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  • Unit finalization order for application, compiled with run-time packages?

    - by Alexander
    I need to execute my code after finalization of SysUtils unit. I've placed my code in separate unit and included it first in uses clause of dpr-file, like this: project Project1; uses MyUnit, // <- my separate unit SysUtils, Classes, SomeOtherUnits; procedure Test; begin // end; begin SetProc(Test); end. MyUnit looks like this: unit MyUnit; interface procedure SetProc(AProc: TProcedure); implementation var Test: TProcedure; procedure SetProc(AProc: TProcedure); begin Test := AProc; end; initialization finalization Test; end. Note that MyUnit doesn't have any uses. This is usual Windows exe, no console, without forms and compiled with default run-time packages. MyUnit is not part of any package (but I've tried to use it from package too). I expect that finalization section of MyUnit will be executed after finalization section of SysUtils. This is what Delphi's help tells me. However, this is not always the case. I have 2 test apps, which differs a bit by code in Test routine/dpr-file and units, listed in uses. MyUnit, however, is listed first in all cases. One application is run as expected: Halt0 - FinalizeUnits - ...other units... - SysUtils's finalization - MyUnit's finalization - ...other units... But the second is not. MyUnit's finalization is invoked before SysUtils's finalization. The actual call chain looks like this: Halt0 - FinalizeUnits - ...other units... - SysUtils's finalization (skipped) - MyUnit's finalization - ...other units... - SysUtils's finalization (executed) Both projects have very similar settings. I tried a lot to remove/minimize their differences, but I still do not see a reason for this behaviour. I've tried to debug this and found out that: it seems that every unit have some kind of reference counting. And it seems that InitTable contains multiply references to the same unit. When SysUtils's finalization section is called first time - it change reference counter and do nothing. Then MyUnit's finalization is executed. And then SysUtils is called again, but this time ref-count reaches zero and finalization section is executed: Finalization: // SysUtils' finalization 5003B3F0 55 push ebp // here and below is some form of stub 5003B3F1 8BEC mov ebp,esp 5003B3F3 33C0 xor eax,eax 5003B3F5 55 push ebp 5003B3F6 688EB50350 push $5003b58e 5003B3FB 64FF30 push dword ptr fs:[eax] 5003B3FE 648920 mov fs:[eax],esp 5003B401 FF05DCAD1150 inc dword ptr [$5011addc] // here: some sort of reference counter 5003B407 0F8573010000 jnz $5003b580 // <- this jump skips execution of finalization for first call 5003B40D B8CC4D0350 mov eax,$50034dcc // here and below is actual SysUtils' finalization section ... Can anyone can shred light on this issue? Am I missing something?

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  • Best Practices / Patterns for Enterprise Protection/Remediation of SSNs (Social Security Numbers)

    - by Erik Neu
    I am interested in hearing about enterprise solutions for SSN handling. (I looked pretty hard for any pre-existing post on SO, including reviewing the terriffic SO automated "Related Questions" list, and did not find anything, so hopefully this is not a repeat.) First, I think it is important to enumerate the reasons systems/databases use SSNs: (note—these are reasons for de facto current state—I understand that many of them are not good reasons) Required for Interaction with External Entities. This is the most valid case—where external entities your system interfaces with require an SSN. This would typically be government, tax and financial. SSN is used to ensure system-wide uniqueness. SSN has become the default foreign key used internally within the enterprise, to perform cross-system joins. SSN is used for user authentication (e.g., log-on) The enterprise solution that seems optimum to me is to create a single SSN repository that is accessed by all applications needing to look up SSN info. This repository substitutes a globally unique, random 9-digit number (ASN) for the true SSN. I see many benefits to this approach. First of all, it is obviously highly backwards-compatible—all your systems "just" have to go through a major, synchronized, one-time data-cleansing exercise, where they replace the real SSN with the alternate ASN. Also, it is centralized, so it minimizes the scope for inspection and compliance. (Obviously, as a negative, it also creates a single point of failure.) This approach would solve issues 2 and 3, without ever requiring lookups to get the real SSN. For issue #1, authorized systems could provide an ASN, and be returned the real SSN. This would of course be done over secure connections, and the requesting systems would never persist the full SSN. Also, if the requesting system only needs the last 4 digits of the SSN, then that is all that would ever be passed. Issue #4 could be handled the same way as issue #1, though obviously the best thing would be to move away from having users supply an SSN for log-on. There are a couple of papers on this: UC Berkely: http://bit.ly/bdZPjQ Oracle Vault: bit.ly/cikbi1

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  • Reliable and fast way to convert a zillion ODT files in PDF?

    - by Marco Mariani
    I need to pre-produce a million or two PDF files from a simple template (a few pages and tables) with embedded fonts. Usually, I would stay low level in a case like this, and compose everything with a library like ReportLab, but I joined late in the project. Currently, I have a template.odt and use markers in the content.xml files to fill with data from a DB. I can smoothly create the ODT files, they always look rigth. For the ODT to PDF conversion, I'm using openoffice in server mode (and PyODConverter w/ named pipe), but it's not very reliable: in a batch of documents, there is eventually a point after which all the processed files are converted into garbage (wrong fonts and letters sprawled all over the page). Problem is not predictably reproducible (does not depend on the data), happens in OOo 2.3 and 3.2, in Ubuntu, XP, Server 2003 and Windows 7. My Heisenbug detector is ticking. I tried to reduce the size of batches and restarting OOo after each one; still, a small percentage of the documents are messed up. Of course I'll write about this on the Ooo mailing lists, but in the meanwhile, I have a delivery and lost too much time already. Where do I go? Completely avoid the ODT format and go for another template system. Suggestions? Anything that takes a few seconds to run is way too slow. OOo takes around a second and it sums to 15 days of processing time. I had to write a program for clustering the jobs over several clients. Keep the format but go for another tool/program for the conversion. Which one? There are many apps in the shareware or commercial repositories for windows, but trying each one is a daunting task. Some are too slow, some cannot be run in batch without buying it first, some cannot work from command line, etc. Open source tools tend not to reinvent the wheel and often depend on openoffice. Converting to an intermediate .DOC format could help to avoid the OOo bug, but it would double the processing time and complicate a task that is already too hairy. Try to produce the PDFs twice and compare them, discarding the whole batch if there's something wrong. Although the documents look equal, I know of no way to compare the binary content. Restart OOo after processing each document. it would take a lot more time to produce them it would lower the percentage of the wrong files, and make it very hard to identify them. Go for ReportLab and recreate the pages programmatically. This is the approach I'm going to try in a few minutes. Learn to properly format bulleted lists Thanks a lot.

