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  • Using a large list of terms, search through page text and replace words with links

    - by dunc
    A while ago I posted this question asking if it's possible to convert text to HTML links if they match a list of terms from my database. I have a fairly huge list of terms - around 6000. The accepted answer on that question was superb, but having never used XPath, I was at a loss when problems started occurring. At one point, after fiddling with code, I somehow managed to add over 40,000 random characters to our database - the majority of which required manual removal. Since then I've lost faith in that idea and the more simple PHP solutions simply weren't efficient enough to deal with the amount of data and the quantity of terms. My next attempt at a solution is to write a JS script which, once the page has loaded, retrieves the terms and matches them against the text on a page. This answer has an idea which I'd like to attempt. I would use AJAX to retrieve the terms from the database, to build an object such as this: var words = [ { word: 'Something', link: 'http://www.something.com' }, { word: 'Something Else', link: 'http://www.something.com/else' } ]; When the object has been built, I'd use this kind of code: //for each array element $.each(words, function() { //store it ("this" is gonna become the dom element in the next function) var search = this; $('.message').each( function() { //if it's exactly the same if ($(this).text() === search.word) { //do your magic tricks $(this).html('<a href="' + search.link + '">' + search.link + '</a>'); } } ); } ); Now, at first sight, there is a major issue here: with 6,000 terms, will this code be in any way efficient enough to do what I'm trying to do?. One option would possibly be to perform some of the overhead within the PHP script that the AJAX communicates with. For instance, I could send the ID of the post and then the PHP script could use SQL statements to retrieve all of the information from the post and match it against all 6,000 terms.. then the return call to the JavaScript could simply be the matching terms, which would significantly reduce the number of matches the above jQuery would make (around 50 at most). I have no problem with the script taking a few seconds to "load" on the user's browser, as long as it isn't impacting their CPU usage or anything like that. So, two questions in one: Can I make this work? What steps can I take to make it as efficient as possible? Thanks in advance,

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  • passing back answers in prolog

    - by AhmadAssaf
    i have this code than runs perfectly .. returns a true .. when tracing the values are ok .. but its not returning back the answer .. it acts strangely when it ends and always return empty list .. uninstantiated variable .. test :- extend(4,12,[4,3,1,2],[[1,5],[3,4],[6]],_ExtendedBins). %printing basic information about the extend(NumBins,Capacity,RemainingNumbers,BinsSoFar,_ExtendedBins) :- getNumberofBins(BinsSoFar,NumberOfBins), msort(RemainingNumbers,SortedRemaining),nl, format("Current Number of Bins is :~w\n",[NumberOfBins]), format("Allowed Capacity is :~w\n",[Capacity]), format("maximum limit in bin is :~w\n",[NumBins]), format("Trying to fit :~w\n\n",[SortedRemaining]), format("Possible Solutions :\n\n"), fitElements(NumBins,NumberOfBins, Capacity,SortedRemaining,BinsSoFar,[]). %this is were the creation for possibilities will start %will check first if the number of bins allowed is less than then %we create a new list with all the possible combinations %after that we start matching to other bins with capacity constraint fitElements(NumBins,NumberOfBins, Capacity,RemainingNumbers,Bins,ExtendedBins) :- ( NumberOfBins < NumBins -> print('Creating new set: '); print('Sorry, Cannot create New Sets')), createNewList(Capacity,RemainingNumbers,Bins,ExtendedBins). createNewList(Capacity,RemainingNumbers,Bins,ExtendedBins) :- createNewList(Capacity,RemainingNumbers,Bins,[],ExtendedBins), print(ExtendedBins). createNewList(0,Bins,Bins,ExtendedBins,ExtendedBins). createNewList(_,[],_,ExtendedBins,ExtendedBins). createNewList(Capacity,[Element|Rest],Bins,Temp,ExtendedBins) :- conjunct_to_list(Element,ListedElement), append(ListedElement,Temp,NewList), sumlist(NewList,Sum), (Sum =< Capacity, append(ListedElement,ExtendedBins,Result); Capacity = 0), createNewList(Capacity,Rest,Bins,NewList,Result). fit(0,[],ExtendedBins,ExtendedBins). fit(Capacity,[Element|Rest],Bin,ExtendedBins) :- conjunct_to_list(Element,Listed), append(Listed,Bin,NewBin), sumlist(NewBin,Sum), (Sum =< Capacity -> fit(Capacity,Rest,NewBin,ExtendedBins); Capacity = 0, append(NewBin,ExtendedBins,NewExtendedBins), print(NewExtendedBins), fit(0,[],NewBin,ExtendedBins)). %get the number of bins provided getNumberofBins(List,NumberOfBins) :- getNumberofBins(List,0,NumberOfBins). getNumberofBins([],NumberOfBins,NumberOfBins). getNumberofBins([_List|Rest],TempCount,NumberOfBins) :- NewCount is TempCount + 1, %calculate the count getNumberofBins(Rest,NewCount,NumberOfBins). %recursive call %Convert set of terms into a list - used when needed to append conjunct_to_list((A,B), L) :- !, conjunct_to_list(A, L0), conjunct_to_list(B, L1), append(L0, L1, L). conjunct_to_list(A, [A]). Greatly appreciate the help

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  • Solution Output Directory

    - by L.E.O
    The project that I'm currently working on is being developed by multiple teams where each team is responsible for different part of the project. They all have set up their own C# projects and solutions with configuration settings specific to their own needs. However, now we need to create another, global solution, which will combine and build all projects into the same output directory. The problem that I have encountered though, is that I have found only one way to make all projects build into the same output directory - I need to modify configurations for all of them. That is what we would like to avoid. We would prefer that all these projects had no knowledge about this "global" solution. Each team must retain possibility to work just with their own sub-solution. One possible workaround is to create a special configuration for all projects just for this "global" solution, but that could create extra problems since now you have to constantly sync this configuration settings with the regular one, used by that specific team. Last thing we want to do is to spend hours trying to figure out why something doesn't work when building under global solution just because of some check box that developers have checked in their configuration, but forgot to do so in the global configuration. So, to simplify, we need some sort of output directory setting or post build event that would only be present when building from that global, all-inclusive solution. Is there any way to achieve this without changing something in projects configurations? Update 1: Some extra details I guess I need to mention: We need this global solution to be as close as possible to what the end user gets when he installs our application, since we intend to use it for debugging of the entire application when we need to figure out which part of the application isn't working before sending this bug to the team working on that part. This means that when building under global solution the output directory hierarchy should be the same as it would be in Program Files after installation, so that if, for example, we have Program Files/MyApplication/Addins folder which contains all the addins developed by different teams, we need the global solution to copy the binaries from addins projects and place them in the output directory accordingly. The thing is, the team developing an addin doest necessary know that it is an addin and that it should be placed in that folder, so they cannot change their relative output directory to be build/bin/Debug/Addins.

