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• Move options between multiple lists

  - by Martha

  We currently have a form with the standard multi-select functionality of "here are the available options, here are the selected options, here are some buttons to move stuff back and forth." However, the client now wants the ability to not just select certain items, but to also categorize them. For example, given a list of books, they want to not just select the ones they own, but also the ones they've read, the ones they would like to read, and the ones they've heard about. (All examples fictional.) Thankfully, a selected item can only be in one category at a time. I can find many examples of moving items between listboxes, but not a single one for moving items between multiple listboxes. To add to the complication, the form needs to have two sets of list+categories, e.g. a list of movies that need to be categorized in addition to the aforementioned books. EDIT: Having now actually sat down to try to code the non-javascripty bits, I need to revise my question, because I realized that multiple select lists won't really work from the "how do I inform the server about all this lovely new information" standpoint. So the html code is now a pseudo-listbox, i.e. an unordered list (ul) displayed in a box with a scrollbar, and each list item (<li>) has a set of five radio buttons (unselected/own/read/like/heard). My task is still roughly the same: how to take this one list and make it easy to categorize the items, in such a way that the user can tell at a glance what is in what category. (The pseudo-listbox has some of the same disadvantages as a multi-select listbox, namely it's hard to tell what's selected if the list is long enough to scroll.) The dream solution would be a drag-and-drop type thing, but at this point even buttons would be OK. Another modification (a good one) is that the client has revised the lists, so the longest is now "only" 62 items long (instead of the many hundreds they had before). The categories will still mostly contain zero, one, or two selected items, possibly a couple more if the user was overzealous. As far as OS and stuff, the site is in classic asp (quit snickering!), the server-side code is VBScript, and so far we've avoided the various Javascript libraries by the simple expedient of almost never using client-side scripting. This one form for this one client is currently the big exception. Give 'em an inch and they want a mile... Oh, and I have to add: I suck at Javascript, or really at any C-descendant language. Curly braces give me hives. I'd really, really like something I can just copy & paste into my page, maybe tweak some variable names, and never look at it again. A girl can dream, can't she? :) [existing code deleted because it's largely irrelevant.]

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• VB6 ADO Command to SQL Server

  - by Emtucifor

  I'm getting an inexplicable error with an ADO command in VB6 run against a SQL Server 2005 database. Here's some code to demonstrate the problem: Sub ADOCommand() Dim Conn As ADODB.Connection Dim Rs As ADODB.Recordset Dim Cmd As ADODB.Command Dim ErrorAlertID As Long Dim ErrorTime As Date Set Conn = New ADODB.Connection Conn.ConnectionString = "Provider=SQLOLEDB.1;Integrated Security=SSPI;Initial Catalog=database;Data Source=server" Conn.CursorLocation = adUseClient Conn.Open Set Rs = New ADODB.Recordset Rs.CursorType = adOpenStatic Rs.LockType = adLockReadOnly Set Cmd = New ADODB.Command With Cmd .Prepared = False .CommandText = "ErrorAlertCollect" .CommandType = adCmdStoredProc .NamedParameters = True .Parameters.Append .CreateParameter("@ErrorAlertID", adInteger, adParamOutput) .Parameters.Append .CreateParameter("@CreateTime", adDate, adParamOutput) Set .ActiveConnection = Conn Rs.Open Cmd ErrorAlertID = .Parameters("@ErrorAlertID").Value ErrorTime = .Parameters("@CreateTime").Value End With Debug.Print Rs.State ' Shows 0 - Closed Debug.Print Rs.RecordCount ' Of course this fails since the recordset is closed End Sub So this code was working not too long ago but now it's failing on the last line with the error: Run-time error '3704': Operation is not allowed when the object is closed Why is it closed? I just opened it and the SP returns rows. I ran a trace and this is what the ADO library is actually submitting to the server: declare @p1 int set @p1=1 declare @p2 datetime set @p2=''2010-04-22 15:31:07:770'' exec ErrorAlertCollect @ErrorAlertID=@p1 output,@CreateTime=@p2 output select @p1, @p2 Running this as a separate batch from my query editor yields: Msg 102, Level 15, State 1, Line 4 Incorrect syntax near '2010'. Of course there's an error. Look at the double single quotes in there. What the heck could be causing that? I tried using adDBDate and adDBTime as data types for the date parameter, and they give the same results. When I make the parameters adParamInputOutput, then I get this: declare @p1 int set @p1=default declare @p2 datetime set @p2=default exec ErrorAlertCollect @ErrorAlertID=@p1 output,@CreateTime=@p2 output select @p1, @p2 Running that as a separate batch yields: Msg 156, Level 15, State 1, Line 2 Incorrect syntax near the keyword 'default'. Msg 156, Level 15, State 1, Line 4 Incorrect syntax near the keyword 'default'. What the heck? SQL Server doesn't support this kind of syntax. You can only use the DEFAULT keyword in the actual SP execution statement. I should note that removing the extra single quotes from the above statement makes the SP run fine. ... Oh my. I just figured it out. I guess it's worth posting anyway.

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• My form php is not working and I can't figure out where I went wrong

  - by user1081524

  I'm fairly new to all this, but I've created a form, and this is what I've written to send it. I've used "[email protected]" here instead of the real address <?php /* Set e-mail recipient */ $myemail = "[email protected]"; /* Check all form inputs using check_input function */ $names = check_input($_POST['names'], "Please return to our Application Form and enter your and your future spouse's names."); $weddingtype = check_input($_POST['weddingtype'], "Please return to our Application Form and fill in what kind of wedding you will be having."); $religioussect = check_input($_POST['religioussect'], "Please return to our Application Form and tell us about your religion and wedding traditions."); $dateone = check_input($_POST['dateone'], "Please return to our Application Form and give us the date for at least one event."); $eventone = check_input($_POST['eventone'], "Please return to our Application Form and list at least one event."); $locationone = check_input($_POST['locationone'], "Please return to our Application Form and give us the location for at least one event."); $durationone = check_input($_POST['durationone'], "Please return to our Application Form and give us the duration of at least one event."); $typeone = check_input($_POST['typeone'], "Please return to our Application Form and tell us whether you would like video, photo or both for at least one event."); $datetwo = $_POST['datetwo']; $eventtwo = $_POST['eventtwo']; $locationtwo = $_POST['locationtwo']; $durationtwo = $_POST['durationtwo']; $typetwo = $_POST['typetwo']; $datethree = $_POST['datethree']; $eventthree = $_POST['eventthree']; $locationthree = $_POST['locationthree']; $durationthree = $_POST['durationthree']; $typethree = $_POST['typethree']; $datefour = $_POST['datefour']; $eventfour = $_POST['eventfour']; $locationfour = $_POST['locationfour']; $durationfour = $_POST['durationfour']; $typefour = $_POST['typefour']; $guests1 = check_input($_POST['guests1'], "Please return to our Application Form and tell us how many guests will attend at least one event."); $guests2 = $_POST['guests2']; $guests3 = $_POST['guests3']; $guests4 = $_POST['guests4']; $concerns = $_POST['concerns']; if(!isset($_POST['submit'])){ $subject = "Quote Application"; /*Message for the e-mail */ $message = "Hello! Another happy couple has filled out a Quote Application Form :D Hooray! Their names are $names. What sort of wedding are they having? '$weddingtype'. What religious sect and wedding traditions are they following? '$religioussect'. Now for their wedding events... Ooh boy! 1. $dateone $eventone $durationone $typeone Estimated guests: $guests1 $locationone 2. $datetwo $eventtwo $durationtwo $typetwo Estimated guests: $guests2 $locationtwo 3. $datethree $eventthree $durationthree $typethree Estimated guests: $guests3 $locationthree 4. $datefour $eventfour $durationfour $typefour Estimated guests: $guests4 $locationfour Any concerns the couple have follow here: '$concerns' You better be ready to get to work now! And also, have a really good day :) "; /* Functions used */ function check_input($data, $problem='') { $data = trim($data); $data = stripslashes($data); $data = htmlspecialchars($data); if ($problem && strlen($data) == 0) { show_error($problem); } return $data; } function show_error($myError) { ?> <b>We apologize for the inconvenience, an error occurred.</b><br /> <?php echo $myError; ?> <?php exit(); } /* Send the message using mail() function */ mail($myemail, $subject, $message); /* Redirect visitor to the thank you page */ header('Location: thanks.html'); exit(); ?> Please help me find what I'm doing wrong, I'm barely a beginner here. Thanks in advance!

