanth/.ssh/id_ecdsa-cert type -1 debug1: identity file /home/sumanth/.ssh/id_ed25519 type -1 debug1: identity file /home/sumanth/.ssh/id_ed25519-cert type -1 debug1: Enabling compatibility mode for protocol 2.0 debug1: Local version string SSH-2.0-OpenSSH_6.6.1p1 Ubuntu-2ubuntu2 debug1: Remote protocol version 2.0, remote software version OpenSSH_5.9p1 Debian-5ubuntu1 debug1: match: OpenSSH_5.9p1 Debian-5ubuntu1 pat OpenSSH_5* compat 0x0c000000 debug2: fd 3 setting O_NONBLOCK debug3: load_hostkeys: loading entries for host "172.16.2.156" from file "/home/sumanth/.ssh/known_hosts" debug3: load_hostkeys: found key type ECDSA in file /home/sumanth/.ssh/known_hosts:5 debug3: load_hostkeys: loaded 1 keys debug3: order_hostkeyalgs: prefer hostkeyalgs: [email protected],[email protected],[email protected],ecdsa-sha2-nistp256,ecdsa-sha2-nistp384,ecdsa-sha2-nistp521 debug1: SSH2_MSG_KEXINIT sent debug1: SSH2_MSG_KEXINIT received debug2: kex_parse_kexinit: [email protected],ecdh-sha2-nistp256,ecdh-sha2-nistp384,ecdh-sha2-nistp521,diffie-hellman-group-exchange-sha256,diffie-hellman-group-exchange-sha1,diffie-hellman-group14-sha1,diffie-hellman-group1-sha1 debug2: kex_parse_kexinit: [email protected],[email protected],[email protected],ecdsa-sha2-nistp256,ecdsa-sha2-nistp384,ecdsa-sha2-nistp521,[email protected],[email protected],[email protected],[email protected],[email protected],ssh-ed25519,ssh-rsa,ssh-dss debug2: kex_parse_kexinit: aes128-ctr,aes192-ctr,aes256-ctr,arcfour256,arcfour128,[email protected],[email protected],[email protected],aes128-cbc,3des-cbc,blowfish-cbc,cast128-cbc,aes192-cbc,aes256-cbc,arcfour,[email protected] debug2: kex_parse_kexinit: aes128-ctr,aes192-ctr,aes256-ctr,arcfour256,arcfour128,[email protected],[email protected],[email protected],aes128-cbc,3des-cbc,blowfish-cbc,cast128-cbc,aes192-cbc,aes256-cbc,arcfour,[email protected] debug2: kex_parse_kexinit: [email protected],[email protected],[email protected],[email protected],[email protected],[email protected],[email protected],[email protected],[email protected],hmac-md5,hmac-sha1,[email protected],[email protected],hmac-sha2-256,hmac-sha2-512,hmac-ripemd160,[email protected],hmac-sha1-96,hmac-md5-96 debug2: kex_parse_kexinit: [email protected],[email protected],[email protected],[email protected],[email protected],[email protected],[email protected],[email protected],[email protected],hmac-md5,hmac-sha1,[email protected],[email protected],hmac-sha2-256,hmac-sha2-512,hmac-ripemd160,[email protected],hmac-sha1-96,hmac-md5-96 debug2: kex_parse_kexinit: none,[email protected],zlib debug2: kex_parse_kexinit: none,[email protected],zlib debug2: kex_parse_kexinit: debug2: kex_parse_kexinit: debug2: kex_parse_kexinit: first_kex_follows 0 debug2: kex_parse_kexinit: reserved 0 debug2: kex_parse_kexinit: ecdh-sha2-nistp256,ecdh-sha2-nistp384,ecdh-sha2-nistp521,diffie-hellman-group-exchange-sha256,diffie-hellman-group-exchange-sha1,diffie-hellman-group14-sha1,diffie-hellman-group1-sha1 debug2: kex_parse_kexinit: ssh-rsa,ssh-dss,ecdsa-sha2-nistp256 debug2: kex_parse_kexinit: aes128-ctr,aes192-ctr,aes256-ctr,arcfour256,arcfour128,aes128-cbc,3des-cbc,blowfish-cbc,cast128-cbc,aes192-cbc,aes256-cbc,arcfour,[email protected] debug2: kex_parse_kexinit: aes128-ctr,aes192-ctr,aes256-ctr,arcfour256,arcfour128,aes128-cbc,3des-cbc,blowfish-cbc,cast128-cbc,aes192-cbc,aes256-cbc,arcfour,[email