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  • Why does SFINAE not apply to this?

    - by Simon Buchan
    I'm writing some simple point code while trying out Visual Studio 10 (Beta 2), and I've hit this code where I would expect SFINAE to kick in, but it seems not to: template<typename T> struct point { T x, y; point(T x, T y) : x(x), y(y) {} }; template<typename T, typename U> struct op_div { typedef decltype(T() / U()) type; }; template<typename T, typename U> point<typename op_div<T, U>::type> operator/(point<T> const& l, point<U> const& r) { return point<typename op_div<T, U>::type>(l.x / r.x, l.y / r.y); } template<typename T, typename U> point<typename op_div<T, U>::type> operator/(point<T> const& l, U const& r) { return point<typename op_div<T, U>::type>(l.x / r, l.y / r); } int main() { point<int>(0, 1) / point<float>(2, 3); } This gives error C2512: 'point<T>::point' : no appropriate default constructor available Given that it is a beta, I did a quick sanity check with the online comeau compiler, and it agrees with an identical error, so it seems this behavior is correct, but I can't see why. In this case some workarounds are to simply inline the decltype(T() / U()), to give the point class a default constructor, or to use decltype on the full result expression, but I got this error while trying to simplify an error I was getting with a version of op_div that did not require a default constructor*, so I would rather fix my understanding of C++ rather than to just do what works. Thanks! *: the original: template<typename T, typename U> struct op_div { static T t(); static U u(); typedef decltype(t() / u()) type; }; Which gives error C2784: 'point<op_div<T,U>::type> operator /(const point<T> &,const U &)' : could not deduce template argument for 'const point<T> &' from 'int', and also for the point<T> / point<U> overload.

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  • jQuery 1.4.2 and IE6: change event not firing first time using keyboard

    - by macca1
    I've done a good amount of research on this and found a bunch of reported problems and solutions but the general consensus seems that all change problems in IE6 were fixed in jQuery 1.4.2. I'm having an issue where a change event is not firing in jQuery 1.4.2, but it did fire successfully in jQuery 1.3.2. This is in IE6. I'm about to submit a bug for this, but for my sanity I wanted to post it here first to see if there's something dumb I'm missing. I don't understand why this is working this way... <HTML> <HEAD> <TITLE>jQuery 1.4.2 Problem </TITLE> <script src="jquery-1.4.2.min.js" type="text/javascript"></script> <script> $(document).ready( function() { $("#firstBox").change(function() { alert("CHANGE"); }); // ONLOAD of document autofocus into the first element... $("form").find(":input:visible:first").focus() }); </script> </HEAD> <BODY> <form> <select id="firstBox"> <option value="" selected="selected">--</option> <option value="1">One</option> <option value="2">Two</option> </select> <br><br> <input size="10" id="secondBox"> </form> </BODY> </HTML> Simple enough, right? Onload of the page, give the first element focus. Onchange of the first element, alert. If you use the mouse, it works as expected. The page loads, the focus is in the drop down, you change the option, you get the alert. The problem is if you use the keyboard. The page loads, the focus is in the drop down, you press the down arrow. The option changes. Tab off the field, no alert. Weird. To make it even weirder, if you tab back into the field and change it again (all using the keyboard), the change event DOES fire after tab out this time. Any ideas?

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  • Separate specific #ifdef branches

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

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  • Question about Architecture for Viewing Images in ASP.NET MVC App

    - by Charlie Flowers
    I have an approach in mind for an image viewer in a web app, and want to get a sanity check and any thoughts you stackoverflowers might have. Here's the whirlwind nutshell summary: I'm working on an ASP.NET MVC application that will run in my company's retail stores. Even though it is a web application, we own the store machines and have control over them. We have a "windows agent" running on the store machine which we can talk to from the browser via javascript (it is a WCF service, and our web app has permission to talk to it from the browser). One of the web pages needs to be an "image viewer" page with some common things like Rotate & Zoom. Now, there are some WebForms controls that offer Rotate and Zoom. However, they take up server resources and generate a good bit of traffic between the server and the browser. For example, the Rotate function would cause an ajax call to the server, which would then generate a new image written to a .NET Canvas object, which would then be written to a file on the server, which would then be returned from the ajax call and refreshed inside the browser. Normally, that's a pretty good way of doing things. But in our case, we have code running on the store machine that we can communicate with. This leads me to consider the following approach: When the user asks to view an image, we tell our "windows agent" to download it from our image server to the store machine. We then redirect our browser to our image viewer page, which will pull the image from the local file we just wrote to the store machine. When the user clicks "Rotate", we cause JavaScript code in the browser to call our "windows agent" software, asking it to perform the "Rotate" function. The "windows agent" does the rotation using the same kind of imaging control that would formerly have been used on the server, but it does so now on the store machine. Javascript in the browser then refreshes the image on the page to show the newly rotated image. Zoom and similar features would be implemented the same way. This seems to be much more efficient, scalable, and responsive for the end-users. However, I've never heard of anything like it being done, mostly because it's rare to have this combination of a web app plus a "windows agent" on the client machine. What do you think? Feasible? Reasonable? Any pitfalls I overlooked or improvements / suggestions you can see? Has anyone done anything like this who would like to offer the wisdom of experience? Thanks!

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  • Cepstral Analysis for pitch detection

    - by Ohmu
    Hi! I'm looking to extract pitches from a sound signal. Someone on IRC just explain to me how taking a double FFT achieves this. Specifically: take FFT take log of square of absolute value (can be done with lookup table) take another FFT take absolute value I am attempting this using vDSP I can't understand how I didn't come across this technique earlier. I did a lot of hunting and asking questions; several weeks worth. More to the point, I can't understand why I didn't think of it. I am attempting to achieve this with vDSP library. it looks as though it has functions to handle all of these tasks. However, I'm wondering about the accuracy of the final result. I have previously used a technique which scours the frequency bins of a single FFT for local maxima. when it encounters one, it uses a cunning technique (the change in phase since the last FFT) to more accurately place the actual peak within the bin. I am worried that this precision will be lost with this technique I'm presenting here. I guess the technique could be used after the second FFT to get the fundamental accurately. But it kind of looks like the information is lost in step 2. as this is a potentially tricky process, could someone with some experience just look over what I'm doing and check it for sanity? also, I've heard there is an alternative technique involving fitting a quadratic over neighbouring bins. Is this of comparable accuracy? if so, I would favour it, as it doesn't involve remembering bin phases. so questions: does this approach makes sense? Can it be improved? I'm a bit worried about And the log square component; there seems to be a vDSP function to do exactly that: vDSP_vdbcon however, there is no indication it precalculates a log-table -- I assume it doesn't, as the FFT function requires an explicit pre-calculation function to be called and passed into it. and this function doesn't. Is there some danger of harmonics being picked up? is there any cunning way of making vDSP pull out the maxima, biggest first? Can anyone point me towards some research or literature on this technique? the main question: is it accurate enough? Can the accuracy be improved? I have just been told by an expert that the accuracy IS INDEED not sufficient. Is this the end of the line? Pi PS I get SO annoyed (npi) when I want to create tags, but cannot. :| I have suggested to the maintainers that SO keep track of attempted tags, but I'm sure I was ignored. we need tags for vDSP, accelerate framework, cepstral analysis

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  • Check my anagram code from a job interview in the past.

    - by Michael Dorgan
    Had the following as an interview question a while ago and choked so bad on basic syntax that I failed to advance (once the adrenalin kicks in, coding goes out the window.) Given a list of string, return a list of sets of strings that are anagrams of the input set. i.e. "dog","god", "foo" should return {"dog","god"}. Afterward, I created the code on my own as a sanity check and it's been around now for a bit. I'd welcome input on it to see if I missed anything or if I could have done it much more efficiently. Take it as a chance to improve myself and learn other techniques: void Anagram::doWork(list input, list &output) { typedef list SortType; SortType sortedInput; // sort each string and pair it with the original for(list<string>::iterator i = input.begin(); i != input.end(); ++i) { string tempString(*i); std::sort(tempString.begin(), tempString.end()); sortedInput.push_back(make_pair(*i, tempString)); } // Now step through the new sorted list for(SortType::iterator i = sortedInput.begin(); i != sortedInput.end();) { set<string> newSet; // Assume (hope) we have a match and pre-add the first. newSet.insert(i->first); // Set the secondary iterator one past the outside to prevent // matching the original SortType::iterator j = i; ++j; while(j != sortedInput.end()) { if(i->second == j->second) { // If the string matches, add it to the set and remove it // so that future searches need not worry about it newSet.insert(j->first); j = sortedInput.erase(j); } else { // else, next element ++j; } } // If size is bigger than our original push, we have a match - save it to the output if(newSet.size() > 1) { output.push_back(newSet); } // erase this element and update the iterator i = sortedInput.erase(i); } }

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  • Windows splash screen using GDI+

