Search Results

Search found 12887 results on 516 pages for 'hard boiled wonderland'.

Page 490/516 | < Previous Page | 486 487 488 489 490 491 492 493 494 495 496 497  | Next Page >

  • Undefined template methods trick ?

    - by Matthieu M.
    A colleague of mine told me about a little piece of design he has used with his team that sent my mind boiling. It's a kind of traits class that they can specialize in an extremely decoupled way. I've had a hard time understanding how it could possibly work, and I am still unsure of the idea I have, so I thought I would ask for help here. We are talking g++ here, specifically the versions 3.4.2 and 4.3.2 (it seems to work with both). The idea is quite simple: 1- Define the interface // interface.h template <class T> struct Interface { void foo(); // the method is not implemented, it could not work if it was }; // // I do not think it is necessary // but they prefer free-standing methods with templates // because of the automatic argument deduction // template <class T> void foo(Interface<T>& interface) { interface.foo(); } 2- Define a class, and in the source file specialize the interface for this class (defining its methods) // special.h class Special {}; // special.cpp #include "interface.h" #include "special.h" // // Note that this specialization is not visible outside of this translation unit // template <> struct Interface<Special> { void foo() { std::cout << "Special" << std::endl; } }; 3- To use, it's simple too: // main.cpp #include "interface.h" class Special; // yes, it only costs a forward declaration // which helps much in term of dependencies int main(int argc, char* argv[]) { Interface<Special> special; foo(special); return 0; }; It's an undefined symbol if no translation unit defined a specialization of Interface for Special. Now, I would have thought this would require the export keyword, which to my knowledge has never been implemented in g++ (and only implemented once in a C++ compiler, with its authors advising anyone not to, given the time and effort it took them). I suspect it's got something to do with the linker resolving the templates methods... Do you have ever met anything like this before ? Does it conform to the standard or do you think it's a fortunate coincidence it works ? I must admit I am quite puzzled by the construct...

    Read the article

  • C++ Template const char array to int

    - by Levi Schuck
    So, I'm wishing to be able to have a static const compile time struct that holds some value based on a string by using templates. I only desire up to four characters. I know that the type of 'abcd' is int, and so is 'ab','abc', and although 'a' is of type char, it works out for a template<int v> struct What I wish to do is take sizes of 2,3,4,5 of some const char, "abcd" and have the same functionality as if they used 'abcd'. Note that I do not mean 1,2,3, or 4 because I expect the null terminator. cout << typeid("abcd").name() << endl; tells me that the type for this hard coded string is char const [5], which includes the null terminator on the end. I understand that I will need to twiddle the values as characters, so they are represented as an integer. I cannot use constexpr since VS10 does not support it (VS11 doesn't either..) So, for example with somewhere this template defined, and later the last line template <int v> struct something { static const int value = v; }; //Eventually in some method cout << typeid(something<'abcd'>::value).name() << endl; works just fine. I've tried template<char v[5]> struct something2 { static const int value = v[0]; } template<char const v[5]> struct something2 { static const int value = v[0]; } template<const char v[5]> struct something2 { static const int value = v[0]; } All of them build individually, though when I throw in my test, cout << typeid(something2<"abcd">::value).name() << endl; I get 'something2' : invalid expression as a template argument for 'v' 'something2' : use of class template requires template argument list Is this not feasible or am I misunderstanding something?

    Read the article

  • How do I add "Press any key to boot from usb" when installing Windows from a flash drive? (Grub4dos question / how to remove a bootloader)

    - by Vincent
    Hi there! I've been struggling with this problem for a while now and finially decided to ask for help. Let me first explain what the main purpose of the app is: to provide the a very easy to use way of backing up files, after which I format the drive and start Windows 7 setup. I do this by booting WinPE, which runs a script to detect Windows installations and then opens a file browser. After the file browser is closed, the script continues and formats the drive that contains the Windows installation, and starts an unattended Windows 7 install. Now here is the problem: When you start Windows setup or WinPE from a dvd, you get a nice option to "Press any key to boot from DVD". This is to prevent the computer from booting the DVD when the first phase of the installation is complete and the computer reboots. However, when booting from a flash drive, Windows does not provide this option: it simply boots the flash drive every reboot. To replicate the "press any key" function, I installed Grub4Dos, which works great. It provides a small menu, the first standard item being "Continue installation", the second being "start installation". After quite a lot of tweaking, I got everything working: Start installation starts WinPE, which in turn starts the Windows installation. At first reboot, the Grub4Dos menu comes up, counts 5 seconds and boots the second stage of the installation. Here, I am greeted with the error: "Windows setup could not configure windows to run on this computer's hardware." When I boot into WinPE the normal way (put the bootmgr on the stick root) and change my bios to boot from the primary hdd after first reboot, I don't get this error. I've been looking around, and the only thing I could find was that the BIOS automatically names the boot device hd0, and that Windows can only be run / installed to hd 0. I'm not sure if this is the problem. I read about remapping to solve this problem, but to do that you have to know the phisical location of the hard drive and partition, like hd(0,1). I want this flash drive to work on any PC, regardless of where the OS is installed, so that's not really a possibility. A possible fix I thought of is removing the bootloader from the flash drive when I'm in WinPE. That way, when the pc reboots the BIOS will not see the flash drive as a boot drive and instead boot the primary hdd. I have yet to find a way to do this. Thank you for reading my question, and if you have any suggestion, please do.