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  • Calling compiled C from R with .C()

    - by Sarah
    I'm trying to call a program (function getNBDensities in the C executable measurementDensities_out) from R. The function is passed several arrays and the variable double runsum. Right now, the getNBDensities function basically does nothing: it prints to screen the values of passed parameters. My problem is the syntax of calling the function: array(.C("getNBDensities", hr = as.double(hosp.rate), # a vector (s x 1) sp = as.double(samplingProbabilities), # another vector (s x 1) odh = as.double(odh), # another vector (s x 1) simCases = as.integer(x[c("xC1","xC2","xC3")]), # another vector (s x 1) obsCases = as.integer(y[c("yC1","yC2","yC3")]), # another vector (s x 1) runsum = as.double(runsum), # double DUP = TRUE, NAOK = TRUE, PACKAGE = "measurementDensities_out")$f, dim = length(y[c("yC1","yC2","yC3")]), dimnames = c("yC1","yC2","yC3")) The error I get, after proper execution of the function (i.e., the right output is printed to screen), is Error in dim(data) <- dim : attempt to set an attribute on NULL I'm unclear what the dimensions are that I should be passing the function: should it be s x 5 + 1 (five vectors of length s and one double)? I've tried all sorts of combinations (including sx5+1) and have found only seemingly conflicting descriptions/examples online of what's supposed to happen here. For those who are interested, the C code is below: #include <R.h> #include <Rmath.h> #include <math.h> #include <Rdefines.h> #include <R_ext/PrtUtil.h> #define NUM_STRAINS 3 #define DEBUG void getNBDensities( double *hr, double *sp, double *odh, int *simCases, int *obsCases, double *runsum ); void getNBDensities( double *hr, double *sp, double *odh, int *simCases, int *obsCases, double *runsum ) { #ifdef DEBUG for ( int s = 0; s < NUM_STRAINS; s++ ) { Rprintf("\nFor strain %d",s); Rprintf("\n\tHospitalization rate = %lg", hr[s]); Rprintf("\n\tSimulation probability = %lg",sp[s]); Rprintf("\n\tSimulated cases = %d",simCases[s]); Rprintf("\n\tObserved cases = %d",obsCases[s]); Rprintf("\n\tOverdispersion parameter = %lg",odh[s]); } Rprintf("\nRunning sum = %lg",runsum[0]); #endif } naive solution While better (i.e., potentially faster or syntactically clearer) solutions may exist (see Dirk's answer below), the following simplification of the code works: out<-.C("getNBDensities", hr = as.double(hosp.rate), sp = as.double(samplingProbabilities), odh = as.double(odh), simCases = as.integer(x[c("xC1","xC2","xC3")]), obsCases = as.integer(y[c("yC1","yC2","yC3")]), runsum = as.double(runsum)) The variables can be accessed in >out.

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  • Is it OK to put a standard, pure C header #include directive inside a namespace?

    - by mic_e
    I've got a project with a class log in the global namespace (::log). So, naturally, after #include <cmath>, the compiler gives an error message each time I try to instantiate an object of my log class, because <cmath> pollutes the global namespace with lots of three-letter methods, one of them being the logarithm function log(). So there are three possible solutions, each having their unique ugly side-effects. Move the log class to it's own namespace and always access it with it's fully qualified name. I really want to avoid this because the logger should be as convenient as possible to use. Write a mathwrapper.cpp file which is the only file in the project that includes <cmath>, and makes all the required <cmath> functions available through wrappers in a namespace math. I don't want to use this approach because I have to write a wrapper for every single required math function, and it would add additional call penalty (cancelled out partially by the -flto compiler flag) The solution I'm currently considering: Replace #include <cmath> by namespace math { #include "math.h" } and then calculating the logarithm function via math::log(). I have tried it out and it does, indeed, compile, link and run as expected. It does, however, have multiple downsides: It's (obviously) impossible to use <cmath>, because the <cmath> code accesses the functions by their fully qualified names, and it's deprecated to use in C++. I've got a really, really bad feeling about it, like I'm gonna get attacked and eaten alive by raptors. So my question is: Is there any recommendation/convention/etc that forbid putting include directives in namespaces? Could anything go wrong with diferent C standard library implementations (I use glibc), different compilers (I use g++ 4.7, -std=c++11), linking? Have you ever tried doing this? Are there any alternate ways to banish the math functions from the global namespace? I've found several similar questions on stackoverflow, but most were about including other C++ headers, which obviously is a bad idea, and those that weren't made contradictory statements about linking behaviour for C libraries. Also, would it be beneficial to additionally put the #include <math.h> inside extern "C" {}?

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  • C++ Euler-Problem 14 Program Freezing

    - by Tim
    I'm working on Euler Problem 14: http://projecteuler.net/index.php?section=problems&id=14 I figured the best way would be to create a vector of numbers that kept track of how big the series was for that number... for example from 5 there are 6 steps to 1, so if ever reach the number 5 in a series, I know I have 6 steps to go and I have no need to calculate those steps. With this idea I coded up the following: #include <iostream> #include <vector> #include <iomanip> using namespace std; int main() { vector<int> sizes(1); sizes.push_back(1); sizes.push_back(2); int series, largest = 0, j; for (int i = 3; i <= 1000000; i++) { series = 0; j = i; while (j > (sizes.size()-1)) { if (j%2) { j=(3*j+1)/2; series+=2; } else { j=j/2; series++; } } series+=sizes[j]; sizes.push_back(series); if (series>largest) largest=series; cout << setw(7) << right << i << "::" << setw(5) << right << series << endl; } cout << largest << endl; return 0; } It seems to work relatively well for smaller numbers but this specific program stalls at the number 113382. Can anyone explain to me how I would go about figuring out why it freezes at this number? Is there some way I could modify my algorithim to be better? I realize that I am creating duplicates with the current way I'm doing it: for example, the series of 3 is 3,10,5,16,8,4,2,1. So I already figured out the sizes for 10,5,16,8,4,2,1 but I will duplicate those solutions later. Thanks for your help!

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  • Problems Embedding Video using FCK Editor

    - by Exline
    Hello, I am using FCK Editor 2.6.4 and having problems trying to embed a (non-YouTube) video into a content area. I found this previous question / post: [EDIT -- as a new user, I am only able to post one link in this post. The post in question is titled, "Can I embed video using FCK Editor?") and have investigated all of the proposed solutions, but none of them work properly: 1 -- Using the "Embed Flash" button in the control panel almost works. However, the video I am attempting to add contains a querystring with parameters, like this: http://static.animoto.com/swf/w.swf?w=swf/vp1&e=1275795594&f=mGQklEgxXKs9vfEIdGnWsA&d=132&m=p&r=w&i=m&ct=Homes%20in%20Eagle%20Creek&cu=http://hometoindy.com/eagle-creek-real-estate.php&options= and in using the Flash embed tool, it encodes all of the "&" characters to "&", thus breaking them. If it were just for me, I could manually change them back, but clients who use this will not know how to do that. 2 -- I have installed the YouTube video plugin, and it works great... for YouTube. But it cannot be used to embed non-YouTube videos (it automatically changes the URL to YouTube, no matter what). 3 -- I have installed the EmbedMovies plugin, but it throws a javascript error when attempting to add a video file (such as the above) to a page. (The EmbedMovies plugin page on SourceForge says it has been updated for FCK Editor 2.6, but it does not work.) 4 -- Pasting directly into the editor window (of course) does not work. The only way I've been able to make this work is by pasting into the Source panel, and this is not a good option for clients who are not familiar with HTML. So, is there a good, working plugin for FCK editor that will allow me to quickly and easily embed a video such as the one above into a content area? I don't need to be able to see or preview it in the editor window; I just need it to work when the page is loaded on the front end. Thanks!

