Search Results

Search found 388 results on 16 pages for 'clarity'.

Page 9/16 | < Previous Page | 5 6 7 8 9 10 11 12 13 14 15 16  | Next Page >

  • Projected Results: Sound project management practices, combined with a complete technology platform, have an immediate and lasting impact on an organization’s bottom line.

    - by Melissa Centurio Lopes
    Normal 0 false false false EN-US X-NONE X-NONE MicrosoftInternetExplorer4 /* Style Definitions */ table.MsoNormalTable {mso-style-name:"Table Normal"; mso-tstyle-rowband-size:0; mso-tstyle-colband-size:0; mso-style-noshow:yes; mso-style-priority:99; mso-style-qformat:yes; mso-style-parent:""; mso-padding-alt:0in 5.4pt 0in 5.4pt; mso-para-margin-top:0in; mso-para-margin-right:0in; mso-para-margin-bottom:10.0pt; mso-para-margin-left:0in; line-height:115%; mso-pagination:widow-orphan; font-size:11.0pt; font-family:"Calibri","sans-serif"; mso-ascii-font-family:Calibri; mso-ascii-theme-font:minor-latin; mso-fareast-font-family:"Times New Roman"; mso-fareast-theme-font:minor-fareast; mso-hansi-font-family:Calibri; mso-hansi-theme-font:minor-latin; mso-bidi-font-family:"Times New Roman"; mso-bidi-theme-font:minor-bidi;} Article By: Alan Joch, is a business and technology writer who specializes in enterprise applications, cloud computing, mobile computing, and the Web. It’s no secret that complex, large-scale projects need close management controls to ensure that they’re delivered on time and on budget. But now there’s growing evidence that failing to meet these goals can have far-reaching consequences, not only for the reputations and value of individual organizations but also for the tenure of their top executives. Government watchdogs forced one large contractor to suspend a multibillion-dollar defense program—and delay payment receipts—until a better management system was launched to more accurately track spending, project milestones, and other fundamental metrics. Significant delays in the opening of the £4.3 billion Terminal 5 at Heathrow Airport impaired an airline’s operations and contributed to a drop in its share prices. These real-world examples are noteworthy because of the huge financial risks they created. They’re also far from being isolated cases. Research by the Economist Intelligence Unit found that only 11 percent of companies claimed they delivered expected ROI on major capital projects 90 percent of the time or more. In addition, 12 percent of respondents said they achieved planned ROI less than half the time. According to Phil Thornton, lead consultant at the analyst firm Clarity Economics, the numbers demonstrate obvious challenges related to managing risks, accurately predicting ROI, and consistently delivering bottom-line growth for major capital investments “Portfolio management is a path to improve your organization’s competitive advantage. It helps make sure your organization is investing in the right things and not spending its time on things that are not delivering the intended results for the firm.” Read the full article here

    Read the article

  • Say goodbye to System.Reflection.Emit (any dynamic proxy generation) in WinRT

    - by mbrit
    tl;dr - Forget any form of dynamic code emitting in Metro-style. It's not going to happen.Over the past week or so I've been trying to get Moq (the popular open source TDD mocking framework) to work on WinRT. Irritatingly, the day before Release Preview was released it was actually working on Consumer Preview. However in Release Preview (RP) the System.Reflection.Emit namespace is gone. Forget any form of dynamic code generation and/or MSIL injection.This kills off any project based on the popular Castle Project Dynamic Proxy component, of which Moq is one example. You can at this point in time not perform any form of mocking using dynamic injection in your Metro-style unit testing endeavours.So let me take you through my journey on this, so that other's don't have to...The headline fact is that you cannot load any assembly that you create at runtime. WinRT supports one Assembly.Load method, and that takes the name of an assembly. That has to be placed within the deployment folder of your app. You cannot give it a filename, or stream. The methods are there, but private. Try to invoke them using Reflection and you'll be met with a caspol exception.You can, in theory, use Rotor to replace SRE. It's all there, but again, you can't load anything you create.You can't write to your deployment folder from within your Metro-style app. But, can you use another service on the machine to move a file that you create into the deployment folder and load it? Not really.The networking stack in Metro-style is intentionally "damaged" to prevent socket communication from Metro-style to any end-point on the local machine. (It just times out.) This militates against an approach where your Metro-style app can signal a properly installed service on the machine to create proxies on its behalf. If you wanted to do this, you'd have to route the calls through a C&C server somewhere. The reason why Microsoft has done this is obvious - taking out SRE know means they don't have to do it in an emergency later. The collateral damage in removing SRE is that you can't do mocking in test mode, but you also can't do any form of injection in production mode. There are plenty of reasons why enterprise apps might want to do this last point particularly. At CP, the assumption was that their inspection tools would prevent SRE being used as a malware vector - it now seems they are less confident about that. (For clarity, the risk here is in allowing a nefarious program to download instructions from a C&C server and make up executable code on the fly to run, getting around the marketplace restrictions.)So, two things:- System.Reflection.Emit is gone in Metro-style/WinRT. Get over it - dynamic, on-the-fly code generation is not going to to happen.- I've more or less got a version of Moq working in Metro-style. This is based on the idea of "baking" the dynamic proxies before you use them. You can find more information here: https://github.com/mbrit/moqrt

    Read the article

  • Why learn Flash Builder 4 (Flex) when I can just use Flash Professional?