    Read the article

  • make img height 100% of td

    - by kristina childs
    I'm creating an HTML email and since background images can't be used on anything but <body> thought I could get around this by making a border image 100% height within a cell. Perhaps it was wishful thinking? I've searched at the solutions that worked in the past no longer work in modern browsers. Is there any special trick to making this happen without setting a hard height for the cell? Here are the things I've tried so far: <td width="25" style="margin:0; padding:0;"> <img src="http://www.mysite.com/images/side-left.jpg" width="25" height="100%" alt="border" style="margin:0; padding:0; display: block;" /> </td> stretches the image to 100% height of the entire table (even though the table is nested in a <td width="25" height="100%" style="margin:0; padding:0;"> <div style="height:100%; diplay: block;"> <img src="http://www.mysite.com/images/side-left.jpg" width="25" height="100%" alt="border" style="margin:0; padding:0; display: block;" /> </div> </td> ditto <td width="25" height="1" style="margin:0; padding:0;"> <div style="height:100%; diplay: block;"> <img src="http://www.mysite.com/images/side-left.jpg" width="25" height="100%" alt="border" style="margin:0; padding:0; display: block;" /> </div> </td> setting a smaller td size does not force it to strectch as expected. bummer.

    Read the article

  • NVIDIA graphics driver in Ubuntu 12.04

    - by user924501
    So my overall goal is that I want to be able to code with CUDA enabled applications. However, upon many days of searching, using installation walkthroughs, and reinstalling countless times after driver failure... I'm now here as a last resort. I cannot get Ubuntu 12.04 LTS to install the NVIDIA 295.59 driver for my GeForce GT 540M NVIDIA graphics card. My main system specs is as follows... (I believe having the Intel processor may be the problem) DELL Laptop XPS L502X Intel® Core™ i7-2620M CPU @ 2.70GHz × 4 Intel Integrated Graphics 64 bit NVIDIA GeForce GT 540M Ubuntu 12.04 LTS All other specs are irrelevant unless I forgot something? Methods Tried: aptitude install nvidia-current (all packages) Results: Nothing really happened. Nothing in the additional drivers menu appeared, nor was the NVIDIA X Server settings application allowing access because it thought there was no NVIDIA X Server installed. Downloaded driver from nvidia.com. Set nomodeset in the grub boot menu through /boot/grub/grub.cfg Went to console and turned off lightdm. Installed the driver, but it said the pre-install failed? (mean anything?) Started up lightdm again. Results: NVIDIA X Server settings still didn't notice it. Even tried to do nvidia-xconfig multiple times. I also went into the config file to make sure the driver setting said "nvidia". aptitude install nvidia-173 (all packages) Results: Couldn't find the xorg-video-abi-10 virtual package. It was nowhere to be found and the ubuntu forums everywhere had no answers. Lots of people were having this problem. This is easily done in windows, simply download the driver and debug in visual studio with no problems at all. I'd like clear step-by-step instructions on how I should go about this. I'm relatively new to linux but I can find my way around pretty well so you aren't talking to a straight-up beginner. Also, if you think another thread may have the answer please post because I was having a hard time looking for my specific type of problem. TL;DR I want to have access to my GPU so I can code with CUDA while in Ubuntu 12.04 on my 64 bit laptop that also has Intel integrated graphics on the processor. Solution: sudo apt-add-repository ppa:ubuntu-x-swat/x-updates && sudo apt-get update && sudo apt-get upgrade && sudo apt-get install nvidia-current

    Read the article

  • SQL Design Question regarding schema and if Name value pair is the best solution

    - by Aur
    I am having a small problem trying to decide on database schema for a current project. I am by no means a DBA. The application parses through a file based on user input and enters that data in the database. The number of fields that can be parsed is between 1 and 42 at the current moment. The current design of the database is entirely flat with there being 42 columns; some have repeated columns such as address1, address2, address3, etc... This says that I should normalize the data. However, data integrity is not needed at this moment and the way the data is shaped I'm looking at several joins. Not a bad thing but the data is still in a 1 to 1 relationship and I still see a lot of empty fields per row. So my concerns are that this does not allow the database or the application to be very extendable. If they want to add more fields to be parsed (which they do) than I'd need to create another table and add another foreign key to the linking table. The third option is I have a table where the fields are defined and a table for each record. So what I was thinking is to make a table that stores the value and then links to those two tables. The problem is I can picture the size of that table growing large depending on the input size. If someone gives me a file with 300,000 records than 300,000 x 40 = 12 million so I have some reservations. However I think if I get to that point than I should be happy it is being used. This option also allows for more custom displaying of information albeit a bit more work but little rework even if you add more fields. So the problem boils down to: 1. Current design is a flat file which makes extending it hard and it is not normalized. 2. Normalize the tables although no real benefits for the moment but requirements change. 3. Normalize it down into the name value pair and hope size doesn't hurt. There are a large number of inserts, updates, and selects against that table. So performance is a worry but I believe the saying is design now, performance testing later? I'm probably just missing something practical so any comments would be appreciated even if it’s a quick sanity check. Thank you for your time.

