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• Is it possible to reliably auto-decode user files to Unicode? [C#]

  - by NVRAM

  I have a web application that allows users to upload their content for processing. The processing engine expects UTF8 (and I'm composing XML from multiple users' files), so I need to ensure that I can properly decode the uploaded files. Since I'd be surprised if any of my users knew their files even were encoded, I have very little hope they'd be able to correctly specify the encoding (decoder) to use. And so, my application is left with task of detecting before decoding. This seems like such a universal problem, I'm surprised not to find either a framework capability or general recipe for the solution. Can it be I'm not searching with meaningful search terms? I've implemented BOM-aware detection (http://en.wikipedia.org/wiki/Byte_order_mark) but I'm not sure how often files will be uploaded w/o a BOM to indicate encoding, and this isn't useful for most non-UTF files. My questions boil down to: Is BOM-aware detection sufficient for the vast majority of files? In the case where BOM-detection fails, is it possible to try different decoders and determine if they are "valid"? (My attempts indicate the answer is "no.") Under what circumstances will a "valid" file fail with the C# encoder/decoder framework? Is there a repository anywhere that has a multitude of files with various encodings to use for testing? While I'm specifically asking about C#/.NET, I'd like to know the answer for Java, Python and other languages for the next time I have to do this. So far I've found: A "valid" UTF-16 file with Ctrl-S characters has caused encoding to UTF-8 to throw an exception (Illegal character?) (That was an XML encoding exception.) Decoding a valid UTF-16 file with UTF-8 succeeds but gives text with null characters. Huh? Currently, I only expect UTF-8, UTF-16 and probably ISO-8859-1 files, but I want the solution to be extensible if possible. My existing set of input files isn't nearly broad enough to uncover all the problems that will occur with live files. Although the files I'm trying to decode are "text" I think they are often created w/methods that leave garbage characters in the files. Hence "valid" files may not be "pure". Oh joy. Thanks.

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• What container type provides better (average) performance than std::map?

  - by Truncheon

  In the following example a std::map structure is filled with 26 values from A - Z (for key) and 0 - 26 for value. The time taken (on my system) to lookup the last entry (10000000 times) is roughly 250 ms for the vector, and 125 ms for the map. (I compiled using release mode, with O3 option turned on for g++ 4.4) But if for some odd reason I wanted better performance than the std::map, what data structures and functions would I need to consider using? I apologize if the answer seems obvious to you, but I haven't had much experience in the performance critical aspects of C++ programming. UPDATE: This example is rather trivial and hides the true complexity of what I'm trying to achieve. My real world project is a simple scripting language that uses a parser, data tree, and interpreter (instead of a VM stack system). I need to use some kind of data structure (perhaps map) to store the variables names created by script programmers. These are likely to be pretty randomly named, so I need a lookup method that can quickly find a particular key within a (probably) fairly large list of names. #include <ctime> #include <map> #include <vector> #include <iostream> struct mystruct { char key; int value; mystruct(char k = 0, int v = 0) : key(k), value(v) { } }; int find(const std::vector<mystruct>& ref, char key) { for (std::vector<mystruct>::const_iterator i = ref.begin(); i != ref.end(); ++i) if (i->key == key) return i->value; return -1; } int main() { std::map<char, int> mymap; std::vector<mystruct> myvec; for (int i = 'a'; i < 'a' + 26; ++i) { mymap[i] = i - 'a'; myvec.push_back(mystruct(i, i - 'a')); } int pre = clock(); for (int i = 0; i < 10000000; ++i) { find(myvec, 'z'); } std::cout << "linear scan: milli " << clock() - pre << "\n"; pre = clock(); for (int i = 0; i < 10000000; ++i) { mymap['z']; } std::cout << "map scan: milli " << clock() - pre << "\n"; return 0; }

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• How to combine designable components with dependency injection

