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  • Subroutine & GoTo design

    - by sub
    I have a strange question concerning subroutines: As I'm creating a minimal language and I don't want to add high-level loops like while or for I was planning on just adding gotos to keep it Turing-Complete. Now I thought, eww - gotos - I wouldn't want to program in that language if I had to use gotos so often. So I thought about adding subroutines instead. I see the difference as the following: gotos Go to (captain obvious) a previously defined point and continue executing the program from there. Leads to hardly understandable and buggy code, I think that's a fact. subroutines Similiar: You define their starting point somewhere, as you call them the program jumps there - but the subroutine can go back to the point it was called from with return. Okay. Why didn't I just add the more function-like, nice looking subroutines? Because: In order to make return work if I call subroutines from within subroutines from within other subroutines, I'd have to use a stack containing the point where the currently running subroutine came from at top. That would then mean that I would, if I create loops using the subroutines, end up with an extremely memory-eating, overflowing stack with return locations. Not good. Don't think of my subroutines as functions. They are just gotos that return to the point they were called from, they don't actually give back values like the return x; statement in nearly all today's languages. Now to my actual questions: How can I solve the above problem with the stack overflow on loops with subroutines? Do I have to add a separate goto language construct without the return option? Assembler doesn't have loops but as I have seen myJumpPoint:, jnz, jz, retn. That means to me that there must also be a stack containing all the return locations. Am I right with that? What about long running loops then? Don't they overflow the stack/eat memory then? Am I getting the retn symbol in assembler totally wrong? If yes, please explain it to me.

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  • Evaluating points in time by months, but without referencing years in Rails

    - by MikeH
    FYI, There is some overlap in the initial description of this question with a question I asked yesterday, but the question is different. My app has users who have seasonal products. When a user selects a product, we allow him to also select the product's season. We accomplish this by letting him select a start date and an end date for each product. We're using date_select to generate two sets of drop-downs: one for the start date and one for the end date. Including years doesn't make sense for our model. So we're using the option: discard_year => true When you use discard_year => true, Rails sets a year in the database, it just doesn't appear in the views. Rails sets all the years to either 0001 or 0002 in our app. Yes, we could make it 2009 and 2010 or any other pair. But the point is that we want the months and days to function independent of a particular year. If we used 2009 and 2010, then those dates would be wrong next year because we don't expect these records to be updated every year. My problem is that we need to dynamically evaluate the availability of products based on their relationship to the current month. For example, assume it's March 15. Regardless of the year, I need a method that can tell me that a product available from October to January is not available right now. If we were using actual years, this would be pretty easy. For example, in the products model, I can do this: def is_available? (season_start.past? && season_end.future?) end I can also evaluate a start_date and an end_date against current_date However, in setup I've described above where we have arbitrary years that only make sense relative to each other, these methods don't work. For example, is_available? would return false for all my products because their end date is in the year 0001 or 0002. What I need is a method just like the ones I used as examples above, except that they evaluate against current_month instead of current_date, and past? and future months instead of years. I have no idea how to do this or whether Rails has any built in functionality that could help. I've gone through all the date and time methods/helpers in the API docs, but I'm not seeing anything equivalent to what I'm describing. Thanks.

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  • Possible to use Python with Intel's Atom Developer SDK (C/C++)?

    - by Jordan Magnuson
    So I've made a game in Python and PyGame. Now I'm interested in submitting the game to Intel's March Developer Challenge. However, the developer challenge requires use of Intel's Atom Developer SDK (http://appdeveloper.intel.com/en-us/sdk), which only has API's for C and C++. I'm new to Python and PyGame, and have no experience in C or C++. My question is, would it be possible to somehow implement Intel's Atom SDK through/with/from a Python application (as the first link above suggests)? I've read up a little bit on embedding/extending Python into/with C, but I'm not entirely sure what to embed or where. I mean, I know I can do things like this in C: #include <Python.h> int main(int argc, char *argv[]) { Py_Initialize(); PyRun_SimpleString("from time import time,ctime\n" "print 'Today is',ctime(time())\n"); Py_Finalize(); return 0; } But what do I do about all my dependencies on Python and Pygame, for people that don't have those installed on their machines? Normally Py2Exe takes care of compacting the required dependencies (I've managed to package my game into an exe/zip), but how do I take care of that stuff in the context of embedding within C? Can I somehow work with py2exe on this, or do I need to do something entirely different for embedding within C? It seems like it would be a lot easier to go the route of extending Python with the C validation code, rather than trying to embed my whole game within C, but I think that's not an option, "because the library provided is currently only available as a Visual Studio 2008 '.lib'", meaning the application has to be compiled with Visual Studio...? Any help, thoughts, or ideas are much appreciated! You can find the complete SDK Developer's Guide on the intel site above, but here is their "Hello World" using the C Language API: #include <stdio.h> #include “adpcore.h” int main( int argc, char* argv[] ) { ADP_RET_CODE ret_code; const ADP_APPLICATIONID myApplicationID = {{ 0x12345678,0x11112222,0x33331234,0x567890ab}}; if ((ret_code = ADP_Initialize()) != ADP_SUCCESS ){ printf( “ERROR: exiting” ); exit( -1 ); } if (( ret_code = ADP_IsAuthorized( myApplicationId )) == ADP_AUTHORIZED ) printf( “Hello World” ); else printf( “Not authorized to run” ); exit 0; } 35 Page SDK Developer Guide: http:// appdeveloper.intel.com/sites/files/pages/SDK%20Developer%20Guide.pdf

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  • shopify_app syntax error