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• Separate specific #ifdef branches

  - by detly

  In short: I want to generate two different source trees from the current one, based only on one preprocessor macro being defined and another being undefined, with no other changes to the source. If you are interested, here is my story... In the beginning, my code was clean. Then we made a new product, and yea, it was better. But the code saw only the same peripheral devices, so we could keep the same code. Well, almost. There was one little condition that needed to be changed, so I added: #if defined(PRODUCT_A) condition = checkCat(); #elif defined(PRODUCT_B) condition = checkCat() && checkHat(); #endif ...to one and only one source file. In the general all-source-files-include-this header file, I had: #if !(defined(PRODUCT_A)||defined(PRODUCT_B)) #error "Don't make me replace you with a small shell script. RTFM." #endif ...so that people couldn't compile it unless they explicitly defined a product type. All was well. Oh... except that modifications were made, components changed, and since the new hardware worked better we could significantly re-write the control systems. Now when I look upon the face of the code, there are more than 60 separate areas delineated by either: #ifdef PRODUCT_A ... #else ... #endif ...or the same, but for PRODUCT_B. Or even: #if defined(PRODUCT_A) ... #elif defined(PRODUCT_B) ... #endif And of course, sometimes sanity took a longer holiday and: #ifdef PRODUCT_A ... #endif #ifdef PRODUCT_B ... #endif These conditions wrap anywhere from one to two hundred lines (you'd think that the last one could be done by switching header files, but the function names need to be the same). This is insane. I would be better off maintaining two separate product-based branches in the source repo and porting any common changes. I realise this now. Is there something that can generate the two different source trees I need, based only on PRODUCT_A being defined and PRODUCT_B being undefined (and vice-versa), without touching anything else (ie. no header inclusion, no macro expansion, etc)?

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• creating a JQuery function

  - by Russell Parrott

  Sorry to bother you guys & girls again on Christmas eve, but I need help creating a reusable JQuery function. I have "badly crafted" this set of code that all works. But I would really like to put it as a function so I don't have to keep repeating everything for each form. I am not too sure about how all the if statements can be combined etc. that is why I wrote it as it is. Any help much appreciated - Oh I suppose it could also be some kind of plugin but that might be the next step if I can understand how the function works. $(':input:visible').live('blur',function(){ if($(this).attr('required')) { if($(this).val() == '' ) { $(this).css({'background-color':'#FFEEEE' }); $(this).parent('form').children('input[type=submit]').hide(); $(this).next('.errormsg').html('OOPs ... '+$(this).prev('label').html()+' is required'); $(this).focus(); $(this).attr('placeholder').hide(); } else { $(this).css({'background-color':'#FFF' , 'border-color':'#999999'}); $(this).next('.errormsg').empty(); $(this).parent('form').children('input[type=submit]').show(); } } return false; }); $(':input[max]').live('blur',function(){ if($(this).attr('max') < parseInt($(this).val()) ){ $(this).next('.errormsg').html( 'OOPs ... the maximum value is '+$(this).attr('max') ); $(this).parent('form').children('input[type=submit]').hide(); $(this).focus(); } else {} return false; }); $(':input[min]').live('blur',function(){ if($(this).attr('min') > parseInt($(this).val()) ){ $(this).next('.errormsg').html( 'OOPs ... the minimum value is '+$(this).attr('min') ); $(this).parent('form').children('input[type=submit]').hide(); $(this).focus(); } else {} return false; }); $(':input[maxlength]').live('keyup',function(){ if($(this).val()==''){ } else { $(this).next('.errormsg').html( $(this).attr('maxlength')- $(this).val().length +' chars remaining'); } return false; }); As said, help much appreciated with one small (I hope) thing, how can I break out of any function IF there are no error messages to actually submit the form?

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• mdadm: Win7-install created a boot partition on one of my RAID6 drives. How to rebuild?

  - by EXIT_FAILURE

  My problem happened when I attempted to install Windows 7 on it's own SSD. The Linux OS I used which has knowledge of the software RAID system is on a SSD that I disconnected prior to the install. This was so that windows (or I) wouldn't inadvertently mess it up. However, and in retrospect, foolishly, I left the RAID disks connected, thinking that windows wouldn't be so ridiculous as to mess with a HDD that it sees as just unallocated space. Boy was I wrong! After copying over the installation files to the SSD (as expected and desired), it also created an ntfs partition on one of the RAID disks. Both unexpected and totally undesired! . I changed out the SSDs again, and booted up in linux. mdadm didn't seem to have any problem assembling the array as before, but if I tried to mount the array, I got the error message: mount: wrong fs type, bad option, bad superblock on /dev/md0, missing codepage or helper program, or other error In some cases useful info is found in syslog - try dmesg | tail or so dmesg: EXT4-fs (md0): ext4_check_descriptors: Block bitmap for group 0 not in group (block 1318081259)! EXT4-fs (md0): group descriptors corrupted! I then used qparted to delete the newly created ntfs partition on /dev/sdd so that it matched the other three /dev/sd{b,c,e}, and requested a resync of my array with echo repair > /sys/block/md0/md/sync_action This took around 4 hours, and upon completion, dmesg reports: md: md0: requested-resync done. A bit brief after a 4-hour task, though I'm unsure as to where other log files exist (I also seem to have messed up my sendmail configuration). In any case: No change reported according to mdadm, everything checks out. mdadm -D /dev/md0 still reports: Version : 1.2 Creation Time : Wed May 23 22:18:45 2012 Raid Level : raid6 Array Size : 3907026848 (3726.03 GiB 4000.80 GB) Used Dev Size : 1953513424 (1863.02 GiB 2000.40 GB) Raid Devices : 4 Total Devices : 4 Persistence : Superblock is persistent Update Time : Mon May 26 12:41:58 2014 State : clean Active Devices : 4 Working Devices : 4 Failed Devices : 0 Spare Devices : 0 Layout : left-symmetric Chunk Size : 4K Name : okamilinkun:0 UUID : 0c97ebf3:098864d8:126f44e3:e4337102 Events : 423 Number Major Minor RaidDevice State 0 8 16 0 active sync /dev/sdb 1 8 32 1 active sync /dev/sdc 2 8 48 2 active sync /dev/sdd 3 8 64 3 active sync /dev/sde Trying to mount it still reports: mount: wrong fs type, bad option, bad superblock on /dev/md0, missing codepage or helper program, or other error In some cases useful info is found in syslog - try dmesg | tail or so and dmesg: EXT4-fs (md0): ext4_check_descriptors: Block bitmap for group 0 not in group (block 1318081259)! EXT4-fs (md0): group descriptors corrupted! I'm a bit unsure where to proceed from here, and trying stuff "to see if it works" is a bit too risky for me. This is what I suggest I should attempt to do: Tell mdadm that /dev/sdd (the one that windows wrote into) isn't reliable anymore, pretend it is newly re-introduced to the array, and reconstruct its content based on the other three drives. I also could be totally wrong in my assumptions, that the creation of the ntfs partition on /dev/sdd and subsequent deletion has changed something that cannot be fixed this way. My question: Help, what should I do? If I should do what I suggested , how do I do that? From reading documentation, etc, I would think maybe: mdadm --manage /dev/md0 --set-faulty /dev/sdd mdadm --manage /dev/md0 --remove /dev/sdd mdadm --manage /dev/md0 --re-add /dev/sdd However, the documentation examples suggest /dev/sdd1, which seems strange to me, as there is no partition there as far as linux is concerned, just unallocated space. Maybe these commands won't work without. Maybe it makes sense to mirror the partition table of one of the other raid devices that weren't touched, before --re-add. Something like: sfdisk -d /dev/sdb | sfdisk /dev/sdd Bonus question: Why would the Windows 7 installation do something so st...potentially dangerous? Update I went ahead and marked /dev/sdd as faulty, and removed it (not physically) from the array: # mdadm --manage /dev/md0 --set-faulty /dev/sdd # mdadm --manage /dev/md0 --remove /dev/sdd However, attempting to --re-add was disallowed: # mdadm --manage /dev/md0 --re-add /dev/sdd mdadm: --re-add for /dev/sdd to /dev/md0 is not possible --add, was fine. # mdadm --manage /dev/md0 --add /dev/sdd mdadm -D /dev/md0 now reports the state as clean, degraded, recovering, and /dev/sdd as spare rebuilding. /proc/mdstat shows the recovery progress: md0 : active raid6 sdd[4] sdc[1] sde[3] sdb[0] 3907026848 blocks super 1.2 level 6, 4k chunk, algorithm 2 [4/3] [UU_U] [>....................] recovery = 2.1% (42887780/1953513424) finish=348.7min speed=91297K/sec nmon also shows expected output: ¦sdb 0% 87.3 0.0| > |¦ ¦sdc 71% 109.1 0.0|RRRRRRRRRRRRRRRRRRRRRRRRRRRRRRRRRRRR > |¦ ¦sdd 40% 0.0 87.3|WWWWWWWWWWWWWWWWWWWW > |¦ ¦sde 0% 87.3 0.0|> || It looks good so far. Crossing my fingers for another five+ hours :) Update 2 The recovery of /dev/sdd finished, with dmesg output: [44972.599552] md: md0: recovery done. [44972.682811] RAID conf printout: [44972.682815] --- level:6 rd:4 wd:4 [44972.682817] disk 0, o:1, dev:sdb [44972.682819] disk 1, o:1, dev:sdc [44972.682820] disk 2, o:1, dev:sdd [44972.682821] disk 3, o:1, dev:sde Attempting mount /dev/md0 reports: mount: wrong fs type, bad option, bad superblock on /dev/md0, missing codepage or helper program, or other error In some cases useful info is found in syslog - try dmesg | tail or so And on dmesg: [44984.159908] EXT4-fs (md0): ext4_check_descriptors: Block bitmap for group 0 not in group (block 1318081259)! [44984.159912] EXT4-fs (md0): group descriptors corrupted! I'm not sure what do do now. Suggestions? Output of dumpe2fs /dev/md0: dumpe2fs 1.42.8 (20-Jun-2013) Filesystem volume name: Atlas Last mounted on: /mnt/atlas Filesystem UUID: e7bfb6a4-c907-4aa0-9b55-9528817bfd70 Filesystem magic number: 0xEF53 Filesystem revision #: 1 (dynamic) Filesystem features: has_journal ext_attr resize_inode dir_index filetype extent flex_bg sparse_super large_file huge_file uninit_bg dir_nlink extra_isize Filesystem flags: signed_directory_hash Default mount options: user_xattr acl Filesystem state: clean Errors behavior: Continue Filesystem OS type: Linux Inode count: 244195328 Block count: 976756712 Reserved block count: 48837835 Free blocks: 92000180 Free inodes: 243414877 First block: 0 Block size: 4096 Fragment size: 4096 Reserved GDT blocks: 791 Blocks per group: 32768 Fragments per group: 32768 Inodes per group: 8192 Inode blocks per group: 512 RAID stripe width: 2 Flex block group size: 16 Filesystem created: Thu May 24 07:22:41 2012 Last mount time: Sun May 25 23:44:38 2014 Last write time: Sun May 25 23:46:42 2014 Mount count: 341 Maximum mount count: -1 Last checked: Thu May 24 07:22:41 2012 Check interval: 0 (<none>) Lifetime writes: 4357 GB Reserved blocks uid: 0 (user root) Reserved blocks gid: 0 (group root) First inode: 11 Inode size: 256 Required extra isize: 28 Desired extra isize: 28 Journal inode: 8 Default directory hash: half_md4 Directory Hash Seed: e177a374-0b90-4eaa-b78f-d734aae13051 Journal backup: inode blocks dumpe2fs: Corrupt extent header while reading journal super block