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  • rake db:create not working for legacy rails app (2.3.5) using MySQL (5.5.28)

    - by ridicter
    I'm a new Rails Developer, and I'm working on a legacy Rails app. Whenever I run the rake db:create command, I get an error that the database couldn't be created. I have found many StackOverflow questions related to this, but in troubleshooting nearly all permutations of solutions, I couldn't resolve the issue. I created the three Dbs (dev, prod, test), created the user with all access privileges to these dbs, and ran rake db:create. I'm running Mac OS X Lion, MySQL 5.5.28, Rails 2.3.5, Ruby 1.8.7. Here are my settings development: adapter: mysql encoding: utf8 database: adva_development username: adva password: **** host: localhost socket: /tmp/mysql.sock Here's the error: Couldn't create database for {"adapter"=>"mysql", "username"=>"adva", "host"=>"localhost", "encoding"=>"utf8", "database"=>"adva_development", "socket"=>"/tmp/mysql.sock", "password"=>"****"}, charset: utf8, collation: utf8_unicode_ci (if you set the charset manually, make sure you have a matching collation) I have done the following troubleshooting: Verified user and password are correct, and the user has access to the DB. (Double checked user access with SELECT * FROM mysql.db WHERE Db = 'adva_development' \G; User has all privileges.) Verify the socket is correct. I don't really understand sockets, but I can plainly see it at /tmp/mysql.sock. Checked collation and character set. I found out I had created the DB in latin charset and collation, so I recreated them. I ran show variables like "collation_database"; and show variables like "character_set_database"; and came back with utf8 and utf8_unicode_ci respectively. I followed the instructions in this question. After uninstalling mysql gem, I ran the following but came up with the same error: gem install --no-rdoc --no-ri mysql -- --with-mysql-dir=/usr/local/mysql-5.5.28-osx10.6-x86_64/bin --with-mysql-config=/usr/local/mysql-5.5.28-osx10.6-x86_64/bin/mysql_config Following Matt's suggestion, here's what a rake --trace db:create reveals: ** Invoke db:create (first_time) ** Invoke db:load_config (first_time) ** Invoke rails_env (first_time) ** Execute rails_env ** Execute db:load_config ** Execute db:create Couldn't create database for {"database"=>"adva_development", "adapter"=>"mysql", "host"=>"127.0.0.1", "password"=>"woof2adva", "username"=>"adva", "encoding"=>"utf8"}, charset: utf8, collation: utf8_unicode_ci (if you set the charset manually, make sure you have a matching collation) After 3 days and six or seven hours, I have pretty much run out of options. I tried various random things, like replacing localhost with 127.0.0.1 to no avail. Could there be something wrong related to my specific environment? Mac OS X Lion + MySQL 5.5.28? I plan on trying on setting up everything in a Linux environment. Thanks!

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  • C++ dynamic type construction and detection

    - by KneLL
    There was an interesting problem in C++, but it concerns more likely architecture. There are many (10, 20, 40, etc) classes that describe some characteristics (mix-in classes), for exmaple: struct Base { virtual ~Base() {} }; struct A : virtual public Base { int size; }; struct B : virtual public Base { float x, y; }; struct C : virtual public Base { bool some_bool_state; }; struct D : virtual public Base { string str; } // .... Primary module declares and exports a function (for simplicity just function declarations without classes): // .h file void operate(Base *pBase); // .cpp file void operate(Base *pBase) { // .... } Any other module can has a code like this: #include "mixins.h" #include "primary.h" class obj1_t : public A, public C, public D {}; class obj2_t : public B, public D {}; // ... void Pass() { obj1_t obj1; obj2_t obj2; operate(&obj1); operate(&obj2); } The question is how to know what the real type of given object in operate() without dynamic_cast and any type information in classes (constants, etc)? Function operate() is used with big array of objects in small time periods and dynamic_cast is too slow for it. And I don't want to include constants (enum obj_type { ... }) because this is not OOP-way. // module operate.cpp void some_operate(Base *pBase) { processA(pBase); processB(pBase); } void processA(A *pA) { } void processB(B *pB) { } I cannot directly pass a pBase to these functions. And it's impossible to have all possible combinations of classes, because I can add new classes just by including new .h files. As one of solutions that comed to mind, in editor application I can use a composite container: struct CompositeObject { vector<Base *pBase> parts; }; But editor does not need a time optimization and can use dynamic_cast for parts to determine the exact type. In operate() I cannot use this solution. So, is it possible to not use a dynamic_cast and type information to solve this problem? Or maybe I should use another architecture?

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  • session fixation

    - by markiv
    Hi All, I am new to web development, and trying to get a hold on security issues. I went through this article on http://guides.rubyonrails.org/security.html these are some of the steps the author has mentioned how an attacker fixes session. 1. The attacker creates a valid session id: He loads the login page of the web application where he wants to fix the session, and takes the session id in the cookie from the response (see number 1 and 2 in the image). 2. He possibly maintains the session. Expiring sessions, for example every 20 minutes, greatly reduces the time-frame for attack. Therefore he accesses the web application from time to time in order to keep the session alive. 3. Now the attacker will force the user’s browser into using this session id (see number 3 in the image). As you may not change a cookie of another domain (because of the same origin policy), the attacker has to run a JavaScript from the domain of the target web application. Injecting the JavaScript code into the application by XSS accomplishes this attack. Here is an example: <script>?document.cookie="_session_id=16d5b78abb28e3d6206b60f22a03c8d9";?</script>. Read more about XSS and injection later on. 4. The attacker lures the victim to the infected page with the JavaScript code. By viewing the page, the victim’s browser will change the session id to the trap session id. 5. As the new trap session is unused, the web application will require the user to authenticate. 6. From now on, the victim and the attacker will co-use the web application with the same session: The session became valid and the victim didn’t notice the attack. I dont understand couple of points. i) why is user made to login in step5, since session is sent through. ii) I saw possible solutions on wiki, like user properties check and others why cant we just reset the session for the user whoever is login in when they enter username and password in step5? Thanks in advance Markiv

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  • JQuery within a partial view not being called

    - by XN16
    I have a view that has some jQuery to load a partial view (via a button click) into a div in the view. This works without a problem. However within the partial view I have a very similar bit of jQuery that will load another partial view into a div in the first partial view, but this isn't working, it almost seems like the jQuery in the first partial view isn't being loaded. I have tried searching for solutions, but I haven't managed to find an answer. I have also re-created the jQuery function in a @Ajax.ActionLink which works fine, however I am trying to avoid the Microsoft helpers as I am trying to learn jQuery. Here is the first partial view which contains the jQuery that doesn't seem to work, it also contains the @Ajax.ActionLink that does work: @model MyProject.ViewModels.AddressIndexViewModel <script> $(".viewContacts").click(function () { $.ajax({ url: '@Url.Action("CustomerAddressContacts", "Customer")', type: 'POST', data: { addressID: $(this).attr('data-addressid') }, cache: false, success: function (result) { $("#customerAddressContactsPartial-" + $(this).attr('data-addressid')) .html(result); }, error: function () { alert("error"); } }); return false; }); </script> <table class="customers" style="width: 100%"> <tr> <th style="width: 25%"> Name </th> <th style="width: 25%"> Actions </th> </tr> </table> @foreach (Address item in Model.Addresses) { <table style="width: 100%; border-top: none"> <tr id="[email protected]"> <td style="width: 25%; border-top: none"> @Html.DisplayFor(modelItem => item.Name) </td> <td style="width: 25%; border-top: none"> <a href="#" class="viewContacts standardbutton" data-addressid="@item.AddressID">ContactsJQ</a> @Ajax.ActionLink("Contacts", "CustomerAddressContacts", "Customer", new { addressID = item.AddressID }, new AjaxOptions { UpdateTargetId = "customerAddressContactsPartial-" + @item.AddressID, HttpMethod = "POST" }, new { @class = "standardbutton"}) </td> </tr> </table> <div id="[email protected]"></div> } If someone could explain what I am doing wrong here and how to fix it then I would be very grateful. Thanks very much.