    - by Jason McKenna
    I want to learn Flash Builder 4 (Flex) because I see sooo many jobs requesting experience with it. i also just like knowing stuff. I am also very interested in focusing on RIA development now. BUT... can anyone tell me CLEARLY why the heck I would ever use FLEX over Flash Pro?? it is a time investment, so is it worth it? All I read are misguided posts about how Flash Pro is for games and banner ads, and Flex is for programmers and RIAs blah blah... this simply isn't so from my 9 years of contracting experience. I'm 99.9% certain that I can build anything a flex developer can build, but using Flash Pro. I can build powerful AS3-driven apps for the desktop, mobile device, or browser, and I can link to databases with XML and I can import text files and communicate with ColdFusion and everything. The advantage with Flash Pro is that I can also easily and clearly animate transitions and build custom elements that look the way I want/need them to look for my specific client. Why would I want to use a bunch of pre-built components that drive my file sizes to the moon?? Who is happy with a drag-n-drop button?? Is Flex just a thing made for programmer people with no artistic inclination? What is the advantage of using it?? It takes me back to Visual Basic class. Seems like a pain to have to use multiple tools to import crap from Flash Pro into Flex and yada yada... why when I can do it all nicely in Flash Pro to begin with. Am I clueless, or missing some major piece of the puzzle? Thanks for any clarity. PS, I couldn't care less about the code editors. It aint that bad people. They make it out like the thing doesn't even respond to keyboard input or something. Does everthing I need it do anyways. Please help out here. If I just dont need to learn it, I dont want to waste the time. Jase

    Read the article

  • Script For Detecting Availability of XMLHttp in Internet Explorer

    - by Duncan Mills
    Having the XMLHttpRequest API available is key to any ADF Faces Rich Client application. Unfortunately, it is possible for users to switch off this option in Internet Explorer as a Security setting. Without XMLHttpRequest available, your ADF Faces application will simply not work correctly, but rather than giving the user a bad user experience wouldn't it be nicer to tell them that they need to make some changes in order to use the application?  Thanks to Blake Sullivan in the ADF Faces team we now have a little script that can do just this. The script is available from https://samplecode.oracle.com here - The attached file browserCheck.js is what you'll need to add to your project.The best way to use this script is to make changes to whatever template you are using for the entry points to your application. If you're not currently using template then you'll have to make the same change in each of your JSPX pages. Save the browserCheck.js file into a /js/ directory under your HTML root within your UI project (e.g. ViewController)In the template or page, select the <af:document> object in the Structure window. From the right mouse (context) menu choose Facet and select the metaContainer facet.Switch to the source code view and locate the metaContainer facet. Then insert the following lines (I've included the facet tag for clarity but you'll already have that):      <f:facet name="metaContainer">        <af:resource type="javascript"                      source="/js/browserCheck.js"/>        <af:resource type="javascript">           xmlhttpNativeCheck(                     "help/howToConfigureYourBrowser.html");        </af:resource>      </f:facet>Note that the argument to the xmlhttpNativeCheck function is a page that you want to show to the user if they need to change their browser configuration. So build this page in the appropriate place as well. You can also just call the function without any arguments e.g. xmlhttpNativeCheck(); in which case it will pop up default instructions for the user to follow, but not redirect to any other page.

    Read the article

  • Patterns for Handling Changing Property Sets in C++

    - by Bhargav Bhat
    I have a bunch "Property Sets" (which are simple structs containing POD members). I'd like to modify these property sets (eg: add a new member) at run time so that the definition of the property sets can be externalized and the code itself can be re-used with multiple versions/types of property sets with minimal/no changes. For example, a property set could look like this: struct PropSetA { bool activeFlag; int processingCount; /* snip few other such fields*/ }; But instead of setting its definition in stone at compile time, I'd like to create it dynamically at run time. Something like: class PropSet propSetA; propSetA("activeFlag",true); //overloading the function call operator propSetA("processingCount",0); And the code dependent on the property sets (possibly in some other library) will use the data like so: bool actvFlag = propSet["activeFlag"]; if(actvFlag == true) { //Do Stuff } The current implementation behind all of this is as follows: class PropValue { public: // Variant like class for holding multiple data-types // overloaded Conversion operator. Eg: operator bool() { return (baseType == BOOLEAN) ? this->ToBoolean() : false; } // And a method to create PropValues various base datatypes static FromBool(bool baseValue); }; class PropSet { public: // overloaded[] operator for adding properties void operator()(std::string propName, bool propVal) { propMap.insert(std::make_pair(propName, PropVal::FromBool(propVal))); } protected: // the property map std::map<std::string, PropValue> propMap; }; This problem at hand is similar to this question on SO and the current approach (described above) is based on this answer. But as noted over at SO this is more of a hack than a proper solution. The fundamental issues that I have with this approach are as follows: Extending this for supporting new types will require significant code change. At the bare minimum overloaded operators need to be extended to support the new type. Supporting complex properties (eg: struct containing struct) is tricky. Supporting a reference mechanism (needed for an optimization of not duplicating identical property sets) is tricky. This also applies to supporting pointers and multi-dimensional arrays in general. Are there any known patterns for dealing with this scenario? Essentially, I'm looking for the equivalent of the visitor pattern, but for extending class properties rather than methods. Edit: Modified problem statement for clarity and added some more code from current implementation.

    Read the article

  • Why learn Flash Builder 4 (Flex) when I can just use Flash Professional?