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• Passing password value through URL

  - by Steven Wright

  OK I see a lot of people asking about passing other values, URLS, random stuff through a URL, but don't find anything about sending a password to a password field. Here is my situation: I have a ton of sites I use on a daily basis with my work and oh about 90% require logins. Obviously remembering 80 bajillion logins for each site is dumb, especially when there are more than one user name I use for each site. So to make life easier, I drew up a nifty JSP app that stores all of my logins in a DB table and creates a user interface for the specific page I want to visit. Each page has a button that sends a username, password into the id parameters of the html inputs. Problem: I can get the usernames and other info to show up just dandy, but when I try and send a password to a password field, it seems that nothing gets received by the page I'm trying to hit. Is there some ninja stuff I need to be doing here or is it just not easily possible? Basically this is what I do now: http://addresshere/support?loginname=steveoooo&loginpass=passwordhere and some of my html looks like this: <form name="userform" method="post" action="index.jsp" > <input type="hidden" name="submit_login" value="y"> <table width="100%"> <tr class="main"> <td width="100" nowrap>Username:</td> <td><input type="text" name="loginname" value="" size="30" maxlength="64"></td> </tr> <tr class="main"> <td>Password: </font></td> <td><input type="password" name="loginpass" value="" size="30" maxlength="64"></td> </tr> <tr class="main"> <td><center><input type="submit" name="submit" value="Login"></center></td> </tr> </table> </form> Any suggestions?

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• Which network protocol to use for lightweight notification of remote apps?

  - by Chris Thornton

  I have this situation.... Client-initiated SOAP 1.1 communication between one server and let's say, tens of thousands of clients. Clients are external, coming in through our firewall, authenticated by certificate, https, etc.. They can be anywhere, and usually have their own firewalls, NAT routers, etc... They're truely external, not just remote corporate offices. They could be in a corporate/campus network, DSL/Cable, even Dialup. Client uses Delphi (2005 + SOAP fixes from 2007), and the server is C#, but from an architecture/design standpoint, that shouldn't matter. Currently, clients push new data to the server and pull new data from the server on 15-minute polling loop. The server currently does not push data - the client hits the "messagecount" method, to see if there is new data to pull. If 0, it sleeps for another 15 min and checks again. We're trying to get that down to 7 seconds. If this were an internal app, with one or just a few dozen clients, we'd write a cilent "listener" soap service, and would push data to it. But since they're external, sit behind their own firewalls, and sometimes private networks behind NAT routers, this is not practical. So we're left with polling on a much quicker loop. 10K clients, each checking their messagecount every 10 seconds, is going to be 1000/sec messages that will mostly just waste bandwidth, server, firewall, and authenticator resources. So I'm trying to design something better than what would amount to a self-inflicted DoS attack. I don't think it's practical to have the server send soap messages to the client (push) as this would require too much configuration at the client end. But I think there are alternatives that I don't know about. Such as: 1) Is there a way for the client to make a request for GetMessageCount() via Soap 1.1, and get the response, and then perhaps, "stay on the line" for perhaps 5-10 minutes to get additional responses in case new data arrives? i.e the server says "0", then a minute later in response to some SQL trigger (the server is C# on Sql Server, btw), knows that this client is still "on the line" and sends the updated message count of "5"? 2) Is there some other protocol that we could use to "ping" the client, using information gathered from their last GetMessageCount() request? 3) I don't even know. I guess I'm looking for some magic protocol where the client can send a GetMessageCount() request, which would include info for "oh by the way, in case the answer changes in the next hour, ping me at this address...". Also, I'm assuming that any of these "keep the line open" schemes would seriously impact the server sizing, as it would need to keep many thousands of connections open, simultaneously. That would likely impact the firewalls too, I think. Is there anything out there like that? Or am I pretty much stuck with polling? TIA, Chris

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• mysql: can't set max_allowed_package to anything grater than 16MB