    - by Luther
    The eventual aim of this is to have a splash screen in windows that uses transparency but that's not what I'm stuck on at the moment. In order to create a transparent window, I'm first trying to composite the splash screen and text on an off screen buffer using GDI+. At the moment I'm just trying to composite the buffer and display it in response to a 'WM_PAINT' message. This isn't working out at the moment; all I see is a black window. I imagine I've misunderstood something with regards to setting up render targets in GDI+ and then rendering them (I'm trying to render the screen using straight forward GDI blit) Anyway, here's the code so far: //my window initialisation code void MyWindow::create_hwnd(HINSTANCE instance, const SIZE &dim) { DWORD ex_style = WS_EX_LAYERED ; //eventually I'll be making use of this layerd flag m_hwnd = CreateWindowEx( ex_style, szFloatingWindowClass , L"", WS_POPUP , 0, 0, dim.cx, dim.cy, null, null, instance, null); SetWindowLongPtr(m_hwnd ,0, (__int3264)(LONG_PTR)this); m_display_dc = GetDC(NULL); //This was sanity check test code - just loading a standard HBITMAP and displaying it in WM_PAINT. It worked fine //HANDLE handle= LoadImage(NULL , L"c:\\test_image2.bmp", IMAGE_BITMAP, 0, 0, LR_LOADFROMFILE); m_gdip_offscreen_bm = new Gdiplus::Bitmap(dim.cx, dim.cy); m_gdi_dc = Gdiplus::Graphics::FromImage(m_gdip_offscreen_bm);//new Gdiplus::Graphics(m_splash_dc );//window_dc ;m_splash_dc //this draws the conents of my splash screen - this works if I create a GDI+ context for the window, rather than for an offscreen bitmap. //For all I know, it might actually be working but when I try to display the contents on screen, it shows a black image draw_all(); //this is just to show that drawing something simple on the offscreen bit map seems to have no effect Gdiplus::Pen pen(Gdiplus::Color(255, 0, 0, 255)); m_gdi_dc->DrawLine(&pen, 0,0,100,100); DWORD last_error = GetLastError(); //returns '0' at this stage } And here's the snipit that handles the WM_PAINT message: ---8<----------------------- //Paint message snippit case WM_PAINT: { BITMAP bm; PAINTSTRUCT ps; HDC hdc = BeginPaint(vg->m_hwnd, &ps); //get the HWNDs DC HDC hdcMem = vg->m_gdi_dc->GetHDC(); //get the HDC from our offscreen GDI+ object unsigned int width = vg->m_gdip_offscreen_bm->GetWidth(); //width and height seem fine at this point unsigned int height = vg->m_gdip_offscreen_bm->GetHeight(); BitBlt(hdc, 0, 0, width, height, hdcMem, 0, 0, SRCCOPY); //this blits a black rectangle DWORD last_error = GetLastError(); //this was '0' vg->m_gdi_dc->ReleaseHDC(hdcMem); EndPaint(vg->m_hwnd, &ps); //end paint return 1; } ---8<----------------------- My apologies for the long post. Does anybody know what I'm not quite understanding regarding how you write to an offscreen buffer using GDI+ (or GDI for that matter)and then display this on screen? Thank you for reading.

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  • How to prevent google chrome from caching my inputs, esp hidden ones when user click back?

    - by melaos
    hi there, i have an asp.net mvc app which have quite a few hidden inputs to keep values around and formatting their names so that i can use the Model binding later when i submit the form. i stumble into a weird bug with chrome which i don't have with IE or Firefox when the user submits the form and click on the back button, i find that chrome will keep my hidden input values as well. this whole chunk is generated via javascript hence i believe chrome is caching this. function addProductRow(productId, productName) { if (productName != "") { //use guid to ensure that the row never repeats var guid = $.Guid.New(); var temp = parseFloat($(".tboProductCount").val()); //need the span to workaround for chrome var szHTML = "<tr valign=\"top\" id=\"productRow\"><td class=\"productIdCol\"><input type=\"hidden\" id=productRegsID" + temp + "\" name=\"productRegs[" + temp + "].productId\" value=\"" + productId + "\"/>" + "<span id=\"spanProdID" + temp + "\" name=\"spanProdID" + temp + "\" >" + productId + "</span>" + "</td>" //+ "<td><input type=\"text\" id=\"productRegName\" name=\"productRegs[" + temp + "].productName\" value=\"" + productName + "\" class=\"productRegName\" size=\"50\" readonly=\"readonly\"/></td>" + "<td><span id=\"productRegName\" name=\"productRegs[" + temp + "].productName\" class=\"productRegName\">"+ productName + "<\span></td>" + "<td id=\"" + guid + "\" class=\"productrowguid\" \>" + "<input type=\"text\" size=\"20\" id=\"productSerialNo" + temp + "\" name=\"productRegs[" + temp + "].serialNo\" value=\"" + "\" class=\"productSerialNo\" maxlength=\"18\" />" + "<a class=\"fancybox\" id=\"btnImgSerialNo" + temp + "\" href=\"#divSerialNo" + temp + "\"><img class=\"btnImgSerialNo\" src=\"Images/landing_14.gif\" /></a>" + "<span id=\"snFlag" + temp + "\" class=\"redWarning\"></span></td>" + "<td><input type=\"text\" id=\"productRegDate" + temp + "\" name=\"productRegs[" + temp + "].PurchaseDate\" readonly=\"readonly\" />" + "<span id=\"snRegDate" + temp + "\" class=\"redWarning\"></span></td>" + "<td align=\"center\"><img style=\"cursor:pointer\" id=\"btnImgDelete\" src=\"Images/btn_remove.gif\" onclick=\"javascript:removeProductRow('" + guid + "')\" /><div style=\"display:none;\"><div id=\"divSerialNo" + temp + "\" style=\"font-family:verdana;font-size:11px;width:600px\">" + serialnumbergeneral + "<br /><br />" + getSNImageByCategory(productId) + "</div></div></td>" + "</tr>"; $(".ProductRegistrationTable").append(szHTML); $("a.fancybox").fancybox(); //initialization $("#productRegDate" + temp).datepicker({ minDate: new Date(1996, 1 - 1, 1), maxDate: 0 }); //sanity check //s7test alert('1 '+$("#spanProdID" + temp)); alert('2 '+$("#productRegsID" + temp)); } //end function addNewProductRow i need the id to be refreshed when the user select a new product, but putting another span tag beside it shows that the span will have the new id will the hidden input will still have the previous id. is there an elegant way to workaround this issue? thanks

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

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

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  • Securing paths in PHP

    - by tjm
    I'm writing some PHP which takes some paths to different content directories, and uses these to include various parts of pages later. I'm trying to ensure that the paths are as they seem, and none of them break the rules of the application. I have PRIVATEDIR which must lie above DOCUMENT_ROOT (aka) PUBLICDIR. CONTENTDIR which must lie within PRIVATEDIR and not go back below PUBLICDIR and some other *DIR's which must remain within CONTENTDIR. Currently I set up some defaults, and then override the ones the user specifies and then sanity check them with the following. private function __construct($options) { error_reporting(0); if(is_array($options)) { $this->opts = array_merge($this->opts, $options); } if($this->opts['STATUS']==='debug') { error_reporting(E_ALL | E_NOTICE | E_STRICT); } $this->opts['PUBLICDIR'] = realpath($_SERVER['DOCUMENT_ROOT']) .DIRECTORY_SEPARATOR; $this->opts['PRIVATEDIR'] = realpath($this->opts['PUBLICDIR'] .$this->opts['PRIVATEDIR']) .DIRECTORY_SEPARATOR; $this->opts['CONTENTDIR'] = realpath($this->opts['PRIVATEDIR'] .$this->opts['CONTENTDIR']) .DIRECTORY_SEPARATOR; $this->opts['CACHEDIR'] = realpath($this->opts['PRIVATEDIR'] .$this->opts['CACHEDIR']) .DIRECTORY_SEPARATOR; $this->opts['ERRORDIR'] = realpath($this->opts['CONTENTDIR'] .$this->opts['ERRORDIR']) .DIRECTORY_SEPARATOR; $this->opts['TEMPLATEDIR' = realpath($this->opts['CONTENTDIR'] .$this->opts['TEMPLATEDIR']) .DIRECTORY_SEPARATOR; // then here I have to check that PRIVATEDIR is above PUBLICDIR // and that all the rest remain within private dir and don't drop // down into (or below) PUBLICDIR again. And die with an error if // they don't conform. } The thing is this seems like a lot of work to do, especially as it must be run, every time a page is accessed, before I can do anything else, e.g check for a cached version of the page I'm serving. Part of me is thinking, since all of these paths are predefined by the maintainer of the site, they SHOULD be aware of what paths they are allowing access to and ensuring they are secure. But, I think I'm thinking that because currently I am said maintainer, and I KNOW my paths conform to the rules. That said, I do want to secure this thing from any accidental errors by future maintainers (and I bet, now I've said above "I KNOW...", probably from myself somewhere down the line). This just feels like a suboptimal solution. I wonder how fast this would really be and what you would suggest to improve it or as an alternative? Thanks.

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  • rpm rollback ignoring rpms - no error output

    - by John H
    Issue rpm rollback is not working with a set of repackaged rpms created in the last couple days, but does work with more recent ones. [root@host1 repackage]# ls -l zsh-4.2.6-* -rw-r--r-- 1 root root 1788283 Apr 10 2011 zsh-4.2.6-3.el5.i386.rpm -rw-r--r-- 1 root root 1788691 Aug 18 04:38 zsh-4.2.6-5.el5.i386.rpm [root@host1 repackage]# rpm -q zsh zsh-4.2.6-6.el5 [root@host1 repackage]# rpm --test -Uvh --rollback 'Aug 18 01:00' [root@host1 repackage]# rpm -e zsh [root@host1 repackage]# [root@host1 repackage]# ls -l zsh* -rw-r--r-- 1 root root 1788283 Apr 10 2011 zsh-4.2.6-3.el5.i386.rpm -rw-r--r-- 1 root root 1788691 Aug 18 04:38 zsh-4.2.6-5.el5.i386.rpm -rw-r--r-- 1 root root 1789064 Aug 20 09:06 zsh-4.2.6-6.el5.i386.rpm [root@host1 repackage]# cp zsh-4.2.6-6.el5.i386.rpm /tmp [root@host1 repackage]# rpm --test -Uvh --rollback 'Aug 18 01:00' Rollback packages (+1/-0) to Mon Aug 20 09:02:16 2012 (0x50323558): Preparing... ########################################### [100%] Cleaning up repackaged packages: Removing /var/spool/repackage/zsh-4.2.6-6.el5.i386.rpm: [root@host1 repackage]# ls -l zsh-4.2.6-* -rw-r--r-- 1 root root 1788283 Apr 10 2011 zsh-4.2.6-3.el5.i386.rpm -rw-r--r-- 1 root root 1788691 Aug 18 04:38 zsh-4.2.6-5.el5.i386.rpm [root@host1 repackage]# cp /tmp/zsh-4.2.6-6.el5.i386.rpm . [root@host1 repackage]# rpm -Uvh --rollback 'Aug 18 01:00' Rollback packages (+1/-0) to Mon Aug 20 09:06:05 2012 (0x5032363d): Preparing... ########################################### [100%] 1:zsh ########################################### [ 50%] Cleaning up repackaged packages: Removing /var/spool/repackage/zsh-4.2.6-6.el5.i386.rpm: [root@host1 repackage]# rpm --test -Uvh --rollback 'April 9' [root@host1 repackage]# Now, if I run my test commands with -Uvvh I get debug messages to stderror which shows me that rpm reads each of the rpm files in /var/spool/repackage. The only interesting bit is the "expected size" but after searching, the expected size should be different, as it records the files as they are on the filesystem. D: opening db environment /var/lib/rpm/Packages joinenv D: opening db index /var/lib/rpm/Packages rdonly mode=0x0 D: locked db index /var/lib/rpm/Packages D: opening db index /var/lib/rpm/Installtid rdonly mode=0x0 D: opening db index /var/lib/rpm/Pubkeys rdonly mode=0x0 D: read h# 769 Header sanity check: OK D: ========== DSA pubkey id 53268101 37017186 (h#769) D: read h# 32 Header V3 DSA signature: OK, key ID 37017186 D: read h# 40 Header V3 DSA signature: OK, key ID 37017186 ... D: read h# 1753 Header V3 DSA signature: OK, key ID 37017186 D: Expected size: 3628918 = lead(96)+sigs(344)+pad(0)+data(3628478) D: Actual size: 3583695 D: /var/spool/repackage/Deployment_Guide-en-US-5.2-11.noarch.rpm: Header V3 DSA signature: OK, key ID 37017186 D: Expected size: 1100789 = lead(96)+sigs(344)+pad(0)+data(1100349) D: Actual size: 1109281 D: /var/spool/repackage/NetworkManager-0.7.0-10.el5_5.2.i386.rpm: Header V3 DSA signature: OK, key ID 37017186 D: Expected size: 1098167 = lead(96)+sigs(344)+pad(0)+data(1097727) D: Actual size: 1106179 D: /var/spool/repackage/NetworkManager-0.7.0-9.el5.i386.rpm: Header V3 DSA signature: OK, key ID 37017186 D: Expected size: 84351 = lead(96)+sigs(344)+pad(0)+data(83911) D: Actual size: 85378 ... D: Expected size: 1788276 = lead(96)+sigs(344)+pad(0)+data(1787836) D: Actual size: 1788691 D: /var/spool/repackage/zsh-4.2.6-5.el5.i386.rpm: Header V3 DSA signature: OK, key ID 37017186 D: --- erase h#1758 D: closed db index /var/lib/rpm/Pubkeys D: closed db index /var/lib/rpm/Installtid D: closed db index /var/lib/rpm/Packages D: closed db environment /var/lib/rpm/Packages D: May free Score board((nil)) I am able to copy these rpms out of the repackage directory and if I run them through cpio, extract the files. I also tried backing up and rebuilding the rpm database - no change. System Information: RHEL 5.8 rpm 4.4.2.3 /etc/yum.conf tsflags=repackage /etc/rpm/macros %_repackage_all_erasures 1