    - by Pete171
    Edit: Debugging has got me further. Question clarified. We have installed Ruby, RubyGems and Rails and have forked the shopify_app project. We have created a new rails applications and added three items to the Gemfile: execjs, therubyracer and shopify_app. Running rails s in order to start our rails application returns this trace: root@ubuntu:/usr/local/pete-shopify/cart# rails s Faraday: you may want to install system_timer for reliable timeouts /var/lib/gems/1.8/gems/shopify_app-4.1.0/lib/shopify_app.rb:15:in `require': /var/lib /gems/1.8/gems/shopify_app-4.1.0/lib/shopify_app/login_protection.rb:5: syntax error, unexpected ':', expecting kEND (SyntaxError) ...rce::UnauthorizedAccess, with: :close_session ^ from /var/lib/gems/1.8/gems/shopify_app-4.1.0/lib/shopify_app.rb:15 from /var/lib/gems/1.8/gems/bundler-1.2.1/lib/bundler/runtime.rb:68:in `require' from /var/lib/gems/1.8/gems/bundler-1.2.1/lib/bundler/runtime.rb:68:in `require' from /var/lib/gems/1.8/gems/bundler-1.2.1/lib/bundler/runtime.rb:66:in `each' from /var/lib/gems/1.8/gems/bundler-1.2.1/lib/bundler/runtime.rb:66:in `require' from /var/lib/gems/1.8/gems/bundler-1.2.1/lib/bundler/runtime.rb:55:in `each' from /var/lib/gems/1.8/gems/bundler-1.2.1/lib/bundler/runtime.rb:55:in `require' from /var/lib/gems/1.8/gems/bundler-1.2.1/lib/bundler.rb:128:in `require' from /usr/local/pete-shopify/cart/config/application.rb:7 from /var/lib/gems/1.8/gems/railties-3.2.8/lib/rails/commands.rb:53:in `require' from /var/lib/gems/1.8/gems/railties-3.2.8/lib/rails/commands.rb:53 from /var/lib/gems/1.8/gems/railties-3.2.8/lib/rails/commands.rb:50:in `tap' from /var/lib/gems/1.8/gems/railties-3.2.8/lib/rails/commands.rb:50 from script/rails:6:in `require' from script/rails:6 I haven't modified any files since forking from Github. Lines 1 - 6 of login_protection.rb are as follows: module ShopifyApp::LoginProtection extend ActiveSupport::Concern included do rescue from ActiveResource::UnauthorizedAccess, with: :close_session end I've looked into this and it seems that the error is caused by a new-style hash syntax between Ruby 1.8 and 1.9; key : value instead of key => value. Running ruby -v from the command line returns ruby 1.9.3p0 (2011-10-30 revision 33570) [x86_64-linux]. This would seem to be OK... but I did some debugging, and inside the file /var/lib/gems/1.8/gems/shopify_app-4.1.0/lib/shopify_app.rb (at the top) by putting this: puts RUBY_VERSION exit It printed 1.8.7. **Why are ruby -v and RUBY_VERSION giving me different results? And am I correct in assuming this is the cause of my problems? Note: To upgrade Ruby I installed the later version with apt-get and then switched to it by using update-alternatives --config ruby and selecting option 2 like this: root@ubuntu:/usr/local/pete-shopify/cart# update-alternatives --config ruby There are 2 choices for the alternative ruby (providing /usr/bin/ruby). Selection Path Priority Status ------------------------------------------------------------ 0 /usr/bin/ruby1.8 50 auto mode 1 /usr/bin/ruby1.8 50 manual mode * 2 /usr/bin/ruby1.9.1 10 manual mode Also note: We're PHP/Python developers so this is all new to us! Summary: 1 - Am I right in determining the cause of the syntax error? 2 - Why does RUBY_VERSION and ruby -v give me different results?

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  • Function lfit in numerical recipes, providing a test function

    - by Simon Walker
    Hi I am trying to fit collected data to a polynomial equation and I found the lfit function from Numerical Recipes. I only have access to the second edition, so am using that. I have read about the lfit function and its parameters, one of which is a function pointer, given in the documentation as void (*funcs)(float, float [], int)) with the help The user supplies a routine funcs(x,afunc,ma) that returns the ma basis functions evaluated at x = x in the array afunc[1..ma]. I am struggling to understand how this lfit function works. An example function I found is given below: void fpoly(float x, float p[], int np) /*Fitting routine for a polynomial of degree np-1, with coe?cients in the array p[1..np].*/ { int j; p[1]=1.0; for (j=2;j<=np;j++) p[j]=p[j-1]*x; } When I run through the source code for the lfit function in gdb I can see no reference to the funcs pointer. When I try and fit a simple data set with the function, I get the following error message. Numerical Recipes run-time error... gaussj: Singular Matrix ...now exiting to system... Clearly somehow a matrix is getting defined with all zeroes. I am going to involve this function fitting in a large loop so using another language is not really an option. Hence why I am planning on using C/C++. For reference, the test program is given here: int main() { float x[5] = {0., 0., 1., 2., 3.}; float y[5] = {0., 0., 1.2, 3.9, 7.5}; float sig[5] = {1., 1., 1., 1., 1.}; int ndat = 4; int ma = 4; /* parameters in equation */ float a[5] = {1, 1, 1, 0.1, 1.5}; int ia[5] = {1, 1, 1, 1, 1}; float **covar = matrix(1, ma, 1, ma); float chisq = 0; lfit(x,y,sig,ndat,a,ia,ma,covar,&chisq,fpoly); printf("%f\n", chisq); free_matrix(covar, 1, ma, 1, ma); return 0; } Also confusing the issue, all the Numerical Recipes functions are 1 array-indexed so if anyone has corrections to my array declarations let me know also! Cheers

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  • Is there a way to handle the dynamic change of a dropdown for a single row in a grid-based datawindo

    - by TomatoSandwich
    Is there a way to handle the dynamic change of a dropdown for a single row in a grid-based datawindow? Example: NAME LIKABILITY PURCHASED IN COLOUR (Text) (DropDown*) (Text) (Text) Bananas [Good] Hands Yellow [Bad] [Bananas are good] Apples [Good] Bags Red [Bad] Given the above is a grid-based datawindow, where the fields 'NAME','PURCHASED IN' and 'COLOUR' are text fields, where as the 'LIKABILITY' field is a dropdown*. I say dropdown* because the same visual representation can be created by using a DropDownList (hardcoded within the datawindow element at design time), or a DropDownDW (or DDDW, a select statement that can be based on other elements in the datawindow). However, there is no way I can get 'Bananas' having it's 3 dropdowns, while Apples has only 2. If I enter multiple rows of 'Bananas', then all rows have 3 dropdowns, but as soon as I add an Apples row, all dropdowns revert to 2 selections. To attempt to achieve this functionality, I have tried the following options: -- 1) dw_1.Object.likability.values("Good~tG/Bad~tB/Bananas are good~tDRWHO") on ue_itemchange when editing NAME. FAILS: edits all instances of LIKABILITY instead of the current row. -- 2) Duplicate Dropdowns, having one filtered, one unfiltered selection list per row, visible based on NAME selection. FAILS: can't set visibility/overlapping columns on grid-based datawindow. (Source) -- 3) Hard-code display value as Database value, or Vice Versa. Have 'GOOD','BAD','BANANASAREGOOD' as the display and database values, and change handling of options from G, B, DRWHO to these new values. FAILS: 3rd option appears for all rows, still selectable on Apple rows, which is wrong. -- 4) DDDW retrieve list of options for dropdown. Create a DDDW that uses the value of NAME to determine what selections it should have for the dropdown. FAILS: edits all instances of the dropdown, not just the current row. -- 5) DDDW retrieve counter of options available (if B then 3 else 2), then have duplicate dropdown columns that protect/unprotect based on DDDW counter. FAILS: Can't autoselect dddw value to populate column to cause protect on other two columns, ugly solution in any case. -- There is now a bounty on this question for anyone who can give me a solution that will enable me to edit a dropdown column for a single row on a grid-based datawindow in PB 10.5

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  • Sending multiline message via sockets without closing the connection