    Read the article

  • Finding open contiguous blocks of time for every day of a month, fast

    - by Chris
    I am working on a booking availability system for a group of several venues, and am having a hard time generating the availability of time blocks for days in a given month. This is happening server-side in PHP, but the concept itself is language agnostic -- I could be doing this in JS or anything else. Given a venue_id, month, and year (6/2012 for example), I have a list of all events occurring in that range at that venue, represented as unix timestamps start and end. This data comes from the database. I need to establish what, if any, contiguous block of time of a minimum length (different per venue) exist on each day. For example, on 6/1 I have an event between 2:00pm and 7:00pm. The minimum time is 5 hours, so there's a block open there from 9am - 2pm and another between 7pm and 12pm. This would continue for the 2nd, 3rd, etc... every day of June. Some (most) of the days have nothing happening at all, some have 1 - 3 events. The solution I came up with works, but it also takes waaaay too long to generate the data. Basically, I loop every day of the month and create an array of timestamps for each 15 minutes of that day. Then, I loop the time spans of events from that day by 15 minutes, marking any "taken" timeslot as false. Remaining, I have an array that contains timestamp of free time vs. taken time: //one day's array after processing through loops (not real timestamps) array( 12345678=>12345678, // <--- avail 12345878=>12345878, 12346078=>12346078, 12346278=>false, // <--- not avail 12346478=>false, 12346678=>false, 12346878=>false, 12347078=>12347078, // <--- avail 12347278=>12347278 ) Now I would need to loop THIS array to find continuous time blocks, then check to see if they are long enough (each venue has a minimum), and if so then establish the descriptive text for their start and end (i.e. 9am - 2pm). WHEW! By the time all this looping is done, the user has grown bored and wandered off to Youtube to watch videos of puppies; it takes ages to so examine 30 or so days. Is there a faster way to solve this issue? To summarize the problem, given time ranges t1 and t2 on day d, how can I determine the remaining time left in d that is longer than the minimum time block m. This data is assembled on demand via AJAX as the user moves between calendar months. Results are cached per-page-load, so if the user goes to July a second time, the data that was generated the first time would be reused. Any other details that would help, let me know. Edit Per request, the database structure (or the part that is relevant here) *events* id (bigint) title (varchar) *event_times* id (bigint) event_id (bigint) venue_id (bigint) start (bigint) end (bigint) *venues* id (bigint) name (varchar) min_block (int) min_start (varchar) max_start (varchar)

    Read the article

  • Jquery: Incrimentation for each set of elements in more than 1 div

    - by Jack
    I'm making a Jquery slideshow. It lists thumbnails, that when clicked on, reveal the large image as an overlay. To match up the thumbs with the large images I'm adding attributes to each thumbnail and large image. The attributes contain a number which matches each thumb to its corresponding large image. It works when one slideshow is present on a page. But I want it to work if more than one slideshow is present. Here's the code for adding attributes to thumbs and large images: thumbNo = 0; largeNo = 0; $(this + '.slideshow_content .thumbs img').each(function() { thumbNo++; $(this).attr('thumbimage', thumbNo); $(this).attr("title", "Enter image gallery"); }); $(this + '.slideshow_content .image_container .holder img').each(function() { largeNo++; $(this).addClass('largeImage' + largeNo); }); This works. To make the incrementation work when there are two slideshows on a page, I thought I could stick this code in an each function... $('.slideshow').each(function() { thumbNo = 0; largeNo = 0; $(this + '.slideshow_content .thumbs img').each(function() { thumbNo++; $(this).attr('thumbimage', thumbNo); $(this).attr("title", "Enter image gallery"); }); $(this + '.slideshow_content .image_container .holder img').each(function() { largeNo++; $(this).addClass('largeImage' + largeNo); }); }); The problem with this is that the incrimenting operator does not reset for the second slideshow div (.slideshow), so I end up with thumbs in the first .slideshow div being numbered 1,2,3 etc.. and thumbs in the second .slideshow div being numbered 4,5,6 etc. How do I make the numbers in the second .slideshow div reset and start from 1 again? This is really hard to explain concisely. I hope someone gets the gist.

    Read the article

  • Unusual Template Behavior with XSL

    - by bobber205
    Experiencing some very odd behavior with, what should be, a very simple use of XSL and XSLT. Here's a code sample. <xsl:template match="check"> <div class="check"> <xsl:apply-templates mode="check"> <xsl:with-param name="checkName">testVariable</xsl:with-param> </xsl:apple-templates> </div> </xsl:template> The template called above <xsl:template match="option" mode="check"> <xsl:param name="checkName" /> <div class="option"> <input type="checkbox"> </input> <label> testText </label> </div> </xsl:template> Pretty simple right? It should, for each instance of a instance in the XML create a checkbox in a with a hard coded label. However, what I'm getting is <div class="check"></div> <div class="option>Checkbox stuff here</div> <div class="option>Checkbox stuff here</div> <div class="option>Checkbox stuff here</div> <div class="option>Checkbox stuff here</div> <div class="check"></div> <div class="option>Checkbox stuff here</div> <div class="option>Checkbox stuff here</div> <div class="option>Checkbox stuff here</div> <div class="option>Checkbox stuff here</div> Here's some sample XML <check><option key="1"/><option key="0"/><option key="0"/><option key="0"/><option key="0"/></check> Anyone know what's going on? :D

    Read the article

  • Complex multiple join query across 3 tables

    - by Keir Simmons
    I have 3 tables: shops, PRIMARY KEY cid,zbid shop_items, PRIMARY KEY id shop_inventory, PRIMARY KEY id shops a is related to shop_items b by the following: a.cid=b.cid AND a.zbid=b.szbid shops is not directly related to shop_inventory shop_items b is related to shop_inventory c by the following: b.cid=c.cid AND b.id=c.iid Now, I would like to run a query which returns a.* (all columns from shops). That would be: SELECT a.* FROM shops a WHERE a.cid=1 AND a.zbid!=0 Note that the WHERE clause is necessary. Next, I want to return the number of items in each shop: SELECT a.*, COUNT(b.id) items FROM shops a LEFT JOIN shop_items b ON b.cid=a.cid AND b.szbid=a.zbid WHERE a.cid=1 GROUP BY b.szbid,b.cid As you can see, I have added a GROUP BY clause for this to work. Next, I want to return the average price of each item in the shop. This isn't too hard: SELECT a.*, COUNT(b.id) items, AVG(COALESCE(b.price,0)) average_price FROM shops a LEFT JOIN shop_items b ON b.cid=a.cid AND b.szbid=a.zbid WHERE a.cid=1 GROUP BY b.szbid,b.cid My next criteria is where it gets complicated. I also want to return the unique buyers for each shop. This can be done by querying shop_inventory c, getting the COUNT(DISTINCT c.zbid). Now remember how these tables are related; this should only be done for the rows in c which relate to an item in b which is owned by the respective shop, a. I tried doing the following: SELECT a.*, COUNT(b.id) items, AVG(COALESCE(b.price,0)) average_price, COUNT(DISTINCT c.zbid) FROM shops a LEFT JOIN shop_items b ON b.cid=a.cid AND b.szbid=a.zbid LEFT JOIN shop_inventory c ON c.cid=b.cid AND c.iid=b.id WHERE a.cid=1 GROUP BY b.szbid,b.cid However, this did not work as it messed up the items value. What is the proper way to achieve this result? I also want to be able to return the total number of purchases made in each shop. This would be done by looking at shop_inventory c and adding up the c.quantity value for each shop. How would I add that in as well?