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  • Extend base class properties

    - by user1888033
    I need your help to extend my base class, here is the similar structure i have. public class ShowRoomA { public audi AudiModelA { get; set; } public benz benzModelA { get; set; } } public class audi { public string Name { get; set; } public string AC { get; set; } public string PowerStearing { get; set; } } public class benz { public string Name { get; set; } public string AC { get; set; } public string AirBag { get; set; } public string MusicSystem { get; set; } } //My Implementation class like this class Main() { private void UpdateDetails() { ShowRoomA ojbMahi = new ShowRoomA(); GetDetails( ojbMahi ); // this works fine } private void GetDetails(ShowRoomA objShowRoom) { objShowRoom = new objShowRoom(); objShowRoom.audi = new audi(); objShowRoom.audi.Name = "AUDIMODEL94CD698"; objShowRoom.audi.AC = "6 TON"; objShowRoom.audi.PowerStearing = "Electric"; objShowRoom.benz= new benz(); objShowRoom.audi.Name = "BENZMODEL34LCX"; objShowRoom.audi.AC = "8 TON"; objShowRoom.audi.AirBag = "Two (1+1)"; objShowRoom.audi.MusicSystem = "Poineer 3500W"; } } // Till this cool. // Now I got requirement for ShowRoomB with replacement of old audi and benz with new models and new other brand cars also added. // I don't want to modify GetDetails() method. by reusing this method additional logic i want to apply to my new extended model. // Here I struck in designing my new model of ShowRoomB (base of ShowRoomA) ... I have tried some thing like... but not sure. public class audiModelB:audi { public string JetEngine { get; set; } } public class benzModelB:benz { public string JetEngine { get; set; } } public class ShowRoomB { public audiModelB AudiModelB { get; set; } public benzModelB benzModelB { get; set; } } // My new code to Implementation class like this class Main() { private void UpdateDetails() { ShowRoomB ojbNahi = new ShowRoomB(); GetDetails( ojbNahi ); // this is NOT working! I know this object does not contain base class directly, still some what i want fill my new model with old properties. Kindly suggest here } } Can any one please give me solutions how to achieve my extending requirement for base class "ShowroomA" Really appreciated your time and suggestions. Thanks in advance,

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  • C# Different class objects in one list

    - by jeah_wicer
    I have looked around some now to find a solution to this problem. I found several ways that could solve it but to be honest I didn't realize which of the ways that would be considered the "right" C# or OOP way of solving it. My goal is not only to solve the problems but also to develop a good set of code standards and I'm fairly sure there's a standard way to handle this problem. Let's say I have 2 types of printer hardwares with their respective classes and ways of communicating: PrinterType1, PrinterType2. I would also like to be able to later on add another type if neccessary. One step up in abstraction those have much in common. It should be possible to send a string to each one of them as an example. They both have variables in common and variables unique to each class. (One for instance communicates via COM-port and has such an object, while the other one communicates via TCP and has such an object). I would however like to just implement a List of all those printers and be able to go through the list and perform things as "Send(string message)" on all Printers regardless of type. I would also like to access variables like "PrinterList[0].Name" that are the same for both objects, however I would also at some places like to access data that is specific to the object itself (For instance in the settings window of the application where the COM-port name is set for one object and the IP/port number for another). So, in short something like: In common: Name Send() Specific to PrinterType1: Port Specific to PrinterType2: IP And I wish to, for instance, do Send() on all objects regardless of type and the number of objects present. I've read about polymorphism, Generics, interfaces and such, but I would like to know how this, in my eyes basic, problem typically would be dealt with in C# (and/or OOP in general). I actually did try to make a base class, but it didn't quite seem right to me. For instance I have no use of a "string Send(string Message)" function in the base class itself. So why would I define one there that needs to be overridden in the derived classes when I would never use the function in the base class ever in the first place? I'm really thankful for any answers. People around here seem very knowledgeable and this place has provided me with many solutions earlier. Now I finally have an account to answer and vote with too. EDIT: To additionally explain, I would also like to be able to access the objects of the actual printertype. For instance the Port variable in PrinterType1 which is a SerialPort object. I would like to access it like: PrinterList[0].Port.Open() and have access to the full range of functionality of the underlaying port. At the same time I would like to call generic functions that work in the same way for the different objects (but with different implementations): foreach (printer in Printers) printer.Send(message)

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  • Parsing ip addresses in php

    - by user2938780
    I am trying to get the number of active connections (Real Time) from a log file by IP connected and having a Play status but instead, it's giving me the total number of IP with status play. The number doesn't decrease at all. Keeps on increasing as soon as a new ip is added. How can I fix it? Here my code: $stringToParse = file_get_contents('wowzamediaserver_access.log'); preg_match_all('/\d{1,3}\.\d{1,3}\.\d{1,3}\.\d{1,3}/', $stringToParse, $matchOP); echo "Number of connections: ".sizeof(array_unique($matchOP[0])); HERE IS THE LOG: 2013-10-30 14:54:36 CET stop stream INFO 200 account1 - _defaultVHost_ account1 _definst_ 149.21 streamURL 1935 fullStreamURL IP_ADDRESS_1 http (cupertino) - 2013-10-30 14:56:12 CET play stream INFO 200 account2 - _defaultVHost_ account1 _definst_ 149.21 streamURL 1935 fullStreamURL IP_ADDRESS_2 rtmp (cupertino) - 2013-10-30 14:58:23 CET stop stream INFO 200 account2 - _defaultVHost_ account1 _definst_ 149.21 streamURL 1935 fullStreamURL IP_ADDRESS_2 rtmp (cupertino) - 2013-10-30 14:58:39 CET play stream INFO 200 account1 - _defaultVHost_ account1 _definst_ 149.21 streamURL 1935 fullStreamURL IP_ADDRESS_1 http (cupertino) - 2013-10-30 14:59:12 CET play stream INFO 200 account2 - _defaultVHost_ account1 _definst_ 149.21 streamURL 1935 fullStreamURL IP_ADDRESS_2 rtmp (cupertino) - I want to be able to count the IP whenever it has a "PLAY" status and don't count it whenever it's "STOP" 2013-10-30 14:59:00 CET play stream INFO 200 tv2vielive - _defaultVHost_ tv2vielive _definst_ 0.315 [any] 1935 rtmp://tv2vie.zion3cloud.com:1935/tv2vielive 78.247.255.186 rtmp http://www.tv2vie.org/swf/flowplayer-3.2.16.swf WIN 11,7,700,202 92565864 3576 3455 1 0 0 0 tv2vielive - - - - - rtmp://tv2vie.zion3cloud.com:1935/tv2vielive/tv2vielive rtmp://tv2vie.zion3cloud.com:1935/tv2vielive/tv2vielive - rtmp://tv2vie.zion3cloud.com:1935/tv2vielive - 2013-10-30 14:59:04 CET stop stream INFO 200 tv2vielive - _defaultVHost_ tv2vielive _definst_ 4.75 [any] 1935 rtmp://tv2vie.zion3cloud.com:1935/tv2vielive 78.247.255.186 rtmp http://www.tv2vie.org/swf/flowplayer-3.2.16.swf WIN 11,7,700,202 92565864 3576 512571 1 7222 0 503766 tv2vielive - - - - - rtmp://tv2vie.zion3cloud.com:1935/tv2vielive/tv2vielive rtmp://tv2vie.zion3cloud.com:1935/tv2vielive/tv2vielive - rtmp://tv2vie.zion3cloud.com:1935/tv2vielive - Any solutions? I have even tried the first answer solution but getting "0" play connections. $stringToParse = file_get_contents('wowzamediaserver_access.log'); $pattern = '~^.* play.* ( ([0-9]{1,3}+\.){3,3}[0-9]{1,3}).*$~m'; preg_match_all($pattern, $stringToParse, $matches); echo count($matches[1]) . ' play actions'; But whenever I use my code, I am getting "Number of connections: xxxxx(actual count of IPs). My concern is that I only need the count of IPs with PLAY status. If that IP changes to STOP then it wont count.