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  • Network Restructure Method for Double-NAT network

    - by Adrian
    Due to a series of poor network design decisions (mostly) made many years ago in order to save a few bucks here and there, I have a network that is decidedly sub-optimally architected. I'm looking for suggestions to improve this less-than-pleasant situation. We're a non-profit with a Linux-based IT department and a limited budget. (Note: None of the Windows equipment we have runs does anything that talks to the Internet nor do we have any Windows admins on staff.) Key points: We have a main office and about 12 remote sites that essentially double NAT their subnets with physically-segregated switches. (No VLANing and limited ability to do so with current switches) These locations have a "DMZ" subnet that are NAT'd on an identically assigned 10.0.0/24 subnet at each site. These subnets cannot talk to DMZs at any other location because we don't route them anywhere except between server and adjacent "firewall". Some of these locations have multiple ISP connections (T1, Cable, and/or DSLs) that we manually route using IP Tools in Linux. These firewalls all run on the (10.0.0/24) network and are mostly "pro-sumer" grade firewalls (Linksys, Netgear, etc.) or ISP-provided DSL modems. Connecting these firewalls (via simple unmanaged switches) is one or more servers that must be publically-accessible. Connected to the main office's 10.0.0/24 subnet are servers for email, tele-commuter VPN, remote office VPN server, primary router to the internal 192.168/24 subnets. These have to be access from specific ISP connections based on traffic type and connection source. All our routing is done manually or with OpenVPN route statements Inter-office traffic goes through the OpenVPN service in the main 'Router' server which has it's own NAT'ing involved. Remote sites only have one server installed at each site and cannot afford multiple servers due to budget constraints. These servers are all LTSP servers several 5-20 terminals. The 192.168.2/24 and 192.168.3/24 subnets are mostly but NOT entirely on Cisco 2960 switches that can do VLAN. The remainder are DLink DGS-1248 switches that I am not sure I trust well enough to use with VLANs. There is also some remaining internal concern about VLANs since only the senior networking staff person understands how it works. All regular internet traffic goes through the CentOS 5 router server which in turns NATs the 192.168/24 subnets to the 10.0.0.0/24 subnets according to the manually-configured routing rules that we use to point outbound traffic to the proper internet connection based on '-host' routing statements. I want to simplify this and ready All Of The Things for ESXi virtualization, including these public-facing services. Is there a no- or low-cost solution that would get rid of the Double-NAT and restore a little sanity to this mess so that my future replacement doesn't hunt me down? Basic Diagram for the main office: These are my goals: Public-facing Servers with interfaces on that middle 10.0.0/24 network to be moved in to 192.168.2/24 subnet on ESXi servers. Get rid of the double NAT and get our entire network on one single subnet. My understanding is that this is something we'll need to do under IPv6 anyway, but I think this mess is standing in the way.

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  • Determining the required depth and specifications for a server cabinet

    - by Bingu Bingme
    I'm trying to understand the considerations ("why") that go into determining the specifications ("what") for a rackmount server cabinet, in order to determine what sort of rack I should purchase for my home use. Since this is for home use, I won't be following certain best practices (eg. hot/cold aisle, not even air conditioning) and may be willing to sacrifice in various areas in order to reduce cost and footprint - but please advise if there are safety concerns or other considerations to note. The most basic specs for a server cabinet are the dimensions (external width x external depth x usable height). Width: commonly 600mm or 800mm (if the use case requires extra clearance around the sides, such as if there is lots of cabling). In my case and most common cases, I'm going to stick with 600mm. Height: Select a sufficiently tall rack to fit my equipment. But how much may I stuff into it? Eg, if there is a 15U rack, can I really populate it with 15U of servers, or should I leave 1U at top and bottom for air circulation? Depth: Racks commonly have external depth of 600mm (network equipment), 800mm, 1000mm, or even longer. I'm trying to see how to fit into the 800mm depth. With reference to http://www.server-racks.com/rack-mount-depth.html, I'm hoping to have the front and rear posts mounted ~ 28.5" (72cm) apart, which would leave only 8cm for front space and rear space. How much rear space (from rear posts to back of rack) do I really need? I won't use cable management arms, so can I mount a 72cm depth server since the power, KVM, network cables won't take up much depth? My most important equipment are all < 60cm depth (4U chassis) and should comfortably fit within the 800mm cabinet. The rest of the equipment are very old 1U servers that range from 65-72cm depth. I might still want to make further use of them, or I might discard them since they are so old. Even if the 72cm servers cannot be powered on in an 800mm rack, I should be able to use them as 1U shelves. But, what server depth can I expect to be able to operate? Or am I forced to upgrade to 1000mm depth racks in order to use any servers deeper than 60cm? With reference to best practices for HP racks, some other specs and installation considerations: There aren't any minimum recommendations for clearance on the sides of the rack. It is recommended to leave 48" front clearance. The 48" front clearance is based on 32" chassis depth, 13" to extend the rack rails and mate the inner/outer rails, and 3" for movement. If I don't use such rails (eg, use shelves instead), it should be sufficient to leave front clearance of chassis depth + 3". It is recommended to leave 30" rear clearance "to provide space for servicing the rack". I'm planning to back the rack into a corner of the room, and wheel it slightly out when I need to access the rear. If the wheeling plan is ok, I still need to know how much rear clearance is required for air circulation and ventilation purposes. Castor wheels and stabilising feet. Since I'm backing the rack into a corner of the room, I'll only be able to set the stabilising feet on the front corners. Thoughts on safety? The rack that I'm considering has front glass doors with side ventilation slits and fully perforated rear doors. I'm hoping this will be a good balance between temperature and noise (only ventilation slits facing out the front, while the rear is facing the walls). Or is the sound of high-rpm fans going to escape through the front slits anyway and destroy my sanity?

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  • What's up with LDoms: Part 2 - Creating a first, simple guest