    - by Yasir Arsanukaev
    Hello folks. Currently I have this code of my client-side Haskell application: import Network.Socket import Network.BSD import System.IO hiding (hPutStr, hPutStrLn, hGetLine, hGetContents) import System.IO.UTF8 connectserver :: HostName -- ^ Remote hostname, or localhost -> String -- ^ Port number or name -> IO Handle connectserver hostname port = withSocketsDo $ do -- withSocketsDo is required on Windows -- Look up the hostname and port. Either raises an exception -- or returns a nonempty list. First element in that list -- is supposed to be the best option. addrinfos <- getAddrInfo Nothing (Just hostname) (Just port) let serveraddr = head addrinfos -- Establish a socket for communication sock <- socket (addrFamily serveraddr) Stream defaultProtocol -- Mark the socket for keep-alive handling since it may be idle -- for long periods of time setSocketOption sock KeepAlive 1 -- Connect to server connect sock (addrAddress serveraddr) -- Make a Handle out of it for convenience h <- socketToHandle sock ReadWriteMode -- Were going to set buffering to LineBuffering and then -- explicitly call hFlush after each message, below, so that -- messages get logged immediately hSetBuffering h LineBuffering return h sendid :: Handle -> String -> IO String sendid h id = do hPutStr h id -- Make sure that we send data immediately hFlush h -- Retrieve results hGetLine h The code portions in connectserver are from this chapter of Real World Haskell book where they say: When dealing with TCP data, it's often convenient to convert a socket into a Haskell Handle. We do so here, and explicitly set the buffering – an important point for TCP communication. Next, we set up lazy reading from the socket's Handle. For each incoming line, we pass it to handle. After there is no more data – because the remote end has closed the socket – we output a message about that. Since hGetContents blocks until the server closes the socket on the other side, I used hGetLine instead. It satisfied me before I decided to implement multiline output to client. I wouldn't like the server to close a socket every time it finishes sending multiline text. The only simple idea I have at the moment is to count the number of linefeeds and stop reading lines after two subsequent linefeeds. Do you have any better suggestions? Thanks.

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  • Coldfusion Report Builder - How can you set different datasources externally between prod/staging/de

    - by Smooth Operator
    Coldfusion Report Builder is great. One small issue. We use ANT+CFANT to deploy. When we create the report, say in a datasource called MyApp_dev on a dev box. Everything works great when the report is created. We deploy the report to our staging server, which has a datasource of MyApp_Staging. That server also, may or may not, have the live app working under MyApp_Live. Ant pushes the update to Staging just great. Run the report, crashes and burns. Why? It seems the report is looking for the MyApp_Dev data_source, even though the application is using the MyApp_Staging datasource. In digging around I found a few approaches, I would like to do this one, final, ideal way from the beginning instead of having to go back to do dozens of reports differently when I have a new Aha! moment. 1) Obvious: Pass in the datasource in to the cfreport tag. Doesn't work for ColdFusion Builder Reports as of v8, or v9 as tested on Linux. 2) Most realistic option (but painful) so far: Pass in the query as an object into the ColdFusion Builder report. Let's think about this: Create the Report with the report builder to my heart's content using the RDS, etc on my local box. When I'm done, copy the query into a snippet of code, or into a database column to be dynamically be injected at runtime with correct datasource. Modify my "run report" event to find the query from the database column, insert it into another dynamic cfquery and potentially... evaluate (!?!) it? Fun side is I can set the cfquery datasource to what I would need for each environment. When I modify the report's columns in CF Report Builder, I always have to update the query in the database. Is there a snippet of code that can extract this for me? Hmm. 3) Less than ideal. Suck it up and let all the reports in staging run off the live server. Maybe copy the live data into staging (sans structural changes) to let it seem similar. Are there any eloquent ways to accomplish the above? Thanks in Advance!

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  • What is the relationship between WebProxy & IWebProxy with respect to WebClient?

    - by Streamline
    I am creating an app (.NET 2.0) that uses WebClient to connect (downloaddata, etc) to/from a http web service. I am adding a form now to handle allowing proxy information to either be stored or set to use the defaults. I am a little confused about some things. First, some of the methods & properties available in either WebProxy or IWebProxy are not in both. What is the difference here with respect to setting up how WebClient will be have when it is called? Secondly, do I have to tell WebClient to use the proxy information if I set it using either WebProxy or IWebProxy class elsewhere? Or is it automatically inherited? Thirdly, when giving the option for the user to use the default proxy (whatever is set in IE) and using the default credentials (I assume also whatever is set in IE) are these two mutually exclusive? Or you only use default credentials when you have also used default proxy? This gets me to the whole difference between WebProxy and IWebProxy. WebRequest.DefaultProxy is a IWebPRoxy class but UseDefaultCredentials is not a method on the IWebProxy class, rather it is only on WebProxy and in turn, How to set the proxy to the WebRequest.DefautlProxy if they are two different classes? Here is my current method to read the stored form settings by the user - but I am not sure if this is correct, not enough, overkill, or just wrong because of the mix of WebProxy and IWebProxy: private WebProxy _proxyInfo = new WebProxy(); private WebProxy SetProxyInfo() { if (UseProxy) { if (UseIEProxy) { // is doing this enough to set this as default for WebClient? IWebProxy iProxy = WebRequest.DefaultWebProxy; if (UseIEProxyCredentials) { _proxyInfo.UseDefaultCredentials = true; } else { // is doing this enough to set this as default credentials for WebClient? WebRequest.DefaultWebProxy.Credentials = new NetworkCredential(ProxyUsername, ProxyPassword); } } else { // is doing this enough to set this as default for WebClient? WebRequest.DefaultWebProxy = new WebProxy(ProxyAddress, ParseLib.StringToInt(ProxyPort)); if (UseIEProxyCredentials) { _proxyInfo.UseDefaultCredentials = true; } else { WebRequest.DefaultWebProxy.Credentials = new NetworkCredential(ProxyUsername, ProxyPassword); } } } // Do I need to WebClient to absorb this returned proxy info if I didn't set or use defaults? return _proxyInfo; } Is there any reason to not just scrap storing app specific proxy information and only allow the app the ability to use the default proxy information & credentials for the logged in user? Will this ever not be enough if using HTTP? Part 2 Question: How can I test that the WebClient instance is using the proxy information or not?