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  • VB.Net 2008 IDE hanging - MSVB7.dll eating 100% CPU when editing code

    - by Andrew Backer
    I am having a problem with msvb7.dll eating 50%+ cpu on my dual core system. This usually lasts 10-30 seconds or so, during which time the IDE is non-responsive. This occurs when I do pretty much anything in the text editor, and can be replicated by simply adding blank lines to a function, and then deleting them. Or pasting some code. Or... lotsa stuff. SP1 installed I had DevExpress' refactor/coderush, components, and codeit.right installed, but have removed all 3 of them. (I had installed the latest version of Refactor Pro! (9.3.4), perhaps the day before) I have tried a VS.NET Repair. There is a kb that referenced some cpu destroying with vb, but it was included in SP1 Also: The solution consists of ~30 VB projects and 2 C# projects 8 other developers aren't having any issues with this (or at least not the SAME issues, we all have em) Clean get from TFS was done Project builds properly, can can even debug. This doesn't seem to happen on really small solutions, but perhaps it does and it just goes away super quick. Any clues at all as to what might be causing this, or how to fix it? I REALLY don't want to lose another day uninstalling and reinstalling and patching and so on =) If that even fixes it. Here is the stack trace (process explorer) that I get from the threads window when the msvb7.dll is churning. --- title in process explorer [threads] tab for process -------- cpu:49.28% cswitch delta: 300 to 3500 startaddress: [msvb7.dll+0x4218c] msvb7.dll version: 9.0.30729.1 --- actual stack trace ------- ntkrnlpa.exe!KiUnexpectedInterrupt+0x121 ntkrnlpa.exe!ZwYieldExecution+0x1c56 ntkrnlpa.exe!KiDispatchInterrupt+0x72e NDIS.sys!NdisFreeToBlockPool+0x15e1 // shortened stack trace. all of these are from msvb7, msvb7.dll+0x46ce7 <- 0x2676a <- 0x2698e <- 0x38031 <- 0x2659f <- 0x26644 msvb7.dll+0x25f29 <- 0x2ac7a <- 0x27522 <- 0x274a0 <- 0x2b5ce <- 0x2b6e4 msvb7.dll+0x67d0a <- 0x68551 <- 0x6817b <- 0x681f0 <- 0x67c38 <- 0x65fa8 msvb7.dll+0x666c6 <- 0x6672c <- 0x6673d <- 0x6677c <- 0x667b4 <- 0x63c77 msvb7.dll+0x63e97 <- 0x42c3a <- 0x42bc1 <- 0x41bd7 kernel32.dll!GetModuleFileNameA+0x1b4 This is the list of stuff from "copy info" in help-about, shortened to a resonable length. Microsoft Visual Studio 2008 | Version 9.0.30729.1 SP Microsoft Visual Studio 2008 Professional Edition - ENU Service Pack 1 (KB945140) KB945140 Microsoft .NET Framework | Version 3.5 SP1 Microsoft Visual Basic 2008 Microsoft Visual C# 2008 Microsoft Visual F# for Visual Studio 2008 Microsoft Visual Studio 2008 Team Explorer | Version 9.0.30729.1 Microsoft Visual Studio 2008 Tools for Office Microsoft Visual Web Developer 2008 Hotfix for Microsoft Visual Studio 2008 Professional Edition - ENU KB944899, KB945282, KB946040, KB946308, KB946344, KB946581, KB947171 KB947173, KB947180, KB947540, KB947789, KB948127, KB946260, KB946458, KB948816 Microsoft Recipe Framework Package 8.0 Process Editor WIT Designer 1.4.0.0 Process Editor for Microsoft Visual Studio Team Foundation Server, Version 1.4.0.0 tangible T4 Editor 9.0 tangible T4 Text Template Editor - T4 Editor tangibleprojectsystem 1.0 Team Foundation Server Power Tools October 2008 SQL Prompt 4.0 (disabled)

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  • SUSE EC2 Problem - zypper - Permission denied

    - by phuu
    I'm trying to use zypper to install gcc on my Amazon EC2 instance running SUSE.When I try:zypper in gcc I get: Retrieving repository 'SLE11-SDK-SP1' metadata [] Permission to access 'http://eu-west-1-ec2-update.susecloud.net/repo/install/SLE11-SDK-SP1/sle-11-i586/media.1/media' denied. Abort, retry, ignore? [a/r/i/?] (a): i Retrieving repository 'SLE11-SDK-SP1' metadata [error] Repository 'SLE11-SDK-SP1' is invalid. Can't provide /media.1/media : User-requested skipping of a file Please check if the URIs defined for this repository are pointing to a valid repository. Warning: Disabling repository 'SLE11-SDK-SP1' because of the above error. Retrieving repository 'SLE11-SDK-SP1-Updates' metadata [|] Permission to access 'http://eu-west-1-ec2-update.susecloud.net/repo/update/SLE11-SDK-SP1-Updates/sle-11-i586/repodata/repomd.xml' denied. Abort, retry, ignore? [a/r/i/?] (a): i Retrieving repository 'SLE11-SDK-SP1-Updates' metadata [error] Repository 'SLE11-SDK-SP1-Updates' is invalid. Can't provide /repodata/repomd.xml : User-requested skipping of a file Please check if the URIs defined for this repository are pointing to a valid repository. Warning: Disabling repository 'SLE11-SDK-SP1-Updates' because of the above error. Retrieving repository 'SLES11-Extras' metadata [/] Permission to access 'http://eu-west-1-ec2-update.susecloud.net/repo/update/SLES11-Extras/sle-11-i586/repodata/repomd.xml' denied. Abort, retry, ignore? [a/r/i/?] (a): r Permission to access 'http://eu-west-1-ec2-update.susecloud.net/repo/update/SLES11-Extras/sle-11-i586/repodata/repomd.xml' denied. Abort, retry, ignore? [a/r/i/?] (a): zypper in gcc Invalid answer 'zypper in gcc'. [a/r/i/?] (a): a Retrieving repository 'SLES11-Extras' metadata [error] Repository 'SLES11-Extras' is invalid. Can't provide /repodata/repomd.xml : Please check if the URIs defined for this repository are pointing to a valid repository. Warning: Disabling repository 'SLES11-Extras' because of the above error. Retrieving repository 'SLES11-SP1' metadata [-] Permission to access 'http://eu-west-1-ec2-update.susecloud.net/repo/install/SLES11-SP1/sle-11-i586/media.1/media' denied. Abort, retry, ignore? [a/r/i/?] (a): a Retrieving repository 'SLES11-SP1' metadata [error] Repository 'SLES11-SP1' is invalid. Can't provide /media.1/media : Please check if the URIs defined for this repository are pointing to a valid repository. Warning: Disabling repository 'SLES11-SP1' because of the above error. Retrieving repository 'SLES11-SP1-Updates' metadata [] Permission to access 'http://eu-west-1-ec2-update.susecloud.net/repo/update/SLES11-SP1-Updates/sle-11-i586/repodata/repomd.xml' denied. I've search for the problem and this thread came up, but offered no solutions.I've triedsces-activate. Am I doing something wrong? I should say I'm very new to this, and I admit I don't really know what I'm doing, but I'm trying to learn about setting up and running a server and so I thought I'd throw myself in at the deep(ish) end. Thanks for reading.

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  • Dealing with HTTP w00tw00t attacks

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

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  • Bypass BIOS password set by faulty Toshiba firmware on Satellite A55 laptop?