  - by Wim Coenen

  When creating a designable .NET component, you are required to provide a default constructor. From the IComponent documentation: To be a component, a class must implement the IComponent interface and provide a basic constructor that requires no parameters or a single parameter of type IContainer. This makes it impossible to do dependency injection via constructor arguments. (Extra constructors could be provided, but the designer would ignore them.) Some alternatives we're considering: Service Locator Don't use dependency injection, instead use the service locator pattern to acquire dependencies. This seems to be what IComponent.Site.GetService is for. I guess we could create a reusable ISite implementation (ConfigurableServiceLocator?) which can be configured with the necessary dependencies. But how does this work in a designer context? Dependency Injection via properties Inject dependencies via properties. Provide default instances if they are necessary to show the component in a designer. Document which properties need to be injected. Inject dependencies with an Initialize method This is much like injection via properties but it keeps the list of dependencies that need to be injected in one place. This way the list of required dependencies is documented implicitly, and the compiler will assists you with errors when the list changes. Any idea what the best practice is here? How do you do it? edit: I have removed "(e.g. a WinForms UserControl)" since I intended the question to be about components in general. Components are all about inversion of control (see section 8.3.1 of the UMLv2 specification) so I don't think that "you shouldn't inject any services" is a good answer. edit 2: It took some playing with WPF and the MVVM pattern to finally "get" Mark's answer. I see now that visual controls are indeed a special case. As for using non-visual components on designer surfaces, I think the .NET component model is fundamentally incompatible with dependency injection. It appears to be designed around the service locator pattern instead. Maybe this will start to change with the infrastructure that was added in .NET 4.0 in the System.ComponentModel.Composition namespace.

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• symfony/zend integration - blank screen

  - by user142176

  Hi, I need to use ZendAMF on a symfony project and I'm currently working on integrating the two. I have a frontend app with two modules, one of which is 'gateway' - the AMF gateway. In my frontend app config, I have the following in the configure function: // load symfony autoloading first parent::initialize(); // Integrate Zend Framework require_once('[MY PATH TO ZEND]\Loader.php'); spl_autoload_register(array('Zend_Loader', 'autoload')); The executeIndex function my the gateway actions.class.php looks like this // No Layout $this->setLayout(false); // Set MIME Type $this->getResponse()->setContentType('application/x-amf; charset='.sfConfig::get('sf_charset')); // Disable cause this is a non-html page sfConfig::set('sf_web_debug', false); // Create AMF Server $server = new Zend_Amf_Server(); $server->setClass('MYCLASS'); echo $server->handle(); return sfView::NONE; Now when I try to visit the url for the gateway module, or even the other module which was working perfectly fine until this attempt, I only see a blank screen, with not even the symfony dev bar loaded. Oddly enough, my symfony logs are not being updated as well, which suggests that Synfony is not even being 'reached'. So presumably the error has something to do with Zend, but I have no idea how to figure out what the error could be. One thing I do know for sure is that this is not a file path error, because if I change the path in the following line (a part of frontendConfiguration as shown above), I get a Zend_Amf_Server not found error. So the path must be correct. Also if I comment out this very same line, the second module resumes to normality, and my gateway broadcasts a blank x-amf stream. spl_autoload_register(array('Zend_Loader', 'autoload')); Does anyone have any tips on how I could attach this problem? Thanks P.S. I'm currently running an older version of Zend, which is why I am using Zend_Loader instead of Zend_autoLoader (I think). But I've tried switching to the new lib, but the error still remains. So it's not a version problem as well.