  - by sas

  I'm not sure if this is the right place to post these kind of questions, if it's not so, please (politely) let me know... :-) I need to save files greater than 16MB on a mysql database from a php site... I've already changed the c:\xampp\mysql\bin\my.cnf and set max_allowed_packet to 16 MB, and everything worked fine then I set it to 32 MB but there´s no way I can handle a file bigger than 16 MB I get the following error: 'MySQL server has gone away' (the same error I had when max_allowed_packet was set to 1MB) there must be some other setting that doesn´t allow me to handle files bigger than 16MB maybe the php client, I guess, but I don't know where to edit it this is the code I'm running when file.txt is smaller than 16.776.192 bytes long, it works fine, but if file.txt has 16.777.216 bytes i get the aforementioned error oh, and the field download.content is a longblob... $file = 'file.txt'; $file_handle = fopen( $file, 'r' ); $content = fread( $file_handle, filesize( $file ) ); fclose( $file_handle ); db_execute( 'truncate table download', true ); $sql = "insert into download( code, title, name, description, original_name, mime_type, size, content, user_insert_id, date_insert, user_update_id, date_update ) values ( 'new file', 'new file', 'sas.jpg', 'new file', '$file', 'mime', " . filesize( $file ) . ", '" . addslashes( $content ) . "', 0, " . db_char_to_sql( now_char(), 'datetime' ) . ", 0, " . db_char_to_sql( now_char(), 'datetime' ) . " )"; db_execute( $sql, true ); (the db_execute funcion just opens the connections and executes the sql stuff) running on windows XP sp2 server version: 5.0.67-community PHP Version 4.4.9 mysql client API version: 3.23.49 using: ApacheFriends XAMPP (Basispaket) version 1.6.8 that comes with + Apache 2.2.9 + MySQL 5.0.67 (Community Server) + PHP 5.2.6 + PHP 4.4.9 + PEAR + phpMyAdmin 2.11.9.2 ... this is part of the content of c:\xampp\mysql\bin\my.cnf # The MySQL server [mysqld] port= 3306 socket= "C:/xampp/mysql/mysql.sock" basedir="C:/xampp/mysql" tmpdir="C:/xampp/tmp" datadir="C:/xampp/mysql/data" skip-locking key_buffer = 16M # max_allowed_packet = 1M max_allowed_packet = 32M table_cache = 128 sort_buffer_size = 512K net_buffer_length = 8K read_buffer_size = 256K read_rnd_buffer_size = 512K myisam_sort_buffer_size = 8M

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• Primary language - C++/Qt, C#, Java?

  - by Airjoe

  I'm looking for some input, but let me start with a bit of background (for tl;dr skip to end). I'm an IT major with a concentration in networking. While I'm not a CS major nor do I want to program as a vocation, I do consider myself a programmer and do pretty well with the concepts involved. I've been programming since about 6th grade, started out with a proprietary game creation language that made my transition into C++ at college pretty easy. I like to make programs for myself and friends, and have been paid to program for local businesses. A bit about that- I wrote some programs for a couple local businesses in my senior year in high school. I wrote management systems for local shops (inventory, phone/pos orders, timeclock, customer info, and more stuff I can't remember). It definitely turned out to be over my head, as I had never had any formal programming education. It was a great learning experience, but damn was it crappy code. Oh yeah, by the way, it was all vb6. So, I've used vb6 pretty extensively, I've used c++ in my classes (intro to programming up to algorithms), used Java a little bit in another class (had to write a ping client program, pretty easy) and used Java for some simple Project Euler problems to help learn syntax and such when writing the program for the class. I've also used C# a bit for my own simple personal projects (simple programs, one which would just generate an HTTP request on a list of websites and notify if one responded unexpectedly or not at all, and another which just held a list of things to do and periodically reminded me to do them), things I would've written in vb6 a year or two ago. I've just started using Qt C++ for some undergrad research I'm working on. Now I've had some formal education, I [think I] understand organization in programming a lot better (I didn't even use classes in my vb6 programs where I really should have), how it's important to structure code, split into functions where appropriate, document properly, efficiency both in memory and speed, dynamic and modular programming etc. I was looking for some input on which language to pick up as my "primary". As I'm not a "real programmer", it will be mostly hobby projects, but will include some 'real' projects I'm sure. From my perspective: QtC++ and Java are cross platform, which is cool. Java and C# run in a virtual machine, but I'm not sure if that's a big deal (something extra to distribute, possibly a bit slower? I think Qt would require additional distributables too, right?). I don't really know too much more than this, so I appreciate any help, thanks! TL;DR Am an avocational programmer looking for a language, want quick and straight forward development, liked vb6, will be working with database driven GUI apps- should I go with QtC++, Java, C#, or perhaps something else?

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• asp:Button is not calling server-side function

  - by Richard Neil Ilagan

  Hi guys, I know that there has been much discussion here about this topic, but none of the threads I got across helped me solve this problem. I'm hoping that mine is somewhat unique, and may actually merit a different solution. I'm instantiating an asp:Button inside a data-bound asp:GridView through template fields. Some of the buttons are supposed to call a server-side function, but for some weird reason, it doesn't. All the buttons do when you click them is fire a postback to the current page, doing nothing, effectively just reloading the page. Below is a fragment of the code: <asp:GridView ID="gv" runat="server" AutoGenerateColumns="false" CssClass="l2 submissions" ShowHeader="false"> <Columns> <asp:TemplateField> <ItemTemplate><asp:Panel ID="swatchpanel" CssClass='<%# Bind("status") %>' runat="server"></asp:Panel></ItemTemplate> <ItemStyle Width="50px" CssClass="sw" /> </asp:TemplateField> <asp:BoundField DataField="description" ReadOnly="true"> </asp:BoundField> <asp:BoundField DataField="owner" ReadOnly="true"> <ItemStyle Font-Italic="true" /> </asp:BoundField> <asp:BoundField DataField="last-modified" ReadOnly="true"> <ItemStyle Width="100px" /> </asp:BoundField> <asp:TemplateField> <ItemTemplate> <asp:Button ID="viewBtn" cssclass='<%# Bind("sid") %>' runat="server" Text="View" OnClick="viewBtnClick" /> </ItemTemplate> </asp:TemplateField> </Columns> </asp:GridView> The viewBtn above should call the viewBtnClick() function on server-side. I do have that function defined, along with a proper signature (object,EventArgs). One thing that may be of note is that this code is actually inside an ASCX, which is loaded in another ASCX, finally loaded into an ASPX. Any help or insight into the matter will be SO appreciated. Thanks! (oh, and please don't mind my trashy HTML/CSS semantics - this is still in a very,very early stage :p)

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• template pass by const reference

  - by 7vies

  Hi, I've looked over a few similar questions, but I'm still confused. I'm trying to figure out how to explicitly (not by compiler optimization etc) and C++03-compatible avoid copying of an object when passing it to a template function. Here is my test code: #include <iostream> using namespace std; struct C { C() { cout << "C()" << endl; } C(const C&) { cout << "C(C)" << endl; } ~C() { cout << "~C()" << endl; } }; template<class T> void f(T) { cout << "f<T>" << endl; } template<> void f(C c) { cout << "f<C>" << endl; } // (1) template<> void f(const C& c) { cout << "f<C&>" << endl; } // (2) int main() { C c; f(c); return 0; } (1) accepts the object of type C, and makes a copy. Here is the output: C() C(C) f<C> ~C() ~C() So I've tried to specialize with a const C& parameter (2) to avoid this, but this simply doesn't work (apparently the reason is explained in this question). Well, I could "pass by pointer", but that's kind of ugly. So is there some trick that would allow to do that somehow nicely? EDIT: Oh, probably I wasn't clear. I already have a templated function template<class T> void f(T) {...} But now I want to specialize this function to accept a const& to another object: template<> void f(const SpecificObject&) {...} But it only gets called if I define it as template<> void f(SpecificObject) {...}

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• Can I take the voice data (f.e. in mp3 format) from speech recognition? [closed]

  - by Ersin Gulbahar

  Possible Duplicate: Android: Voice Recording and saving audio I mean ; I use voice recognition classes on android and I succeed voice recognition. But I want to real voice data not words instead of it. For example I said 'teacher' and android get you said teacher.Oh ok its good but I want to my voice which include 'teacher'.Where is it ? Can I take it and save another location? I use this class to speech to text : package net.viralpatel.android.speechtotextdemo; import java.util.ArrayList; import android.app.Activity; import android.content.ActivityNotFoundException; import android.content.Intent; import android.os.Bundle; import android.speech.RecognizerIntent; import android.view.Menu; import android.view.View; import android.widget.ImageButton; import android.widget.TextView; import android.widget.Toast; public class MainActivity extends Activity { protected static final int RESULT_SPEECH = 1; private ImageButton btnSpeak; private TextView txtText; @Override public void onCreate(Bundle savedInstanceState) { super.onCreate(savedInstanceState); setContentView(R.layout.activity_main); txtText = (TextView) findViewById(R.id.txtText); btnSpeak = (ImageButton) findViewById(R.id.btnSpeak); btnSpeak.setOnClickListener(new View.OnClickListener() { @Override public void onClick(View v) { Intent intent = new Intent( RecognizerIntent.ACTION_RECOGNIZE_SPEECH); intent.putExtra(RecognizerIntent.EXTRA_LANGUAGE_MODEL, "en-US"); try { startActivityForResult(intent, RESULT_SPEECH); txtText.setText(""); } catch (ActivityNotFoundException a) { Toast t = Toast.makeText(getApplicationContext(), "Ops! Your device doesn't support Speech to Text", Toast.LENGTH_SHORT); t.show(); } } }); } @Override public boolean onCreateOptionsMenu(Menu menu) { getMenuInflater().inflate(R.menu.activity_main, menu); return true; } @Override protected void onActivityResult(int requestCode, int resultCode, Intent data) { super.onActivityResult(requestCode, resultCode, data); switch (requestCode) { case RESULT_SPEECH: { if (resultCode == RESULT_OK && null != data) { ArrayList<String> text = data .getStringArrayListExtra(RecognizerIntent.EXTRA_RESULTS); txtText.setText(text.get(0)); } break; } } } } Thanks.