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  • Logging in with WebFinger and OpenID

    - by Ryan
    I would like to apologize in advance for the ugly formatting. In order to talk about the problem, I need to be posting a bunch of URLs, but the excessive URLs and my lack of reputation makes StackOverflow think I could be a spammer. Any instance of 'ht~tp' is supposed to be 'http'. '{dot}' is supposed to be '.' and '{colon}' is supposed to be ':'. Also, my lack of reputation has prevented me from tagging my question with 'webfinger' and 'google-profiles'. Onto my question: I am messing around with WebFinger and trying to create a small rails app that enables a user to log in using nothing but their WebFinger account. I can succesfully finger myself, and I get back an XRD file with the following snippet: Link rel="ht~tp://specs{dot}openid{dot}net/auth/2.0/provider" href="ht~tp://www{dot}google{dot}com/profiles/{redacted}"/ Which, to me, reads, "I have an OpenID 2.0 login at the url: ht~tp://www{dot}google{dot}com/profiles/{redacted}". But when I try to use that URL to log in, I get the following error OpenID::DiscoveryFailure (Failed to fetch identity URL ht~tp://www{dot}google{dot}com/profiles/{redacted} : Error encountered in redirect from ht~tp://www{dot}google{dot}com/profiles/{redacted}: Error fetching /profiles/{Redacted}: Connection refused - connect(2)): When I replace the profile URL with 'ht~tps://www{dot}google{dot}com/accounts/o8/id', the login works perfectly. here is the code that I am using (I'm using RedFinger as a plugin, and JanRain's ruby-openid, installed without the gem) require "openid" require 'openid/store/filesystem.rb' class SessionsController < ApplicationController def new @session = Session.new #render a textbox requesting a webfinger address, and a submit button end def create ####################### # # Pay Attention to this section right here # ####################### #use given webfinger address to retrieve openid login finger = Redfinger.finger(params[:session][:webfinger_address]) openid_url = finger.open_id.first.to_s #openid_url is now: ht~tp://www{dot}google{dot}com/profiles/{redacted} #Get needed info about the acquired OpenID login file_store = OpenID::Store::Filesystem.new("./noncedir/") consumer = OpenID::Consumer.new(session,file_store) response = consumer.begin(openid_url) #ERROR HAPPENS HERE #send user to OpenID login for verification redirect_to response.redirect_url('ht~tp://localhost{colon}3000/','ht~tp://localhost{colon}3000/sessions/complete') end def complete #interpret return parameters file_store = OpenID::Store::Filesystem.new("./noncedir/") consumer = OpenID::Consumer.new(session,file_store) response = consumer.complete params case response.status when OpenID::SUCCESS session[:openid] = response.identity_url #redirect somehwere here end end end Is it possible for me to use the URL I received from my WebFinger to log in with OpenID?

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  • How can I make Swig correctly wrap a char* buffer that is modified in C as a Java Something-or-other

    - by Ukko
    I am trying to wrap some legacy code for use in Java and I was quite happy to see that Swig was able to handle the header file and it generate a great wrapper that almost works. Now I am looking for the deep magic that will make it really work. In C I have a function that looks like this DLL_IMPORT int DustyVoodoo(char *buff, int len, char *curse); This integer returned by this function is an error code in case it fails. The arguments are buff is a character buffer len is the length of the data in the buffer curse the another character buffer that contains the result of calling DustyVoodoo So, you can see where this is going, the result is actually coming back via the third argument. Also len is confusing since it may be the length of both buffers, they are always allocated as being the same size in calling code but given what DustyVoodoo does I don't think that they need be the same. To be safe both buffers should be the same size in practice, say 512 chars. The C code generated for the binding is as follows: SWIGEXPORT jint JNICALL Java_pemapiJNI_DustyVoodoo(JNIEnv *jenv, jclass jcls, jstring jarg1, jint jarg2, jstring jarg3) { jint jresult = 0 ; char *arg1 = (char *) 0 ; int arg2 ; char *arg3 = (char *) 0 ; int result; (void)jenv; (void)jcls; arg1 = 0; if (jarg1) { arg1 = (char *)(*jenv)->GetStringUTFChars(jenv, jarg1, 0); if (!arg1) return 0; } arg2 = (int)jarg2; arg3 = 0; if (jarg3) { arg3 = (char *)(*jenv)->GetStringUTFChars(jenv, jarg3, 0); if (!arg3) return 0; } result = (int)PemnEncrypt(arg1,arg2,arg3); jresult = (jint)result; if (arg1) (*jenv)->ReleaseStringUTFChars(jenv, jarg1, (const char *)arg1); if (arg3) (*jenv)->ReleaseStringUTFChars(jenv, jarg3, (const char *)arg3); return jresult; } It is correct for what it does; however, it misses the fact that cursed is not just an input, it is altered by the function and should be returned as an output. It also does not know that the java Strings are really buffers and should be backed by a suitably sized array. I think that Swig can do the right thing here, I just can't figure out from the documentation how to tell Swig what it needs to know. Any typemap masers in the house?