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  • Problem with jQuery.ajax with 'delete' method in ie

    - by Max Williams
    I have a page where the user can edit various content using buttons and selects that trigger ajax calls. In particular, one action causes a url to be called remotely, with some data and a 'put' request, which (as i'm using a restful rails backend) triggers my update action. I also have a delete button which calls the same url but with a 'delete' request. The 'update' ajax call works in all browsers but the 'delete' one doesn't work in IE. I've got a vague memory of encountering something like this before...can anyone shed any light? here's my ajax calls: //update action - works in all browsers jQuery.ajax({ async:true, data:data, dataType:'script', type:'put', url:"/quizzes/"+quizId+"/quiz_questions/"+quizQuestionId, success: function(msg){ initializeQuizQuestions(); setPublishButtonStatus(); } }); //delete action - fails in ie function deleteQuizQuestion(quizQuestionId, quizId){ //send ajax call to back end to change the difficulty of the quiz question //back end will then refresh the relevant parts of the page (progress bars, flashes, quiz status) jQuery.ajax({ async:true, dataType:'script', type:'delete', url:"/quizzes/"+quizId+"/quiz_questions/"+quizQuestionId, success: function(msg){ alert("success"); initializeQuizQuestions(); setSelectStatus(quizQuestionId, true); jQuery("tr[id*='quiz_question_"+quizQuestionId+"']").removeClass('selected'); }, error: function(msg){ alert("error:" + msg); } }); } I put the alerts in success and error in the delete ajax just to see what happens, and the 'error' part of the ajax call is triggered, but WITH NO CALL BEING MADE TO THE BACK END (i know this by watching my back end server logs). So, it fails before it even makes the call. I can't work out why - the 'msg' i get back from the error block is blank. Any ideas anyone? Is this a known problem? I've tested it in ie6 and ie8 and it doesn't work in either. thanks - max EDIT - the solution - thanks to Nick Craver for pointing me in the right direction. Rails (and maybe other frameworks?) has a subterfuge for the unsupported put and delete requests: a post request with the parameter "_method" (note the underscore) set to 'put' or 'delete' will be treated as if the actual request type was that string. So, in my case, i made this change - note the 'data' option': jQuery.ajax({ async:true, data: {"_method":"delete"}, dataType:'script', type:'post', url:"/quizzes/"+quizId+"/quiz_questions/"+quizQuestionId, success: function(msg){ alert("success"); initializeQuizQuestions(); setSelectStatus(quizQuestionId, true); jQuery("tr[id*='quiz_question_"+quizQuestionId+"']").removeClass('selected'); }, error: function(msg){ alert("error:" + msg); } }); } Rails will now treat this as if it were a delete request, preserving the REST system. The reason my PUT example worked was just because in this particular case IE was happy to send a PUT request, but it officially does not support them so it's best to do this for PUT requests as well as DELETE requests.

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  • Linux: How to find all serial devices (ttyS, ttyUSB, ..) without opening them?

    - by Thomas Tempelmann
    What is the proper way to get a list of all available serial ports/devices on a Linux system? In other words, when I iterate over all devices in /dev/, how do I tell which ones are serial ports in the classic way, i.e. those usually supporting baud rates and RTS/CTS flow control? The solution would be coded in C. I ask because I am using a 3rd party library that does this clearly wrong: It appears to only iterate over /dev/ttyS*. The problem is that there are, for instance, serial ports over USB (provided by USB-RS232 adapters), and those are listed under /dev/ttyUSB*. And reading the Serial-HOWTO at Linux.org, I get the idea that there'll be other name spaces as well, as time comes. So I need to find the official way to detect serial devices. Problem is that there appears none documented, or I can't find it. I imagine one way would be to open all files from /dev/tty* and call a specific ioctl() on them that is only available on serial devices. Would that be a good solution, though? Update hrickards suggested to look at the source for "setserial". Its code does exactly what I had in mind: First, it opens a device with: fd = open (path, O_RDWR | O_NONBLOCK) Then it invokes: ioctl (fd, TIOCGSERIAL, &serinfo) If that call returns no error, then it's a serial dev, apparently. I found similar code here, which suggested to also add the O_NOCTTY option. There is one problem with this approach, though: When I tested this code on BSD Unix (i.e. OSX), it worked as well, however serial devices that are provided thru Bluetooth cause the system (driver) to try to connect to the bluetooth device, which takes a while before it'll return with a timeout error. This is caused by just opening the device. And I can imagine that similar things can happen on Linux as well - ideally, I should not need to open the device to figure out its type. I wonder if there's also a way to invoke ioctl functions without an open, or open a device in a way that it does not cause connections to be made? Any ideas?

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  • MySQL table organization and optimization (Rails)

    - by aguynamedloren
    I've been learning Ruby on Rails over the past few months with no prior programming experience. Lately, I've been thinking about database optimization and table organization. I know there are great books on the subject, but I typically learn by example / as I go. Here's a hypothetical situation: Let's say I am building a social network for a niche community with 250,000 members (users). The users have the ability to attend events. Let's say there are 50,000 past/present/future events. Much like Facebook events, a user can attend any number of events and an event can have any number of attendees. In the database, there would be a table for users and a table for events. Somehow I would have to create an association between the users and events. I could create an "events" column in the users table such that each user row would contain a hash of event IDs, or I could create an "attendees" column in the events table such that each event row would contain a hash of user IDs. Neither of these solutions seem ideal, however. On a users profile page, I want to display the list of events they are associated with, which would require scanning the 50,000 event rows for the user ID of said user if I include an "attendees" column in the events table. Likewise, on an event page, I want to display a list of attendees for the event, which would require scanning the 250,000 user rows for the event ID of said event if I include an "events" column in the users table. Option 3 would be to create a third table that contains the attendee information for each and every event - but I don't see how this would solve any problems. Are these non-issues? Rails makes accessing all of this information easy, but I guess I'm worried about scale. It is entirely possible that I am under-estimating the speed and processing power of modern databases / servers / etc. How long would it take to scan 250,000 user rows for specific event IDs - 10ms? 100ms? 1,000ms? I guess that's not that bad. Am I just over-thinking this?

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  • Another IKImageView Question: copying a region

    - by Brian Postow
    I'm trying to use the select and copy feature of the IKImageView. If all you want to do is have an app with an image, select a portion and copy it to the clipboard, it's easy. You set the copy menu pick to the first responder's copy:(id) method and magically everything works. However, if you want something more complicated, like you want to copy as part of some other operation, I can't seem to find the method to do this. IKImageView doesn't seem to have a copy method, it doesn't seem to have a method that will even tell you the selected rectangle! I have gone through Hillegass' book, so I understand how the clipboard works, just not how to get the portion of the image out of the view... Now, I'm starting to think that I made a mistake in basing my project on IKImageView, but it's what Preview is built on (or so I've read), so I figured it had to be stable... and anyway, now it's too late, I'm too deep in this to start over... So, other than not using IKImageView, any suggestions on how to copy the select region to the clipboard manually? EDIT actually, I have found the copy(id) method, but when I call it, I get <Error>: CGBitmapContextCreate: unsupported parameter combination: 8 integer bits/component; 16 bits/pixel; 1-component color space; kCGImageAlphaPremultipliedLast; 2624 bytes/row. Which obviously doesn't happen when I do a normal copy through the first-responder... I understand the error message, but I'm not sure where it's getting those parameters from... Is there any way to trace through this and see how this is happening? A debugger won't help for obvious reasons, as well as the fact that I'm doing this in Mozilla, so a debugger isn't an option anyway... EDIT 2 It occurs to me that the copy:(id) method I found may be copying the VIEW rather than copying a chunk of the image to the clipboard, which is what I need. The reason I thought it was the clipboard copy is that in another project, where I'm copying from an IKImageView to the clipboard straight from the edit menu, it just sends a copy:(id) to the firstResponder, but I'm not actually sure what the firstresponder does with it...