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• Minutia on Objective-C Categories and Extensions.

  - by Matt Wilding

  I learned something new while trying to figure out why my readwrite property declared in a private Category wasn't generating a setter. It was because my Category was named: // .m @interface MyClass (private) @property (readwrite, copy) NSArray* myProperty; @end Changing it to: // .m @interface MyClass () @property (readwrite, copy) NSArray* myProperty; @end and my setter is synthesized. I now know that Class Extension is not just another name for an anonymous Category. Leaving a Category unnamed causes it to morph into a different beast: one that now gives compile-time method implementation enforcement and allows you to add ivars. I now understand the general philosophies underlying each of these: Categories are generally used to add methods to any class at runtime, and Class Extensions are generally used to enforce private API implementation and add ivars. I accept this. But there are trifles that confuse me. First, at a hight level: Why differentiate like this? These concepts seem like similar ideas that can't decide if they are the same, or different concepts. If they are the same, I would expect the exact same things to be possible using a Category with no name as is with a named Category (which they are not). If they are different, (which they are) I would expect a greater syntactical disparity between the two. It seems odd to say, "Oh, by the way, to implement a Class Extension, just write a Category, but leave out the name. It magically changes." Second, on the topic of compile time enforcement: If you can't add properties in a named Category, why does doing so convince the compiler that you did just that? To clarify, I'll illustrate with my example. I can declare a readonly property in the header file: // .h @interface MyClass : NSObject @property (readonly, copy) NSString* myString; @end Now, I want to head over to the implementation file and give myself private readwrite access to the property. If I do it correctly: // .m @interface MyClass () @property (readonly, copy) NSString* myString; @end I get a warning when I don't synthesize, and when I do, I can set the property and everything is peachy. But, frustratingly, if I happen to be slightly misguided about the difference between Category and Class Extension and I try: // .m @interface MyClass (private) @property (readonly, copy) NSString* myString; @end The compiler is completely pacified into thinking that the property is readwrite. I get no warning, and not even the nice compile error "Object cannot be set - either readonly property or no setter found" upon setting myString that I would had I not declared the readwrite property in the Category. I just get the "Does not respond to selector" exception at runtime. If adding ivars and properties is not supported by (named) Categories, is it too much to ask that the compiler play by the same rules? Am I missing some grand design philosophy?

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• Primary language - QtC++, C#, Java?

  - by Airjoe

  I'm looking for some input, but let me start with a bit of background (for tl;dr skip to end). I'm an IT major with a concentration in networking. While I'm not a CS major nor do I want to program as a vocation, I do consider myself a programmer and do pretty well with the concepts involved. I've been programming since about 6th grade, started out with a proprietary game creation language that made my transition into C++ at college pretty easy. I like to make programs for myself and friends, and have been paid to program for local businesses. A bit about that- I wrote some programs for a couple local businesses in my senior year in high school. I wrote management systems for local shops (inventory, phone/pos orders, timeclock, customer info, and more stuff I can't remember). It definitely turned out to be over my head, as I had never had any formal programming education. It was a great learning experience, but damn was it crappy code. Oh yeah, by the way, it was all vb6. So, I've used vb6 pretty extensively, I've used c++ in my classes (intro to programming up to algorithms), used Java a little bit in another class (had to write a ping client program, pretty easy) and used Java for some simple Project Euler problems to help learn syntax and such when writing the program for the class. I've also used C# a bit for my own simple personal projects (simple programs, one which would just generate an HTTP request on a list of websites and notify if one responded unexpectedly or not at all, and another which just held a list of things to do and periodically reminded me to do them), things I would've written in vb6 a year or two ago. I've just started using Qt C++ for some undergrad research I'm working on. Now I've had some formal education, I [think I] understand organization in programming a lot better (I didn't even use classes in my vb6 programs where I really should have), how it's important to structure code, split into functions where appropriate, document properly, efficiency both in memory and speed, dynamic and modular programming etc. I was looking for some input on which language to pick up as my "primary". As I'm not a "real programmer", it will be mostly hobby projects, but will include some 'real' projects I'm sure. From my perspective: QtC++ and Java are cross platform, which is cool. Java and C# run in a virtual machine, but I'm not sure if that's a big deal (something extra to distribute, possibly a bit slower? I think Qt would require additional distributables too, right?). I don't really know too much more than this, so I appreciate any help, thanks! TL;DR Am an avocational programmer looking for a language, want quick and straight forward development, liked vb6, will be working with database driven GUI apps- should I go with QtC++, Java, C#, or perhaps something else?

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• Two collections and a for loop. (Urgent help needed) Checking an object variable against an inputted

  - by Elliott

  Hi there, I'm relatively new to java, I'm certain the error is trivial. But can't for the life of me spot it. I have an end of term exam on monday and currently trying to get to grips with past papers! Anyway heregoes, in another method (ALGO_1) I search over elements of and check the value H_NAME equals a value entered in the main. When I attempt to run the code I get a null pointer exception, also upon trying to print (with System.out.println etc) the H_NAME value after each for loop in the snippet I also get a null statement returned to me. I am fairly certain that the collection is simply not storing the data gathered up by the Scanner. But then again when I check the collection size with size() it is about the right size. Either way I'm pretty lost and would appreciate the help. Main questions I guess to ask are: from the readBackground method is the data.add in the wrong place? is the snippet simply structured wrongly? oh and another point when I use System.out.println to check the Background object values name, starttime, increment etc they print out fine. Thanks in advance.(PS im guessing the formatting is terrible, apologies.) snippet of code: for(Hydro hd: hydros){ System.out.println(hd.H_NAME); for(Background back : backgs){ System.out.println(back.H_NAME); if(back.H_NAME.equals(hydroName)){ //get error here public static Collection<Background> readBackground(String url) throws IOException { URL u = new URL(url); InputStream is = u.openStream(); InputStreamReader isr = new InputStreamReader(is); BufferedReader b = new BufferedReader(isr); String line =""; Vector<Background> data = new Vector<Background>(); while((line = b.readLine())!= null){ Scanner s = new Scanner(line); String name = s.next(); double starttime = Double.parseDouble(s.next()); double increment = Double.parseDouble(s.next()); double sum = 0; double p = 0; double nterms = 0; while((s.hasNextDouble())){ p = Double.parseDouble(s.next()); nterms++; sum += p; } double pbmean = sum/nterms; Background SAMP = new Background(name, starttime, increment, pbmean); data.add(SAMP); } return data; } Edit/Delete Message

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• Which network protocol to use for lightweight notification of remote apps (Delphi 2005)

  - by Chris Thornton

  I have this situation.... Client-initiated SOAP 1.1 communication between one server and let's say, tens of thousands of clients. Clients are external, coming in through our firewall, authenticated by certificate, https, etc.. They can be anywhere, and usually have their own firewalls, NAT routers, etc... They're truely external, not just remote corporate offices. They could be in a corporate/campus network, DSL/Cable, even Dialup. Currently, clients push new data to the server and pull new data from the server on 15-minute polling loop. The server currently does not push data - the client hits the "messagecount" method, to see if there is new data to pull. If 0, it sleeps for another 15 min and checks again. We're trying to get that down to 7 seconds. If this were an internal app, with one or just a few dozen clients, we'd write a cilent "listener" soap service, and would push data to it. But since they're external, sit behind their own firewalls, and sometimes private networks behind NAT routers, this is not practical. So we're left with polling on a much quicker loop. 10K clients, each checking their messagecount every 10 seconds, is going to be 1000/sec messages that will mostly just waste bandwidth, server, firewall, and authenticator resources. So I'm trying to design something better than what would amount to a self-inflicted DoS attack. I don't think it's practical to have the server send soap messages to the client (push) as this would require too much configuration at the client end. But I think there are alternatives that I don't know about. Such as: 1) Is there a way for the client to make a request for GetMessageCount() via Soap 1.1, and get the response, and then perhaps, "stay on the line" for perhaps 5-10 minutes to get additional responses in case new data arrives? i.e the server says "0", then a minute later in response to some SQL trigger (the server is C# on Sql Server, btw), knows that this client is still "on the line" and sends the updated message count of "5"? 2) Is there some other protocol that we could use to "ping" the client, using information gathered from their last GetMessageCount() request? 3) I don't even know. I guess I'm looking for some magic protocol where the client can send a GetMessageCount() request, which would include info for "oh by the way, in case the answer changes in the next hour, ping me at this address...". Also, I'm assuming that any of these "keep the line open" schemes would seriously impact the server sizing, as it would need to keep many thousands of connections open, simultaneously. That would likely impact the firewalls too, I think. Is there anything out there like that? Or am I pretty much stuck with polling? TIA, Chris

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• Printing a variable only when it changes?