    - by Jason McKenna
    I want to learn Flash Builder 4 (Flex) because I see so many jobs requesting experience with it. I also just like knowing stuff. I am also very interested in focusing on RIA development now. BUT... can anyone tell me CLEARLY why the heck I would ever use FLEX over Flash Pro? It is a time investment, so is it worth it? All I read are misguided posts about how Flash Pro is for games and banner ads, and Flex is for programmers and RIAs blah blah... this simply isn't so from my 9 years of contracting experience. I'm 99.9% certain that I can build anything a flex developer can build, but using Flash Pro. I can build powerful AS3-driven apps for the desktop, mobile device, or browser, and I can link to databases with XML and I can import text files and communicate with ColdFusion and everything. The advantage with Flash Pro is that I can also easily and clearly animate transitions and build custom elements that look the way I want/need them to look for my specific client. Why would I want to use a bunch of pre-built components that drive my file sizes to the moon? Who is happy with a drag-n-drop button? Is Flex just a thing made for programmer people with no artistic inclination? What is the advantage of using it? It takes me back to Visual Basic class. Seems like a pain to have to use multiple tools to import crap from Flash Pro into Flex and yada yada... why when I can do it all nicely in Flash Pro to begin with. Am I clueless, or missing some major piece of the puzzle? Thanks for any clarity. PS, I couldn't care less about the code editors. It ain't that bad people. They make it out like the thing doesn't even respond to keyboard input or something. Does everything I need it do anyways. Please help out here. If I just don't need to learn it, I don't want to waste the time.

    Read the article

  • WPF: Reloading app parts to handle persistence as well as memory management.

    - by Ingó Vals
    I created a app using Microsoft's WPF. It mostly handles data reading and input as well as associating relations between data within specific parameters. As a total beginner I made some bad design decision ( not so much decisions as using the first thing I got to work ) but now understanding WPF better I'm getting the urge to refactor my code with better design principles. I had several problems but I guess each deserves it's own question for clarity. Here I'm asking for proper ways to handle the data itself. In the original I wrapped each row in a object when fetched from database ( using LINQ to SQL ) somewhat like Active Record just not active or persistence (each app instance had it's own data handling part). The app has subunits handling different aspects. However as it was setup it loaded everything when started. This creates several problems, for example often it wouldn't be neccesary to load a part unless we were specifically going to work with that part so I wan't some form of lazy loading. Also there was problem with inner persistance because you might create a new object/row in one aspect and perhaps set relation between it and different object but the new object wouldn't appear until the program was restarted. Persistance between instances of the app won't be huge problem because of the small amount of people using the program. While I could solve this now using dirty tricks I would rather refactor the program and do it elegantly, Now the question is how. I know there are several ways and a few come to mind: 1) Each aspect of the program is it's own UserControl that get's reloaded/instanced everytime you navigate to it. This ensures you only load up the data you need and you get some persistancy. DB server located on same LAN and tables are small so that shouldn't be a big problem. Minor drawback is that you would have to remember the state of each aspect so you wouldn't always start at beginners square. 2) Having a ViewModel type object at the base level of the app with lazy loading and some kind of timeout. I would then propegate this object down the visual tree to ensure every aspect is getting it's data from the same instance 3) Semi active record data layer with static load methods. 4) Some other idea What in your opinion is the most practical way in WPF, what does MVVM assume?

    Read the article

  • What are the arguments against parsing the Cthulhu way?

    - by smarmy53
    I have been assigned the task of implementing a Domain Specific Language for a tool that may become quite important for the company. The language is simple but not trivial, it already allows nested loops, string concatenation, etc. and it is practically sure that other constructs will be added as the project advances. I know by experience that writing a lexer/parser by hand -unless the grammar is trivial- is a time consuming and error prone process. So I was left with two options: a parser generator à la yacc or a combinator library like Parsec. The former was good as well but I picked the latter for various reasons, and implemented the solution in a functional language. The result is pretty spectacular to my eyes, the code is very concise, elegant and readable/fluent. I concede it may look a bit weird if you never programmed in anything other than java/c#, but then this would be true of anything not written in java/c#. At some point however, I've been literally attacked by a co-worker. After a quick glance at my screen he declared that the code is uncomprehensible and that I should not reinvent parsing but just use a stack and String.Split like everybody does. He made a lot of noise, and I could not convince him, partially because I've been taken by surprise and had no clear explanation, partially because his opinion was immutable (no pun intended). I even offered to explain him the language, but to no avail. I'm positive the discussion is going to re-surface in front of management, so I'm preparing some solid arguments. These are the first few reasons that come to my mind to avoid a String.Split-based solution: you need lot of ifs to handle special cases and things quickly spiral out of control lots of hardcoded array indexes makes maintenance painful extremely difficult to handle things like a function call as a method argument (ex. add( (add a, b), c) very difficult to provide meaningful error messages in case of syntax errors (very likely to happen) I'm all for simplicity, clarity and avoiding unnecessary smart-cryptic stuff, but I also believe it's a mistake to dumb down every part of the codebase so that even a burger flipper can understand it. It's the same argument I hear for not using interfaces, not adopting separation of concerns, copying-pasting code around, etc. A minimum of technical competence and willingness to learn is required to work on a software project after all. (I won't use this argument as it will probably sound offensive, and starting a war is not going to help anybody) What are your favorite arguments against parsing the Cthulhu way?* *of course if you can convince me he's right I'll be perfectly happy as well

    Read the article

  • SQL SERVER – QUOTED_IDENTIFIER ON/OFF Explanation and Example – Question on Real World Usage