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  • Running same powershell script multiple asynchronous times with separate runspace

    - by teqnomad
    Hi, I have a powershell script which is called by a batch script which is called by Trap Receiver (which also passes environment variables) (running on windows 2008). The traps are flushed out at times in sets of 2-4 trap events, and the batch script will echo the trap details for each message to a logfile, but the powershell script on the next line of the batch script will only appear to process the first trap message (the powershell script writes to the same logfile). My interpetation is that the defaultrunspace is common to all iterations of the script running and this is why the others appear to be ignored. I've tried adding "-sta" when I invoke the powershell script using "powershell.exe -command", but this didn't help. I've researched and found a method using C# but I don't know this language, and busy enough learning powershell, so hoping to find a more direct solution especially as interleaving a "wrapper" between batch and powershell will involve passing the environment variables. http://www.codeproject.com/KB/threads/AsyncPowerShell.aspx I've hunted through stackoverflow, and again the only question of similar vein was using C#. Any suggestions welcome. Some script background: The powershell script is actually a modification of a great script found at gregorystrike website - cant post the link as I'm limited to one link but its the one for Lefthand arrays. Lots of mods so it can do multiple targets from one .ini file, taking in the environment variables, and options to run portions of the script interactively with winform. But you can see the gist of the original script. The batch script is pretty basic. The keys things are I'm trying to filter out trap noise using the :~ operator, and I tried -sta option to see if this would compartmentalise the powershell script. set debug=off set CMD_LINE_ARGS="%*" set LHIPAddress="%2" set VARBIND8="%8" shift shift shift shift shift shift shift set CHASSIS="%9" echo %DATE% %TIME% "Trap Received: %LHIPAddress% %CHASSIS% %VARBIND8%" >> C:\Logs\trap_out.txt set ACTION="%VARBIND8:~39,18%" echo %DATE% %TIME% "Action substring is %ACTION%" 2>&1 >> C:\Logs\trap_out.txt if %ACTION%=="Remote Copy Volume" ( echo Prepostlefthand_env_v2.9 >> C:\Logs\trap_out.txt c:\Windows\System32\WindowsPowerShell\v1.0\PowerShell.exe -sta -executionpolicy unrestricted -command " & 'C:\Scripts\prepostlefthand_env_v2.9.ps1' Backupsettings.ini ALL" 2>&1 >> C:\Logs\trap_out.txt ) ELSE ( echo %DATE% %TIME% Action substring is %ACTION% so exiting" 2>&1 >> C:\Logs\trap.out.txt ) exit

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  • Can't make my WCF extension work

    - by Sergio Romero
    I have a WCF solution that consists of the following class libraries: Exercise.Services: Contains the implementation classes for the services. Exercise.ServiceProxy: Contains the classes that are instantiated in the client. Exercise.HttpHost: Contains the services (*.svc files). I'm calling the service from a console application and the "first version" works really well so I took the next step which is to create a custom ServiceHostFactory, ServiceHost, and InstanceProvider so I can use constructor injection in my services as it is explained in this article. These classes are implemented in yet another class library: 4. Exercise.StructureMapWcfExtension Now even though I've modified my service this: <%@ ServiceHost Language="C#" Debug="true" Factory="Exercise.StructureMapWcfExtension.StructureMapServiceHostFactory" Service="Exercise.Services.PurchaseOrderService" %> I always get the following exception: System.ServiceModel.CommunicationException Security negotiation failed because the remote party did not send back a reply in a timely manner. This may be because the underlying transport connection was aborted. It fails in this line of code: public class PurchaseOrderProxy : ClientBase<IPurchaseOrderService>, IPurchaseOrderService { public PurchaseOrderResponse CreatePurchaseOrder(PurchaseOrderRequest purchaseOrderRequest) { return base.Channel.CreatePurchaseOrder(purchaseOrderRequest); //Fails here } } But that is not all, I added a trace to the web.config file and this is the error that appears in the log file: System.InvalidOperationException The service type provided could not be loaded as a service because it does not have a default (parameter-less) constructor. To fix the problem, add a default constructor to the type, or pass an instance of the type to the host. So this means that my ServiceHostFactory is never being hit, I even set a breakpoint in both its constructor and its method and they never get hit. I've added a reference of the StructureMapWcfExtension library to all the other ones (even the console client), one by one to no avail. I also tried to use the option in the host's web.config file to configure the factory like so: <serviceHostingEnvironment> <serviceActivations> <add service="Exercise.Services.PurchaseOrderService" relativeAddress="PurchaseOrderService.svc" factory="Exercise.StructureMapWcfExtension.StructureMapServiceHostFactory"/> </serviceActivations> </serviceHostingEnvironment> That didn't work either. Please I need help in getting this to work so I can incorporate it to our project. Thank you. UPDATE: Here's the service host factory's code: namespace Exercise.StructureMapWcfExtension { public class StructureMapServiceHostFactory : ServiceHostFactory { private readonly Container Container; public StructureMapServiceHostFactory() { Container = new Container(); new ContainerConfigurer().Configure(Container); } protected override ServiceHost CreateServiceHost(Type serviceType, Uri[] baseAddresses) { return new StructureMapServiceHost(Container, serviceType, baseAddresses); } } public class ContainerConfigurer { public void Configure(Container container) { container.Configure(r => r.For<IPurchaseOrderFacade>().Use<PurchaseOrderFacade>()); } } }

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  • An IOCP documentation interpretation question - buffer ownership ambiguity

    - by Poni
    Since I'm not a native English speaker I might be missing something so maybe someone here knows better than me. Taken from WSASend's doumentation at MSDN: lpBuffers [in] A pointer to an array of WSABUF structures. Each WSABUF structure contains a pointer to a buffer and the length, in bytes, of the buffer. For a Winsock application, once the WSASend function is called, the system owns these buffers and the application may not access them. This array must remain valid for the duration of the send operation. Ok, can you see the bold text? That's the unclear spot! I can think of two translations for this line (might be something else, you name it): Translation 1 - "buffers" refers to the OVERLAPPED structure that I pass this function when calling it. I may reuse the object again only when getting a completion notification about it. Translation 2 - "buffers" refer to the actual buffers, those with the data I'm sending. If the WSABUF object points to one buffer, then I cannot touch this buffer until the operation is complete. Can anyone tell what's the right interpretation to that line? And..... If the answer is the second one - how would you resolve it? Because to me it implies that for each and every data/buffer I'm sending I must retain a copy of it at the sender side - thus having MANY "pending" buffers (in different sizes) on an high traffic application, which really going to hurt "scalability". Statement 1: In addition to the above paragraph (the "And...."), I thought that IOCP copies the data to-be-sent to it's own buffer and sends from there, unless you set SO_SNDBUF to zero. Statement 2: I use stack-allocated buffers (you know, something like char cBuff[1024]; at the function body - if the translation to the main question is the second option (i.e buffers must stay as they are until the send is complete), then... that really screws things up big-time! Can you think of a way to resolve it? (I know, I asked it in other words above).