  - by user1781639

  First off, my question was a little vague or confusing since I'm not really sure how to phrase my question to be specific. I'm trying to query a database of stockists for a Knitting company (school project using PHP) but I'm looking to print the city as a heading instead of with each stockists information. Here is what I have at the moment: $sql = "SELECT * FROM mc16korustockists where locale = 'south'"; $result = pg_exec($sql); $nrows = pg_numrows($result); print $nrows; $items = pg_fetch_all($result); print_r($items); for ($i=0; $i<$nrows2; $i++) { print "<h2>"; print $items[$i]['city']; print "</h2>"; print $items[$i]['name']; print $items[$i]['address']; print $items[$i]['city']; print $items[$i]['phone']; print "<br />"; print "<br />"; } I'm querying the database for all of the data in it, the rows being ref, name, address, city and phone, and executing it. Querying the number of rows then using that to determine how many iterations for the loop to run is all fine but what I'd like to have is for the h2 heading to appear above the for ($i=0;) line. Trying just breaks my page so that might be out of the question. I figure I'd have to count the number of entries in 'city' until it detects a change then change the heading to that name I think? That or make a heap of queries and set a variable for each name but at point, I might as well do it manually (and I highly doubt it would be best practice). Oh, and I'd welcome any critiques to my PHP since I'm just starting out. Thanks and if you need any more information, just ask! P.S. Our class is learning with PostgreSQL instead of MySQL as you can see in the tags.

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• CKEdtior not displaying

  - by user1708468

  I am trying to integrate CKEditor into a MVC application. As far as I can tell all I should really have to do is. Add the following to my master page. <script type="text/javascript" src="../../ckeditor/ckeditor.js"></script> <script type="text/javascript" src="../../ckeditor/adapters/jquery.js"></script> <script type="text/jscript" src="../../Scripts/jquery-1.3.2.js"></script> Then on my view itself. I have the following code: <script type="text/javascript"> $(document).ready(function() { $('#news').ckeditor(); }); </script> <fieldset> <legend>Fields</legend> <p> <label for="title">Title:</label> <%=Html.TextBox("title")%> <%= Html.ValidationMessage("title", "*") %> </p> <p> <label for="news">News:</label> <%=Html.TextArea("news")%> <%= Html.ValidationMessage("news", "*") %> </p> <p> <label for="publishedDate">Publication Date:</label> <%= Html.TextBox("publishedDate") %> <%= Html.ValidationMessage("publishedDate", "*") %> </p> <p> <input type="submit" value="Create" /> </p> </fieldset> Please bear in mind I am not trying to get this to actually DO anything postback wise. Just to actually render in the first place. Can someone point out exactly what it is I am doing wrong? Oh and if it helps any VS is also giving me the following warning: Warning 1 Error updating JScript IntelliSense: ..Cut to Protect the innocent..\ckeditor\ckeditor.js: 'getFirst()' is null or not an object @ 15:180 ..Cut to Protect the innocent..\Views\Shared\Admin.Master 1 1 ilaTraining

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• How should I smooth the transition between these two states in flex/flashbuilder

  - by Joshua

  I have an item in which has two states, best described as open and closed, and they look like this: and And what I'd like to do is is smooth the transition between one state and the other, effectively by interpolating between the two points in a smooth manner (sine) to move the footer/button-block and then fade in the pie chart. However this is apparently beyond me and after wrestling with my inability to do so for an hour+ I'm posting it here :D So my transition block looks as follows <s:transitions> <s:Transition id="TrayTrans" fromState="*" toState="*"> <s:Sequence> <s:Move duration="400" target="{footer}" interpolator="{Sine}"/> <s:Fade duration="300" targets="{body}"/> </s:Sequence> </s:Transition> <s:Transition> <s:Rotate duration="3000" /> </s:Transition> </s:transitions> where {body} refers to the pie chart and {footer} refers to the footer/button-block. However this doesn't work so I don't really know what to do... Additional information which may be beneficial: The body block is always of fixed height (perhaps of use for the Xby variables in some effects?). It needs to work in both directions. Oh and the Sine block is defined above in declarations just as <s:Sine id="Sine">. Additionally! How would I go about setting the pie chart to rotate continually using these transition blocks? (this would occur without the labels on) Or is that the wrong way to go about it as it's not a transition as such? The effect I'm after is one where the pie chart rotates slowly without labels prior to a selection of a button below, but on selection the rotation stops and labels appear... Thanks a lot in advance! And apologies on greyscale, but I can't really decide on a colour scheme. Any suggestions welcome.

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• PHP, PEAR, and oci8 configuration

  - by zack_falcon

  I'll make this quick. I installed Oracle 11g (with appropriate database, users, etc), Apache 2.4.6, and PHP 5.5.4 on a Fedora 19 system. I wanted to connect PHP to Oracle. What I really wanted to do was to download MDB2_Driver_oci8, which I thought would be easy, but before I can do such a thing, PHP needs to have that plug-in enabled, so here's what I did: Tried to install oci8 via the following: pecl install oci8 When that didn't exactly work the first few times, I figured out I, for some reason, needed "Development tools" - via yum groupinstall "Development Tools" Then I figured out later that PHP actually doesn't do oci8 - it's PHP Devel. So, I had to install that too, via yum install php-devel. And then, I finally got to install oci8. It asked for the Oracle Directory, and that was that. But it said the following: Configuration option 'php_ini' is not set to php.ini location You should add 'extensions=oci8.so' to php.ini First, I did a locate oci8.so - found it in /usr/lib64/php/modules/ Second, I added what it told me to, to the php.ini file. Third, I checked the usual php_info() test page - no mention of OCI8. Uh-oh. Fourth, running both php -i and php -m listed oci8 as one of the modules. Weird. In desperation, I went ahead and downloaded the MDB2_Driver_oci8. Maybe that will fix things. Nope. When I loaded my PHP Webpage, it returned the following: Error message: extension oci8 is not compiled into PHP As well as: MDB2 error: not found Strange. And then I decided to check the error logs: PHP Startup - unable to load dynamic library '/usr/lib64/php/modules/oci8.so' - libclntsh.so.11.1: cannot open shared object file: No such file or directory in Unknown on line 0 And now I'm stuck. I tried going into the php.ini, and found that the extension_dir was commented out. I put it back in, which only seemed to break stuff. Things of note: I followed this (link) guide on how to configure PHP and install oci8. ./configure --with-oci8 doesn't work. Fedora says no such directory. As both the webpage files and the actual server reside on the same PC, I did not install the Oracle Client files. The extension_dir is commented out by default in the php.ini. This is just one of my problems in a long line of problems concerning the replication of an already existing and working, but dying, setup. It seems whenever I want to solve a problem, I have to do X first. And by doing X, I uncover another problem, which I have to solve by doing Y, which has its own problems, etc, etc. Any help would be much appreciated. Thanks.