protected] debug2: kex_parse_kexinit: hmac-md5,hmac-sha1,[email protected],hmac-sha2-256,hmac-sha2-256-96,hmac-sha2-512,hmac-sha2-512-96,hmac-ripemd160,[email protected],hmac-sha1-96,hmac-md5-96 debug2: kex_parse_kexinit: hmac-md5,hmac-sha1,[email protected],hmac-sha2-256,hmac-sha2-256-96,hmac-sha2-512,hmac-sha2-512-96,hmac-ripemd160,[email protected],hmac-sha1-96,hmac-md5-96 debug2: kex_parse_kexinit: none,[email protected] debug2: kex_parse_kexinit: none,[email protected] debug2: kex_parse_kexinit: debug2: kex_parse_kexinit: debug2: kex_parse_kexinit: first_kex_follows 0 debug2: kex_parse_kexinit: reserved 0 debug2: mac_setup: setup hmac-md5 debug1: kex: server->client aes128-ctr hmac-md5 none debug2: mac_setup: setup hmac-md5 debug1: kex: client->server aes128-ctr hmac-md5 none debug1: sending SSH2_MSG_KEX_ECDH_INIT debug1: expecting SSH2_MSG_KEX_ECDH_REPLY debug1: Server host key: ECDSA ea:4e:15:52:15:dd:6b:09:d4:36:cb:14:2d:c3:1b:7a debug3: load_hostkeys: loading entries for host "172.16.2.156" from file "/home/sumanth/.ssh/known_hosts" debug3: load_hostkeys: found key type ECDSA in file /home/sumanth/.ssh/known_hosts:5 debug3: load_hostkeys: loaded 1 keys debug1: Host '172.16.2.156' is known and matches the ECDSA host key. debug1: Found key in /home/sumanth/.ssh/known_hosts:5 debug1: ssh_ecdsa_verify: signature correct debug2: kex_derive_keys debug2: set_newkeys: mode 1 debug1: SSH2_MSG_NEWKEYS sent debug1: expecting SSH2_MSG_NEWKEYS debug2: set_newkeys: mode 0 debug1: SSH2_MSG_NEWKEYS received debug1: Roaming not allowed by server debug1: SSH2_MSG_SERVICE_REQUEST sent debug2: service_accept: ssh-userauth debug1: SSH2_MSG_SERVICE_ACCEPT received debug2: key: /home/sumanth/.ssh/id_rsa ((nil)), debug2: key: /home/sumanth/.ssh/id_dsa ((nil)), debug2: key: /home/sumanth/.ssh/id_ecdsa ((nil)), debug2: key: /home/sumanth/.ssh/id_ed25519 ((nil)),

    Read the article

  • Installing Windows on HP Proliant Servers without SmartStart

    - by Fitzroy
    I have a PXE server for deploying Windows XP and Windows 7 to workstations. The process is as follows: Boot the workstation from the NIC. Workstation sends a DHCP request. DHCP server responds with an IP address and the location of the PXE server. Workstation downloads WinPE image file from PXE server via TFTP Workstation stores WinPE image file in memory and executes it. Once booted into WinPE, I connect to a network share to gain access to either the Windows XP or Windows 7 installation files. A custom script is launched to guide you through the process of formatting and partitioning the hard drive(s) (using DISKPART and FORMAT). Another custom script asks for details such as the hostname to assign to the workstation. The answers provided are used to build an unattended answer file (SIF [Setup Information File] for WinXP and XML for Win7). The Windows setup EXE is launched, passing the unattended answer file to it as a parameter. The Windows XP and Windows 7 installation sources have been customised to include the drivers for our Dell workstations. They also run a number of scripts upon first booting up to install software packages. This process works very well for our workstations and I would now like to use it for building our servers too. The vast majority of our servers are HP Proliant DL360 G6, DL380 G5 and DL380 