    Read the article

  • Inventory count in CakePHP

    - by metrobalderas
    We are developing an inventory tracking system. Basically we've got an order table in which orders are placed. When an order is payed, the status changes from 0 to 1. This table has multiple children in another table order_items. This is the main structure. CREATE TABLE order( id INT UNSIGNED AUTO_INCREMENT PRIMARY KEY, user_id INT UNSIGNED, status INT(1), total INT UNSIGNED ); CREATE TABLE order_items( id INT UNSIGNED AUTO_INCREMENT PRIMARY KEY, order_id INT UNSIGNED, article_id INT UNSIGNED, size enum('s', 'm', 'l', 'xl'), quantity INT UNSIGNED ); Now, we've got a stocks table with similar architecture for the acquisitions. This is the structure. CREATE TABLE stock( id INT UNSIGNED AUTO_INCREMENT PRIMARY KEY, article_id INT UNSIGNED ); CREATE TABLE stock_items( id INT UNSIGNED AUTO_INCREMENT PRIMARY KEY, stock_id INT UNSIGNED, size enum('s', 'm', 'l', 'xl'), quantity INT(2) ); The main difference is that stocks has no status field. What we are looking for is a way to sum each article size from stock_items, then sum each article size from order_items where Order.status = 1 and substract both these items to find our current inventory. This is the table we want to get from a single query: Size | Stocks | Sales | Available s | 10 | 3 | 7 m | 15 | 13 | 2 l | 7 | 4 | 3 Initially we thought abouth using complex find conditions, but perhaps that's the wrong approach. Also, since it's not a direct join, it turns out to be quite hard. This is the code we have to retrieve the stock's total for each item. function stocks_total($id){ $find = $this->StockItem->find('all', array( 'conditions' => array( 'StockItem.stock_id' => $this->find('list', array('conditions' => array('Stock.article_id' => $id))) ), 'fields' => array_merge( array( 'SUM(StockItem.cantidad) as total' ), array_keys($this->StockItem->_schema) ), 'group' => 'StockItem.size', 'order' => 'FIELD(StockItem.size, \'s\', \'m\' ,\'l\' ,\'xl\') ASC' )); return $find; } Thanks.

    Read the article

  • jquery animation problem using stop

    - by Flanders
    Hi! When running a Jquery animation like slideDown(), it looks like a number of element-specific css properties is set to be updated at a specific interval and when the animation is complete these properties are unset and the display property is simply set to auto or whatever. At least in firebug you can't see those temporary properties any more. The problem I've encountered is the scenario where we stop the slide down with stop(). The element is then left with the current temporary css values. Which is fine because it has to, but let us say that I stoped the slidedown because I have decided to slide it back up again a bit prematurely. It would look something like this: $(this).slideDown(2000) //The below events is not in queue but will rather start execute almost simultaneously as the above line. (dont remember the exact syntax) $(this).delay(1000).stop().slideUp(2000) The above code might not make much sense, but the point is: After 1 second of sliding down the animation is stopped and it starts to slide back up. Works like a charm. BUT!!! And here is the problem. Once it it has slid back up the elements css properties are reset to the exact values it had 1000ms into the slideDown() animation (when stop() was called). If we now try to run the following: $(this).slideDown(2000) It will slide down to the very point the prior slideDown was aborted and not further at half the speed (since it uses the same time for approximately half the height). This is because the css properties were saved as I see it. But it is not especially wished for. Of course I want it to slide all the way down this time. Due to UI interaction that is hard to predict everything might soon break. The longer animations we use increases the risk of something like this happening. Is this to be considered a bug, or am I doing something wrong? Or maybe it's just a feature that is not supported? I guess I can use a callback function to reset the css properties, but depending on the animation used, different css properties are used to render it, and covering your back would result in quite a not-so-fancy solution.

    Read the article

  • pdo connection scope

    - by Scarface
    Hey guys I have a connection class I found for pdo. I am calling the connection method on the page that the file is included on. The problem is that within functions the $conn variable is not defined even though I stated the method was public (bare with me I am very new to OOP), and I was wondering if anyone had an elegant solution other then using global in every function. Any suggestions are greatly appreciated. CONNECTION class PDOConnectionFactory{ // receives the connection public $con = null; // swich database? public $dbType = "mysql"; // connection parameters // when it will not be necessary leaves blank only with the double quotations marks "" public $host = "localhost"; public $user = "user"; public $senha = "password"; public $db = "database"; // arrow the persistence of the connection public $persistent = false; // new PDOConnectionFactory( true ) <--- persistent connection // new PDOConnectionFactory() <--- no persistent connection public function PDOConnectionFactory( $persistent=false ){ // it verifies the persistence of the connection if( $persistent != false){ $this->persistent = true; } } public function getConnection(){ try{ // it carries through the connection $this->con = new PDO($this->dbType.":host=".$this->host.";dbname=".$this->db, $this->user, $this->senha, array( PDO::ATTR_PERSISTENT => $this->persistent ) ); // carried through successfully, it returns connected return $this->con; // in case that an error occurs, it returns the error; }catch ( PDOException $ex ){ echo "We are currently experiencing technical difficulties. We have a bunch of monkies working really hard to fix the problem. Check back soon: ".$ex->getMessage(); } } // close connection public function Close(){ if( $this->con != null ) $this->con = null; } } PAGE USED ON include("includes/connection.php"); $db = new PDOConnectionFactory(); $conn = $db->getConnection(); function test(){ try{ $sql = 'SELECT * FROM topic'; $stmt = $conn->prepare($sql); $result=$stmt->execute(); } catch(PDOException $e){ echo $e->getMessage(); } } test();

    Read the article

  • Why doesen't the number 2 work in this for-loop?