    - by Stefan Hinker
    Welcome back! In the first part, we discussed the basic concepts of LDoms and how to configure a simple control domain.  We saw how resources were put aside for guest systems and what infrastructure we need for them.  With that, we are now ready to create a first, very simple guest domain.  In this first example, we'll keep things very simple.  Later on, we'll have a detailed look at things like sizing, IO redundancy, other types of IO as well as security. For now,let's start with this very simple guest.  It'll have one core's worth of CPU, one crypto unit, 8GB of RAM, a single boot disk and one network port.  CPU and RAM are easy.  The network port we'll create by attaching a virtual network port to the vswitch we created in the primary domain.  This is very much like plugging a cable into a computer system on one end and a network switch on the other.  For the boot disk, we'll need two things: A physical piece of storage to hold the data - this is called the backend device in LDoms speak.  And then a mapping between that storage and the guest domain, giving it access to that virtual disk.  For this example, we'll use a ZFS volume for the backend.  We'll discuss what other options there are for this and how to chose the right one in a later article.  Here we go: root@sun # ldm create mars root@sun # ldm set-vcpu 8 mars root@sun # ldm set-mau 1 mars root@sun # ldm set-memory 8g mars root@sun # zfs create rpool/guests root@sun # zfs create -V 32g rpool/guests/mars.bootdisk root@sun # ldm add-vdsdev /dev/zvol/dsk/rpool/guests/mars.bootdisk \ mars.root@primary-vds root@sun # ldm add-vdisk root mars.root@primary-vds mars root@sun # ldm add-vnet net0 switch-primary mars That's all, mars is now ready to power on.  There are just three commands between us and the OK prompt of mars:  We have to "bind" the domain, start it and connect to its console.  Binding is the process where the hypervisor actually puts all the pieces that we've configured together.  If we made a mistake, binding is where we'll be told (starting in version 2.1, a lot of sanity checking has been put into the config commands themselves, but binding will catch everything else).  Once bound, we can start (and of course later stop) the domain, which will trigger the boot process of OBP.  By default, the domain will then try to boot right away.  If we don't want that, we can set "auto-boot?" to false.  Finally, we'll use telnet to connect to the console of our newly created guest.  The output of "ldm list" shows us what port has been assigned to mars.  By default, the console service only listens on the loopback interface, so using telnet is not a large security concern here. root@sun # ldm set-variable auto-boot\?=false mars root@sun # ldm bind mars root@sun # ldm start mars root@sun # ldm list NAME STATE FLAGS CONS VCPU MEMORY UTIL UPTIME primary active -n-cv- UART 8 7680M 0.5% 1d 4h 30m mars active -t---- 5000 8 8G 12% 1s root@sun # telnet localhost 5000 Trying 127.0.0.1... Connected to localhost. Escape character is '^]'. ~Connecting to console "mars" in group "mars" .... Press ~? for control options .. {0} ok banner SPARC T3-4, No Keyboard Copyright (c) 1998, 2011, Oracle and/or its affiliates. All rights reserved. OpenBoot 4.33.1, 8192 MB memory available, Serial # 87203131. Ethernet address 0:21:28:24:1b:50, Host ID: 85241b50. {0} ok We're done, mars is ready to install Solaris, preferably using AI, of course ;-)  But before we do that, let's have a little look at the OBP environment to see how our virtual devices show up here: {0} ok printenv auto-boot? auto-boot? = false {0} ok printenv boot-device boot-device = disk net {0} ok devalias root /virtual-devices@100/channel-devices@200/disk@0 net0 /virtual-devices@100/channel-devices@200/network@0 net /virtual-devices@100/channel-devices@200/network@0 disk /virtual-devices@100/channel-devices@200/disk@0 virtual-console /virtual-devices/console@1 name aliases We can see that setting the OBP variable "auto-boot?" to false with the ldm command worked.  Of course, we'd normally set this to "true" to allow Solaris to boot right away once the LDom guest is started.  The setting for "boot-device" is the default "disk net", which means OBP would try to boot off the devices pointed to by the aliases "disk" and "net" in that order, which usually means "disk" once Solaris is installed on the disk image.  The actual devices these aliases point to are shown with the command "devalias".  Here, we have one line for both "disk" and "net".  The device paths speak for themselves.  Note that each of these devices has a second alias: "net0" for the network device and "root" for the disk device.  These are the very same names we've given these devices in the control domain with the commands "ldm add-vnet" and "ldm add-vdisk".  Remember this, as it is very useful once you have several dozen disk devices... To wrap this up, in this part we've created a simple guest domain, complete with CPU, memory, boot disk and network connectivity.  This should be enough to get you going.  I will cover all the more advanced features and a little more theoretical background in several follow-on articles.  For some background reading, I'd recommend the following links: LDoms 2.2 Admin Guide: Setting up Guest Domains Virtual Console Server: vntsd manpage - This includes the control sequences and commands available to control the console session. OpenBoot 4.x command reference - All the things you can do at the ok prompt

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  • .NET to iOS: From WinForms to the iPad

    - by RobertChipperfield
    One of the great things about working at Red Gate is getting to play with new technology - and right now, that means mobile. A few weeks ago, we decided that a little research into the tablet computing arena was due, and purely from a numbers point of view, that suggested the iPad as a good target device. A quick trip to iPhoneDevCon in San Diego later, and Marine and I came back full of ideas, and with some concept of how iOS development was meant to work. Here's how we went from there to the release of Stacks & Heaps, our geeky take on the classic "Snakes & Ladders" game. Step 1: Buy a Mac I've played with many operating systems in my time: from the original BBC Model B, through DOS, Windows, Linux, and others, but I'd so far managed to avoid buying fruit-flavoured computer hardware! If you want to develop for the iPhone, iPad or iPod Touch, that's the first thing that needs to change. If you've not used OS X before, the first thing you'll realise is that everything is different! In the interests of avoiding a flame war in the comments section, I'll only go so far as to say that a lot of my Windows-flavoured muscle memory no longer worked. If you're in the UK, you'll also realise your keyboard is lacking a # key, and that " and @ are the other way around from normal. The wonderful Ukelele keyboard layout editor restores some sanity here, as long as you don't look at the keyboard when you're typing. I couldn't give up the PC entirely, but a handy application called Synergy comes to the rescue - it lets you share a single keyboard and mouse between multiple machines. There's a few limitations: Alt-Tab always seems to go to the Mac, and Windows 7's UAC dialogs require the local mouse for security reasons, but it gets you a long way at least. Step 2: Register as an Apple Developer You can register as an Apple Developer free of charge, and that lets you download XCode and the iOS SDK. You also get the iPhone / iPad emulator, which is handy, since you'll need to be a paid member before you can deploy your apps to a real device. You can either enroll as an individual, or as a company. They both cost the same ($99/year), but there's a few differences between them. If you register as a company, you can add multiple developers to your team (all for the same $99 - not $99 per developer), and you get to use your company name in the App Store. However, you'll need to send off significantly more documentation to Apple, and I suspect the process takes rather longer than for an individual, where they just need to verify some credit card details. Here's a tip: if you're registering as a company, do so as early as possible. The approval process can take a while to complete, so get the application in in plenty of time. Step 3: Learn to love the square brackets! Objective-C is the language of the iPad. C and C++ are also supported, and if you're doing some serious game development, you'll probably spend most of your time in C++ talking OpenGL, but for forms-based apps, you'll be interacting with a lot of the Objective-C SDK. Like shifting from Ctrl-C to Cmd-C, it feels a little odd at first, with the familiar string.format(.) turning into: NSString *myString = [NSString stringWithFormat:@"Hello world, it's %@", [NSDate date]]; Thankfully XCode's auto-complete is normally passable, if not up to Visual Studio's standards, which coupled with a huge amount of content on Stack Overflow means you'll soon get to grips with the API. You'll need to get used to some terminology changes, though; here's an incomplete approximation: Coming from a .NET background, there's some luxuries you no longer have developing Objective C in XCode: Generics! Remember back in .NET 1.1, when all collections were just objects? Yup, we're back there now. ReSharper. Or, more generally, very much refactoring support. The not-many-keystrokes to rename a class, its file, and al references to it in Visual Studio turns into a much more painful experience in XCode. Garbage collection. This is actually rather less of an issue than you might expect: if you follow the rules, the reference counting provided by Objective C gets you a long way without too much pain. Circular references are their usual problematic self, though. Decent exception handling. You do have exceptions, but they're nowhere near as widely used. Generally, if something goes wrong, you get nil (see translation table above) back. Which brings me on to. Calling a method on a nil object isn't a failure - it just returns nil itself! There's many arguments for and against this, but personally I fall into the "stuff should fail as quickly and explicitly as possible" camp. Less specifically, I found that there's more chance of code failing at runtime rather than getting caught at compile-time: using the @selector(.) syntax to pass a method signature isn't (can't be) checked at compile-time, so the first you know about a typo is a crash when you try and call it. The solution to this is of course lots of great testing, both automated and manual, but I still find comfort in provably correct type safety being enforced in addition to testing. Step 4: Submit to the App Store Assuming you want to distribute to more than a handful of devices, you're going to need to submit your app to the Apple App Store. There's a few gotchas in terms of getting builds signed with the right certificates, and you'll be bouncing around between XCode and iTunes Connect a fair bit, but eventually you get everything checked off the to-do list, and are ready to upload your first binary! With some amount of anticipation, I pressed the Upload button in XCode, ready to release our creation into the world, but was instead greeted by an error informing me my XML file was malformed. Uh. A little Googling later, and it turned out that a simple rename from "Stacks&Heaps.app" to "StacksAndHeaps.app" worked around an XML escaping bug, and we were good to go. The next step is to wait for approval (or otherwise). After a couple of weeks of intensive development, this part is agonising. Did we make it? The Apple jury is still out at the moment, but our fingers are firmly crossed! In the meantime, you can see some screenshots and leave us your email address if you'd like us to get in touch when it does go live at the MobileFoo website. Step 5: Profit! Actually, that wasn't the idea here: Stacks & Heaps is free; there's no adverts, and we're not going to sell all your data either. So why did we do it? We wanted to get an idea of what it's like to move from coding for a desktop environment, to something completely different. We don't know whether in a year's time, the iPad will still be the dominant force, or whether Android will have smoothed out some bugs, tweaked the performance, and polished the UI, but I think it's a fairly sure bet that the tablet form factor is here to stay. We want to meet people who are using it, start chatting to them, and find out about some of the pain they're feeling. What better way to do that than do it ourselves, and get to write a cool game in the process?

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  • Establishing WebLogic Server HTTPS Trust of IIS Using a Microsoft Local Certificate Authority