G6. They’re running Windows Server 2003 (various editions) or 2008 (various editions). To date, we have always built the HP Proliant servers using the SmartStart CD provided. SmartStart does three useful things for us: Setup RAID with HP Array Configuration Utility (ACU). Installs and configures SNMP Installs various HP Tools for Windows (HP Array Configuration Utility, HP Array Diagnostic Utility, HP Proliant Integrated Management Log Viewer, etc) Using SmartStart I have never had to manually download and install Windows drivers for network, sound, video, etc. I'm not sure if this is because SmartStart copies drivers from the CD during setup, or whether Windows just has the drivers natively in its driver CAB. If I abandon the SmartStart CD in favour of my PXE server I would have to do the following: As I wont have access to ACU, I'll configure the RAID (before booting to the PXE server) by pressing F8 (during the boot process) to access Option ROM Configuration for Arrays (ORCA). Installation of SNMP and the HP Tools will have to be installed once the Windows installation is complete using the Proliant Support Pack. Is this method OK? Is there anything that the SmartStart CD does that I'll be unable to do by other means? Are there any disadvantages to not using the SmartStart CD? Many thanks. UPDATE 05/01/12 I’ve been reading through the SmartStart Scripting Toolkit documentation. The scripting toolkit contains command line tools which work within WinPE and can such things as configure BIOS settings, configure an array and setup ILO. I’m personally not too bothered about configuring BIOS settings as I rarely deviate from the defaults (unless the server is to be a Hyper-V host). I’m not too fussed about being able to configure the array from within WinPE, as I’m happy to just press F8 and use Option ROM Configuration for Arrays (ORCA). Although, if it’s easy enough to do, I will explore this further, as it saves time if everything can be configured from within WinPE. One of the nice features all the tools possess is that you can pass input files to them. EG. Configure one server to your requirements, capture its configuration to a file (using the appropriate tool), you can then use the tool on other servers passing the input file with the captured configuration. Array controller drivers appear to be included with the toolkit along with example of how to incorporate them within a WinPE build. I suppose WinPE won’t be able to see logical volumes (I.E 2x physical disks in a RAID 1 configuration) without the array controller drivers? I mentioned in my post that SmartStart normally installs a bunch of Windows HP tools for you. I’ve had a look today, and if you run the SmartStart CD from within Windows all the tools can be installed. Therefore I can do this after the Windows installation is complete. The SmartStart CD appears to contain a lot Windows drivers. I can customise my Windows 2008 source to incorporate these drivers. However, I understand that incorporating an array controller driver is a little different to most drivers. I believe that you have to provide the driver during the very early stages of the Windows setup. I’m working through the Scripting Toolkit documentation to try and work this out...

    Read the article

  • fd partitions gone from 2 discs, md happy with it and resyncs. How to recover ?