    - by Pinal Dave
    This is a follow up blog post of SQL SERVER – QUOTED_IDENTIFIER ON/OFF and ANSI_NULL ON/OFF Explanation. I wrote that blog six years ago and I had plans that I will write a follow up blog post of the same. Today, when I was going over my to-do list and I was surprised that I had an item there which was six years old and I never got to do that. In the earlier blog post I wrote about exploitation of the Quoted Identifier and ANSI Null. In this blog post we will see a quick example of Quoted Identifier. However, before we continue this blog post, let us see a refresh what both of Quoted Identifider do. QUOTED IDENTIFIER ON/OFF This option specifies the setting for use of double quotes. When this is on, double quotation mark is used as part of the SQL Server identifier (object name). This can be useful in situations in which identifiers are also SQL Server reserved words. In simple words when we have QUOTED IDENTIFIER ON, anything which is wrapped in double quotes becomes an object. E.g. -- The following will work SET QUOTED_IDENTIFIER ON GO CREATE DATABASE "Test1" GO -- The following will throw an error about Incorrect syntax near 'Test2'. SET QUOTED_IDENTIFIER OFF GO CREATE DATABASE "Test2" GO This feature is particularly helpful when we are working with reserved keywords in SQL Server. For example if you have to create a database with the name VARCHAR or INT or DATABASE you may want to put double quotes around your database name and turn on quoted identifiers to create a database with the such name. Personally, I do not think so anybody will ever create a database with the reserve keywords intentionally, as it will just lead to confusion. Here is another example to give you further clarity about how Quoted Idenifier setting works with SELECT statement. -- The following will throw an error about Invalid column name 'Column'. SET QUOTED_IDENTIFIER ON GO SELECT "Column" GO -- The following will work SET QUOTED_IDENTIFIER OFF GO SELECT "Column" GO Personally, I always use the following method to create database as it works irrespective of what is the quoted identifier’s status. It always creates objects with my desire name whenever I would like to create. CREATE DATABASE [Test3] I believe the future of the quoted identifier on or off is useful in the real world when we have script generated from another database where this setting was ON and we have to now execute the same script again in our environment again. Question to you - I personally have never used this feature as I mentioned earlier. I believe this feature is there to support the scripts which are generated in another SQL Database or generate the script for other database. Do you have a real world scenario where we need to turn on or off Quoted Identifiers. Click to Download Scripts Reference: Pinal Dave (http://blog.sqlauthority.com) Filed under: PostADay, SQL, SQL Authority, SQL Query, SQL Server, SQL Tips and Tricks, T SQL, Technology

    Read the article

  • 5 ways to stop code thrashing&hellip;

    - by MarkPearl
    A few days ago I was programming on a personal project and hit a roadblock. I was applying the MVVM pattern and for some reason my view model was not updating the view when the state changed??? I had applied this pattern many times before and had never had this problem. It just didn’t make sense. So what did I do… I did what anyone would have done in my situation and looked to pass the blame to someone or something else. I tried to blame one of the inherited base classes, but it looked fine, then to Visual Studio, but it seemed to be fine and eventually to any random segment of code I came across. My elementary problem had now mushroomed into one that had lost any logical basis and I was in thrashing mode! So what to do when you begin to thrash? 1) Do a general code cleanup – Now there is a difference between cleaning code and changing code . When you thrash you change code and you want to avoid this. What you really want to do is things like rename variables to have better meaning and go over your comments. 2) Do a proof of concept – if cleaning code doesn’t help. The  you want to isolate the problem and identify the key concepts. When you isolate code you ideally want it to be in a totally separate project with as little complexity as possible. Make the building blocks and try and replicate the functionality that you are getting in your current application. 3) Phone a friend – I have found speaking to someone else about the problem generally helps me solve any thrashing issues I am having. Usually they don’t even have to say anything to solve the problem, just you talking them through the problem helps you get clarity of mind. 4) Let the dust settle – Sometimes time is the best solution. I have had a few problems that no matter who I discussed them with and no matter how much code cleaning I had done I just couldn’t seem to fix it. My brain just seemed to be going in circles. A good nights rest has always helped and often just the break away from the problem has helped me find a solution. 5) Stack overflow it – So similar to phone a friend. I am really surprised to see what a melting pot stack overflow has been and what a help it has been in solving technology specific problems. Just be considerate to those using the site and explain clearly exactly what problem you are having and the technologies you are using or else you will probably not get any useful help…

    Read the article

  • JSR 355 Final Release, and moves JCP to version 2.9

    - by heathervc
    JSR 355, JCP EC Merge, passed the JCP EC Final Approval Ballot on 13 August 2012, with 14 Yes votes, 1 abstain (1 member did not vote) on the SE/EE EC, and 12 yes votes (2 members were not eligible to vote) on the ME EC.  JSR 355 posted a Final Release this week, moving the JCP program version to JCP 2.9.  The transition to a merged EC will happen after the 2012 EC Elections, as defined in the Appendix B of the JCP (pasted below), and the EC will operate under the new EC Standing Rules. In the previous version (2.8) of this Process Document there were two separate Executive Committees, one for Java ME and one for Java SE and Java EE combined. The single Executive Committee described in this version of the Process Document will be implemented through the following process: The 2012 annual elections will be held as defined in JCP 2.8, but candidates will be informed that if they are elected their term will be for only a single year, since all candidates must stand for re-election in 2013. Immediately after the 2012 election the two ECs will be merged. Oracle and IBM's second seats will be eliminated, resulting in a single EC with 30 members. All subsequent JSR ballots (even for in-progress JSRs) will then be voted on by the merged EC. For the 2013 annual elections three Ratified and two Elected Seats will be eliminated, thereby reducing the EC to 25 members. All 25 seats will be up for re-election in 2013. Members elected in 2013 will be ranked to determine whether their initial term will be one or two years. The 50% of Ratified and 50% of Elected members who receive the most votes will serve an initial two-year term, while all others will serve an initial one year term. All members elected in 2014 and subsequently will serve a two-year term. For clarity, note that the provisions specified in this version of the Process Document regarding a merged EC will apply to subsequent ballots on all existing JSRs, whether or not the Spec Leads of those JSRs chose to adopt this version of the Process Document in its entirety. <end of Appendix> Also of note:  the materials and minutes from the July EC meeting and the June EC Meeting are now available--following the July EC Meeting, Samsung and SK Telecom lost their EC seats. The June EC meeting also had a public portion--the audio from the public portion of the EC meeting are now posted online.  For Spec Leads there is also the recording of the EG Nominations call.