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  • Complex relationship between tables in NHibernate

    - by Ilya Kogan
    Hi all, I'm writing a Fluent NHibernate mapping for a legacy Oracle database. The challenge is that the tables have composite primary keys. If I were at total freedom, I would redesign the relationships and auto-generate primary keys, but other applications must write to the same database and read from it, so I cannot do it. These are the two tables I'll focus on: Example data Trips table: 1, 10:00, 11:00 ... 1, 12:00, 15:00 ... 1, 16:00, 19:00 ... 2, 12:00, 13:00 ... 3, 9:00, 18:00 ... Faults table: 1, 13:00 ... 1, 23:00 ... 2, 12:30 ... In this case, vehicle 1 made three trips and has two faults. The first fault happened during the second trip, and the second fault happened while the vehicle was resting. Vehicle 2 had one trip, during which a fault happened. Constraints Trips of the same vehicle never overlap. So the tables have an optional one-to-many relationship, because every fault either happens during a trip or it doesn't. If I wanted to join them in SQL, I would write: select ... from Faults left outer join Trips on Faults.VehicleId = Trips.VehicleId and Faults.FaultTime between Trips.TripStartTime and Trips.TripEndTime and then I'd get a dataset where every fault appears exactly once (one-to-many as I said). Note that there is no Vehicles table, and I don't need one. But I did create a view that contains all VehicleIds from both tables, so I can use it as a junction table. What am I actually looking for? The tables are huge because they cover years of data, and every time I only need to fetch a range of a few hours. So I need a mapping and a criteria that will run something like the following SQL underneath: select ... from Faults left outer join Trips on Faults.VehicleId = Trips.VehicleId and Faults.FaultTime between Trips.TripStartTime and Trips.TripEndTime where Faults.FaultTime between :p0 and :p1 Do you have any ideas how to achieve it? Note 1: Currently the application shouldn't write to the database, so persistence is not a must, although if the mapping supports persistence, it may help at some point in the future. Note 2: I know it's a tough one, so if you give me a great answer, you will be properly rewarded :) Thank you for reading this long question, and now I only hope for the best :)

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  • overloading friend operator<< for template class

    - by starcorn
    Hello, I have read couple of the question regarding my problem on stackoverflow now, and none of it seems to solve my problem. Or I maybe have done it wrong... The overloaded << if I make it into an inline function. But how do I make it work in my case? warning: friend declaration std::ostream& operator<<(std::ostream&, const D<classT>&)' declares a non-template function warning: (if this is not what you intended, make sure the function template has already been declared and add <> after the function name here) -Wno-non-template-friend disables this warning /tmp/cc6VTWdv.o:uppgift4.cc:(.text+0x180): undefined reference to operator<<(std::basic_ostream<char, std::char_traits<char> >&, D<int> const&)' collect2: ld returned 1 exit status template <class T> T my_max(T a, T b) { if(a > b) return a; else return b; } template <class classT> class D { public: D(classT in) : d(in) {}; bool operator>(const D& rhs) const; classT operator=(const D<classT>& rhs); friend ostream& operator<< (ostream & os, const D<classT>& rhs); private: classT d; }; int main() { int i1 = 1; int i2 = 2; D<int> d1(i1); D<int> d2(i2); cout << my_max(d1,d2) << endl; return 0; } template <class classT> ostream& operator<<(ostream &os, const D<classT>& rhs) { os << rhs.d; return os; }

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  • How to find relation between change in latitudes at centre of map and top/bottom

    - by Imran
    Hi, Im having little trouble finding a relation between the movement at centre and edge of a circle, Im doing for panning world map,my map extent is 180,89:-180,-89, my map pans by adding change(dx,dY) to its extents and not its centre. Now a situation has arrrised where I have to move the map to a specific centre, to calculate the change in longitudes is very easy and simple, but its the change in lattitudes that has caused problem. It seems the change in centreY of map is more than the change at edge of the mapY, or simply if I have to move the map centre from 0long,0lat to 73long,33lat, for dX I simply get 73, but for dY apparently it looks 33 but if i add 33 to top of map that is 89 , it will be 122 which is incorrect since Latitudes are between 90 and -90 . It seems a case a projection of a circle on 2D plane where the edge of circle since is moving backward due to angle expereinces less change and the centre expereinces more change, now is there a relation between these two factors? I tried converting the difference between OriginY and destinationY into radians and then add to Top and Bottom of Map, but it did'nt really work for me. Please note that the map is project on a virtual canvas whose width starts from 256 and increases by 256*2^z , z=0 is default and whole world is visible at that extent of canvas code: public void moveMapTo(double destinationLongitude,double destinationLattitude) // moves map to the new centre { double dXLong=destinationLongitude-centreLongitude; double atanhsinO = atanh(Math.sin(destinationLattitude * Math.PI / 180.00)); double atanhsinD = atanh(Math.sin(centreLatitude * Math.PI / 180.00)); double atanhCentre = (atanhsinD + atanhsinO) / 2; double latitudeSpan =destinationLattitude - centreLatitude; double radianOfCentreLatitude = Math.atan(Math.sinh(atanhCentre)); double dXLat=latitudeSpan / Math.cos(radianOfCentreLatitude); dXLat*=getLattitudeSpan()*(Math.PI/180); <--- HERE IS THE PORBLEM System.out.println("dxLong:"+dXLong+"_dxLat:"+dXLat); mapLeft+=dXLong; mapRight+=dXLong; mapTop+=dXLat; mapBottom+=dXLat; } ////latitude span function private double getLattitudeSpan() { double latitudeSpan = mapTop - mapBottom; latitudeSpan = latitudeSpan / Math.cos(radianOfCentreLatitude); return Math.abs(latitudeSpan); } //ht

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  • A continued saga of C# interoprability with unmanaged C++