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• Twitter traffic might not be what it seems

  - by Piet

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

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• Analytic functions – they’re not aggregates

  - by Rob Farley

  SQL 2012 brings us a bunch of new analytic functions, together with enhancements to the OVER clause. People who have known me over the years will remember that I’m a big fan of the OVER clause and the types of things that it brings us when applied to aggregate functions, as well as the ranking functions that it enables. The OVER clause was introduced in SQL Server 2005, and remained frustratingly unchanged until SQL Server 2012. This post is going to look at a particular aspect of the analytic functions though (not the enhancements to the OVER clause). When I give presentations about the analytic functions around Australia as part of the tour of SQL Saturdays (starting in Brisbane this Thursday), and in Chicago next month, I’ll make sure it’s sufficiently well described. But for this post – I’m going to skip that and assume you get it. The analytic functions introduced in SQL 2012 seem to come in pairs – FIRST_VALUE and LAST_VALUE, LAG and LEAD, CUME_DIST and PERCENT_RANK, PERCENTILE_CONT and PERCENTILE_DISC. Perhaps frustratingly, they take slightly different forms as well. The ones I want to look at now are FIRST_VALUE and LAST_VALUE, and PERCENTILE_CONT and PERCENTILE_DISC. The reason I’m pulling this ones out is that they always produce the same result within their partitions (if you’re applying them to the whole partition). Consider the following query: SELECT     YEAR(OrderDate),     FIRST_VALUE(TotalDue)         OVER (PARTITION BY YEAR(OrderDate)               ORDER BY OrderDate, SalesOrderID               RANGE BETWEEN UNBOUNDED PRECEDING                         AND UNBOUNDED FOLLOWING),     LAST_VALUE(TotalDue)         OVER (PARTITION BY YEAR(OrderDate)               ORDER BY OrderDate, SalesOrderID               RANGE BETWEEN UNBOUNDED PRECEDING                         AND UNBOUNDED FOLLOWING),     PERCENTILE_CONT(0.95)         WITHIN GROUP (ORDER BY TotalDue)         OVER (PARTITION BY YEAR(OrderDate)),     PERCENTILE_DISC(0.95)         WITHIN GROUP (ORDER BY TotalDue)         OVER (PARTITION BY YEAR(OrderDate)) FROM Sales.SalesOrderHeader ; This is designed to get the TotalDue for the first order of the year, the last order of the year, and also the 95% percentile, using both the continuous and discrete methods (‘discrete’ means it picks the closest one from the values available – ‘continuous’ means it will happily use something between, similar to what you would do for a traditional median of four values). I’m sure you can imagine the results – a different value for each field, but within each year, all the rows the same. Notice that I’m not grouping by the year. Nor am I filtering. This query gives us a result for every row in the SalesOrderHeader table – 31465 in this case (using the original AdventureWorks that dates back to the SQL 2005 days). The RANGE BETWEEN bit in FIRST_VALUE and LAST_VALUE is needed to make sure that we’re considering all the rows available. If we don’t specify that, it assumes we only mean “RANGE BETWEEN UNBOUNDED PRECEDING AND CURRENT ROW”, which means that LAST_VALUE ends up being the row we’re looking at. At this point you might think about other environments such as Access or Reporting Services, and remember aggregate functions like FIRST. We really should be able to do something like: SELECT     YEAR(OrderDate),     FIRST_VALUE(TotalDue)         OVER (PARTITION BY YEAR(OrderDate)               ORDER BY OrderDate, SalesOrderID               RANGE BETWEEN UNBOUNDED PRECEDING                         AND UNBOUNDED FOLLOWING) FROM Sales.SalesOrderHeader GROUP BY YEAR(OrderDate) ; But you can’t. You get that age-old error: Msg 8120, Level 16, State 1, Line 5 Column 'Sales.SalesOrderHeader.OrderDate' is invalid in the select list because it is not contained in either an aggregate function or the GROUP BY clause. Msg 8120, Level 16, State 1, Line 5 Column 'Sales.SalesOrderHeader.SalesOrderID' is invalid in the select list because it is not contained in either an aggregate function or the GROUP BY clause. Hmm. You see, FIRST_VALUE isn’t an aggregate function. None of these analytic functions are. There are too many things involved for SQL to realise that the values produced might be identical within the group. Furthermore, you can’t even surround it in a MAX. Then you get a different error, telling you that you can’t use windowed functions in the context of an aggregate. And so we end up grouping by doing a DISTINCT. SELECT DISTINCT     YEAR(OrderDate),         FIRST_VALUE(TotalDue)              OVER (PARTITION BY YEAR(OrderDate)                   ORDER BY OrderDate, SalesOrderID                   RANGE BETWEEN UNBOUNDED PRECEDING                             AND UNBOUNDED FOLLOWING),         LAST_VALUE(TotalDue)             OVER (PARTITION BY YEAR(OrderDate)                   ORDER BY OrderDate, SalesOrderID                   RANGE BETWEEN UNBOUNDED PRECEDING                             AND UNBOUNDED FOLLOWING),     PERCENTILE_CONT(0.95)          WITHIN GROUP (ORDER BY TotalDue)         OVER (PARTITION BY YEAR(OrderDate)),     PERCENTILE_DISC(0.95)         WITHIN GROUP (ORDER BY TotalDue)         OVER (PARTITION BY YEAR(OrderDate)) FROM Sales.SalesOrderHeader ; I’m sorry. It’s just the way it goes. Hopefully it’ll change the future, but for now, it’s what you’ll have to do. If we look in the execution plan, we see that it’s incredibly ugly, and actually works out the results of these analytic functions for all 31465 rows, finally performing the distinct operation to convert it into the four rows we get in the results. You might be able to achieve a better plan using things like TOP, or the kind of calculation that I used in http://sqlblog.com/blogs/rob_farley/archive/2011/08/23/t-sql-thoughts-about-the-95th-percentile.aspx (which is how PERCENTILE_CONT works), but it’s definitely convenient to use these functions, and in time, I’m sure we’ll see good improvements in the way that they are implemented. Oh, and this post should be good for fellow SQL Server MVP Nigel Sammy’s T-SQL Tuesday this month.

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• T4 Template error - Assembly Directive cannot locate referenced assembly in Visual Studio 2010 proje