    Read the article

  • When to use typedef?

    - by futlib
    I'm a bit confused about if and when I should use typedef in C++. I feel it's a balancing act between readability and clarity. Here's a code sample without any typedefs: int sum(std::vector<int>::const_iterator first, std::vector<int>::const_iterator last) { static std::map<std::tuple<std::vector<int>::const_iterator, std::vector<int>::const_iterator>, int> lookup_table; std::map<std::tuple<std::vector<int>::const_iterator, std::vector<int>::const_iterator>, int>::iterator lookup_it = lookup_table.find(lookup_key); if (lookup_it != lookup_table.end()) return lookup_it->second; ... } Pretty ugly IMO. So I'll add some typedefs within the function to make it look nicer: int sum(std::vector<int>::const_iterator first, std::vector<int>::const_iterator last) { typedef std::tuple<std::vector<int>::const_iterator, std::vector<int>::const_iterator> Lookup_key; typedef std::map<Lookup_key, int> Lookup_table; static Lookup_table lookup_table; Lookup_table::iterator lookup_it = lookup_table.find(lookup_key); if (lookup_it != lookup_table.end()) return lookup_it->second; ... } The code is still a bit clumsy, but I get rid of most nightmare material. But there's still the int vector iterators, this variant gets rid of those: typedef std::vector<int>::const_iterator Input_iterator; int sum(Input_iterator first, Input_iterator last) { typedef std::tuple<Input_iterator, Input_iterator> Lookup_key; typedef std::map<Lookup_key, int> Lookup_table; static Lookup_table lookup_table; Lookup_table::iterator lookup_it = lookup_table.find(lookup_key); if (lookup_it != lookup_table.end()) return lookup_it->second; ... } This looks clean, but is it still readable? When should I use a typedef? As soon as I have a nightmare type? As soon as it occurs more than once? Where should I put them? Should I use them in function signatures or keep them to the implementation?

    Read the article

  • Does OO, TDD, and Refactoring to Smaller Functions affect Speed of Code?

    - by Dennis
    In Computer Science field, I have noticed a notable shift in thinking when it comes to programming. The advice as it stands now is write smaller, more testable code refactor existing code into smaller and smaller chunks of code until most of your methods/functions are just a few lines long write functions that only do one thing (which makes them smaller again) This is a change compared to the "old" or "bad" code practices where you have methods spanning 2500 lines, and big classes doing everything. My question is this: when it call comes down to machine code, to 1s and 0s, to assembly instructions, should I be at all concerned that my class-separated code with variety of small-to-tiny functions generates too much extra overhead? While I am not exactly familiar with how OO code and function calls are handled in ASM in the end, I do have some idea. I assume that each extra function call, object call, or include call (in some languages), generate an extra set of instructions, thereby increasing code's volume and adding various overhead, without adding actual "useful" code. I also imagine that good optimizations can be done to ASM before it is actually ran on the hardware, but that optimization can only do so much too. Hence, my question -- how much overhead (in space and speed) does well-separated code (split up across hundreds of files, classes, and methods) actually introduce compared to having "one big method that contains everything", due to this overhead? UPDATE for clarity: I am assuming that adding more and more functions and more and more objects and classes in a code will result in more and more parameter passing between smaller code pieces. It was said somewhere (quote TBD) that up to 70% of all code is made up of ASM's MOV instruction - loading CPU registers with proper variables, not the actual computation being done. In my case, you load up CPU's time with PUSH/POP instructions to provide linkage and parameter passing between various pieces of code. The smaller you make your pieces of code, the more overhead "linkage" is required. I am concerned that this linkage adds to software bloat and slow-down and I am wondering if I should be concerned about this, and how much, if any at all, because current and future generations of programmers who are building software for the next century, will have to live with and consume software built using these practices. UPDATE: Multiple files I am writing new code now that is slowly replacing old code. In particular I've noted that one of the old classes was a ~3000 line file (as mentioned earlier). Now it is becoming a set of 15-20 files located across various directories, including test files and not including PHP framework I am using to bind some things together. More files are coming as well. When it comes to disk I/O, loading multiple files is slower than loading one large file. Of course not all files are loaded, they are loaded as needed, and disk caching and memory caching options exist, and yet still I believe that loading multiple files takes more processing than loading a single file into memory. I am adding that to my concern.

    Read the article

  • Are there software options (preferabbly .NET) for doing distance and speed analysis of footballers moving on video?