    - by Gilad
    After a day of banging my head against the wall both literally and metaphorically, I plead for help: I have an unmanaged C++ project, which is compiled as a DLL. Let's call it CPP Project. It currently works in an unmanaged environment. In addition, I have created a WPF project, that shall be called WPF Project. This project is a simple and currently almost empty project. It contains a single window and I want it to use code from Project 1. For that, I have created a CLR C++ project, which shall be called Interop Project and is also compiled as a DLL. For simplicity I will attach some basic testing code I have boiled down to the basics. CPP Project has the following two testing files: tester.h #pragma once extern "C" class __declspec(dllexport) NativeTester { public: void NativeTest(); }; tester.cpp #include "tester.h" void NativeTester::NativeTest() { int i = 0; } Interop Project has the following file: InteropLib.h #pragma once #include <tester.h> using namespace System; namespace InteropLib { public ref class InteropProject { public: static void Test() { NativeTester nativeTester; nativeTester.NativeTest(); } }; } Lastly, WPF Project has a single window refrencing Interop Project: MainWindow.xaml.cs using System; using System.Windows; using InteropLib; namespace AppGUI { public partial class MainWindow : Window { public MainWindow() { InitializeComponent(); InteropProject.Test(); } } } And the XAML itself has an empty window (default created). Once I am trying to run the WPF project, I get the following error: System.Windows.Markup.XamlParseException: 'The invocation of the constructor on type 'AppGUI.MainWindow' that matches the specified binding constraints threw an exception.' Line number '3' and line position '9'. --- System.IO.FileNotFoundException: Could not load file or assembly 'InteropLib.dll' or one of its dependencies. The specified module could not be found. at AppGUI.MainWindow..ctor() Interestingly enough, if I do not export the class from CPP Project, I do not get this error. Say, if i change tester.h to: #pragma once class NativeTester { public: void NativeTest() { int i = 0; } }; However, in this case I cannot use my more complex classes. If I move my implementation to a cpp file like before, I get unresolved linkage errors due to my not exporting my code. The C++ code I want to actually use is large and has many classes and is object oriented, so I can't just move all my implementation to the h files. Please help me understand this horrific error I've been trying resolve without success. Thanks.

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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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  • Multiline editable textarea in SVG

    - by Timo
    I'm trying to implement multiline editable textfield in SVG. I have the following code in http://jsfiddle.net/ca4d3/ : <svg width="1000" height="1000" overflow="scroll"> <g transform="rotate(5)"> <rect width="300" height="400" fill="#22DD22" fill-opacity="0.5"/> </g> <foreignObject x="10" y="10" overflow="visible" width="10000" height="10000" requiredFeatures="http://www.w3.org/TR/SVG11/feature#Extensibility"> <p style="display:table-cell;padding:10px;border:1px solid red; background-color:white;opacity:0.5;font-family:Verdana; font-size:20px;white-space: pre; word-wrap: normal; overflow: visible; overflow-y: visible; overflow-x:visible;" contentEditable="true" xmlns="http://www.w3.org/1999/xhtml"> Write here some text. Be smart and select some word. If you wanna be really COOL, paste here something cool! </p> </foreignObject> </svg> In newest Chrome, Safari and Firefox the code works in some way, but in Opera and IE 9 not. The goal is that: 0) Works in newest Chrome, Safari, Firefox, Opera and IE and if ever possible in some pads. 1) White-spaces are preserved and text wraps only on newline char (works in Chrome, Safari and Firefox, but not in Opera and IE 9 *). 2) The textfield is editable (in the same reliable and stabile way as textareas and contenteditable p elements in html) and height and width is expanded to fit text (works in Chrome, Safari and Firefox, but not in Opera and IE 9 *). 3) Texfield can be transformed (rotated, skewed, translated) while maintaining text editability (Tested rotation, but not work in any browser *). EDIT: Foreignobject rotation works on Firefox 15.0.1, but not in Safari 5.1.7 (6534.57.2), Chrome 22.0.1229.79, Opera 12.02, IE 9. Tested on Mac OS X 10.6.8. 4) Textfield can be clipped and masked while not necessarily maintaining text editability (not yet tested). *) using above code These all can be achieved using Flash, but Flash has so severe problems that it is not suitable for my purposes (after every little change in code, all have to be compiled again using Flex, which is slow, font size has limits, tracking technique is pixeloriented, not relative to em size etc.) and there still are differences across platforms. And I want to give a try to SVG! GUESTION: Can I achieve my goals 0-4 with current SVG support in browsers? Is coming SVG 2.0 for some help in this case? EDIT: Changed display:table to display:table-cell (and added new jsfiddle), because display:table made the field to loses focus when pressed arrow-up on first text row.

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  • Lazy loading the addthis script? (or lazy loading external js content dependent on already fired eve

    - by Keith Bentrup
    I want to have the addthis widget available for my users, but I want to lazy load it so that my page loads as quickly as possible. However, after trying it via a script tag and then via my lazy loading method, it appears to only work via the script tag. In the obfuscated code, I see something that looks like it's dependent on the DOMContentLoaded event (at least for firefox). Since the DOMContentLoaded event has already fired, the widget doesn't render properly. What to do? I could just use a script tag (slower)... or could I fire (in a cross browser way) the DOMContentLoaded (or equivalent) event? I have a feeling this may not be possible b/c I believe that (like jQuery) there are multiple tests of the content ready event, and so multiple simulated events would have to occur. Nonetheless, this is an interesting problem b/c I have seen a couple widgets now assume that you are including their stuff via static script tags. It would be nice if they wrote code that was more useful to developers concerned about speed, but until then, is there a work around?? And/or are any of my assumptions wrong? Edit: Because the 1st answer to the question seemed to miss the point of my problem, I wanted to clarify the situation. This is about a specific problem. I'm not looking for yet another lazy load script or check if some dependencies are loaded script. Specifically this problem deals with external widgets that you do not have control over and may or may not be obfuscated delaying the load of the external widgets until they are needed or at least, til substantially after everything else has been loaded including other deferred elements b/c of the how the widget was written, precludes existing, typical lazy loading paradigms While it's esoteric, I have seen it happen with a couple widgets - where the widget developers assume that you're just willing to throw in another script tag at the bottom of the page. I'm looking to save those 500-1000 ms** though as numerous studies by yahoo, google, and amazon show it to be important to your user's experience. **My testing with hammerhead and personal experience indicates that this will be my savings in this case.