    - by Brian
    How can the CMOS be cleared on the Toshiba Satellite A55-S1065? I have this 7 year old laptop that has been crippled by a glitch in its BIOS: 'A "Password =" prompt may be displayed when the computer is turned on, even though no power-on password has been set. If this happens, there is no password that will satisfy the password request. The computer will be unusable until this problem is resolved. [..] The occurrence of this problem on any particular computer is unpredictable -- it may never happen, but it could happen any time that the computer is turned on. [..] Toshiba will cover the cost of this repair under warranty until Dec 31, 2010.' -Toshiba As they stated, this machine is "unusable." The escape key does not bypass the prompt (nor does any other key), thus no operating system can be booted and no firmware updates can be installed. After doing some research, I found solutions that have been suggested for various Toshiba Satellite models afflicted by this glitch: "Make arrangements with a Toshiba Authorized Service Provider to have this problem resolved." -Toshiba (same link). Even prior to the expiration of Toshiba's support ("repair under warranty until Dec 31, 2010"), there have been reports that this solution is prohibitively expensive, labor charges accruing even when the laptop is still under warranty, and other reports that are generally discouraging: "They were unable to fix it and the guy who worked on it said he couldn’t find the jumpers on the motherboard to clear the BIOS. I paid $39 for my troubles and still have the password problem." - Steve. Since the costs of the repairs can now exceed the value of the hardware, it would seem this is a DIY solution, or a non-solution (i.e. the hardware is trash). Build a Toshiba parallel loopback by stripping and soldering the wires on a DB25 plug to connect connect these pins: 1-5-10, 2-11, 3-17, 4-12, 6-16, 7-13, 8-14, 9-15, 18-25. -CGSecurity. According to a list of supported models on pwcrack, this will likely not work for my Satellite A55-1065 (as well as many other models of similar age). -pwcrack Disconnect the laptop battery for an extended period of time. Doesn't work, laptop sat in a closet for several years without the battery connected and I forgot about the whole thing for awhile. The poor thing. Clear CMOS by setting the proper jumper setting or by removing the CMOS (RTC) battery, or by short circuiting a (hidden?) jumper that looks like a pair of solder marks -various sources for various Satellite models: Satellite A105: "you will see C88 clearly labeled right next the jack that the wireless card plugs into. There are two little solder squares (approx 1/16") at this location" -kerneltrap Satellite 1800: "Underneath the RAM there is black sticker, peel off the black sticker and you will reveal two little solder marks which are actually 'jumpers'. Very carefully hold a flat-head screwdriver touching both points and power on the unit briefly, effectively 'shorting' this circuit." -shadowfax2020 Satellite L300: "Short the B500 solder pads on the system board." -Lester Escobar Satellite A215: "Short the B500 solder pads on the system board." -fixya Clearing the CMOS could resolve the issue, but I cannot locate a jumper or a battery on this board. Nothing that looks remotely like a battery can be removed (everything is soldered). I have looked closely at the area around the memory and do not see any obvious solder pads that could be a secret jumper. Here are pictures (click for full resolution) : Where is the jumper (or solder pads) to short circuit and wipe the CMOS on this board? Possibly related questions: Remove Toshiba laptop BIOS password? Password Problem Toshiba Satellite..

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  • SUSE EC2 Problem - zypper - Permission denied

    - by phuu
    Hi. I'm trying to use zypper to install gcc on my Amazon EC2 instance running SUSE.When I try:zypper in gcc I get: Retrieving repository 'SLE11-SDK-SP1' metadata [] Permission to access 'http://eu-west-1-ec2-update.susecloud.net/repo/install/SLE11-SDK-SP1/sle-11-i586/media.1/media' denied. Abort, retry, ignore? [a/r/i/?] (a): i Retrieving repository 'SLE11-SDK-SP1' metadata [error] Repository 'SLE11-SDK-SP1' is invalid. Can't provide /media.1/media : User-requested skipping of a file Please check if the URIs defined for this repository are pointing to a valid repository. Warning: Disabling repository 'SLE11-SDK-SP1' because of the above error. Retrieving repository 'SLE11-SDK-SP1-Updates' metadata [|] Permission to access 'http://eu-west-1-ec2-update.susecloud.net/repo/update/SLE11-SDK-SP1-Updates/sle-11-i586/repodata/repomd.xml' denied. Abort, retry, ignore? [a/r/i/?] (a): i Retrieving repository 'SLE11-SDK-SP1-Updates' metadata [error] Repository 'SLE11-SDK-SP1-Updates' is invalid. Can't provide /repodata/repomd.xml : User-requested skipping of a file Please check if the URIs defined for this repository are pointing to a valid repository. Warning: Disabling repository 'SLE11-SDK-SP1-Updates' because of the above error. Retrieving repository 'SLES11-Extras' metadata [/] Permission to access 'http://eu-west-1-ec2-update.susecloud.net/repo/update/SLES11-Extras/sle-11-i586/repodata/repomd.xml' denied. Abort, retry, ignore? [a/r/i/?] (a): r Permission to access 'http://eu-west-1-ec2-update.susecloud.net/repo/update/SLES11-Extras/sle-11-i586/repodata/repomd.xml' denied. Abort, retry, ignore? [a/r/i/?] (a): zypper in gcc Invalid answer 'zypper in gcc'. [a/r/i/?] (a): a Retrieving repository 'SLES11-Extras' metadata [error] Repository 'SLES11-Extras' is invalid. Can't provide /repodata/repomd.xml : Please check if the URIs defined for this repository are pointing to a valid repository. Warning: Disabling repository 'SLES11-Extras' because of the above error. Retrieving repository 'SLES11-SP1' metadata [-] Permission to access 'http://eu-west-1-ec2-update.susecloud.net/repo/install/SLES11-SP1/sle-11-i586/media.1/media' denied. Abort, retry, ignore? [a/r/i/?] (a): a Retrieving repository 'SLES11-SP1' metadata [error] Repository 'SLES11-SP1' is invalid. Can't provide /media.1/media : Please check if the URIs defined for this repository are pointing to a valid repository. Warning: Disabling repository 'SLES11-SP1' because of the above error. Retrieving repository 'SLES11-SP1-Updates' metadata [] Permission to access 'http://eu-west-1-ec2-update.susecloud.net/repo/update/SLES11-SP1-Updates/sle-11-i586/repodata/repomd.xml' denied. I've search for the problem and this thread came up, but offered no solutions.I've triedsces-activate. Am I doing something wrong? I should say I'm very new to this, and I admit I don't really know what I'm doing, but I'm trying to learn about setting up and running a server and so I thought I'd throw myself in at the deep(ish) end. Thanks for reading.

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  • Internal drives vs USB-3 with external SSD or eSata with External SSD

    - by normstorm
    I have a need to carry VMWare Virtual Machines with me for work. These are very large files (each VM is 20GB or more) and I carry around about 40 to 50 VM's to simulate different software configurations for different client needs. Key: they won't fit on the internal hard drive of my current laptop. I currently execute the VM's from an external 7200RPM 2.5" USB-2 drive. I keep copies of the VM's on other 5400 external USB-2 drives. The VM's work from this drive, but they are slow, costing me much time and frustration. It can take upwards of 30 minutes just to make a copy of one of the VM's. They can take upwards of 10-15 minutes to fully launch and then they operate sluggishly. I am buying a new laptop (Core I7, 8GB RAM and other high-end specs). I intend to buy an SSD for the O/S volume (C:). This SSD will not be large enough to hold the VM's. I have always wanted a second internal hard drive to operate the VM's. To have two hard drives, though, I am finding that I will have to go to a 17" laptop which would be bulky/heavy. I am instead considering purchasing a 15" laptop with either an eSATA port or USB-3 ports and then purchasing two external drives. One of the drives might be an external SSD (maybe OCX brand) for operating the VM's and the other a 7400RPM 1TB hard drive for carrying around the VM's not currently in use. The question is which options would give me the biggest bang for the buck and the weight: 1) 2nd Internal SSD hard drive. This would mean buying a 17" laptop with two drive "bays". The first bay would hold an SSD drive for the C: drive. I would leave the first bay empty from the manufacture and then purchase/install an aftermarket SSD drive. This second SSD drive would have to be very large (256 GB), which would be expensive. I would still also need another external hard drive for carrying around the VM's not in use. 2) 2nd internal hard drive - 7400 RPM. Again, a 17" laptop would be required, but there are models available with on SSD drive for the C: drive and a second 7200 RPM hard drives. The second drive could probably be large enough to hold the VM's in use as well as those not in use. But would it be fast enough to drive the VM's? 3) USB-3 with External SSD. I could buy a 15" laptop with an SSD drive for the C: drive and a second hard drive for general files. I would operate the VM's from an external USB-3 SSD drive and have a third USB-3 external 7200 RPM drive for holding the VM's not in use. 4) eSATA with External SSD. Ditto, just eSATA instead of USB-3 5) USB-3 with External 7400 RPM drive. Ditto, but the drive running the VM's would be USB-3 attached 7400 RPM drives rather than SSD. 6) eSATA with External 7400 RPM drive. Dittor, but the drive running the VM's would be eSATA attached 7400 RPM drives rather than SSD. Any thoughts on this and any creative solutions?