    - by user647124
    Everyone agrees that self-signed and demo certificates for SSL and HTTPS should never be used in production and preferred not to be used elsewhere. Most self-signed and demo certificates are provided by vendors with the intention that they are used only to integrate within the same environment. In a vendor’s perfect world all application servers in a given enterprise are from the same vendor, which makes this lack of interoperability in a non-production environment an advantage. For us working in the real world, where not only do we not use a single vendor everywhere but have to make do with self-signed certificates for all but production, testing HTTPS between an IIS ASP.NET service provider and a WebLogic J2EE consumer application can be very frustrating to set up. It was for me, especially having found many blogs and discussion threads where various solutions were described but did not quite work and were all mostly similar but just a little bit different. To save both you and my future (who always seems to forget the hardest-won lessons) all of the pain and suffering, I am recording the steps that finally worked here for reference and sanity. How You Know You Need This The first cold clutches of dread that tells you it is going to be a long day is when you attempt to a WSDL published by IIS in WebLogic over HTTPS and you see the following: <Jul 30, 2012 2:51:31 PM EDT> <Warning> <Security> <BEA-090477> <Certificate chain received from myserver.mydomain.com - 10.555.55.123 was not trusted causing SSL handshake failure.> weblogic.wsee.wsdl.WsdlException: Failed to read wsdl file from url due to -- javax.net.ssl.SSLKeyException: [Security:090477]Certificate chain received from myserver02.mydomain.com - 10.555.55.123 was not trusted causing SSL handshake failure. The above is what started a three day sojourn into searching for a solution. Even people who had solved it before would tell me how they did, and then shrug when I demonstrated that the steps did not end in the success they claimed I would experience. Rather than torture you with the details of everything I did that did not work, here is what finally did work. Export the Certificates from IE First, take the offending WSDL URL and paste it into IE (if you have an internal Microsoft CA, you have IE, even if you don’t use it in favor of some other browser). To state the semi-obvious, if you received the error above there is a certificate configured for the IIS host of the service and the SSL port has been configured properly. Otherwise there would be a different error, usually about the site not found or connection failed. Once the WSDL loads, to the right of the address bar there will be a lock icon. Click the lock and then click View Certificates in the resulting dialog (if you do not have a lock icon but do have a Certificate Error message, see http://support.microsoft.com/kb/931850 for steps to install the certificate then you can continue from the point of finding the lock icon). Figure 1: View Certificates in IE Next, select the Details tab in the resulting dialog Figure 2: Use Certificate Details to Export Certificate Click Copy to File, then Next, then select the Base-64 encoded option for the format Figure 3: Select the Base-64 encoded option for the format For the sake of simplicity, I choose to save this to the root of the WebLogic domain. It will work from anywhere, but later you will need to type in the full path rather than just the certificate name if you save it elsewhere. Figure 4: Browse to Save Location Figure 5: Save the Certificate to the Domain Root for Convenience This is the point where I ran into some confusion. Some articles mentioned exporting the entire chain of certificates. This supposedly works for some types of certificates, or if you have a few other tools and the time to learn them. For the SSL experts out there, they already have these tools, know how to use them well, and should not be wasting their time reading this article meant for folks who just want to get things wired up and back to unit testing and development. For the rest of us, the easiest way to make sure things will work is to just export all the links in the chain individually and let WebLogic Server worry about re-assembling them into a chain (which it does quite nicely). While perhaps not the most elegant solution, the multi-step process is easy to repeat and uses only tools that are immediately available and require no learning curve. So… Next, go to Tools then Internet Options then the Content tab and click Certificates. Go to the Trust Root Certificate Authorities tab and find the certificate root for your Microsoft CA cert (look for the Issuer of the certificate you exported earlier). Figure 6: Trusted Root Certification Authorities Tab Export this one the same way as before, with a different name Figure 7: Use a Unique Name for Each Certificate Repeat this once more for the Intermediate Certificate tab. Import the Certificates to the WebLogic Domain Now, open an command prompt, navigate to [WEBLOGIC_DOMAIN_ROOT]\bin and execute setDomainEnv. You should then be in the root of the domain. If not, CD to the domain root. Assuming you saved the certificate in the domain root, execute the following: keytool -importcert -alias [ALIAS-1] -trustcacerts -file [FULL PATH TO .CER 1] -keystore truststore.jks -storepass [PASSWORD] An example with the variables filled in is: keytool -importcert -alias IIS-1 -trustcacerts -file microsftcert.cer -keystore truststore.jks -storepass password After several lines out output you will be prompted with: Trust this certificate? [no]: The correct answer is ‘yes’ (minus the quotes, of course). You’ll you know you were successful if the response is: Certificate was added to keystore If not, check your typing, as that is generally the source of an error at this point. Repeat this for all three of the certificates you exported, changing the [ALIAS-1] and [FULL PATH TO .CER 1] value each time. For example: keytool -importcert -alias IIS-1 -trustcacerts -file microsftcert.cer -keystore truststore.jks -storepass password keytool -importcert -alias IIS-2 -trustcacerts -file microsftcertRoot.cer -keystore truststore.jks -storepass password keytool -importcert -alias IIS-3 -trustcacerts -file microsftcertIntermediate.cer -keystore truststore.jks -storepass password In the above we created a new JKS key store. You can re-use an existing one by changing the name of the JKS file to one you already have and change the password to the one that matches that JKS file. For the DemoTrust.jks  that is included with WebLogic the password is DemoTrustKeyStorePassPhrase. An example here would be: keytool -importcert -alias IIS-1 -trustcacerts -file microsoft.cer -keystore DemoTrust.jks -storepass DemoTrustKeyStorePassPhrase keytool -importcert -alias IIS-2 -trustcacerts -file microsoftRoot.cer -keystore DemoTrust.jks -storepass DemoTrustKeyStorePassPhrase keytool -importcert -alias IIS-2 -trustcacerts -file microsoftInter.cer -keystore DemoTrust.jks -storepass DemoTrustKeyStorePassPhrase Whichever keystore you use, you can check your work with: keytool -list -keystore truststore.jks -storepass password Where “truststore.jks” and “password” can be replaced appropriately if necessary. The output will look something like this: Figure 8: Output from keytool -list -keystore Update the WebLogic Keystore Configuration If you used an existing keystore rather than creating a new one, you can restart your WebLogic Server and skip the rest of this section. For those of us who created a new one because that is the instructions we found online… Next, we need to tell WebLogic to use the JKS file (truststore.jks) we just created. Log in to the WebLogic Server Administration Console and navigate to Servers > AdminServer > Configuration > Keystores. Scroll down to “Custom Trust Keystore:” and change the value to “truststore.jks” and the value of “Custom Trust Keystore Passphrase:” and “Confirm Custom Trust Keystore Passphrase:” to the password you used when earlier, then save your changes. You will get a nice message similar to the following: Figure 9: To Be Safe, Restart Anyways The “No restarts are necessary” is somewhat of an exaggeration. If you want to be able to use the keystore you may need restart the server(s). To save myself aggravation, I always do. Your mileage may vary. Conclusion That should get you there. If there are some erroneous steps included for your situation in particular, I will offer up a semi-apology as the process described above does not take long at all and if there is one step that could be dropped from it, is still much faster than trying to figure this out from other sources.

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  • 10 tape technology features that make you go hmm.

    - by Karoly Vegh
    A week ago an Oracle/StorageTek Tape Specialist, Christian Vanden Balck, visited Vienna, and agreed to visit customers to do techtalks and update them about the technology boom going around tape. I had the privilege to attend some of his sessions and noted the information and features that took the customers by surprise and made them think. Allow me to share the top 10: I. StorageTek as a brand: StorageTek is one of he strongest names in the Tape field. The brand itself was valued so much by customers that even after Sun Microsystems acquiring StorageTek and the Oracle acquiring Sun the brand lives on with all the Oracle tapelibraries are officially branded StorageTek.See http://www.oracle.com/us/products/servers-storage/storage/tape-storage/overview/index.html II. Disk information density limitations: Disk technology struggles with information density. You haven't seen the disk sizes exploding lately, have you? That's partly because there are physical limits on a disk platter. The size is given, the number of platters is limited, they just can't grow, and are running out of physical area to write to. Now, in a T10000C tape cartridge we have over 1000m long tape. There you go, you have got your physical space and don't need to stuff all that data crammed together. You can write in a reliable pattern, and have space to grow too. III. Oracle has a market share of 62% worldwide in recording head manufacturing. That's right. If you are running LTO drives, with a good chance you rely on StorageTek production. That's two out of three LTO recording heads produced worldwide.  IV. You can store 1 Exabyte data in a single tape library. Yes, an Exabyte. That is 1000 Petabytes. Or, a million Terabytes. A thousand million GigaBytes. You can store that in a stacked StorageTek SL8500 tapelibrary. In one SL8500 you can put 10.000 T10000C cartridges, that store 10TB data (compressed). You can stack 10 of these SL8500s together. Boom. 1000.000 TB.(n.b.: stacking means interconnecting the libraries. Yes, cartridges are moved between the stacked libraries automatically.)  V. EMC: 'Tape doesn't suck after all. We moved on.': Do you remember the infamous 'Tape sucks, move on' Datadomain slogan? Of course they had to put it that way, having only had disk products. But here's a fun fact: on the EMCWorld 2012 there was a major presence of a Tape-tech company - EMC, in a sudden burst of sanity is embracing tape again. VI. The miraculous T10000C: Oracle StorageTek has developed an enterprise-grade tapedrive and cartridge, the T10000C. With awesome numbers: The Cartridge: Native 5TB capacity, 10TB with compression Over a kilometer long tape within the cartridge. And it's locked when unmounted, no rattling of your data.  Replaced the metalparticles datalayer with BaFe (bariumferrite) - metalparticles lose around 7% of magnetism within 30 days. BaFe does not. Yes we employ solid-state physicists doing R&D on demagnetisation in our labs. Can be partitioned, storage tiering within the cartridge!  The Drive: 2GB Cache Encryption implemented in HW - no performance hit 252 MB/s native sustained data rate, beats disk technology by far. Not to mention peak throughput.  Leading the tape while never touching the data side of it, protecting your data physically too Data integritiy checking (CRC recalculation) on tape within the drive without having to read it back to the server reordering data from tape-order, delivering it back in application-order  writing 32 tracks at once, reading them back for CRC check at once VII. You only use 20% of your data on a regular basis. The rest 80% is just lying around for years. On continuously spinning disks. Doubly consuming energy (power+cooling), blocking diskstorage capacity. There is a solution called SAM (Storage Archive Manager) that provides you a filesystem unifying disk and tape, moving data on-demand and for clients transparently between the different storage tiers. You can share these filesystems with NFS or CIFS for clients, and enjoy the low TCO of tape. Tapes don't spin. They sit quietly in their slots, storing 10TB data, using no energy, producing no heat, automounted when a client accesses their data.See: http://www.oracle.com/us/products/servers-storage/storage/storage-software/storage-archive-manager/overview/index.html VIII. HW supported for three decades: Did you know that the original PowderHorn library was released in '87 and has been only discontinued in 2010? That is over two decades of supported operation. Tape libraries are - just like the data carrying on tapecartridges - built for longevity. Oh, and the T10000C cartridge has 30-year archival life for long-term retention.  IX. Tape is easy to manage: Have you heard of Tape Storage Analytics? It is a central graphical tool to summarize, monitor, analyze dataflow, health and performance of drives and libraries, see: http://www.oracle.com/us/products/servers-storage/storage/tape-storage/tape-analytics/overview/index.html X. The next generation: The T10000B drives were able to reuse the T10000A cartridges and write on them even more data. On the same cartridges. We call this investment protection, and this is very important for Oracle for the future too. We usually support two generations of cartridges together. The current drive is a T10000C. (...I know I promised to enlist 10, but I got still two more I really want to mention. Allow me to work around the problem: ) X++. The TallBots, the robots moving around the cartridges in the StorageTek library from tapeslots to the drives are cableless. Cables, belts, chains running to moving parts in a library cause maintenance downtimes. So StorageTek eliminated them. The TallBots get power, commands, even firmwareupgrades through the rails they are running on. Also, the TallBots don't just hook'n'pull the tapes out of their slots, they actually grip'n'lift them out. No friction, no scratches, no zillion little plastic particles floating around in the library, in the drives, on your data. (X++)++: Tape beats SSDs and Disks. In terms of throughput (252 MB/s), in terms of TCO: disks cause around 290x more power and cooling, in terms of capacity: 10TB on a single media and soon more.  So... do you need to store large amounts of data? Are you legally bound to archive it for dozens of years? Would you benefit from automatic storage tiering? Have you got large mediachunks to be streamed at times? Have you got power and cooling issues in the growing datacenters? Do you find EMC's 180° turn of tape attitude interesting, but appreciate it at the same time? With all that, you aren't alone. The most data on this planet is stored on tape. Tape is coming. Big time.