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  • replace XmlSlurper tag with arbitrary XML

    - by Misha Koshelev
    Dear All: I am trying to replace specific XmlSlurper tags with arbitrary XML strings. The best way I have managed to come up with to do this is: #!/usr/bin/env groovy import groovy.xml.StreamingMarkupBuilder def page=new XmlSlurper(new org.cyberneko.html.parsers.SAXParser()).parseText(""" <html> <head></head> <body> <one attr1='val1'>asdf</one> <two /> <replacemewithxml /> </body> </html> """.trim()) import groovy.xml.XmlUtil def closure closure={ bind,node-> if (node.name()=="REPLACEMEWITHXML") { bind.mkp.yieldUnescaped "<replacementxml>sometext</replacementxml>" } else { bind."${node.name()}"(node.attributes()) { mkp.yield node.text() node.children().each { child-> closure(bind,child) } } } } println XmlUtil.serialize( new StreamingMarkupBuilder().bind { bind-> closure(bind,page) } ) However, the only problem is the text() element seems to capture all child text nodes, and thus I get: <?xml version="1.0" encoding="UTF-8"?> <HTML>asdf<HEAD/> <BODY>asdf<ONE attr1="val1">asdf</ONE> <TWO/> <replacementxml>sometext</replacementxml> </BODY> </HTML> Any ideas/help much appreciated. Thank you! Misha p.s. Also, out of curiosity, if I change the above to the "Groovier" notation as follows, the groovy compiler thinks I am trying to access the ${node.name()} member of my test class. Is there a way to specify this is not the case while still not passing the actual builder object? Thank you! :) def closure closure={ node-> if (node.name()=="REPLACEMEWITHXML") { mkp.yieldUnescaped "<replacementxml>sometext</replacementxml>" } else { "${node.name()}"(node.attributes()) { mkp.yield node.text() node.children().each { child-> closure(child) } } } } println XmlUtil.serialize( new StreamingMarkupBuilder().bind { closure(page) } )

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  • Positioning element under another

    - by Cedar Jensen
    I am not an expert web-dev so please bear with me here. I would like to display a banner style header for a page with the top part taken up by an image that is 275x116 and then a horizontal menu bar (styled using ul items) appearing at 70% from the top of the banner. How would I set this up so that the banner appears underneath my navigation? Currently, a portion of the left side of my menu bar sits underneath the image but I'd like it to be the opposite so the menu bar is above the image, some thing like this: ============= <start of header> =========== -------- | img | | | | Horizontal menu | | -------- ============= <end of header> =========== My css: #header { background-color: green; border: 0; margin: 0; padding: 0; overflow: hidden; width: 100%; height: 120px; } #logo { background: green url(images/logo.png) no-repeat scroll 0 0; margin: 0px 0px; border: 1px solid white; left: 20px; top: 20px; width: 275px; height: 116px; position: absolute; z-index: -1000; } .container { border:1px solid grey; margin-left:auto; margin-right:auto; width:960px; } My Html: <body> <div id="header"> <div id="logo"> </div> <div class="container" id="primaryNavbar"> <ul> <li><a href="#">Home</a></li> <li><a href="#">Books</a></li> <li><a href="#">Shows</a></li> <li><a href="#">Movies</a></li> </ul> <div class="clear">&nbsp;</div> </div> <!-- end of container --> </div> <!-- end of header --> </body> I thought that setting the position to "absolute" for the logo element and adding in a very low z-index would achieve this but that isn't the case here. Any suggestions?

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  • Accidental Complexity in OpenSSL HMAC functions

    - by Hassan Syed
    SSL Documentation Analaysis This question is pertaining the usage of the HMAC routines in OpenSSL. Since Openssl documentation is a tad on the weak side in certain areas, profiling has revealed that using the: unsigned char *HMAC(const EVP_MD *evp_md, const void *key, int key_len, const unsigned char *d, int n, unsigned char *md, unsigned int *md_len); From here, shows 40% of my library runtime is devoted to creating and taking down **HMAC_CTX's behind the scenes. There are also two additional function to create and destroy a HMAC_CTX explicetly: HMAC_CTX_init() initialises a HMAC_CTX before first use. It must be called. HMAC_CTX_cleanup() erases the key and other data from the HMAC_CTX and releases any associated resources. It must be called when an HMAC_CTX is no longer required. These two function calls are prefixed with: The following functions may be used if the message is not completely stored in memory My data fits entirely in memory, so I choose the HMAC function -- the one whose signature is shown above. The context, as described by the man page, is made use of by using the following two functions: HMAC_Update() can be called repeatedly with chunks of the message to be authenticated (len bytes at data). HMAC_Final() places the message authentication code in md, which must have space for the hash function output. The Scope of the Application My application generates a authentic (HMAC, which is also used a nonce), CBC-BF encrypted protocol buffer string. The code will be interfaced with various web-servers and frameworks Windows / Linux as OS, nginx, Apache and IIS as webservers and Python / .NET and C++ web-server filters. The description above should clarify that the library needs to be thread safe, and potentially have resumeable processing state -- i.e., lightweight threads sharing a OS thread (which might leave thread local memory out of the picture). The Question How do I get rid of the 40% overhead on each invocation in a (1) thread-safe / (2) resume-able state way ? (2) is optional since I have all of the source-data present in one go, and can make sure a digest is created in place without relinquishing control of the thread mid-digest-creation. So, (1) can probably be done using thread local memory -- but how do I resuse the CTX's ? does the HMAC_final() call make the CTX reusable ?. (2) optional: in this case I would have to create a pool of CTX's. (3) how does the HMAC function do this ? does it create a CTX in the scope of the function call and destroy it ? Psuedocode and commentary will be useful.

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  • Why is cell phone software still so primitive?