    - by d0nd
    Hey gurus, need some help badly with this one. I run a server with a 6Tb md raid5 volume built over 7*1Tb disks. I've had to shut down the server lately and when it went back up, 2 out of the 7 disks used for the raid volume had lost its conf : dmesg : [ 10.184167] sda: sda1 sda2 sda3 // System disk [ 10.202072] sdb: sdb1 [ 10.210073] sdc: sdc1 [ 10.222073] sdd: sdd1 [ 10.229330] sde: sde1 [ 10.239449] sdf: sdf1 [ 11.099896] sdg: unknown partition table [ 11.255641] sdh: unknown partition table All 7 disks have same geometry and were configured alike : dmesg : Disk /dev/sdb: 1000.2 GB, 1000204886016 bytes 255 heads, 63 sectors/track, 121601 cylinders Units = cylinders of 16065 * 512 = 8225280 bytes Disk identifier: 0x1e7481a5 Device Boot Start End Blocks Id System /dev/sdb1 1 121601 976760001 fd Linux raid autodetect All 7 disks (sdb1, sdc1, sdd1, sde1, sdf1, sdg1, sdh1) were used in a md raid5 xfs volume. When booting, md, which was (obviously) out of sync kicked in and automatically started rebuilding over the 7 disks, including the two "faulty" ones; xfs tried to do some shenanigans as well: dmesg : [ 19.566941] md: md0 stopped. [ 19.817038] md: bind<sdc1> [ 19.817339] md: bind<sdd1> [ 19.817465] md: bind<sde1> [ 19.817739] md: bind<sdf1> [ 19.817917] md: bind<sdh> [ 19.818079] md: bind<sdg> [ 19.818198] md: bind<sdb1> [ 19.818248] md: md0: raid array is not clean -- starting background reconstruction [ 19.825259] raid5: device sdb1 operational as raid disk 0 [ 19.825261] raid5: device sdg operational as raid disk 6 [ 19.825262] raid5: device sdh operational as raid disk 5 [ 19.825264] raid5: device sdf1 operational as raid disk 4 [ 19.825265] raid5: device sde1 operational as raid disk 3 [ 19.825267] raid5: device sdd1 operational as raid disk 2 [ 19.825268] raid5: device sdc1 operational as raid disk 1 [ 19.825665] raid5: allocated 7334kB for md0 [ 19.825667] raid5: raid level 5 set md0 active with 7 out of 7 devices, algorithm 2 [ 19.825669] RAID5 conf printout: [ 19.825670] --- rd:7 wd:7 [ 19.825671] disk 0, o:1, dev:sdb1 [ 19.825672] disk 1, o:1, dev:sdc1 [ 19.825673] disk 2, o:1, dev:sdd1 [ 19.825675] disk 3, o:1, dev:sde1 [ 19.825676] disk 4, o:1, dev:sdf1 [ 19.825677] disk 5, o:1, dev:sdh [ 19.825679] disk 6, o:1, dev:sdg [ 19.899787] PM: Starting manual resume from disk [ 28.663228] Filesystem "md0": Disabling barriers, not supported by the underlying device [ 28.663228] XFS mounting filesystem md0 [ 28.884433] md: resync of RAID array md0 [ 28.884433] md: minimum _guaranteed_ speed: 1000 KB/sec/disk. [ 28.884433] md: using maximum available idle IO bandwidth (but not more than 200000 KB/sec) for resync. [ 28.884433] md: using 128k window, over a total of 976759936 blocks. [ 29.025980] Starting XFS recovery on filesystem: md0 (logdev: internal) [ 32.680486] XFS: xlog_recover_process_data: bad clientid [ 32.680495] XFS: log mount/recovery failed: error 5 [ 32.682773] XFS: log mount failed I ran fdisk and flagged sdg1 and sdh1 as fd. I tried to reassemble the array but it didnt work: no matter what was in mdadm.conf, it still uses sdg and sdh instead of sdg1 and sdh1. I checked in /dev and I see no sdg1 and and sdh1, shich explains why it wont use it. I just don't know why those partitions are gone