  - by CodeSniper

  I ran into the following error recently in Visual Studio 2010 while trying to port Phil Haack’s excellent T4CSS template which was originally built for Visual Studio 2008.   The Problem Error Compiling transformation: Metadata file 'dotless.Core' could not be found In “T4 speak”, this simply means that you have an Assembly directive in your T4 template but the T4 engine was not able to locate or load the referenced assembly. In the case of the T4CSS Template, this was a showstopper for making it work in Visual Studio 2010. On a side note: The T4CSS template is a sweet little wrapper to allow you to use DotLessCss to generate static .css files from .less files rather than using their default HttpHandler or command-line tool.    If you haven't tried DotLessCSS yet, go check it out now!  In short, it is a tool that allows you to templatize and program your CSS files so that you can use variables, expressions, and mixins within your CSS which enables rapid changes and a lot of developer-flexibility as you evolve your CSS and UI. Back to our regularly scheduled program… Anyhow, this post isn't about DotLessCss, its about the T4 Templates and the errors I ran into when converting them from Visual Studio 2008 to Visual Studio 2010. In VS2010, there were quite a few changes to the T4 Template Engine; most were excellent changes, but this one bit me with T4CSS: “Project assemblies are no longer used to resolve template assembly directives.” In VS2008, if you wanted to reference a custom assembly in your T4 Template (.tt file) you would simply right click on your project, choose Add Reference and select that assembly.  Afterwards you were allowed to use the following syntax in your T4 template to tell it to look at the local references: <#@ assembly name="dotless.Core.dll" #> This told the engine to look in the “usual place” for the assembly, which is your project references. However, this is exactly what they changed in VS2010.  They now basically sandbox the T4 Engine to keep your T4 assemblies separate from your project assemblies.  This can come in handy if you want to support different versions of an assembly referenced both by your T4 templates and your project. Who broke the build?  Oh, Microsoft Did! In our case, this change causes a problem since the templates are no longer compatible when upgrading to VS 2010 – thus its a breaking change.  So, how do we make this work in VS 2010? Luckily, Microsoft now offers several options for referencing assemblies from T4 Templates: GAC your assemblies and use Namespace Reference or Fully Qualified Type Name Use a hard-coded Fully Qualified UNC path Copy assembly to Visual Studio "Public Assemblies Folder" and use Namespace Reference or Fully Qualified Type Name.  Use or Define a Windows Environment Variable to build a Fully Qualified UNC path. Use a Visual Studio Macro to build a Fully Qualified UNC path. Option #1 & 2 were already supported in Visual Studio 2008, so if you want to keep your templates compatible with both Visual Studio versions, then you would have to adopt one of these approaches. Yakkety Yak, use the GAC! Option #1 requires an additional pre-build step to GAC the referenced assembly, which could be a pain.  But, if you go that route, then after you GAC, all you need is a simple type name or namespace reference such as: <#@ assembly name="dotless.Core" #> Hard Coding aint that hard! The other option of using hard-coded paths in Option #2 is pretty impractical in most situations since each developer would have to use the same local project folder paths, or modify this setting each time for their local machines as well as for production deployment.  However, if you want to go that route, simply use the following assembly directive style: <#@ assembly name="C:\Code\Lib\dotless.Core.dll" #> Lets go Public! Option #3, the Visual Studio Public Assemblies Folder, is the recommended place to put commonly used tools and libraries that are only needed for Visual Studio.  Think of it like a VS-only GAC.  This is likely the best place for something like dotLessCSS and is my preferred solution.  However, you will need to either use an installer or a pre-build action to copy the assembly to the right folder location.   Normally this is located at:  C:\Program Files (x86)\Microsoft Visual Studio 10.0\Common7\IDE\PublicAssemblies Once you have copied your assembly there, you use the type name or namespace syntax again: <#@ assembly name="dotless.Core" #> Save the Environment! Option #4, using a Windows Environment Variable, is interesting for enterprise use where you may have standard locations for files, but less useful for demo-code, frameworks, and products where you don't have control over the local system.  The syntax for including a environment variable in your assembly directive looks like the following, just as you would expect: <#@ assembly name="%mypath%\dotless.Core.dll" #> “mypath” is a Windows environment variable you setup that points to some fully qualified UNC path on your system.  In the right situation this can be a great solution such as one where you use a msi installer for deployment, or where you have a pre-existing environment variable you can re-use. OMG Macros! Finally, Option #5 is a very nice option if you want to keep your T4 template’s assembly reference local and relative to the project or solution without muddying-up your dev environment or GAC with extra deployments.  An example looks like this: <#@ assembly name="$(SolutionDir)lib\dotless.Core.dll" #> In this example, I’m using the “SolutionDir” VS macro so I can reference an assembly in a “/lib” folder at the root of the solution.   This is just one of the many macros you can use.  If you are familiar with creating Pre/Post-build Event scripts, you can use its dialog to look at all of the different VS macros available. This option gives the best solution for local assemblies without the hassle of extra installers or other setup before the build.   However, its still not compatible with Visual Studio 2008, so if you have a T4 Template you want to use with both, then you may have to create multiple .tt files, one for each IDE version, or require the developer to set a value in the .tt file manually.   I’m not sure if T4 Templates support any form of compiler switches like “#if (VS2010)”  statements, but it would definitely be nice in this case to switch between this option and one of the ones more compatible with VS 2008. Conclusion As you can see, we went from 3 options with Visual Studio 2008, to 5 options (plus one problem) with Visual Studio 2010.  As a whole, I think the changes are great, but the short-term growing pains during the migration may be annoying until we get used to our new found power. Hopefully this all made sense and was helpful to you.  If nothing else, I’ll just use it as a reference the next time I need to port a T4 template to Visual Studio 2010.  Happy T4 templating, and “May the fourth be with you!”

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• Multi-tenant ASP.NET MVC - Views

  - by zowens

  Part I – Introduction Part II – Foundation Part III – Controllers   So far we have covered the basic premise of tenants and how they will be delegated. Now comes a big issue with multi-tenancy, the views. In some applications, you will not have to override views for each tenant. However, one of my requirements is to add extra views (and controller actions) along with overriding views from the core structure. This presents a bit of a problem in locating views for each tenant request. I have chosen quite an opinionated approach at the present but will coming back to the “views” issue in a later post. What’s the deal? The path I’ve chosen is to use precompiled Spark views. I really love Spark View Engine and was planning on using it in my project anyways. However, I ran across a really neat aspect of the source when I was having a look under the hood. There’s an easy way to hook in embedded views from your project. There are solutions that provide this, but they implement a special Virtual Path Provider. While I think this is a great solution, I would rather just have Spark take care of the view resolution. The magic actually happens during the compilation of the views into a bin-deployable DLL. After the views are compiled, the are simply pulled out of the views DLL. Each tenant has its own views DLL that just has “.Views” appended after the assembly name as a convention. The list of reasons for this approach are quite long. The primary motivation is performance. I’ve had quite a few performance issues in the past and I would like to increase my application’s performance in any way that I can. My customized build of Spark removes insignificant whitespace from the HTML output so I can some some bandwidth and load time without having to deal with whitespace removal at runtime.   How to setup Tenants for the Host In the source, I’ve provided a single tenant as a sample (Sample1). This will serve as a template for subsequent tenants in your application. The first step is to add a “PostBuildStep” installer into the project. I’ve defined one in the source that will eventually change as we focus more on the construction of dependency containers. The next step is to tell the project to run the installer and copy the DLL output to a folder in the host that will pick up as a tenant. Here’s the code that will achieve it (this belongs in Post-build event command line field in the Build Events tab of settings) %systemroot%\Microsoft.NET\Framework\v4.0.30319\installutil "$(TargetPath)" copy /Y "$(TargetDir)$(TargetName)*.dll" "$(SolutionDir)Web\Tenants\" copy /Y "$(TargetDir)$(TargetName)*.pdb" "$(SolutionDir)Web\Tenants\" The DLLs with a name starting with the target assembly name will be copied to the “Tenants” folder in the web project. This means something like MultiTenancy.Tenants.Sample1.dll and MultiTenancy.Tenants.Sample1.Views.dll will both be copied along with the debug symbols. This is probably the simplest way to go about this, but it is a tad inflexible. For example, what if you have dependencies? The preferred method would probably be to use IL Merge to merge your dependencies with your target DLL. This would have to be added in the build events. Another way to achieve that would be to simply bypass Visual Studio events and use MSBuild.   I also got a question about how I was setting up the controller factory. Here’s the basics on how I’m setting up tenants inside the host (Global.asax) protected void Application_Start() { RegisterRoutes(RouteTable.Routes); // create a container just to pull in tenants var topContainer = new Container(); topContainer.Configure(config => { config.Scan(scanner => { scanner.AssembliesFromPath(Path.Combine(Server.MapPath("~/"), "Tenants")); scanner.AddAllTypesOf<IApplicationTenant>(); }); }); // create selectors var tenantSelector = new DefaultTenantSelector(topContainer.GetAllInstances<IApplicationTenant>()); var containerSelector = new TenantContainerResolver(tenantSelector); // clear view engines, we don't want anything other than spark ViewEngines.Engines.Clear(); // set view engine ViewEngines.Engines.Add(new TenantViewEngine(tenantSelector)); // set controller factory ControllerBuilder.Current.SetControllerFactory(new ContainerControllerFactory(containerSelector)); } The code to setup the tenants isn’t actually that hard. I’m utilizing assembly scanners in StructureMap as a simple way to pull in DLLs that are not in the AppDomain. Remember that there is a dependency on the host in the tenants and a tenant cannot simply be referenced by a host because of circular dependencies.   Tenant View Engine TenantViewEngine is a simple delegator to the tenant’s specified view engine. You might have noticed that a tenant has to define a view engine. public interface IApplicationTenant { .... IViewEngine ViewEngine { get; } } The trick comes in specifying the view engine on the tenant side. Here’s some of the code that will pull views from the DLL. protected virtual IViewEngine DetermineViewEngine() { var factory = new SparkViewFactory(); var file = GetType().Assembly.CodeBase.Without("file:///").Replace(".dll", ".Views.dll").Replace('/', '\\'); var assembly = Assembly.LoadFile(file); factory.Engine.LoadBatchCompilation(assembly); return factory; } This code resides in an abstract Tenant where the fields are setup in the constructor. This method (inside the abstract class) will load the Views assembly and load the compilation into Spark’s “Descriptors” that will be used to determine views. There is some trickery on determining the file location… but it works just fine.   Up Next There’s just a few big things left such as StructureMap configuring controllers with a convention instead of specifying types directly with container construction and content resolution. I will also try to find a way to use the Web Forms View Engine in a multi-tenant way we achieved with the Spark View Engine without using a virtual path provider. I will probably not use the Web Forms View Engine personally, but I’m sure some people would prefer using WebForms because of the maturity of the engine. As always, I love to take questions by email or on twitter. Suggestions are always welcome as well! (Oh, and here’s another link to the source code).

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