    - by Tomislav Nakic-Alfirevic
    I don't do mobile development, but it strikes me as odd that features like this aren't available by default on most phones: full text search: searches all address book contents, messages, anything else being a plus better call management: e.g. a rotating audio call log, meaning you always have the last N calls recorded for your listening pleasure later (your little girl just said her first "da-da" while you were on a business trip, you had a telephone job interview, you received complex instructions to do something etc.) bluetooth remote control (like e.g. anyRemote, but available by default on a bluetooth phone) no multitasking capabilities worth mentioning and in general no e.g. weekly software updates, making the phone much more usable (even if it had to be done over USB, rather than over the network). I'm sure I was dumbfounded by the lack or design of other features as well, but they don't come to mind right now. To clarify, I'm not talking about smartphones here: my plain, 2-year old phone has a CPU an order of magnitude faster than my first PC, about as much storage space and it's ridiculous how bad (slow, unwieldy) the software is and it's not one phone or one manufacturer. What keeps the (to me) obvious software functionality vacuum on a capable hardware platform from being filled up? Edit: I believe a clarification on the multitasking point might be beneficial. I'll use my phone as an example, although the point is much more general. The phone can multitask and in fact does: you can listen to music and do something else at the same time. On the other hand, the way the software has been designed makes multitasking next to useless. (Ditto with the external touch screen: it can take touch commands, but only one application makes use of it, and only with 3 commands.) To take the multitasking example to the extreme, if I plug my phone into my laptop and it registers as an external disk, it doesn't allow any kind of operation: messages, calling, calendar, everything out of reach, although I can receive a call. No "battery life" issue there: it's charging while connected. BTW, another example of design below the current state of the art: I don't see a phone on the horizon which will remember where in an audio or video file you were when you stopped listening/watching it last time (podcasts are a good use case). Simplistic rewind/fast forward functionality only aggravates the problem.

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  • Invoke Python modules from Java

    - by user36813
    I have a Python interface of a graph library written in C - igraph (the name of library). My need is to invoke the python modules pertaining to this graph library from Java code. It goes like this, the core of library is in c. This core has been imported into Python and interfaces to the functions embedded in core are available in Python. My project's rest of the code is in Java and hence I would like to call the graph functions by Java as well. Jython - which lets you invoke python modules with in Java was an option.I went on trying Jython to discover that it will not work in my case as the core code is in C and Jython wont support anything that is imported as a c dll in python code.I also thought of opting for the approach of calling graph routines directly in c. That is without passing through Python code. I am assuming there must be something which lets you call c code from Java, how ever I am not good in C hence I did not go for it. My last resort seems to execute Python interpreter from command line using Java. But that is a dirty and shameless. Also to deal with the results produced by Python code I will have to write the results in a file and read it back in java. Again dirty way. Is there something that any one can suggest me? Thanks to every one giving time. Thanks Igal for answering. I had a look at it. At first glance it appears as if it is simply calling the python script. Jep jep = new Jep(false, SCRIPT_PATH, cl); jep.set("query", query); jep.runScript(SCRIPT_PATH + file); jep.close(); Isnt it very similar to what we would do if called the python interpreter from command line through a Java code. Runtime runtime = Runtime.getRuntime(); Process proc = runtime.exec("python test.py"); Concern is how do I use the results generated by Python script. The naive way is to write them to file and read it back in Java. I am searching for a smarter approach.Thanks for suggestion anyway.

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  • Cocoa NSOutputStream send to a connection

    - by Chuck
    Hi, I am new to Cocoa, but managed to get a connection (to a FTP) up and running, and I've set up an eventhandler for the NSInputStream iStream to alert every response (which also works). What I manage to get is simply the hello message and a connection timeout 60 sec, closing control connection. After searching stackoverflow and finding a lot of NSOutputStream write problems (e.g. http://stackoverflow.com/questions/703729/how-to-use-nsoutputstreams-write-message) and a lot of confusion in my google hits, I figured I'd try to ask my own question: I've tried reading the developer.apple.com doc on OutputStream, but it seems almost impossible for me to send some data (in this case just a string) to the "connection" via the NSOutputStream oStream. - (IBAction) send_something: sender { const char *send_command_char = [@"USER foo" UTF8String]; send_command_buffer = [NSMutableData dataWithBytes:send_command_char length:strlen(send_command_char) + 1]; uint8_t *readBytes = (uint8_t *)[send_command_buffer mutableBytes]; NSInteger byteIndex = 0; readBytes += byteIndex; int data_len = [send_command_buffer length]; unsigned int len = ((data_len - byteIndex >= 1024) ? 1024 : (data_len-byteIndex)); uint8_t buf[len]; (void)memcpy(buf, readBytes, len); len = [oStream write:(const uint8_t *)buf maxLength:len]; byteIndex += len; } the above seems not to result in any useable events. typing it under NSStreamEventHasSpaceAvailable sometimes give a response if I spam the ftp by keep creating new connection instances and keep sending some command whenever oStream has free space. In other words, nothing "right" and so I'm still unclear how to properly send a command to the connection. Should I open - write - close every time i want to write to oStream (and thus to the ftp) and can I then expect a reply (hasBytesAvailable event on iStream)? For some reason I find it very difficult to find any clear tutorials on this matter. Seems like there are more than a few in the same position as me: unclear how to use oStream write? Any little bit that can help clear this up is greatly appreciated! If needed I can write the rest of the code. Chuck

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  • ASP NET MVC (loading data from database)