    - by Anonymous Type
    Editing Question for Clarity Thanks for feedback so far, very insightful. I'm not sure how far along this part of the software community is, and what if any libraries exist for me to leverage from. Heres what I'm trying to do. Problem: Take an existing video of a game of rugby league. The Rugby League field is 100 metres long, 70 metres wide, and has white line markings every 10 metres running along the width of the field, as well as along the sidelines. Each side has 13 players on the field. Players on each team have identical jerseys that normally constrast strongly against background colours (green/brown field colour) and the referee's colour (usually yellow) and the designated water runner (orange). All players have a unique number in thick white lettering on their backs for identification. Video is taken with a high definition camera. Currently only one camera is used (2D) and existing video does not contain a foreground object of fixed spatial dimensions (as suggested in one answer for comparision measurements, however I could add this to future filming sessions if it is worthwhile). The player's do not run in a straight line 50% of the time but will go sideways on on a diagonal to the play the ball. The distance measured always starts from the spot of the previous "tackle", which ends where the player stops forward movement. It is not always possible to determine the players number from the video (facing other direction, sunlight, others standing in the way of the camera). But this isn't important as the software could allow for manual inputting of unknown "runs" at a later point after analysis. Determine the distance between two points (i.e. where the player started his "run" and where he finished it). I'm guessing that this would be quite doable if I manually marked the start and end point in the video. But how would I use landmarks in the background to determine the distance (assuming the person taking the video has kept it from jerking around). Question: Do software packages or libraries exist that are specialised enough to assist with writing analysis software to determine a sports persons distance travelled based on video taken of the performance?

    Read the article

  • Using GitOAuthPlugin for Jenkins - not working as expected

    - by Blundell
    I need some clarity and maybe a fix. I'm using this plugin to authorise who views our Jenkins ci server: https://wiki.jenkins-ci.org/display/JENKINS/Github+OAuth+Plugin As I understand it anyone who is auth'd to view one of our github project's can also login to our Jenkins box. This works I thought it would also allow the person logging in to only view the Project that they have GitHub permission on. For instance. Three projects on GitHub (A,B,C). Three builds on Jenkins. User 1 has Git access to all 3 projects (A B C). User 2 has Git access to only 1 project (A). When logging into Jenkins: User 1 can see all 3 projects ( this works ) User 2 can only see project A The problem is User 2 can also see all 3 projects when they should only see 1! Have I got this correct, and if so is this a bug? I have the settings set in Jenkins configuration Github Authorization Settings. Here we have some admin users. One organization. And none out of the 4 checkboxes ticked. (User 2, is not an admin, is not part of the org). The plugin is open sourced here: https://github.com/mocleiri/github-oauth-plugin I was trying to get Jenkins to print me the Logs from the plugin but I also failed at viewing these (to see if there was an issue). I followed these instructions: https://wiki.jenkins-ci.org/display/JENKINS/Logging It's the same concept as outlined below but using GitHub rather than manually selecting users: https://wiki.jenkins-ci.org/display/JENKINS/2012/01/03/Allow+access+to+specific+projects+for+Users%28Assigning+security+for+projects+in+Jenkins%29 Have I got this right or wrong? Is it possible to auth a Jenkins user to only see one project?

    Read the article

  • Redundant OpenVPN connections with advanced Linux routing over an unreliable network

    - by konrad
    I am currently living in a country that blocks many websites and has unreliable network connections to the outside world. I have two OpenVPN endpoints (say: vpn1 and vpn2) on Linux servers that I use to circumvent the firewall. I have full access to these servers. This works quite well, except for the high package loss on my VPN connections. This packet loss varies between 1% and 30% depending on time and seems to have a low correlation, most of the time it seems random. I am thinking about setting up a home router (also on Linux) that maintains OpenVPN connections to both endpoints and sends all packets twice, to both endpoints. vpn2 would send all packets from home to vpn1. Return trafic would be send both directly from vpn1 to home, and also through vpn2. +------------+ | home | +------------+ | | | OpenVPN | | links | | | ~~~~~~~~~~~~~~~~~~ unreliable connection | | +----------+ +----------+ | vpn1 |---| vpn2 | +----------+ +----------+ | +------------+ | HTTP proxy | +------------+ | (internet) For clarity: all packets between home and the HTTP proxy will be duplicated and sent over different paths, to increase the chances one of them will arrive. If both arrive, the first second one can be silently discarded. Bandwidth usage is not an issue, both on the home side and endpoint side. vpn1 and vpn2 are close to each other (3ms ping) and have a reliable connection. Any pointers on how this could be achieved using the advanced routing policies available in Linux?

    Read the article

  • Determining a realistic measure of requests per second for a web server

    - by Don
    I'm setting up a nginx stack and optimizing the configuration before going live. Running ab to stress test the machine, I was disappointed to see things topping out at 150 requests per second with a significant number of requests taking 1 second to return. Oddly, the machine itself wasn't even breathing hard. I finally thought to ping the box and saw ping times around 100-125 ms. (The machine, to my surprise, is across the country). So, it seems like network latency is dominating my testing. Running the same tests from a machine on the same network as the server (ping times < 1ms) and I see 5000 requests per second, which is more in-line with what I expected from the machine. But this got me thinking: How do I determine and report a "realistic" measure of requests per second for a web server? You always see claims about performance, but shouldn't network latency be taken into consideration? Sure I can serve 5000 request per second to a machine next to the server, but not to a machine across the country. If I have a lot of slow connections, they will eventually impact my server's performance, right? Or am I thinking about this all wrong? Forgive me if this is network engineering 101 stuff. I'm a developer by trade. Update: Edited for clarity.