    - by Emil
    Hello. I have a function that runs trough each element in an array. It's hard to explain, so I'll just paste in the code here: NSLog(@"%@", arraySub); for (NSString *string in arrayFav){ int favoriteLoop = [string intValue] + favCount; NSLog(@"%d", favoriteLoop); id arrayFavObject = [array objectAtIndex:favoriteLoop]; [arrayFavObject retain]; [array removeObjectAtIndex:favoriteLoop]; [array insertObject:arrayFavObject atIndex:0]; [arrayFavObject release]; id arraySubFavObject = [arraySub objectAtIndex:favoriteLoop]; [arraySubFavObject retain]; [arraySub removeObjectAtIndex:favoriteLoop]; [arraySub insertObject:arraySubFavObject atIndex:0]; [arraySubFavObject release]; id arrayLengthFavObject = [arrayLength objectAtIndex:favoriteLoop]; [arrayLengthFavObject retain]; [arrayLength removeObjectAtIndex:favoriteLoop]; [arrayLength insertObject:arrayLengthFavObject atIndex:0]; [arrayLengthFavObject release]; } NSLog(@"%@", arraySub); The array arrayFav contains these strings: "3", "8", "2", "10", "40". Array array contains 92 strings with a name. Array arraySub contains numbers 0 to 91, representing a filename with a title from the array array. Array arrayLength contains 92 strings representing the size of each file from array arraySub. Now, the first NSLog shows, as expected, the numbers 0 to 91. The NSLog-s in the loop shows the numbers 3, 8, 2, 10, 40, also as expected. But here's the odd part: the last NSLog shows these numbers: 40, 10, 0, 8, 3, 1, 2, 4, 5, 6, 7, 9, 11, 12, 13, 14, 15, 16, 17, 18, 19, 20, 21, 22, 23, 24, 25, 26, 27, 28, 29, 30, 31, 32, 33, 34, 35, 36, 37, 38, 39, 41, 42, 43, 44, 45, 46, 47, 48, 49, 50, 51, 52, 53, 54, 55, 56, 57, 58, 59, 60, 61, 62, 63, 64, 65, 66, 67, 68, 69, 70, 71, 72, 73, 74, 75, 76, 77, 78, 79, 80, 81, 82, 83, 84, 85, 86, 87, 88, 89, 90, 91 that is 40, 10, 0, 8, 3, and so on. It was not supposed to be a zero in there, it was supposed to be a 2.. Do you have any idea at why this is happening or a way to fix it? Thank you.

    Read the article

  • JavaScript Coding for Finding Shipping Total

    - by user2913279
    I am having a very hard time with this code. I have been working on it for days and cannot seem to figure it out. Please help!! Here are the specific I need for the code: Many companies normally charge a shipping and handling charge for purchases. Create a Web page that allows a user to enter a purchase price into a text box and includes a JavaScript function that calculates shipping and handling. Add functionality to the script that adds a minimum shipping and handling charge of $1.50 for any purchase that is less than or equal to $25.00. For any orders over $25.00, add 10% to the total purchase price for shipping and handling, but do not include the $1.50 minimum shipping and handling charge. The formula for calculating a percentage is price * percent / 100. For example, the formula for calculating 10% of a $50.00 purchase price is 50 * 10 / 100, which results in a shipping and handling charge of $5.00. After you determine the total cost of the order (purchase plus shipping and handling), display it in an alert dialog box. Here is the code I have: <!DOCTYPE> <head> <title>Calculate Shipping</title> <script type="text/javascript"> function parseInt() { var salesPrice = document.salesForm.Price.value; var minCharge = salesPrice + 1.50; var shipping = salesPrice * 10/100; if (salesPrice <= 25) window.alert('Your sales total including shipping is $' + minCharge); else window.alert('Your sales total including shipping is $' + salesPrice + shipping); } </script> </head> <body> <form name="salesForm"> <div > <p>Enter Your Purchase Price</p> <input type="text" name="Price" /><br /><br /> <input type="button" name="Calculate" value="Calculate Shipping" onclick="parseInt ()" /> </div> </form> </body> </html> Everything works except for the math in the alert box. It will show an incorrect total...