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  • What is the best way to archive data in a relational database?

    - by GenericTypeTea
    I have a bit of an issue with a particular aspect of a program I'm working on. I need the ability to archive (fix) a table so that a change anywhere in the system will not affect the results it returns. This is the basic structure of what I need to fix: Recipe --> Recipe (as sub recipe) Recipe --> Ingredients So, if I fix a Recipe, I need to ensure all the sub recipes (including all the sub recipes sub recipes and so forth) are fixed and all its ingredients are fixed. The problem is that the sub recipe and ingredients still need to be modifiable as they are used by other recipes that are not fixed. I came up with a solution whereby I serialize (with protobuf-net) a master object that deals with the recipe and all the sub recipes and ingredients and save the archive data to a table like follows: Archive{ ReferenceId, (i.e. RecipeId) ReferenceTypeId, (i.e. Recipe) ArchiveData varbinary(max) } Now, this works great and is almost perfect... however I totally forgot (I'd love to blame the agile development mentally, however this was just short sighted) that this information needs to be reported on. As far as I'm aware I can't think how I could inflate the serialized data back into my Recipe Object and use it in a Report. I'm using the standard SQL 2005 report services at the moment. Alternatively, I guess I could do the following: Duplicate every table and tag the word "Archive" on the end of the table name. This would then give me an area of specific archive data... but ignoring my simplified example, there'd actually be about 15 tables duplicated. Add a nullable, non-foreign key property called "CopiedFromId" to every table that contains fixed data and duplicate every record that the recipe (and all it's sub recipes and all their sub recipes) touches. Create some sort of denormalised structure that could be restored from at a later date to the original, unfixed recipe. Although I think this would be like option 1 and involve a lot of extra tables. Anyway, I'm at a total loss and do not like any of the ideas particularly. Can anyone please advise the best course of action? EDIT: Or 4) Create tables specific to what the report requires and populate them with the data when the user clicks the report button? This would cause about 4 extra tables for the report in question.

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  • Delaying emails in PHP to avoid exceeding server limit

    - by Andrew P.
    Okay, so here's my problem: I have a list of members on a website, and periodically one of the admins my site (who are not very web or tech savvy) will send a newsletter to the memberlist. My current memberlist is well over 800 individuals long. So, I wrote an email script that sends the email to the full memberlist, with the members listed in the Bcc header. However, I've discovered that my host server has a limit of 300 emails per hour, which I apparently exceed even though the members are listed in the Bcc field. (I wasn't previously aware that the behaviour of Bcc was to send separate emails for each name on the list...) After some thought, I've come to the conclusion that my only solution is to have my script send only the email to only the first 300 emails, wait an hour, and send a second email to the next three hundred, wait another hour, and so on until I've sent the email to the whole member list. Looking around on the internet, I've seen some other solutions people have come up with for delaying emails in PHP. Sleep() is obviously not an option, because I can't just leave the script open and running for 3 or four hours. I've seen some people suggest cron jobs, but I'm not sure how feasible it would be to create three new cron jobs every time I send an email, use them once, and then delete them afterward. The final (and what I think is the smartest) solution I've seen, is to have a table in my database to temporarily store the emails to be delayed and sent later, and then create a cron job that checks this sql table every hour or so, compares the timestamp of the row to the current timestamp, and then sends the email if an hour has passed. So I'm asking you all which method you would recommend. Is there an easier solution that I've completely looked over (aside from getting a different hosting plan. ha!), or is there a cleaner way to do it than the database / cron job approach? tl;dr: I have 800 emails to send in an hour on a server that limits me to 300/hr. Using PHP, find a way to get around this problem in a way that the person sending the email needs only to click "send."

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  • Is Java assert broken?

    - by BlairHippo
    While poking around the questions, I recently discovered the assert keyword in Java. At first, I was excited. Something useful I didn't already know! A more efficient way for me to check the validity of input parameters! Yay learning! But then I took a closer look, and my enthusiasm was not so much "tempered" as "snuffed-out completely" by one simple fact: you can turn assertions off.* This sounds like a nightmare. If I'm asserting that I don't want the code to keep going if the input listOfStuff is null, why on earth would I want that assertion ignored? It sounds like if I'm debugging a piece of production code and suspect that listOfStuff may have been erroneously passed a null but don't see any logfile evidence of that assertion being triggered, I can't trust that listOfStuff actually got sent a valid value; I also have to account for the possibility that assertions may have been turned off entirely. And this assumes that I'm the one debugging the code. Somebody unfamiliar with assertions might see that and assume (quite reasonably) that if the assertion message doesn't appear in the log, listOfStuff couldn't be the problem. If your first encounter with assert was in the wild, would it even occur to you that it could be turned-off entirely? It's not like there's a command-line option that lets you disable try/catch blocks, after all. All of which brings me to my question (and this is a question, not an excuse for a rant! I promise!): What am I missing? Is there some nuance that renders Java's implementation of assert far more useful than I'm giving it credit for? Is the ability to enable/disable it from the command line actually incredibly valuable in some contexts? Am I misconceptualizing it somehow when I envision using it in production code in lieu of statements like if (listOfStuff == null) barf();? I just feel like there's something important here that I'm not getting. *Okay, technically speaking, they're actually off by default; you have to go out of your way to turn them on. But still, you can knock them out entirely.

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  • How can I model the data in a multi-language data editor in WPF with MVVM?