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  • Exception Handling And Other Contentious Political Topics

    - by Justin Jones
    So about three years ago, around the time of my last blog post, I promised a friend I would write this post. Keeping promises is a good thing, and this is my first step towards easing back into regular blogging. I fully expect him to return from Pennsylvania to buy me a beer over this. However, it’s been an… ahem… eventful three years or so, and blogging, unfortunately, got pushed to the back burner on my priority list, along with a few other career minded activities. Now that the personal drama of the past three years is more or less resolved, it’s time to put a few things back on the front burner. What I consider to be proper exception handling practices is relatively well known these days. There are plenty of blog posts out there already on this topic which more or less echo my opinions on this topic. I’ll try to include a few links at the bottom of the post. Several years ago I had an argument with a co-worker who posited that exceptions should be caught at every level and logged. This might seem like sanity on the surface, but the resulting error log looked something like this: Error: System.SomeException Followed by small stack trace. Error: System.SomeException Followed by slightly bigger stack trace. Error: System.SomeException Followed by slightly bigger stack trace. Error: System.SomeException Followed by slightly bigger stack trace. Error: System.SomeException Followed by slightly bigger stack trace. Error: System.SomeException Followed by slightly bigger stack trace. Error: System.SomeException Followed by slightly bigger stack trace. Error: System.SomeException Followed by slightly bigger stack trace.   These were all the same exception. The problem with this approach is that the error log, if you run any kind of analytics on in, becomes skewed depending on how far up the stack trace your exception was thrown. To mitigate this problem, we came up with the concept of the “PreLoggedException”. Basically, we would log the exception at the very top level and subsequently throw the exception back up the stack encapsulated in this pre-logged type, which our logging system knew to ignore. Now the error log looked like this: Error: System.SomeException Followed by small stack trace. Much cleaner, right? Well, there’s still a problem. When your exception happens in production and you go about trying to figure out what happened, you’ve lost more or less all context for where and how this exception was thrown, because all you really know is what method it was thrown in, but really nothing about who was calling the method or why. What gives you this clue is the entire stack trace, which we’re losing here. I believe that was further mitigated by having the logging system pull a system stack trace and add it to the log entry, but what you’re actually getting is the stack for how you got to the logging code. You’re still losing context about the actual error. Not to mention you’re executing a whole slew of catch blocks which are sloooooooowwwww……… In other words, we started with a bad idea and kept band-aiding it until it didn’t suck quite so bad. When I argued for not catching exceptions at every level but rather catching them following a certain set of rules, my co-worker warned me “do yourself a favor, never express that view in any future interviews.” I suppose this is my ultimate dismissal of that advice, but I’m not too worried. My approach for exception handling follows three basic rules: Only catch an exception if 1. You can do something about it. 2. You can add useful information to it. 3. You’re at an application boundary. Here’s what that means: 1. Only catch an exception if you can do something about it. We’ll start with a trivial example of a login system that uses a file. Please, never actually do this in production code, it’s just concocted example. So if our code goes to open a file and the file isn’t there, we get a FileNotFound exception. If the calling code doesn’t know what to do with this, it should bubble up. However, if we know how to create the file from scratch we can create the file and continue on our merry way. When you run into situations like this though, What should really run through your head is “How can I avoid handling an exception at all?” In this case, it’s a trivial matter to simply check for the existence of the file before trying to open it. If we detect that the file isn’t there, we can accomplish the same thing without having to handle in in a catch block. 2. Only catch an exception if you can do something about it. Continuing with the poorly thought out file based login system we contrived in part 1, if the code calls a Login(…) method and the FileNotFound exception is thrown higher up the stack, the code that calls Login must account for a FileNotFound exception. This is kind of counterintuitive because the calling code should not need to know the internals of the Login method, and the data file is an implementation detail. What makes more sense, assuming that we didn’t implement any of the good advice from step 1, is for Login to catch the FileNotFound exception and wrap it in a new exception. For argument’s sake we’ll say LoginSystemFailureException. (Sorry, couldn’t think of anything better at the moment.) This gives us two stack traces, preserving the original stack trace in the inner exception, and also is much more informative to the calling code. 3. Only catch an exception if you’re at an application boundary. At some point we have to catch all the exceptions, even the ones we don’t know what to do with. WinForms, ASP.Net, and most other UI technologies have some kind of built in mechanism for catching unhandled exceptions without fatally terminating the application. It’s still a good idea to somehow gracefully exit the application in this case if possible though, because you can no longer be sure what state your application is in, but nothing annoys a user more than an application just exploding. These unhandled exceptions need to be logged, and this is a good place to catch them. Ideally you never want this option to be exercised, but code as though it will be. When you log these exceptions, give them a “Fatal” status (e.g. Log4Net) and make sure these bugs get handled in your next release. That’s it in a nutshell. If you do it right each exception will only get logged once and with the largest stack trace possible which will make those 2am emergency severity 1 debugging sessions much shorter and less frustrating. Here’s a few people who also have interesting things to say on this topic:  http://blogs.msdn.com/b/ericlippert/archive/2008/09/10/vexing-exceptions.aspx http://www.codeproject.com/Articles/9538/Exception-Handling-Best-Practices-in-NET I know there’s more but I can’t find them at the moment.

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  • Can't log in via SSH to any accounts set to use /bin/bash as a default shell

    - by Gui Ambros
    I'm trying to install bash as the default shell on a ARM Linux running on an embedded device (Synology DS212+ NAS). But there's something really wrong, and I can't figure out what it is. Symptoms: 1) Root has /bin/bash as default shell, and can log in normally via SSH: $ grep root /etc/passwd root:x:0:0:root:/root:/bin/bash $ ssh root@NAS root@NAS's password: Last login: Sun Dec 16 14:06:56 2012 from desktop # 2) joeuser has /bin/bash as default shell, and receives "Permission denied" when trying to log in via SSH: $ grep joeuser /etc/passwd joeuser:x:1029:100:Joe User:/home/joeuser:/bin/bash $ ssh joeuser@localhost joeuser@NAS's password: Last login: Sun Dec 16 14:07:22 2012 from desktop Permission denied, please try again. Connection to localhost closed. 3) changing joeuser's shell back to /bin/sh: $ grep joeuser /etc/passwd joeuser:x:1029:100:Joe User:/home/joeuser:/bin/sh $ ssh joeuser@localhost Last login: Sun Dec 16 15:50:52 2012 from localhost $ To make things even more strange, I can log in as joeuser using /bin/bash using the serial console (!). Also a su - joeuser as root works fine, so the bash binary itself is working fine. In an act of despair, I changed joeuser's uid to 0 on /etc/passwd, but also didn't work, so it doesn't seem to be anything permission related. Seems that bash is doing some extra checking that sshd didn't like, and blocking the connections for non-root users. Maybe some sort of sanity checking - or terminal emulation - that is triggering the SIGCHLD, but only when called via ssh. I already went through every single item on sshd_config, and also put SSHD in debug mode, but didn't find anything strange. Here's my /etc/ssh/sshd_config: LogLevel DEBUG LoginGraceTime 2m PermitRootLogin yes RSAAuthentication yes PubkeyAuthentication yes AuthorizedKeysFile %h/.ssh/authorized_keys ChallengeResponseAuthentication no UsePAM yes AllowTcpForwarding no ChrootDirectory none Subsystem sftp internal-sftp -f DAEMON -u 000 And here's the output from /usr/syno/sbin/sshd -d, showing the failed attempt of joeuser trying to log in, with /bin/bash as the shell: debug1: Config token is loglevel debug1: Config token is logingracetime debug1: Config token is permitrootlogin debug1: Config token is rsaauthentication debug1: Config token is pubkeyauthentication debug1: Config token is authorizedkeysfile debug1: Config token is challengeresponseauthentication debug1: Config token is usepam debug1: Config token is allowtcpforwarding debug1: Config token is chrootdirectory debug1: Config token is subsystem debug1: HPN Buffer Size: 87380 debug1: sshd version OpenSSH_5.8p1-hpn13v11 debug1: read PEM private key done: type RSA debug1: private host key: #0 type 1 RSA debug1: read PEM private key done: type DSA debug1: private host key: #1 type 2 DSA debug1: read PEM private key done: type ECDSA debug1: private host key: #2 type 3 ECDSA debug1: rexec_argv[0]='/usr/syno/sbin/sshd' debug1: rexec_argv[1]='-d' Set /proc/self/oom_adj from 0 to -17 debug1: Bind to port 22 on ::. debug1: Server TCP RWIN socket size: 87380 debug1: HPN Buffer Size: 87380 Server listening on :: port 22. debug1: Bind to port 22 on 0.0.0.0. debug1: Server TCP RWIN socket size: 87380 debug1: HPN Buffer Size: 87380 Server listening on 0.0.0.0 port 22. debug1: Server will not fork when running in debugging mode. debug1: rexec start in 6 out 6 newsock 6 pipe -1 sock 9 debug1: inetd sockets after dupping: 4, 4 Connection from 127.0.0.1 port 52212 debug1: HPN Disabled: 0, HPN Buffer Size: 87380 debug1: Client protocol version 2.0; client software version OpenSSH_5.8p1-hpn13v11 SSH: Server;Ltype: Version;Remote: 127.0.0.1-52212;Protocol: 2.0;Client: OpenSSH_5.8p1-hpn13v11 debug1: match: OpenSSH_5.8p1-hpn13v11 pat OpenSSH* debug1: Enabling compatibility mode for protocol 2.0 debug1: Local version string SSH-2.0-OpenSSH_5.8p1-hpn13v11 debug1: permanently_set_uid: 1024/100 debug1: MYFLAG IS 1 debug1: list_hostkey_types: ssh-rsa,ssh-dss,ecdsa-sha2-nistp256 debug1: SSH2_MSG_KEXINIT sent debug1: SSH2_MSG_KEXINIT received debug1: AUTH STATE IS 0 debug1: REQUESTED ENC.NAME is 'aes128-ctr' debug1: kex: client->server aes128-ctr hmac-md5 none SSH: Server;Ltype: Kex;Remote: 127.0.0.1-52212;Enc: aes128-ctr;MAC: hmac-md5;Comp: none debug1: REQUESTED ENC.NAME is 'aes128-ctr' debug1: kex: server->client aes128-ctr hmac-md5 none debug1: expecting SSH2_MSG_KEX_ECDH_INIT debug1: SSH2_MSG_NEWKEYS sent debug1: expecting SSH2_MSG_NEWKEYS debug1: SSH2_MSG_NEWKEYS received debug1: KEX done debug1: userauth-request for user joeuser service ssh-connection method none SSH: Server;Ltype: Authname;Remote: 127.0.0.1-52212;Name: joeuser debug1: attempt 0 failures 0 debug1: Config token is loglevel debug1: Config token is logingracetime debug1: Config token is permitrootlogin debug1: Config token is rsaauthentication debug1: Config token is pubkeyauthentication debug1: Config token is authorizedkeysfile debug1: Config token is challengeresponseauthentication debug1: Config token is usepam debug1: Config token is allowtcpforwarding debug1: Config token is chrootdirectory debug1: Config token is subsystem debug1: PAM: initializing for "joeuser" debug1: PAM: setting PAM_RHOST to "localhost" debug1: PAM: setting PAM_TTY to "ssh" debug1: userauth-request for user joeuser service ssh-connection method password debug1: attempt 1 failures 0 debug1: do_pam_account: called Accepted password for joeuser from 127.0.0.1 port 52212 ssh2 debug1: monitor_child_preauth: joeuser has been authenticated by privileged process debug1: PAM: establishing credentials User child is on pid 9129 debug1: Entering interactive session for SSH2. debug1: server_init_dispatch_20 debug1: server_input_channel_open: ctype session rchan 0 win 65536 max 16384 debug1: input_session_request debug1: channel 0: new [server-session] debug1: session_new: session 0 debug1: session_open: channel 0 debug1: session_open: session 0: link with channel 0 debug1: server_input_channel_open: confirm session debug1: server_input_global_request: rtype [email protected] want_reply 0 debug1: server_input_channel_req: channel 0 request pty-req reply 1 debug1: session_by_channel: session 0 channel 0 debug1: session_input_channel_req: session 0 req pty-req debug1: Allocating pty. debug1: session_new: session 0 debug1: session_pty_req: session 0 alloc /dev/pts/1 debug1: server_input_channel_req: channel 0 request shell reply 1 debug1: session_by_channel: session 0 channel 0 debug1: session_input_channel_req: session 0 req shell debug1: Setting controlling tty using TIOCSCTTY. debug1: Received SIGCHLD. debug1: session_by_pid: pid 9130 debug1: session_exit_message: session 0 channel 0 pid 9130 debug1: session_exit_message: release channel 0 debug1: session_by_tty: session 0 tty /dev/pts/1 debug1: session_pty_cleanup: session 0 release /dev/pts/1 Received disconnect from 127.0.0.1: 11: disconnected by user debug1: do_cleanup debug1: do_cleanup debug1: PAM: cleanup debug1: PAM: closing session debug1: PAM: deleting credentials Here you have the full output of sshd -dd, together with ssh -vv. Bash: # bash --version GNU bash, version 3.2.49(1)-release (arm-none-linux-gnueabi) Copyright (C) 2007 Free Software Foundation, Inc. The bash binary was cross compiled from source. I also tried using a pre-compiled binary from the Optware distribution, but had the exact same problem. I checked for missing shared libraries using objdump -x, but they're all there. Any ideas what could be causing this "Permission denied, please try again."? I'm almost diving in the bash source code to investigate, but trying to avoid hours chasing something that may be silly.