    Read the article

  • capistrano initial deployment

    - by Richard G
    I'm trying to set up Capistrano to deploy to an AWS box. This is the first time I've tried to set this up, so please bear with me. Could someone take a look at this and let me know if you can solve this error? The output below is the deploy.rb file, and it's output when it runs. set :application, "apparel1" set :repository, "git://github.com/rgilling/GroceryRun.git" set :scm, :git set :user, "ubuntu" set :scm_passphrase, "pre5ence" # Or: `accurev`, `bzr`, `cvs`, `darcs`, `git`, `mercurial`, `perforce`, `subversion` or `none` ssh_options[:keys] = ["/Users/rgilling/Documents/Projects/Apparel1/abesakey.pem"] ssh_options[:forward_agent] = true set :location, "ec2-107-22-27-42.compute-1.amazonaws.com" role :web, location # Your HTTP server, Apache/etc role :app, location # This may be the same as your `Web` server role :db, location, :primary => true # This is where Rails migrations will run set :deploy_to, "/var/www/#{application}" set :deploy_via, :remote_cache set :use_sudo, true # if you want to clean up old releases on each deploy uncomment this: # after "deploy:restart", "deploy:cleanup" # if you're still using the script/reaper helper you will need # these http://github.com/rails/irs_process_scripts # If you are using Passenger mod_rails uncomment this: namespace :deploy do task :start do ; end task :stop do ; end task :restart, :roles => :app, :except => { :no_release => true } do run "#{try_sudo} touch #{File.join(current_path,'tmp','restart.txt')}" end end Then the execution results in this permission error. I think I"ve set up the SSH etc. correctly... updating the cached checkout on all servers executing locally: "git ls-remote git://github.com/rgilling/GroceryRun.git HEAD" command finished in 1294ms * executing "if [ -d /var/www/apparel1/shared/cached-copy ]; then cd /var/www/apparel1/shared/cached-copy && git fetch -q origin && git fetch --tags -q origin && git reset -q --hard f35dc5868b52649eea86816d536d5db8c915856e && git clean -q -d -x -f; else git clone -q git://github.com/rgilling/GroceryRun.git /var/www/apparel1/shared/cached-copy && cd /var/www/apparel1/shared/cached-copy && git checkout -q -b deploy f35dc5868b52649eea86816d536d5db8c915856e; fi" servers: ["ec2-107-22-27-42.compute-1.amazonaws.com"] [ec2-107-22-27-42.compute-1.amazonaws.com] executing command ** **[ec2-107-22-27-42.compute-1.amazonaws.com :: err] error: cannot open .git/FETCH_HEAD: Permission denied**

    Read the article

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

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

    Read the article

  • Figuring out QuadCurveTo's parameters

    - by Fev
    Could you guys help me figuring out QuadCurveTo's 4 parameters , I tried to find information on http://docs.oracle.com/javafx/2/api/javafx/scene/shape/QuadCurveTo.html, but it's hard for me to understand without picture , I search on google about 'Quadratic Bezier' but it shows me more than 2 coordinates, I'm confused and blind now. I know those 4 parameters draw 2 lines to control the path , but how we know/count exactly which coordinates the object will throught by only knowing those 2 path-controller. Are there some formulas? import javafx.animation.PathTransition; import javafx.animation.PathTransition.OrientationType; import javafx.application.Application; import static javafx.application.Application.launch; import javafx.scene.Group; import javafx.scene.Scene; import javafx.scene.paint.Color; import javafx.scene.shape.MoveTo; import javafx.scene.shape.Path; import javafx.scene.shape.QuadCurveTo; import javafx.scene.shape.Rectangle; import javafx.stage.Stage; import javafx.util.Duration; public class _6 extends Application { public Rectangle r; @Override public void start(final Stage stage) { r = new Rectangle(50, 80, 80, 90); r.setFill(javafx.scene.paint.Color.ORANGE); r.setStrokeWidth(5); r.setStroke(Color.ANTIQUEWHITE); Path path = new Path(); path.getElements().add(new MoveTo(100.0f, 400.0f)); path.getElements().add(new QuadCurveTo(150.0f, 60.0f, 100.0f, 20.0f)); PathTransition pt = new PathTransition(Duration.millis(1000), path); pt.setDuration(Duration.millis(10000)); pt.setNode(r); pt.setPath(path); pt.setOrientation(OrientationType.ORTHOGONAL_TO_TANGENT); pt.setCycleCount(4000); pt.setAutoReverse(true); pt.play(); stage.setScene(new Scene(new Group(r), 500, 700)); stage.show(); } public static void main(String[] args) { launch(args); } } You can find those coordinates on this new QuadCurveTo(150.0f, 60.0f, 100.0f, 20.0f) line, and below is the picture of Quadratic Bezier

    Read the article

  • 'Good' programming form in maintaining / updating / accessing files by entry

    - by zhermes
    Basic Question: If I'm storying/modifying data, should I access elements of a file by index hard-coded index, i.e. targetFile.getElement(5); via a hardcoded identifier (internally translated into index), i.e. target.getElementWithID("Desired Element"), or with some intermediate DESIRED_ELEMENT = 5; ... target.getElement(DESIRED_ELEMENT), etc. Background: My program (c++) stores data in lots of different 'dataFile's. I also keep a list of all of the data-files in another file---a 'listFile'---which also stores some of each one's properties (see below, but i.e. what it's name is, how many lines of information it has etc.). There is an object which manages the data files and the list file, call it a 'fileKeeper'. The entries of a listFile look something like: filename , contents name , number of lines , some more numbers ... Its definitely possible that I may add / remove fields from this list --- but in general, they'll stay static. Right now, I have a constant string array which holds the identification of each element in each entry, something like: const string fileKeeper::idKeys[] = { "FileName" , "Contents" , "NumLines" ... }; const int fileKeeper::idKeysNum = 6; // 6 - for example I'm trying to manage this stuff in 'good' programatic form. Thus, when I want to retrieve the number of lines in a file (for example), instead of having a method which just retrieves the '3'rd element... Instead I do something like: string desiredID = "NumLines"; int desiredIndex = indexForID(desiredID); string desiredElement = elementForIndex(desiredIndex); where the function indexForID() goes through the entries of idKeys until it finds desiredID then returns the index it corresponds to. And elementForIndex(index) actually goes into the listFile to retrieve the index'th element of the comma-delimited string. Problem: This still seems pretty ugly / poor-form. Is there a way I should be doing this? If not, what are some general ways in which this is usually done? Thanks!

    Read the article

  • Are python list comprehensions always a good programming practice?