    - by rah.deex
    hi experts, its me again... i have some code like this.. using System; using System.Collections.Generic; using System.Linq; using System.Web; namespace MvcGridSample.Models { public class CustomerService { private List<SVC> Customers { get { List<SVC> customers; if (HttpContext.Current.Session["Customers"] != null) { customers = (List<SVC>) HttpContext.Current.Session["Customers"]; } else { //Create customer data store and save in session customers = new List<SVC>(); InitCustomerData(customers); HttpContext.Current.Session["Customers"] = customers; } return customers; } } public SVC GetByID(int customerID) { return this.Customers.AsQueryable().First(customer => customer.seq_ == customerID); } public IQueryable<SVC> GetQueryable() { return this.Customers.AsQueryable(); } public void Add(SVC customer) { this.Customers.Add(customer); } public void Update(SVC customer) { } public void Delete(int customerID) { this.Customers.RemoveAll(customer => customer.seq_ == customerID); } private void InitCustomerData(List<SVC> customers) { customers.Add(new SVC { ID = 1, FirstName = "John", LastName = "Doe", Phone = "1111111111", Email = "[email protected]", OrdersPlaced = 5, DateOfLastOrder = DateTime.Parse("5/3/2007") }); customers.Add(new SVC { ID = 2, FirstName = "Jane", LastName = "Doe", Phone = "2222222222", Email = "[email protected]", OrdersPlaced = 3, DateOfLastOrder = DateTime.Parse("4/5/2008") }); customers.Add(new SVC { ID = 3, FirstName = "John", LastName = "Smith", Phone = "3333333333", Email = "[email protected]", OrdersPlaced = 25, DateOfLastOrder = DateTime.Parse("4/5/2000") }); customers.Add(new SVC { ID = 4, FirstName = "Eddie", LastName = "Murphy", Phone = "4444444444", Email = "[email protected]", OrdersPlaced = 1, DateOfLastOrder = DateTime.Parse("4/5/2003") }); customers.Add(new SVC { ID = 5, FirstName = "Ziggie", LastName = "Ziggler", Phone = null, Email = "[email protected]", OrdersPlaced = 0, DateOfLastOrder = null }); customers.Add(new SVC { ID = 6, FirstName = "Michael", LastName = "J", Phone = "666666666", Email = "[email protected]", OrdersPlaced = 5, DateOfLastOrder = DateTime.Parse("12/3/2007") }); } } } those codes is an example that i've got from the internet.. in that case, the data is created and saved in session before its shown.. the things that i want to ask is how if i want to load the data from table? i'am a newbie here.. please help :) thank b4 for advance..

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  • Simple database design and LINQ

    - by Anders Svensson
    I have very little experience designing databases, and now I want to create a very simple database that does the same thing I have previously had in xml. Here's the xml: <services> <service type="writing"> <small>125</small> <medium>100</medium> <large>60</large> <xlarge>30</xlarge> </service> <service type="analysis"> <small>56</small> <medium>104</medium> <large>200</large> <xlarge>250</xlarge> </service> </services> Now, I wanted to create the same thing in a SQL database, and started doing this ( hope this formats ok, but you'll get the gist, four columns and two rows): > ServiceType Small Medium Large > > Writing 125 100 60 > > Analysis 56 104 200 This didn't work too well, since I then wanted to use LINQ to select, say, the Large value for Writing (60). But I couldn't use LINQ for this (as far as I know) and use a variable for the size (see parameters in the method below). I could only do that if I had a column like "Size" where Small, Medium, and Large would be the values. But that doesn't feel right either, because then I would get several rows with ServiceType = Writing (3 in this case, one for each size), and the same for Analysis. And if I were to add more servicetypes I would have to do the same. Simply repetitive... Is there any smart way to do this using relationships or something? Using the second design above (although not good), I could use the following LINQ to select a value with parameters sent to the method: protected int GetHourRateDB(string serviceType, Size size) { CalculatorLinqDataContext context = new CalculatorLinqDataContext(); var data = (from calculatorData in context.CalculatorDatas where calculatorData.Service == serviceType && calculatorData.Size == size.ToString() select calculatorData).Single(); return data.Hours; } But if there is another better design, could you please also describe how to do the same selection using LINQ with that design? Please keep in mind that I am a rookie at database design, so please be as explicit and pedagogical as possible :-) Thanks! Anders

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  • Function-Local Static Const variable Initialization semantics.

    - by Hassan Syed
    The questions are in bold, for those that cannot be bothered reading a question in depth. This is a followup to this question. It is to do with the initialization semantics of static variables in functions. Static variables should be initialized once, and their internal state might be altered later - as I (currently) do in the linked question. However, the code in question does not require the feature to change the state of the variable later. Let me clarrify my position, since I don't require the string object's internal state to change. The code is for a trait class for meta programming, and as such would would benifit from a const char * const ptr -- thus Ideally a local cost static const variable is needed. My educated guess is that in this case the string in question will be optimally placed in memory by the link-loader, and that the code is more secure and maps to the intended semantics. This leads to the semantics of such a variable "The C++ Programming language Third Edition -- Stroustrup" does not have anything (that I could find) to say about this matter. All that is said is that the variable is initialized once when the flow of control of the thread first reaches the code. This leads me to ponder if the following code would be sensible, and if not what are the intended semantics ?. #include <iostream> const char * const GetString(const char * x_in) { static const char * const x = x_in; return x; } int main() { const char * const temp = GetString("yahoo"); std::cout << temp << std::endl; const char * const temp2 = GetString("yahoo2"); std::cout << temp2 << std::endl; } The following compiles on GCC and prints "yahoo" twice. Which is what I want -- However it might not be standards compliant (which is why I post this question). It might be more elegant to have two functions, "SetString" and "String" where the latter forwards to the first. If it is standards compliant does someone know of a templates implementation in boost (or elsewhere) ?

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  • Advantage of creating a generic repository vs. specific repository for each object?

    - by LuckyLindy
    We are developing an ASP.NET MVC application, and are now building the repository/service classes. I'm wondering if there are any major advantages to creating a generic IRepository interface that all repositories implement, vs. each Repository having its own unique interface and set of methods. For example: a generic IRepository interface might look like (taken from this answer): public interface IRepository : IDisposable { T[] GetAll<T>(); T[] GetAll<T>(Expression<Func<T, bool>> filter); T GetSingle<T>(Expression<Func<T, bool>> filter); T GetSingle<T>(Expression<Func<T, bool>> filter, List<Expression<Func<T, object>>> subSelectors); void Delete<T>(T entity); void Add<T>(T entity); int SaveChanges(); DbTransaction BeginTransaction(); } Each Repository would implement this interface (e.g. CustomerRepository:IRepository, ProductRepository:IRepository, etc). The alternate that we've followed in prior projects would be: public interface IInvoiceRepository : IDisposable { EntityCollection<InvoiceEntity> GetAllInvoices(int accountId); EntityCollection<InvoiceEntity> GetAllInvoices(DateTime theDate); InvoiceEntity GetSingleInvoice(int id, bool doFetchRelated); InvoiceEntity GetSingleInvoice(DateTime invoiceDate, int accountId); //unique InvoiceEntity CreateInvoice(); InvoiceLineEntity CreateInvoiceLine(); void SaveChanges(InvoiceEntity); //handles inserts or updates void DeleteInvoice(InvoiceEntity); void DeleteInvoiceLine(InvoiceLineEntity); } In the second case, the expressions (LINQ or otherwise) would be entirely contained in the Repository implementation, whoever is implementing the service just needs to know which repository function to call. I guess I don't see the advantage of writing all the expression syntax in the service class and passing to the repository. Wouldn't this mean easy-to-messup LINQ code is being duplicated in many cases? For example, in our old invoicing system, we call InvoiceRepository.GetSingleInvoice(DateTime invoiceDate, int accountId) from a few different services (Customer, Invoice, Account, etc). That seems much cleaner than writing the following in multiple places: rep.GetSingle(x => x.AccountId = someId && x.InvoiceDate = someDate.Date); The only disadvantage I see to using the specific approach is that we could end up with many permutations of Get* functions, but this still seems preferable to pushing the expression logic up into the Service classes. What am I missing?