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  • Entity Framework and multi-tenancy database design

    - by Junto
    I am looking at multi-tenancy database schema design for an SaaS concept. It will be ASP.NET MVC - EF, but that isn't so important. Below you can see an example database schema (the Tenant being the Company). The CompanyId is replicated throughout the schema and the primary key has been placed on both the natural key, plus the tenant Id. Plugging this schema into the Entity Framework gives the following errors when I add the tables into the Entity Model file (Model1.edmx): The relationship 'FK_Order_Customer' uses the set of foreign keys '{CustomerId, CompanyId}' that are partially contained in the set of primary keys '{OrderId, CompanyId}' of the table 'Order'. The set of foreign keys must be fully contained in the set of primary keys, or fully not contained in the set of primary keys to be mapped to a model. The relationship 'FK_OrderLine_Customer' uses the set of foreign keys '{CustomerId, CompanyId}' that are partially contained in the set of primary keys '{OrderLineId, CompanyId}' of the table 'OrderLine'. The set of foreign keys must be fully contained in the set of primary keys, or fully not contained in the set of primary keys to be mapped to a model. The relationship 'FK_OrderLine_Order' uses the set of foreign keys '{OrderId, CompanyId}' that are partially contained in the set of primary keys '{OrderLineId, CompanyId}' of the table 'OrderLine'. The set of foreign keys must be fully contained in the set of primary keys, or fully not contained in the set of primary keys to be mapped to a model. The relationship 'FK_Order_Customer' uses the set of foreign keys '{CustomerId, CompanyId}' that are partially contained in the set of primary keys '{OrderId, CompanyId}' of the table 'Order'. The set of foreign keys must be fully contained in the set of primary keys, or fully not contained in the set of primary keys to be mapped to a model. The relationship 'FK_OrderLine_Customer' uses the set of foreign keys '{CustomerId, CompanyId}' that are partially contained in the set of primary keys '{OrderLineId, CompanyId}' of the table 'OrderLine'. The set of foreign keys must be fully contained in the set of primary keys, or fully not contained in the set of primary keys to be mapped to a model. The relationship 'FK_OrderLine_Order' uses the set of foreign keys '{OrderId, CompanyId}' that are partially contained in the set of primary keys '{OrderLineId, CompanyId}' of the table 'OrderLine'. The set of foreign keys must be fully contained in the set of primary keys, or fully not contained in the set of primary keys to be mapped to a model. The relationship 'FK_OrderLine_Product' uses the set of foreign keys '{ProductId, CompanyId}' that are partially contained in the set of primary keys '{OrderLineId, CompanyId}' of the table 'OrderLine'. The set of foreign keys must be fully contained in the set of primary keys, or fully not contained in the set of primary keys to be mapped to a model. The question is in two parts: Is my database design incorrect? Should I refrain from these compound primary keys? I'm questioning my sanity regarding the fundamental schema design (frazzled brain syndrome). Please feel free to suggest the 'idealized' schema. Alternatively, if the database design is correct, then is EF unable to match the keys because it perceives these foreign keys as a potential mis-configured 1:1 relationships (incorrectly)? In which case, is this an EF bug and how can I work around it?

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  • How to fix notifyDataSetChanged/ListView problems in dynamic Adapter wrapper Android

    - by ipaterson
    Summary: Trying to dynamically add heading rows to a ListView via a custom adapter wrapper. ListView is having trouble keeping the scroll position in sync. Runnable demo project provided. I would like to dynamically add items to a list based on the values in a CursorAdapter, several positions ahead of what the user is currently viewing. To do this, I have an adapter that wraps the CursorAdapter and keeps the new content indexed in a SparseArray. The ListView needs to be updated when items are added to the custom adapter, but I have met a lot of pitfalls trying to get that to work and would love some advice. The demo project can be downloaded here: http://dl.dropbox.com/u/15334423/DynamicSectionedList.zip In the demo, the headings are added dynamically by looking ahead 10 places to find the correct position where the list items switch to the next letter. Each implementation of notifyDataSetChanged has problems as described: Demo 1 This demo shows the importance of notifyDataSetChanged(). On clicking anything, the app will crash. This is due to some sanity checking in ListView... mItemCount != adapter.getItemCount(). Moral is, we need to notify the list that data has changed. Demo 2 The natural next step is to notify the ListView of changes when changes occur. Unfortunately, doing so while the ListView is scrolling firmly breaks all touch interaction until the app switches out of touch mode. You will need to "fling scroll" far enough to generate new headings in order to notice this. Tapping the screen will not cause the scroll to stop, and once stopped none of the list items will be clickable. This is due to some if (!mDataChanged) { /* do very important stuff */ } code in AbsListView.onTouchEvent(). Demo 3 To fix this, Demo 3 introduces a pendingChanges flag and the custom Adapter gains a notifyDataSetChangedIfNeeded() which can be called by the ListView once it has entered a "safe" state for changes. The first point where changes must be notified is in ListView.layoutChildren(), so I overrode that method to first notify of changes if needed, then call through. Fling past at least one heading then click a list item. This doesn't quite work right, though I'm not totally sure why. Tapping or selecting an item with the keyboard/trackball causes the list to refresh without properly syncing the old position. It scrolls to the top of the list which is not acceptable. Demo 4 The scroll problem in Demo 3 can be conquered, at least in touch mode. By adding a call to notifyDataSetChangedIfNeeded() on touch down, the data change happens to take place at such a time that all touch interaction works as expected and the list position is properly synced. However, I can't find an analog for that when the device is not in touch mode, not to mention the fact that it definitely seems like a hack. The list almost always scrolls back to the top, I can't find out what causes it to occasionally maintain the correct position. Since Android is fighting me at each step of the way, I feel like there should be a better approach. Please try the demo, if any fixes can be applied to get it working that would be great! Many thanks to anyone who can look into this, hopefully if we can get the code working it will be useful for others trying to accomplish the same optimization for lists with headings.