    - by dln385
    To make the question clear, I'll use a specific example. I have a list of college courses, and each course has a few fields (all of which are strings). The user gives me a string of search terms, and I return a list of courses that match all of the search terms. This can be done in a single list comprehension or a few nested for loops. Here's the implementation. First, the Course class: class Course: def __init__(self, date, title, instructor, ID, description, instructorDescription, *args): self.date = date self.title = title self.instructor = instructor self.ID = ID self.description = description self.instructorDescription = instructorDescription self.misc = args Every field is a string, except misc, which is a list of strings. Here's the search as a single list comprehension. courses is the list of courses, and query is the string of search terms, for example "history project". def searchCourses(courses, query): terms = query.lower().strip().split() return tuple(course for course in courses if all( term in course.date.lower() or term in course.title.lower() or term in course.instructor.lower() or term in course.ID.lower() or term in course.description.lower() or term in course.instructorDescription.lower() or any(term in item.lower() for item in course.misc) for term in terms)) You'll notice that a complex list comprehension is difficult to read. I implemented the same logic as nested for loops, and created this alternative: def searchCourses2(courses, query): terms = query.lower().strip().split() results = [] for course in courses: for term in terms: if (term in course.date.lower() or term in course.title.lower() or term in course.instructor.lower() or term in course.ID.lower() or term in course.description.lower() or term in course.instructorDescription.lower()): break for item in course.misc: if term in item.lower(): break else: continue break else: continue results.append(course) return tuple(results) That logic can be hard to follow too. I have verified that both methods return the correct results. Both methods are nearly equivalent in speed, except in some cases. I ran some tests with timeit, and found that the former is three times faster when the user searches for multiple uncommon terms, while the latter is three times faster when the user searches for multiple common terms. Still, this is not a big enough difference to make me worry. So my question is this: which is better? Are list comprehensions always the way to go, or should complicated statements be handled with nested for loops? Or is there a better solution altogether?

    Read the article

  • How to pass multiple PHP variables to a jQuery function?

    - by jcpeden
    I'm working on a Wordpress plugin. I need to pass plugin directories (that can change depending on an individual's installation) to a jquery function. What is the best way of doing this? The version of the plugin that I can to work on had included all the javascript in the PHP file so the functions were parsed along with the rest of the content before being rendered in a browser. I'm looking at AJAX but I think it might be more complicated than I need. I can get away with just two variables in this case (directories, nothing set by the user). As I've read its good practice, I'm trying to keep the js and php separate. When the plugin initializes, it call the js file: //Wordpress calls the .js when the plugin loads wp_enqueue_script( 'wp-backitup-funtions', plugin_dir_url( __FILE__ ) . 'js/wp-backitup.js', array( 'jquery' ) ); Then I'm in the .js file and need to figure out how to generate the following variables: dir = '<?php echo content_url() ."/plugins"; ?>'; dir = '<?php echo content_url() ."/themes"; ?>'; dir = '<?php echo content_url() ."/uploads"; ?>'; And run the parse the following requests: xmlhttp.open("POST","<?php echo plugins_url() .'/wp-backitup/includes/wp-backitup-restore.php'); ?>",true); xmlhttp.open("POST","<?php echo plugins_url() .'/wp-backitup/includes/wp-backitup-start.php'); ?>",true); xmlhttp.open("POST","<?php echo plugins_url() .'/wp-backitup/wp-backitup-directory.php'); ?>",true); xmlhttp.open("POST","<?php echo plugins_url() .'/wp-backitup/wp-backitup-db.php'); ?>",true); window.location = "<?php echo plugins_url() .'/wp-backitup/backitup-project.zip'); ?>"; xmlhttp.open("POST","<?php echo plugins_url() .'/wp-backitup/wp-backitup-delete.php'); ?>",true); Content URL and Plugins URL differ only by /plugins/ so if I was hard pressed, I would only really need to make a single PHP request and then bring this into the JS.

    Read the article

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

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

    Read the article

  • Indexing on only part of a field in MongoDB

    - by Rob Hoare
    Is there a way to create an index on only part of a field in MongoDB, for example on the first 10 characters? I couldn't find it documented (or asked about on here). The MySQL equivalent would be CREATE INDEX part_of_name ON customer (name(10));. Reason: I have a collection with a single field that varies in length from a few characters up to over 1000 characters, average 50 characters. As there are a hundred million or so documents it's going to be hard to fit the full index in memory (testing with 8% of the data the index is already 400MB, according to stats). Indexing just the first part of the field would reduce the index size by about 75%. In most cases the search term is quite short, it's not a full-text search. A work-around would be to add a second field of 10 (lowercased) characters for each item, index that, then add logic to filter the results if the search term is over ten characters (and that extra field is probably needed anyway for case-insensitive searches, unless anybody has a better way). Seems like an ugly way to do it though. [added later] I tried adding the second field, containing the first 12 characters from the main field, lowercased. It wasn't a big success. Previously, the average object size was 50 bytes, but I forgot that includes the _id and other overheads, so my main field length (there was only one) averaged nearer to 30 bytes than 50. Then, the second field index contains the _id and other overheads. Net result (for my 8% sample) is the index on the main field is 415MB and on the 12 byte field is 330MB - only a 20% saving in space, not worthwhile. I could duplicate the entire field (to work around the case insensitive search problem) but realistically it looks like I should reconsider whether MongoDB is the right tool for the job (or just buy more memory and use twice as much disk space). [added even later] This is a typical document, with the source field, and the short lowercased field: { "_id" : ObjectId("505d0e89f56588f20f000041"), "q" : "Continental Airlines", "f" : "continental " } Indexes: db.test.ensureIndex({q:1}); db.test.ensureIndex({f:1}); The 'f" index, working on a shorter field, is 80% of the size of the "q" index. I didn't mean to imply I included the _id in the index, just that it needs to use that somewhere to show where the index will point to, so it's an overhead that probably helps explain why a shorter key makes so little difference. Access to the index will be essentially random, no part of it is more likely to be accessed than any other. Total index size for the full file will likely be 5GB, so it's not extreme for that one index. Adding some other fields for other search cases, and their associated indexes, and copies of data for lower case, does start to add up, which I why I started looking into a more concise index.