    - by Patrick Szalapski
    Are there any good practices to follow when designing a model/ViewModel to represent data in an app that will view/edit that data in multiple languages? Our top-level class--let's call it Course--contains several collection properties, say Books and TopicsCovered, which each might have a collection property among its data. For example, the data needs to represent course1.Books.First().Title in different languages, and course1.TopicsCovered.First().Name in different languages. We want a app that can edit any of the data for one given course in any of the available languages--as well as edit non-language-specific data, perhaps the Author of a Book (i.e. course1.Books.First().Author). We are having trouble figuring out how best to set up the model to enable binding in the XAML view. For example, do we replace (in the single-language model) each String with a collection of LanguageSpecificString instances? So to get the author in the current language: course1.Books.First().Author.Where(Function(a) a.Language = CurrentLanguage).SingleOrDefault If we do that, we cannot easily bind to any value in one given language, only to the collection of language values such as in an ItemsControl. <TextBox Text={Binding Author.???} /> <!-- no way to bind to the current language author --> Do we replace the top-level Course class with a collection of language-specific Courses? So to get the author in the current language: course1.GetLanguage(CurrentLanguage).Books.First.Author If we do that, we can only easily work with one language at a time; we might want a view to show one language and let the user edit the other. <TextBox Text={Binding Author} /> <!-- good --> <TextBlock Text={Binding ??? } /> <!-- no way to bind to the other language author --> Also, that has the disadvantage of not representing language-neutral data as such; every property (such as Author) would seem to be in multiple languages. Even non-string properties would be in multiple languages. Is there an option in between those two? Is there another way that we aren't thinking of? I realize this is somewhat vague, but it would seem to be a somewhat common problem to design for. Note: This is not a question about providing a multilingual UI, but rather about actually editing multi-language data in a flexible way.

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  • GCC: Simple inheritance test fails

    - by knight666
    I'm building an open source 2D game engine called YoghurtGum. Right now I'm working on the Android port, using the NDK provided by Google. I was going mad because of the errors I was getting in my application, so I made a simple test program: class Base { public: Base() { } virtual ~Base() { } }; // class Base class Vehicle : virtual public Base { public: Vehicle() : Base() { } ~Vehicle() { } }; // class Vehicle class Car : public Vehicle { public: Car() : Base(), Vehicle() { } ~Car() { } }; // class Car int main(int a_Data, char** argv) { Car* stupid = new Car(); return 0; } Seems easy enough, right? Here's how I compile it, which is the same way I compile the rest of my code: /home/oem/android-ndk-r3/build/prebuilt/linux-x86/arm-eabi-4.4.0/bin/arm-eabi-g++ -g -std=c99 -Wall -Werror -O2 -w -shared -fshort-enums -I ../../YoghurtGum/src/GLES -I ../../YoghurtGum/src -I /home/oem/android-ndk-r3/build/platforms/android-5/arch-arm/usr/include -c src/Inheritance.cpp -o intermediate/Inheritance.o (Line breaks are added for clarity). This compiles fine. But then we get to the linker: /home/oem/android-ndk-r3/build/prebuilt/linux-x86/arm-eabi-4.4.0/bin/arm-eabi-gcc -lstdc++ -Wl, --entry=main, -rpath-link=/system/lib, -rpath-link=/home/oem/android-ndk-r3/build/platforms/android-5/arch-arm/usr/lib, -dynamic-linker=/system/bin/linker, -L/home/oem/android-ndk-r3/build/prebuilt/linux-x86/arm-eabi-4.4.0/lib/gcc/arm-eabi/4.4.0, -L/home/oem/android-ndk-r3/build/platforms/android-5/arch-arm/usr/lib, -rpath=../../YoghurtGum/lib/GLES -nostdlib -lm -lc -lGLESv1_CM -z /home/oem/android-ndk-r3/build/platforms/android-5/arch-arm/usr/lib/crtbegin_dynamic.o /home/oem/android-ndk-r3/build/platforms/android-5/arch-arm/usr/lib/crtend_android.o intermediate/Inheritance.o ../../YoghurtGum/bin/YoghurtGum.a -o bin/Galaxians.android As you can probably tell, there's a lot of cruft in there that isn't really needed. That's because it doesn't work. It fails with the following errors: intermediate/Inheritance.o:(.rodata._ZTI3Car[typeinfo for Car]+0x0): undefined reference to `vtable for __cxxabiv1::__si_class_type_info' intermediate/Inheritance.o:(.rodata._ZTI7Vehicle[typeinfo for Vehicle]+0x0): undefined reference to `vtable for __cxxabiv1::__vmi_class_type_info' intermediate/Inheritance.o:(.rodata._ZTI4Base[typeinfo for Base]+0x0): undefined reference to `vtable for __cxxabiv1::__class_type_info' collect2: ld returned 1 exit status make: *** [bin/Galaxians.android] Fout 1 These are the same errors I get from my actual application. If someone could explain to me where I went wrong in my test or what option or I forgot in my linker, I would be very, extremely grateful. Thanks in advance. UPDATE: When I make my destructors non-inlined, I get new and more exciting link errors: intermediate/Inheritance.o:(.rodata+0x78): undefined reference to `vtable for __cxxabiv1::__si_class_type_info' intermediate/Inheritance.o:(.rodata+0x90): undefined reference to `vtable for __cxxabiv1::__vmi_class_type_info' intermediate/Inheritance.o:(.rodata+0xb0): undefined reference to `vtable for __cxxabiv1::__class_type_info' collect2: ld returned 1 exit status make: *** [bin/Galaxians.android] Fout 1

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  • How to force VS 2010 to skip "builds" of projects which haven't changed?

    - by Ladislav Mrnka
    Our product's solution has more than 100+ projects (500+ksloc of production code). Most of them are C# projects but we also have few using C++/CLI to bridge communication with native code. Rebuilding the whole solution takes several minutes. That's fine. If I want to rebuilt the solution I expect that it will really take some time. What is not fine is time needed to build solution after full rebuild. Imagine I used full rebuild and know without doing any changes to to the solution I press Build (F6 or Ctrl+Shift+B). Why it takes 35s if there was no change? In output I see that it started "building" of each project - it doesn't perform real build but it does something which consumes significant amount of time. That 35s delay is pain in the ass. Yes I can improve the time by not using build solution but only build project (Shift+F6). If I run build project on particular test project I'm currently working on it will take "only" 8+s. It requires me to run project build on correct project (the test project to ensure dependent tested code is build as well). At least ReSharper test runner correctly recognizes that only this single project must be build and rerunning test usually contains only 8+s compilation. My current coding Kata is: don't touch Ctrl+Shift+B. The test project build will take 8s even if I don't do any changes. The reason why it takes 8s is because it also "builds" dependencies = in my case it "builds" more than 20 projects but I made changes only to unit test or single dependency! I don't want it to touch other projects. Is there a way to simply tell VS to build only projects where some changes were done and projects which are dependent on changed ones (preferably this part as another build option)? I worry you will tell me that it is exactly what VS is doing but in MS way ... I want to improve my TDD experience and reduce the time of compilation (in TDD the compilation can happen twice per minute). To make this even more frustrated I'm working in a team where most of developers used to work on Java projects prior to joining this one. So you can imagine how they are pissed off when they must use VS in contrast to full incremental compilation in Java. I don't require incremental compilation of classes. I expect working incremental compilation of solutions. Especially in product like VS 2010 Ultimate which costs several thousands dollars. I really don't want to get answers like: Make a separate solution Unload projects you don't need etc. I can read those answers here. Those are not acceptable solutions. We're not paying for VS to do such compromises.