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  • C# struct and NuSOAP(php)

    - by opx
    Hello Im trying to build a client in c# that talks with some remote (php)server with SOAP using the NuSOAP library. Here im using a struct/object that will containt the user info of some user: public struct UserProfile { public string username; public string password; public string email; public string site; public string signature; public int age; public int points; And this is the PHP Code: server->wsdl->addComplexType( 'UserProfile', 'complexType', 'struct', 'all', '', array( 'username' => array('name' => 'username', 'type' => 'xsd:string'), 'password' => array('name' => 'password', 'type' => 'xsd:string'), 'email' => array('name' => 'email', 'type' => 'xsd:string'), 'site' => array('name' => 'site', 'type' => 'xsd:string'), 'signature' => array('name' => 'signature', 'type' => 'xsd:string'), 'age' => array('name' => 'age', 'type' => 'xsd:int'), 'points' => array('name' => 'username', 'type' => 'xsd:int'), ) ); $server->wsdl->addComplexType( 'UserProfileArray', 'complexType', 'array', '', 'SOAP-ENC:Array', array(), array(array('ref' => 'SOAP-ENC:arrayType', 'wsdl:arrayType' => 'tns:UserProfile[]')), 'tns:UserProfile' ); $server->register("getUserProfile", array(), array('return' => 'tns:UserProfileArray'), $namespace, false, 'rpc', false, 'Get the user profile object' ); function getUserProfile(){ $profile['username'] = "user"; $profile['password'] = "pass"; $profile['email'] = "usern@ame"; $profile['site'] = "u.com"; $profile['signature'] = "usucsdckme"; $profile['age'] = 111; $profile['points'] = time() / 2444; return $profile; } Now I already have a working login function, and I want to get the info about the logged in user but I dont know howto obtain these. This is what im using to get the userinfo: string user = txtUser.Text; string pass = txtPass.Text; SimpleService.SimpleService service = new SimpleService.SimpleService(); if(service.login(user, pass)){ //logged in } SoapApp.SimpleService.UserProfile[] user = service.getUserProfile(); // THIS LINE GIVES ME AN EXCEPTION MessageBox.Show(user[0].username + "--" + user[0].points); The getUserProfile() function produces an error: System.Web.Services.Protocols.SoapException was unhandled Message="unable to serialize result" Source="System.Web.Services" or I get something like 'cant parse xml' error. The article I used for this was from: http://www.sanity-free.org/125/php_webservices_and_csharp_dotnet_soap_clients.html The difference on what they are doing and what I try to do is that I only want to get one object returned instead of multiple 'MySoapObjects'. I hope someone is familiar with this and could help me, thanks in advance! Regards, opx

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  • SVN: Release branch headaches, how to merge in website revisions as and when cleared to go live?

    - by Pete Duncanson
    I need a sanity check here if we can, any ideas on correcting/changing the following are very welcome! We've been getting ourselves in knots of late with our SVN and are trying to correct it by putting a Trunk/Release system in place. We have a large website that we develop on and we store it all in SVN. Heres what we had in mind: We have trunk and a release branch All work gets checked into Trunk. When a feature is deemed ready for the next release it is merged into a Release branch. We only have one release branch and just tag "Latest" when we do a push to live We hope to be able to get all the files changed from Latest to Head to give us a zip that we can upload (any ideas on an easy way to do this via scripting?) So we set all this up and where very happy with ourselves. Except its not working and heres why. We work on lots a different features/fixes/problems at once and they don't all get nicely checked in feature complete (but always working at least). Then sometimes you have to wait for Clients to sign off. As a result you end up with revisions which are "ready for live" being scattered with ones which are "still being worked on" in trunk. That means that the completed revisions are not getting merged in sequentially but out of order. I thought SVN could handle this, clever little thing it is, but apparently not. Heres an example: Pete changes some CSS to make a new button look pretty (Revision 1) Dave add some CSS to the bottom of the same CSS file as Pete's for a new feature (Revision 2) Dave's mod gets the nod so he merges it into Release and commits it with a log message mentioning revision number and bug tracking id. Pete adds more buttons to finish this mod, no CSS changes here though (Revision 3) Pete then merges his mods (Revision 1 and 3) into the Head of Release (which has Daves merge in it) but this over-writes Daves CSS additions which now dissapear completely. This leads to the site being broken and the Release branch being pretty much useless. So we tried some other ideas like reverting Release back to "Latest" and then just merging in all the Revisions 1,2 and 3 in order. This worked fine until we had Revision 4 which was not ready for live and Revision 5 which was. Suddenly we are getting ourselves in knots again with exactly the same problem! Ok so take three. Revert to Latest, merge in Revision 5, then do any update back to Head. Tree conflicts galore! So thats a no no. I cracked in the end and built it all up manaually but its not something I want to do regular, ideally I want to script our deployment but can't while Release is in such a mess. HELP! What the heck are we doing wrong? I can't seem to find any solutions to this problem of wanting different none sequential Revisions in Release. If its not possible thats fine but how the heck are we meant to get stuff live easily. We can't branch for every single change, the site takes 30 minutes+ to check out it would take too long. Side note, we are using TortoiseSVN so can we keep command line examples to a minimum in any answers? Latest version of TSVN and SVN Version 1.6 so we have the funky merge tracking etc. EDIT: An excellent blog post which deals with the dev/release cycle (although using GIT but still relivant) thought everyone would like to read it if they found this question interesting. (http://nvie.com/git-model) EDIT 2: I wrote a blog post on how to show which branch you are working on in your website which others have asked me about (http://www.offroadcode.com/2010/5/14/which-svn-branch-are-you-working-on.aspx). Hope that helps. In the meantime we are looking at Kiln and hoping to make the switch next month (gulp!)

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  • Weird XPath behavior in libxml2

    - by Josh K
    I have this XML tree that looks like this (I've changed the tag names but if you're really clever you may figure out what I'm actually doing.) <ListOfThings> <Thing foo:action="add"> <Bar>doStuff --slowly</Bar> <Index>1</Index> </Thing> <Thing foo:action="add"> <Bar>ping yourMother.net</Bar> <Index>2</Index> </Thing> </ListOfThings> With libxml2, I want to programmatically insert a new Thing tag into the ListOfThings with the Index being the highest current index, plus one. I do it like this (sanity checking removed for brevity): xpath = "//urn:myformat[@foo='bar']/" "urn:mysection[@name='baz']/" "urn:ListOfThings/urn:Thing/urn:Index"; xpathObj = xmlXPathEvalExpression(xpath, xpathCtx); nodes = xpathObj->nodesetval; /* Find last value and snarf the value of the tag */ highest = atoi(nodes->nodeTab[nodes->nodeNr - 1]->children->content); snprintf(order, sizeof(order), "%d", highest + 1); /* highest index plus one */ /* now move up two levels.. */ cmdRoot = nodes->nodeTab[nodes->nodeNr - 1]; ASSERT(cmdRoot->parent && cmdRoot->parent->parent); cmdRoot = cmdRoot->parent->parent; /* build the child tag */ newTag = xmlNewNode(NULL, "Thing"); xmlSetProp(newTag, "foo:action", "add"); /* set new node values */ xmlNewTextChild(newTag, NULL, "Bar", command); xmlNewChild(newTag, NULL, "Index", order); /* append this to cmdRoot */ xmlAddChild(cmdRoot, newTag); But if I call this function twice (to add two Things), the XPath expression doesn't catch the new entry I made. Is there a function I need to call to kick XPath in the shins and get it to make sure it really looks over the whole xmlDocPtr again? It clearly does get added to the document, because when I save it, I get the new tags I added. To be clear, the output looks like this: <ListOfThings> <Thing foo:action="add"> <Bar>doStuff --slowly</Bar> <Index>1</Index> </Thing> <Thing foo:action="add"> <Bar>ping yourMother.net</Bar> <Index>2</Index> </Thing> <Thing foo:action="add"> <Bar>newCommand1</Bar> <Index>3</Index> </Thing> <Thing foo:action="add"> <Bar>newCommand2</Bar> <Index>3</Index> <!-- this is WRONG! --> </Thing> </ListOfThings> I used a debugger to check what happened after xmlXPathEvalExpression got called and I saw that nodes->nodeNr was the same each time. Help me, lazyweb, you're my only hope!

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  • Coldfusion "Routines cannot be declared more than once"

    - by Nicholas
    We have the following code in our Application.cfc: <cffunction name="onError" returnType="void" output="false"> <cfargument name="exception" required="true"> <cfargument name="eventname" type="string" required="true"> <cfset cfcatch = exception> <cfinclude template="standalone/errors/error.cfm"> </cffunction> Within the error.cfm page we have this code (I didn't write it): <cfscript> function GetCurrentURL() { var theURL = "http"; if (cgi.https EQ "on" ) theURL = "#TheURL#s"; theURL = theURL & "://#cgi.server_name#"; if(cgi.server_port neq 80) theURL = theURL & ":#cgi.server_port#"; theURL = theURL & "#cgi.path_info#"; if(len(cgi.query_string)) theURL = theURL & "?#cgi.query_string#"; return theURL; } </cfscript> This is all part of a script that puts together bunches of details about the error and records it to the database. When an error occurs, we receive the message "The routine GetCurrentURL has been declared twice in different templates." However, I have searched the entire codebase in several different ways and found "GetCurrentURL" used only twice, both times in error.cfm. The first time is the declaration, and the second is actual use. So I'm not sure why CF is saying "in different templates". My next thought was that the problem is a recursive call, and that error.cfm is erroring and calling itself, so I attempted these two changes, either of which should have resolved the issue and unmasked the real error: <cfif StructKeyExists(variables,"GetCurrentURL") IS "NO"> <cfscript> function GetCurrentURL() { var theURL = "http"; if (cgi.https EQ "on" ) theURL = "#TheURL#s"; theURL = theURL & "://#cgi.server_name#"; if(cgi.server_port neq 80) theURL = theURL & ":#cgi.server_port#"; theURL = theURL & "#cgi.path_info#"; if(len(cgi.query_string)) theURL = theURL & "?#cgi.query_string#"; return theURL; } </cfscript> </cfif> And: <cfscript> if (!StructKeyExists(variables,"GetCurrentURL")) { function GetCurrentURL() { var theURL = "http"; if (cgi.https EQ "on" ) theURL = "#TheURL#s"; theURL = theURL & "://#cgi.server_name#"; if(cgi.server_port neq 80) theURL = theURL & ":#cgi.server_port#"; theURL = theURL & "#cgi.path_info#"; if(len(cgi.query_string)) theURL = theURL & "?#cgi.query_string#"; return theURL; } } </cfscript> Neither worked. I also tried adding this to the page just before the function call: <cfoutput>"#StructKeyExists(variables,"GetCurrentURL")#"</cfoutput> It caused the word "YES" to be printed on screen. This indicates that the above should work, as clearly the contents of the if statement will evaluate to "YES", and thus the if statement will evaluate to false, and thus the function will not be declared, and thus I will retain my sanity. But for some reason this problem persists. Any thoughts on what might be occuring or how to troubleshoot next? I'm stuck at this point.

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