    Read the article

  • IP-dependent local port-forwarding on Linux

    - by chronos
    I have configured my server's sshd to listen on a non-standard port 42. However, at work I am behind a firewall/proxy, which only allow outgoing connections to ports 21, 22, 80 and 443. Consequently, I cannot ssh to my server from work, which is bad. I do not want to return sshd to port 22. The idea is this: on my server, locally forward port 22 to port 42 if source IP is matching the external IP of my work's network. For clarity, let us assume that my server's IP is 169.1.1.1 (on eth1), and my work external IP is 169.250.250.250. For all IPs different from 169.250.250.250, my server should respond with an expected 'connection refused', as it does for a non-listening port. I'm very new to iptables. I have briefly looked through the long iptables manual and these related / relevant questions: http://serverfault.com/questions/57872/iptables-question-forwarding-port-x-to-an-ssh-port-of-different-machine-on-the-n http://serverfault.com/questions/140622/how-can-i-port-forward-with-iptables However, those questions deal with more complicated several-host scenarios, and it is not clear to me which tables and chains I should use for local port-forwarding, and if I should have 2 rules (for "question" and "answer" packets), or only 1 rule for "question" packets. So far I have only enabled forwarding via sysctl. I will start testing solutions tomorrow, and will appreciate pointers or maybe case-specific examples for implementing my simple scenario. Is the draft solution below correct? iptables -A INPUT [-m state] [-i eth1] --source 169.250.250.250 -p tcp --destination 169.1.1.1:42 --dport 22 --state NEW,ESTABLISHED,RELATED -j ACCEPT Should I use the mangle table instead of filter? And/or FORWARD chain instead of INPUT?

    Read the article

  • How ZFS handles online replacement in a RAID-Z (theoretical)

    - by Kevin
    This is a somewhat theoretical question about ZFS and RAID-Z. I'll use a three disk single-parity array as an example for clarity, but the problem can be extended to any number of disks and any parity. Suppose we have disks A, B, and C in the pool, and that it is clean. Suppose now that we physically add disk D with the intention of replacing disk C, and that disk C is still functioning correctly and is only being replaced out of preventive maintenance. Some admins might just yank C and install D, which is a little more organized as devices need not change IDs - however this does leave the array degraded temporarily and so for this example suppose we install D without offlining or removing C. Solaris docs indicate that we can replace a disk without first offlining it, using a command such as: zpool replace pool C D This should cause a resilvering onto D. Let us say that resilvering proceeds "downwards" along a "cursor." (I don't know the actual terminology used in the internal implementation.) Suppose now that midways through the resilvering, disk A fails. In theory, this should be recoverable, as above the cursor B and D contain sufficient parity and below the cursor B and C contain sufficient parity. However, whether or not this is actually recoverable depnds upon internal design decisions in ZFS which I am not aware of (and which the manual doesn't say in certain terms). If ZFS continues to send writes to C below the cursor, then we are fine. If, however, ZFS internally treats C as though it were gone, resilvering D only from parity between A and B and only writing A and B below the cursor, then we're toast. Some experimenting could answer this question but I was hoping maybe someone on here already knows which way ZFS handles this situation. Thank you in advance for any insight!

    Read the article

  • Can I connect a Playstation 3's HDMI output to my monitor's DVI-D input? [migrated]

    - by HankJDoomstorm
    I'm attempting to connect my Playstation 3 to my computer monitor. The monitor has a DVI-D (dual link) input, so before distinguishing between the different DVI varieties, I bought a DVI-I (dual link) to HDMI converter that won't fit into the port on the monitor (not only that, there isn't enough physical space in the back of the monitor to fit that much stuff before it hits the bottom of it). So I grabbed a DVI-D (single link) cable and got a female-to-female DVI-I coupler, and plugged the DVI-D cable into the monitor and the whole mess of converters. The end result was HDMI to DVI-D single link, but my monitor isn't receiving a signal on its digital channel. (For clarity's sake: DVI-D DL input on Monitor, DVI-D SL cable, DVI-I DL female-to-female coupler, DVI-I DL to HDMI converter, HDMI output on PS3) I don't know much about this stuff (obviously), but my educated guess is that the bandwidth of the PS3 is too high for the DVI-D Single Link cable, so nothing's getting through. Will replacing the single link cable with dual link resolve this? If not, is it possible at all? Oh, I should mention I'm aware I won't get audio through the monitor. I have an RCA to 3.5mm converter for that.

    Read the article

  • Use subpath internal proxy for subdomains, but redirect external clients if they ask for that subpath?

    - by HostileFork
    I have a VirtualHost that I'd like to have several subdomains on. (For the sake of clarity, let's say my domain is example.com and I'm just trying to get started by making foo.example.com work, and build from there.) The simplest way I found for a subdomain to work non-invasively with the framework I have was to proxy to a sub-path via mod_rewrite. Thus paths would appear in the client's URL bar as http://foo.example.com/(whatever) while they'd actually be served http://foo.example.com/foo/(whatever) under the hood. I've managed to do that inside my VirtualHost config file like this: ServerAlias *.example.com RewriteEngine on RewriteCond %{HTTP_HOST} ^foo\.example\.com [NC] # <--- RewriteCond %{REQUEST_URI} !^/foo/.*$ [NC] # AND is implicit with above RewriteRule ^/(.*)$ /foo/$1 [PT] (Note: It was surprisingly hard to find that particular working combination. Specifically, the [PT] seemed to be necessary on the RewriteRule. I could not get it to work with examples I saw elsewhere like [L] or trying just [P]. It would either not show anything or get in loops. Also some browsers seemed to cache the response pages for the bad loops once they got one... a page reload after fixing it wouldn't show it was working! Feedback welcome—in any case—if this part can be done better.) Now I'd like to make what http://foo.example.com/foo/(whatever) provides depend on who asked. If the request came from outside, I'd like the client to be permanently redirected by Apache so they get the URL http://foo.example.com/(whatever) in their browser. If it came internally from the mod_rewrite, I want the request to be handled by the web framework...which is unaware of subdomains. Is something like that possible?