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  • How can I fix this NavigationController and UIToolbar offset issue in Objective-C?

    - by editor
    I'm adding a couple of buttons to an already-existing NavigationController. The two buttons are added to a UIView, which is pushed onto the NavigationItem. The buttons stop and reload a UIWebView. Problem is that there's a slight offset issue that is making it all look pretty ugly. I wish I could set the UIToolbar to transparent or clear background but that doesn't seem to be an option. Can't seem to use negative offsets either. I've got color matching, but if you look closely there's 1px or 2px of highlighting up top that's causing a visual mismatch and then a slight offset at the bottom. Some relevant code below (based on this, inbound Googlers). What are my options to resolve this? // create a toolbar for the buttons UIToolbar* toolbar = [[UIToolbar alloc] initWithFrame:CGRectMake(0, 0, 100, 45)]; [toolbar setBarStyle: UIBarStyleDefault]; UIColor *colorForBar = [[UIColor alloc] initWithRed:.72 green:0 blue:0 alpha:0]; toolbar.tintColor = colorForBar; [colorForBar release]; //[toolbar setTranslucent:YES]; // create an array for the buttons NSMutableArray* buttons = [[NSMutableArray alloc] initWithCapacity:3]; // create a standard reload button UIBarButtonItem *reloadButton = [[UIBarButtonItem alloc] initWithBarButtonSystemItem:UIBarButtonSystemItemRefresh target:self action:@selector(reload)]; reloadButton.style = UIBarButtonItemStyleBordered; [buttons addObject:reloadButton]; [reloadButton release]; // create a spacer between the buttons UIBarButtonItem *spacer = [[UIBarButtonItem alloc] initWithBarButtonSystemItem:UIBarButtonSystemItemFixedSpace target:nil action:nil]; [buttons addObject:spacer]; [spacer release]; // create a standard delete button with the trash icon UIBarButtonItem *stopButton = [[UIBarButtonItem alloc] initWithBarButtonSystemItem:UIBarButtonSystemItemStop target:self action:@selector(stopLoading)]; stopButton.style = UIBarButtonItemStyleBordered; [buttons addObject:stopButton]; [stopButton release]; // put the buttons in the toolbar and release them [toolbar setItems:buttons animated:NO]; [buttons release]; // place the toolbar into the navigation bar self.navigationItem.rightBarButtonItem = [[UIBarButtonItem alloc] initWithCustomView:toolbar]; [toolbar release];

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  • difference equations in MATLAB - why the need to switch signs?

    - by jefflovejapan
    Perhaps this is more of a math question than a MATLAB one, not really sure. I'm using MATLAB to compute an economic model - the New Hybrid ISLM model - and there's a confusing step where the author switches the sign of the solution. First, the author declares symbolic variables and sets up a system of difference equations. Note that the suffixes "a" and "2t" both mean "time t+1", "2a" means "time t+2" and "t" means "time t": %% --------------------------[2] MODEL proc-----------------------------%% % Define endogenous vars ('a' denotes t+1 values) syms y2a pi2a ya pia va y2t pi2t yt pit vt ; % Monetary policy rule ia = q1*ya+q2*pia; % ia = q1*(ya-yt)+q2*pia; %%option speed limit policy % Model equations IS = rho*y2a+(1-rho)yt-sigma(ia-pi2a)-ya; AS = beta*pi2a+(1-beta)*pit+alpha*ya-pia+va; dum1 = ya-y2t; dum2 = pia-pi2t; MPs = phi*vt-va; optcon = [IS ; AS ; dum1 ; dum2; MPs]; He then computes the matrix A: %% ------------------ [3] Linearization proc ------------------------%% % Differentiation xx = [y2a pi2a ya pia va y2t pi2t yt pit vt] ; % define vars jopt = jacobian(optcon,xx); % Define Linear Coefficients coef = eval(jopt); B = [ -coef(:,1:5) ] ; C = [ coef(:,6:10) ] ; % B[c(t+1) l(t+1) k(t+1) z(t+1)] = C[c(t) l(t) k(t) z(t)] A = inv(C)*B ; %(Linearized reduced form ) As far as I understand, this A is the solution to the system. It's the matrix that turns time t+1 and t+2 variables into t and t+1 variables (it's a forward-looking model). My question is essentially why is it necessary to reverse the signs of all the partial derivatives in B in order to get this solution? I'm talking about this step: B = [ -coef(:,1:5) ] ; Reversing the sign here obviously reverses the sign of every component of A, but I don't have a clear understanding of why it's necessary. My apologies if the question is unclear or if this isn't the best place to ask.

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  • jqtransform and collapsed DIV

    - by Marco
    Hello, I am using a jQuery plugin called: jqtransform This plugin uses JavaScript to apply CSS styles to form elements. The problem that I have consists in the following scenario: I’m building a search page, with a advanced search option. When the page loads, the div called “advancedSearch” is hidden, and it will only show if the user clicks a element. Inside the div#advancedSearch I have several form elements. However, if I hide the div#advancedSearch with the CSS style: “diplay:none;”, the jqtransform plugin doesn’t work correctly with the elements that are hidden. So my solution was to hide the div#advancedsearch with JavaScript. This actually works, and it does not matter if it’s done after the document is ready or not. But… with the JavaScript solution, the div#advancedSearch stays visible for a couple of milliseconds… which is visually annoying. So I was wondering if the solution to this problem would be in the CSS, or in correcting the jqtransform plugin, or even in finding a way to immediately hide the div#advancedSearch with JS making it immediately hidden. UPDATE 1 After jeerose comment I decided to place here my function (please note that the <%= % are ASP.Net tags, that I use to get the images path) $('.toggleAdvancedSearch').click(function() { $('#advancedSearchWrap').slideToggle(250); $('form.jqtransform').jqTransform({ imgPath: '<%= ResolveClientUrl("~/masterpages/img/jqtransform/") %>' }); return false; }); UPDATE 2 To test the problem, I did the following: Added another element to the page, with the ID “applyStyle”, and onClick I call the $('form').jqTransform(); Disabled the the $('form').jqTransform(); from the load of the page. If I press the a#applyStyle, before expanding the div#advancedSearch I get the same problem that I had. But if I expand the the div#advancedSearch and press the the a#applyStyle after, the problem is solved. However, if I run the page with the $('form').jqTransform(); function on the load, I cannot reapply it after with the pressing of the a#applyStyle. I think that the solution could be: disabling the all the elements that are inside the div#advancedSearch, and on the same function that expands the div, make It also apply the styles to the elements that are inside the div#advancedSearch. However, I don’t know how to do this (nether if this will work). PS: This seems to be a known issue with the plugin, but I cannot wait indefinitely for a solution.

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