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  • Advice on software / database design to avoid using cursors when updating database

    - by Remnant
    I have a database that logs when an employee has attended a course and when they are next due to attend the course (courses tend to be annual). As an example, the following employee attended course '1' on 1st Jan 2010 and, as the course is annual, is due to attend next on the 1st Jan 2011. As today is 20th May 2010 the course status reads as 'Complete' i.e. they have done the course and do not need to do it again until next year: EmployeeID CourseID AttendanceDate DueDate Status 123456 1 01/01/2010 01/01/2011 Complete In terms of the DueDate I calculate this in SQL when I update the employee's record e.g. DueDate = AttendanceDate + CourseFrequency (I pull course frequency this from a separate table). In my web based app (asp.net mvc) I pull back this data for all employees and display it in a grid like format for HR managers to review. This allows HR to work out who needs to go on courses. The issue I have is as follows. Taking the example above, suppose today is 2nd Jan 2011. In this case, employee 123456 is now overdue for the course and I would like to set the Status to Incomplete so that the HR manager can see that they need to action this i.e. get employee on the course. I could build a trigger in the database to run overnight to update the Status field for all employees based on the current date. From what I have read I would need to use cursors to loop over each row to amend the status and this is considered bad practice / inefficient or at least something to avoid if you can??? Alternatively, I could compute the Status in my C# code after I have pulled back the data from the database and before I display it on screen. The issue with this is that the Status in the database would not necessarily match what is shown on screen which just feels plain wrong to me. Does anybody have any advice on the best practice approach to such an issue? It helps, if I did use a cursor I doubt I would be looping over more than 1000 records at any given time. Maybe this is such small volume that using cursors is okay?

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  • Job queue manager with RPC interface

    - by admr
    I need a job queue manager that I can control over the Internet. It should be able to execute and stop processes, check on their status (ideally notice and execute some code when a process exits), respond to commands and also be able to report back to a server. Background: I have a GWT application that allows to create jobs to execute on a cloud instance (currently EC2). I want to push a "job packet" (data for a process to operate on etc) to S3, start a Linux EC2 instance (or use one that's already running), and tell a job manager on the instance to execute that job (possibly parallel to other jobs). It should then pull the "job packet" from S3, run a process that operates on that data and report back to the server that is running the server part of my GWT application with some information (e.g. exit code, stdout, stderr). If I have to write e.g. stdour/err to a file from the process and read that file, that's OK too. I would really like the manager to be "close" to the processes it runs, meaning I want to avoid using something like Runtime.exec from the JDK. It seems like I would have to do that if I used Quartz for example. I'm fine with the calls in both directions being asynchronous. I'm fine with any reasonable technology for the calls as long as I can easily build an interface for that in my GWT server side (e.g. HTTP requests to a servlet over SSL would be nice and trivial). The job manager does not need to have a very sophisticated queueing system. Running several processes either sequentially or in parallel should be fine. Determining how much compute time a process received during its lifetime would be nice (AFAIK, this might be challenging). I did not yet find any existing software that does this, including http://java-source.net/open-source/job-schedulers. I suspect I might have to build an RPC interface (with authentication etc, of course) around a job manager; maybe use something like Apache Commons Exec. In that case, I would prefer Java or Python for the job manager part. I would be happy to hear suggestions for either the former or latter scenario!

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  • gcc/g++: error when compiling large file

    - by Alexander
    Hi, I have a auto-generated C++ source file, around 40 MB in size. It largely consists of push_back commands for some vectors and string constants that shall be pushed. When I try to compile this file, g++ exits and says that it couldn't reserve enough virtual memory (around 3 GB). Googling this problem, I found that using the command line switches --param ggc-min-expand=0 --param ggc-min-heapsize=4096 may solve the problem. They, however, only seem to work when optimization is turned on. 1) Is this really the solution that I am looking for? 2) Or is there a faster, better (compiling takes ages with these options acitvated) way to do this? Best wishes, Alexander Update: Thanks for all the good ideas. I tried most of them. Using an array instead of several push_back() operations reduced memory usage, but as the file that I was trying to compile was so big, it still crashed, only later. In a way, this behaviour is really interesting, as there is not much to optimize in such a setting -- what does the GCC do behind the scenes that costs so much memory? (I compiled with deactivating all optimizations as well and got the same results) The solution that I switched to now is reading in the original data from a binary object file that I created from the original file using objcopy. This is what I originally did not want to do, because creating the data structures in a higher-level language (in this case Perl) was more convenient than having to do this in C++. However, getting this running under Win32 was more complicated than expected. objcopy seems to generate files in the ELF format, and it seems that some of the problems I had disappeared when I manually set the output format to pe-i386. The symbols in the object file are by standard named after the file name, e.g. converting the file inbuilt_training_data.bin would result in these two symbols: binary_inbuilt_training_data_bin_start and binary_inbuilt_training_data_bin_end. I found some tutorials on the web which claim that these symbols should be declared as extern char _binary_inbuilt_training_data_bin_start;, but this does not seem to be right -- only extern char binary_inbuilt_training_data_bin_start; worked for me.

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