    Read the article

  • Why isn't my phone charging with some micro usb cables?

    - by Jacxel
    I ordered 3 microUSB cables over ebay. My phone was at about 50% battery and I wanted to use it as a hot spot for some browsing on my laptop, so I plugged it in, a charging icon appeared on the phone, my laptop showed it as a connected usb device and so I went about my business. About 30 minutes later I checked the phone and to my dismay saw 45% battery. But ahh, I thought, I have been putting the poor little thing under too much pressure, acting as a WiFi hotspot must drain the battery quicker than it can charge via usb, perhaps even using my laptops usb port wouldn't output enough power. Unscathed I continued on and when I was going to bed I plugged the usb cable into a mains adapter and switched everything battery consuming off and content, went to sleep. The next morning I was awoken by my phones alarm which got cut off unexpectedly. I attempted to unlock my phone which showed no more signs of life. Why isn't my phone charging with these new USB cables? For clarity: They transfer data with no problems The phone appears to be charging, showing all the signs and lights it normally would, the cable that came with the phone works as you would expect, so its not a fault with the phone, I think they slow the discharging of the phone, but I could be wrong. Are these just bad quality cables? Is there a way to fix this issue?

    Read the article

  • Sending mail results in "Sender address rejected: Domain not found"

    - by user1281413
    The setup: WHM/CPanel CentOS 5 server running Exim and Courier for mail services, and BIND for domain name services. I recently moved servers. The old server was running a HIGHLY similar configuration, and all accounts were ported via WHM. However, the server is unable to send, and sometimes receive email. Errors I am seeing (when I do get an error mail back) state: 450 4.1.8 : Sender address rejected: Domain not found Edit for clarity: this is the error response from remote mail servers. Numerous independent mail servers come back with the same error. (Email address is merely one valid example) My first instinct of course was to check the domain records. However, k-t.org appears to have a valid record (including an MX record), even after running it through domain checks on a completely different server elsewhere and online. Note that the issue appears to happen with all the domains hosted on the server, not just k-t.org I have also ensured that a PTR was created. My Googling has only lead me to people who had fairly basic DNS mistakes, but either I'm blind/dumb (possible, DNS is not my strong suite), or it's something that is a bit more archaic. I've run out of ideas, and I can't seem to find anything that could explain why servers are unable to resolve the domains. There doesn't seem to be anything missing or incorrect.

    Read the article

  • ext3: maximum recommended partition size / handling large partitions

    - by Hansi
    Hi! I would like to do an encrypted install of Ubuntu on a 2 Terabyte drive (i.e., using LUKS/DMcrypt). In order to not have to type in passwords too often, the partitioning scheme will be 50 GB for / and about 1 TB for /home (and the rest for Windows 7), just for clarity. Even though by now LVM is regarded as being stable, I don't want to bother having more room for errors by introducing unnecessary layers of complexity. For both Ubuntu partitions I want encrypted ext3 with the default blocksize of ext3 (4k?). Thoughts: When I look at most partition schemes here on this site or elsewhere, I usually see at most about 400 or 500 GB partitions (maybe I didn't see enough). There may be different reasons for this, but is reliability an issue here? Are larger ext3 partitions, like about 1 TB, harder to handle for the OS or filesystem driver or at some other level? If I make the partition too large, will it be harder to repair in case of corruptions? Are there some default settings for ext3 that I should change for 1 TB partitions? Question: What maximum partition size for ext3 do you recommend and why? Thanks!

    Read the article

  • ImageMagick installation confusion

    - by Codemonkey
    Well this is turning out to be a pain. I'm on CentOS 6.5. I saw on the PECL ImageMagick changelog that they added a load of constants for new filters, such as LANCZOS2SHARP. Some testing I did earlier suggested that PhotoShop 6.5 was able to downsize photos with better clarity than my current ImageMagick's best effort of LANCZOS, so I thought I'd try to upgrade to test out the newer filters. So, first port of call - get source from pecl for 3.2.0RC1. Installed with no problems. But, ah-ha. Although it says it only requires IM 6.2.4, the filters I'm after don't work unless you have 6.6.6+ So I go http://www.imagemagick.org/script/install-source.php to install the newest version. This is the bit that's puzzled me. It all appears to have installed fined. The tests work fine, passing 40 out of 40. If I run identify -version on the command line it outputs 6.8.9 But if I echo Imagick::getVersion() in PHP it shows 6.5.4, even after restarting php-fpm. rpm -qa | grep ImageMagick shows that I still have 6.5.4 installed locate ImageMagic also only seems to show the 6.5.4 one I feel that the missing link here is ImageMagick-devel, do I need to install that too? How do I go about doing that? Or do I just need to reinstall the pecl-imagemagick 3.2.0RC1 now that I have the latest IM installed?

    Read the article

< Previous Page | 5 6 7 8 9 10 11 12 13 14 15 16  | Next Page >