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  • How Do I Enable My Ubuntu Server To Host Various SSL-Enabled Websites?

    - by Andy Ibanez
    Actually, I Have looked around for a few hours now, but I can't get this to work. The main problem I'm having is that only one out of two sites works. I have my website which will mostly be used for an app. It's called atajosapp.com . atajosapp.com will have three main sites: www.atajosapp.com <- Homepage for the app. auth.atajosapp.com <- Login endpoint for my API (needs SSL) api.atajosapp.com <- Main endpoint for my API (needs SSL). If you attempt to access api.atajosapp.com it works. It will throw you a 403 error and a JSON output, but that's fully intentional. If you try to access auth.atajosapp.com however, the site simply doesn't load. Chrome complains with: The webpage at https://auth.atajosapp.com/ might be temporarily down or it may have moved permanently to a new web address. Error code: ERR_TUNNEL_CONNECTION_FAILED But the website IS there. If you try to access www.atajosapp.com or any other HTTP site, it connects fine. It just doesn't like dealing with more than one HTTPS websites, it seems. The VirtualHost for api.atajosapp.com looks like this: <VirtualHost *:443> DocumentRoot /var/www/api.atajosapp.com ServerName api.atajosapp.com SSLEngine on SSLCertificateFile /certificates/STAR_atajosapp_com.crt SSLCertificateKeyFile /certificates/star_atajosapp_com.key SSLCertificateChainFile /certificates/PositiveSSLCA2.crt </VirtualHost> auth.atajosapp.com Looks very similar: <VirtualHost *:443> DocumentRoot /var/www/auth.atajosapp.com ServerName auth.atajosapp.com SSLEngine on SSLCertificateFile /certificates/STAR_atajosapp_com.crt SSLCertificateKeyFile /certificates/star_atajosapp_com.key SSLCertificateChainFile /certificates/PositiveSSLCA2.crt </VirtualHost> Now I have found many websites that talk about possible solutions. At first, I was getting a message like this: _default_ VirtualHost overlap on port 443, the first has precedence But after googling for hours, I managed to solve it by editing both apache2.conf and ports.conf. This is the last thing I added to ports.conf: <IfModule mod_ssl.c> NameVirtualHost *:443 # SSL name based virtual hosts are not yet supported, therefore no # NameVirtualHost statement here NameVirtualHost *:443 Listen 443 </IfModule> Still, right now only api.atajosapp.com and www.atajosapp.com are working. I still can't access auth.atajosapp.com. When I check the error log, I see this: Init: Name-based SSL virtual hosts only work for clients with TLS server name indication support (RFC 4366) I don't know what else to do to make both sites work fine on this. I purchased a Wildcard SSL certificate from Comodo that supposedly secures *.atajosapp.com, so after hours trying and googling, I don't know what's wrong anymore. Any help will be really appreciated. EDIT: I just ran the apachectl -t -D DUMP_VHOSTS command and this is the output. Can't make much sense of it...: root@atajosapp:/# apachectl -t -D DUMP_VHOSTS apache2: Could not reliably determine the server's fully qualified domain name, using atajosapp.com for ServerName [Thu Nov 07 02:01:24 2013] [warn] NameVirtualHost *:443 has no VirtualHosts VirtualHost configuration: wildcard NameVirtualHosts and _default_ servers: *:443 is a NameVirtualHost default server api.atajosapp.com (/etc/apache2/sites-enabled/api.atajosapp.com:1) port 443 namevhost api.atajosapp.com (/etc/apache2/sites-enabled/api.atajosapp.com:1) port 443 namevhost auth.atajosapp.com (/etc/apache2/sites-enabled/auth.atajosapp.com:1) *:80 is a NameVirtualHost default server atajosapp.com (/etc/apache2/sites-enabled/000-default:1) port 80 namevhost atajosapp.com (/etc/apache2/sites-enabled/000-default:1)

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  • Network authentication + roaming home directory - which technology should I look into using?

    - by Brian
    I'm looking into software which provides a user with a single identity across multiple computers. That is, a user should have the same permissions on each computer, and the user should have access to all of his or her files (roaming home directory) on each computer. There seem to be many solutions for this general idea, but I'm trying to determine the best one for me. Here are some details along with requirements: The network of machines are Amazon EC2 instances running Ubuntu. We access the machines with SSH. Some machines on this LAN may have different uses, but I am only discussing machines for a certain use (running a multi-tenancy platform). The system will not necessarily have a constant amount of machines. We may have to permanently or temporarily alter the amount of machines running. This is the the reason why I'm looking into centralized authentication/storage. The implementation of this effect should be a secure one. We're unsure if users will have direct shell access, but their software will potentially be running (under restricted Linux user names, of course) on our systems, which is as good as direct shell access. Let's assume that their software could potentially be malicious for the sake of security. I have heard of several technologies/combinations to achieve my goal, but I'm unsure of the ramifications of each. An older ServerFault post recommended NFS & NIS, though the combination has security problems according to this old article by Symantec. The article suggests moving to NIS+, but, as it is old, this Wikipedia article has cited statements suggesting a trending away from NIS+ by Sun. The recommended replacement is another thing I have heard of... LDAP. It looks like LDAP can be used to save user information in a centralized location on a network. NFS would still need to be used to cover the 'roaming home folder' requirement, but I see references of them being used together. Since the Symantec article pointed out security problems in both NIS and NFS, is there software to replace NFS, or should I heed that article's suggestions for locking it down? I'm tending toward LDAP because another fundamental piece of our architecture, RabbitMQ, has a authentication/authorization plugin for LDAP. RabbitMQ will be accessible in a restricted manner to users on the system, so I would like to tie the security systems together if possible. Kerberos is another secure authentication protocol that I have heard of. I learned a bit about it some years ago in a cryptography class but don't remember much about it. I have seen suggestions online that it can be combined with LDAP in several ways. Is this necessary? What are the security risks of LDAP without Kerberos? I also remember Kerberos being used in another piece of software developed by Carnegie Mellon University... Andrew File System, or AFS. OpenAFS is available for use, though its setup seems a bit complicated. At my university, AFS provides both requirements... I can log in to any machine, and my "AFS folder" is always available (at least when I acquire an AFS token). Along with suggestions for which path I should look into, does anybody have any guides which were particularly helpful? As the bold text pointed out, LDAP looks to be the best choice, but I'm particularly interested in the implementation details (Keberos? NFS?) with respect to security.

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  • I want to change DPI with Imagemagick without changing the actual byte-size of the image data

    - by user1694803
    I feel so horribly sorry that I have to ask this question here, but after hours of researching how to do an actually very simple task I'm still failing... In Gimp there is a very simple way to do what I want. I only have the German dialog installed but I'll try to translate it. I'm talking about going to "Picture-PrintingSize" and then adjusting the Values "X-Resolution" and "Y-Resolution" which are known to me as so called DPI values. You can also choose the format which by default is "Pixel/Inch". (In German the dialog is "Bild-Druckgröße" and there "X-Auflösung" and "Y-Auflösung") Ok, the values there are often "72" by default. When I change them to e.g. "300" this has the effect that the image stays the same on the computer, but if I print it, it will be smaller if you look at it, but all the details are still there, just smaller - it has a higher resolution on the printed paper (but smaller size... which is fine for me). I am often doing that when I am working with LaTeX, or to be exact with the command "pdflatex" on a recent Ubuntu-Machine. When I'm doing the above process with Gimp manually everything works just fine. The images will appear smaller in the resulting PDF but with high printing quality. What I am trying to do is to automate the process of going into Gimp and adjusting the DPI values. Since Imagemagick is known to be superb and I used it for many other tasks I tried to achieve my goal with this tool. But it does just not do what I want. After trying a lot of things I think this actually is be the command that should be my friend: convert input.png -density 300 output.png This should set the DPI to 300, as I can read everywhere in the web. It seems to work. When I check the file it stays the same. file input.png output.png input.png: PNG image data, 611 x 453, 8-bit grayscale, non-interlaced output.png: PNG image data, 611 x 453, 8-bit grayscale, non-interlaced When I use this command, it seems like it did what I wanted: identify -verbose output.png | grep 300 Resolution: 300x300 PNG:pHYs : x_res=300, y_res=300, units=0 (Funny enough, the same output comes for input.png which confuses me... so this might be the wrong parameters to watch?) But when I now render my TeX with "pdflatex" the image is still big and blurry. Also when I open the image with Gimp again the DPI values are set to "72" instead of "300". So there actually was no effect at all. Now what is the problem here. Am I getting something completely wrong? I can't be that wrong since everything works just fine with Gimp... Thanks for any help in this. I am also open to other automated solutions which are easily done on a Linux system...