    Read the article

  • PHP array performance

    - by dfo
    Hi, this is my first question on Stackoverflow, please bear with me. I'm testing an algorithm for 2d bin packing and I've chosen PHP to mock it up as it's my bread-and-butter language nowadays. As you can see on http://themworks.com/pack_v0.2/oopack.php?ol=1 it works pretty well, but you need to wait around 10-20 seconds for 100 rectangles to pack. For some hard to handle sets it would hit the php's 30s runtime limit. I did some profiling and it shows that most of the time my script goes through different parts of a small 2d array with 0's and 1's in it. It either checks if certain cell equals to 0/1 or sets it to 0/1. It can do such operations million times and each times it takes few microseconds. I guess I could use an array of booleans in a statically typed language and things would be faster. Or even make an array of 1 bit values. I'm thinking of converting the whole thing to some compiled language. Is PHP just not good for it? If I do need to convert it to let's say C++, how good are the automatic converters? My script is just a lot of for loops with basic arrays and objects manipulations. Thank you! Edit. This function gets called more than any other. It reads few properties of a very simple object, and goes through a very small part of a smallish array to check if there's any element not equal to 0. function fits($bin, $file, $x, $y) { $flag = true; $xw = $x + $file->get_width();; $yh = $y + $file->get_height(); for ($i = $x; $i < $xw; $i++) { for ($j = $y; $j < $yh; $j++) { if ($bin[$i][$j] !== 0) { $flag = false; break; } } if (!$flag) break; } return $flag; }

    Read the article

  • Dealing with HTTP w00tw00t attacks

    - by Saif Bechan
    I have a server with apache and I recently installed mod_security2 because I get attacked a lot by this: My apache version is apache v2.2.3 and I use mod_security2.c This were the entries from the error log: [Wed Mar 24 02:35:41 2010] [error] [client 88.191.109.38] client sent HTTP/1.1 request without hostname (see RFC2616 section 14.23): /w00tw00t.at.ISC.SANS.DFind:) [Wed Mar 24 02:47:31 2010] [error] [client 202.75.211.90] client sent HTTP/1.1 request without hostname (see RFC2616 section 14.23): /w00tw00t.at.ISC.SANS.DFind:) [Wed Mar 24 02:47:49 2010] [error] [client 95.228.153.177] client sent HTTP/1.1 request without hostname (see RFC2616 section 14.23): /w00tw00t.at.ISC.SANS.DFind:) [Wed Mar 24 02:48:03 2010] [error] [client 88.191.109.38] client sent HTTP/1.1 request without hostname (see RFC2616 section 14.23): /w00tw00t.at.ISC.SANS.DFind:) Here are the errors from the access_log: 202.75.211.90 - - [29/Mar/2010:10:43:15 +0200] "GET /w00tw00t.at.ISC.SANS.DFind:) HTTP/1.1" 400 392 "-" "-" 211.155.228.169 - - [29/Mar/2010:11:40:41 +0200] "GET /w00tw00t.at.ISC.SANS.DFind:) HTTP/1.1" 400 392 "-" "-" 211.155.228.169 - - [29/Mar/2010:12:37:19 +0200] "GET /w00tw00t.at.ISC.SANS.DFind:) HTTP/1.1" 400 392 "-" "-" I tried configuring mod_security2 like this: SecFilterSelective REQUEST_URI "w00tw00t\.at\.ISC\.SANS\.DFind" SecFilterSelective REQUEST_URI "\w00tw00t\.at\.ISC\.SANS" SecFilterSelective REQUEST_URI "w00tw00t\.at\.ISC\.SANS" SecFilterSelective REQUEST_URI "w00tw00t\.at\.ISC\.SANS\.DFind:" SecFilterSelective REQUEST_URI "w00tw00t\.at\.ISC\.SANS\.DFind:\)" The thing in mod_security2 is that SecFilterSelective can not be used, it gives me errors. Instead I use a rule like this: SecRule REQUEST_URI "w00tw00t\.at\.ISC\.SANS\.DFind" SecRule REQUEST_URI "\w00tw00t\.at\.ISC\.SANS" SecRule REQUEST_URI "w00tw00t\.at\.ISC\.SANS" SecRule REQUEST_URI "w00tw00t\.at\.ISC\.SANS\.DFind:" SecRule REQUEST_URI "w00tw00t\.at\.ISC\.SANS\.DFind:\)" Even this does not work. I don't know what to do anymore. Anyone have any advice? Update 1 I see that nobody can solve this problem using mod_security. So far using ip-tables seems like the best option to do this but I think the file will become extremely large because the ip changes serveral times a day. I came up with 2 other solutions, can someone comment on them on being good or not. The first solution that comes to my mind is excluding these attacks from my apache error logs. This will make is easier for me to spot other urgent errors as they occur and don't have to spit trough a long log. The second option is better i think, and that is blocking hosts that are not sent in the correct way. In this example the w00tw00t attack is send without hostname, so i think i can block the hosts that are not in the correct form. Update 2 After going trough the answers I came to the following conclusions. To have custom logging for apache will consume some unnecessary recourses, and if there really is a problem you probably will want to look at the full log without anything missing. It is better to just ignore the hits and concentrate on a better way of analyzing your error logs. Using filters for your logs a good approach for this. Final thoughts on the subject The attack mentioned above will not reach your machine if you at least have an up to date system so there are basically no worries. It can be hard to filter out all the bogus attacks from the real ones after a while, because both the error logs and access logs get extremely large. Preventing this from happening in any way will cost you resources and they it is a good practice not to waste your resources on unimportant stuff. The solution i use now is Linux logwatch. It sends me summaries of the logs and they are filtered and grouped. This way you can easily separate the important from the unimportant. Thank you all for the help, and I hope this post can be helpful to someone else too.

    Read the article

< Previous Page | 486 487 488 489 490 491 492 493 494 495 496 497  | Next Page >