from /dev and how to readd those... blkid : /dev/sda1: LABEL="boot" UUID="519790ae-32fe-4c15-a7f6-f1bea8139409" TYPE="ext2" /dev/sda2: TYPE="swap" /dev/sda3: LABEL="root" UUID="91390d23-ed31-4af0-917e-e599457f6155" TYPE="ext3" /dev/sdb1: UUID="2802e68a-dd11-c519-e8af-0d8f4ed72889" TYPE="mdraid" /dev/sdc1: UUID="2802e68a-dd11-c519-e8af-0d8f4ed72889" TYPE="mdraid" /dev/sdd1: UUID="2802e68a-dd11-c519-e8af-0d8f4ed72889" TYPE="mdraid" /dev/sde1: UUID="2802e68a-dd11-c519-e8af-0d8f4ed72889" TYPE="mdraid" /dev/sdf1: UUID="2802e68a-dd11-c519-e8af-0d8f4ed72889" TYPE="mdraid" /dev/sdg: UUID="2802e68a-dd11-c519-e8af-0d8f4ed72889" TYPE="mdraid" /dev/sdh: UUID="2802e68a-dd11-c519-e8af-0d8f4ed72889" TYPE="mdraid" fdisk -l : Disk /dev/sda: 40.0 GB, 40020664320 bytes 255 heads, 63 sectors/track, 4865 cylinders Units = cylinders of 16065 * 512 = 8225280 bytes Disk identifier: 0x8c878c87 Device Boot Start End Blocks Id System /dev/sda1 * 1 12 96358+ 83 Linux /dev/sda2 13 134 979965 82 Linux swap / Solaris /dev/sda3 135 4865 38001757+ 83 Linux Disk /dev/sdb: 1000.2 GB, 1000204886016 bytes 255 heads, 63 sectors/track, 121601 cylinders Units = cylinders of 16065 * 512 = 8225280 bytes Disk identifier: 0x1e7481a5 Device Boot Start End Blocks Id System /dev/sdb1 1 121601 976760001 fd Linux raid autodetect Disk /dev/sdc: 1000.2 GB, 1000204886016 bytes 255 heads, 63 sectors/track, 121601 cylinders Units = cylinders of 16065 * 512 = 8225280 bytes Disk identifier: 0xc9bdc1e9 Device Boot Start End Blocks Id System /dev/sdc1 1 121601 976760001 fd Linux raid autodetect Disk /dev/sdd: 1000.2 GB, 1000204886016 bytes 255 heads, 63 sectors/track, 121601 cylinders Units = cylinders of 16065 * 512 = 8225280 bytes Disk identifier: 0xcc356c30 Device Boot Start End Blocks Id System /dev/sdd1 1 121601 976760001 fd Linux raid autodetect Disk /dev/sde: 1000.2 GB, 1000204886016 bytes 255 heads, 63 sectors/track, 121601 cylinders Units = cylinders of 16065 * 512 = 8225280 bytes Disk identifier: 0xe87f7a3d Device Boot Start End Blocks Id System /dev/sde1 1 121601 976760001 fd Linux raid autodetect Disk /dev/sdf: 1000.2 GB, 1000204886016 bytes 255 heads, 63 sectors/track, 121601 cylinders Units = cylinders of 16065 * 512 = 8225280 bytes Disk identifier: 0xb17a2d22 Device Boot Start End Blocks Id System /dev/sdf1 1 121601 976760001 fd Linux raid autodetect Disk /dev/sdg: 1000.2 GB, 1000204886016 bytes 255 heads, 63 sectors/track, 121601 cylinders Units = cylinders of 16065 * 512 = 8225280 bytes Disk identifier: 0x8f3bce61 Device Boot Start End Blocks Id System /dev/sdg1 1 121601 976760001 fd Linux raid autodetect Disk /dev/sdh: 1000.2 GB, 1000204886016 bytes 255 heads, 63 sectors/track, 121601 cylinders Units = cylinders of 16065 * 512 = 8225280 bytes Disk identifier: 0xa98062ce Device Boot Start End Blocks Id System /dev/sdh1 1 121601 976760001 fd Linux raid autodetect I really dont know what happened nor how to recover from this mess. Needless to say the 5TB or so worth of data sitting on those disks are very valuable to me... Any idea any one? Did anybody ever experienced a similar situation or know how to recover from it ? Can someone help me? I'm really desperate... :x

    Read the article

< Previous Page | 437 438 439 440 441 442 443 444 445 446 447 448  | Next Page >