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  • SQL Server Issue: Could not allocate space for object ... primary filegroup is full

    - by Luke
    Trying to figure out a problem at an office that has SQL Server 2005 installed on Windows SBS Server 2008. Here's the setup: It's an office, and the person who set this all up is nowhere to be found. I'm the best hope they have... One of the programs they use on a workstation gives them an error of "Could not allocate space for object 'Billing' in database "MyDatabase" because primary filegroup is full" when trying to save an entry in their software. I searched around for hours, looking for possible solutions. One was to check for available disk space, and another was to defrag. I checked the hard drives on the server, and there is plenty of space free. I also defragged, which may have helped the problem somewhat. It's hard to say, because it seems like with the nature of the error, if you try over and over you might get it to actually save. My next step was to try to see if autogrowth was enabled on the database. This would seem to be a likely / possible solution, but I can't access the database! If I run the SQL Management Studio, I can log in as my Windows user and view the list of databases. However, if I try to do anything (actually view the database, view the properties, add or edit users), I get errors that I don't have permission. For what it's worth, I also tried runing Management Studio as Administrator, in case that would help. No difference, though. Now, what I'm guessing is going on -- from my limited knowledge of SQL and from reading online -- is that though I'm logged in as a Windows administrator, that account does NOT have SQL access. I do see a list of SQL users, including SA, but I again don't have permission to add one or to change the password on an existing one. And nobody at the office has any idea what the SQL passwords could be. So... here's my thinking thus far: 1 - The "Could not allocate" error likely points to a database that needs to be allowed to autogrow. Especially since I verified there is plenty of free space and the HD has been defragmented. 2 - Enabling autogrow would be very easy to do if I had the proper access within SQL Management Stuido. That leads me to this link: http://blogs.technet.com/b/sqlman/archive/2011/06/14/tips-amp-tricks-you-have-lost-access-to-sql-server-now-what.aspx It sounds like it's a step-by-step guide for giving me the access I need to SQL. I'm guessing that if I followed this guide, I would be able to then log in to the SQL server via Management Studio with the proper permissions, and would be able to enable autogrow (or simply view the status of the existing database), and hopefully solve the "Could not allocate space" problem! So I guess I have a few questions: 1 - Would you guys agree with my "diagnosis"? Think I'm barking up the right tree? 2 - Is there any risk at all in hurting / disabling / wrecking the current SQL database or setup with me going through the guide to regain SQL access? I understand that per the guide, I would have to temporarily shut down SQL, so obviously it wouldn't be accessible during that time. But it wouldn't be worth the risk if there's a chance I could mess anything up... Like I said, the workstations ARE currently accessing the database somehow, but nobody knows with what login info or anything. Basically, it's set up, it works (usually), but if they had to reload the software, nobody would know how. Any feedback would be appreciated!! The problem is such that it's not an emergency for them, but an annoyance. If I could fix it, it would be wonderful. But if not, I think they'll manage, especially as they are going to eventually stop using this software. Thank you so much for your time! Luke

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  • Prevent auto mounting Android sdcard under Linux Mint

    - by BullShark
    I recently obtained an older Android phone, so that I could test Android Apps on it. I've needed it because I have a Nexus 7 but not older Android versions, hardware, etc. to test on. I'm having a problem with it under Linux Mint with Cinnamon. When I plug the phone in, or remove and plug the sdcard from the phone back to it while the phone is plugged in, Linux automatically mounts the sdcard. This is a problem because once it is mounted under Linux, it dismounts from the phone running Android 2.3.5, and I can no longer test Android Apps I write that require the sdcard to be present, writable. I went to Menu System Tools System Settings System Details Removable Media, and it brings up this window. I have changed the settings to always "Ask what to do" on "Select how media should be handled". However, the sdcard still gets mounted and then I am asked how I want to open these files (media players, photo importers, file browser, etc.). If I click the checkbox for "Never prompt or start programs on media insertion", then the sdcard is mounted, and I am not asked how to open these files. Eject is just a noob word for Ubuntu users that means umount (unmount) like "Adminstrator" is another ubuntu noob word for the root user. And if I unmount the sdcard, the phone doesn't recognize it again until I take the sdcard out and plug it back in. The phone sees it for a brief moment until Linux Mint takes it over. There are 2 possible solutions and maybe more: 1) Prevent Linux from automounting sdcards some how 2) Tell Android not to allow the computer it is plugged into to take over the sdcard, HOW? Edit: I found out how to prevent the sdcard from being automatically mounted: Now it gets recognized by Linux: bullshark@beastlinux ~ $ dmesg | tail -n 25 [597212.218323] sd 21:0:0:0: [sde] Attached SCSI removable disk [597212.218639] sr 21:0:0:1: Attached scsi CD-ROM sr2 [597212.218910] sr 21:0:0:1: Attached scsi generic sg7 type 5 [597217.139373] sd 21:0:0:0: [sde] 3862528 512-byte logical blocks: (1.97 GB/1.84 GiB) [597217.140726] sd 21:0:0:0: [sde] No Caching mode page present [597217.140735] sd 21:0:0:0: [sde] Assuming drive cache: write through [597217.143595] sd 21:0:0:0: [sde] No Caching mode page present [597217.143602] sd 21:0:0:0: [sde] Assuming drive cache: write through [597217.152240] sde: sde1 [597389.751008] 4:2:1: cannot get freq at ep 0x84 [597390.238742] 4:2:1: cannot get freq at ep 0x84 [597624.903132] sde: detected capacity change from 1977614336 to 0 [597637.677763] sd 21:0:0:0: [sde] 3862528 512-byte logical blocks: (1.97 GB/1.84 GiB) [597637.679616] sd 21:0:0:0: [sde] No Caching mode page present [597637.679626] sd 21:0:0:0: [sde] Assuming drive cache: write through [597637.682508] sd 21:0:0:0: [sde] No Caching mode page present [597637.682515] sd 21:0:0:0: [sde] Assuming drive cache: write through [597637.692758] sde: sde1 [597661.857979] sde: detected capacity change from 1977614336 to 0 [597688.775455] sd 21:0:0:0: [sde] 3862528 512-byte logical blocks: (1.97 GB/1.84 GiB) [597688.776814] sd 21:0:0:0: [sde] No Caching mode page present [597688.776823] sd 21:0:0:0: [sde] Assuming drive cache: write through [597688.780055] sd 21:0:0:0: [sde] No Caching mode page present [597688.780062] sd 21:0:0:0: [sde] Assuming drive cache: write through [597688.788639] sde: sde1 bullshark@beastlinux ~ $ However, the phone still unmounts the sdcard upon being detected by Linux. Linux detects but does not mount, and a few seconds later: Edit #2 (Solution): I solved this one by changing the usb connection type (was usb mass storage) :

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