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• Excel UDF calculation should return 'original' value

  - by LeChe

  Hi all, I've been struggling with a VBA problem for a while now and I'll try to explain it as thoroughly as possible. I have created a VSTO plugin with my own RTD implementation that I am calling from my Excel sheets. To avoid having to use the full-fledged RTD syntax in the cells, I have created a UDF that hides that API from the sheet. The RTD server I created can be enabled and disabled through a button in a custom Ribbon component. The behavior I want to achieve is as follows: If the server is disabled and a reference to my function is entered in a cell, I want the cell to display Disabled If the server is disabled, but the function had been entered in a cell when it was enabled (and the cell thus displays a value), I want the cell to keep displaying that value If the server is enabled, I want the cell to display Loading Sounds easy enough. Here is an example of the - non functional - code: Public Function RetrieveData(id as Long) Dim result as String // This returns either 'Disabled' or 'Loading' result = Application.Worksheet.Function.RTD("SERVERNAME", "", id) RetrieveData = result If(result = "Disabled") Then // Obviously, this recurses (and fails), so that's not an option If(Not IsEmpty(Application.Caller.Value2)) Then // So does this RetrieveData = Application.Caller.Value2 End If End If End Function The function will be called in thousands of cells, so storing the 'original' values in another data structure would be a major overhead and I would like to avoid it. Also, the RTD server does not know the values, since it also does not keep a history of it, more or less for the same reason. I was thinking that there might be some way to exit the function which would force it to not change the displayed value, but so far I have been unable to find anything like that. Any ideas on how to solve this are greatly appreciated! Thanks, Che EDIT: By popular demand, some additional info on why I want to do all this: As I said, the function will be called in thousands of cells and the RTD server needs to retrieve quite a bit of information. This can be quite hard on both network and CPU. To allow the user to decide for himself whether he wants this load on his machine, he or she can disable the updates from the server. In that case, he or she should still be able to calculate the sheets with the values currently in the fields, yet no updates are pushed into them. Once new data is required, the server can be enabled and the fields will be updated. Again, since we are talking about quite a bit of data here, I would rather not store it somewhere in the sheet. Plus, the data should be usable even if the workbook is closed and loaded again.

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• Linux Kernel crash mutex_lock_slowpath "blocked for more than 120 seconds". What to do?

  - by Roddick

  I have out-of-the box Debian Lenny with non-custom kernel 2.6.26-2-amd64. Brand new server that is used to 5% of it's potential, CPU and Disk-wise. Meaning it probably not crashing because of overload. every few days it freezes with hundreds of these messages in console log: : [284847.828428] INFO: task apache2:12473 blocked for more than 120 seconds. : [284847.868468] "echo 0 > /proc/sys/kernel/hung_task_timeout_secs" disables this message. : [284847.912759] apache2 D ffff8101bc6b7ab0 0 12473 14358 : [284847.912763] ffff810160d5bc50 0000000000000082 ffff8101c0002e40 0000000000000000 : [284847.912766] ffff8101a7c42950 ffff810327d92810 ffff8101a7c42bd8 0000000400000044 : [284847.912770] ffff8101c0002e40 00000000000612d0 0000000000000000 00000040000612d0 : [284847.912773] Call Trace: : [284847.912786] [<ffffffff80429b0d>] __mutex_lock_slowpath+0x64/0x9b : [284847.912790] [<ffffffff80429972>] mutex_lock+0xa/0xb : [284847.912794] [<ffffffff802a20b9>] do_lookup+0x82/0x1c1 : [284847.912800] [<ffffffff802a4271>] __link_path_walk+0x87a/0xd19 : [284847.912805] [<ffffffff80295844>] kmem_getpages+0x96/0x15f : [284847.912808] [<ffffffff80295fb7>] ____cache_alloc_node+0x6d/0x106 : [284847.912814] [<ffffffff802a4756>] path_walk+0x46/0x8b : [284847.912819] [<ffffffff802a4a82>] do_path_lookup+0x158/0x1cf : [284847.912822] [<ffffffff802a3879>] getname+0x140/0x1a7 : [284847.912827] [<ffffffff802a53f1>] __user_walk_fd+0x37/0x4c : [284847.912831] [<ffffffff8029e381>] vfs_lstat_fd+0x18/0x47 : [284847.912840] [<ffffffff8029e3c9>] sys_newlstat+0x19/0x31 : [284847.912848] [<ffffffff8020beda>] system_call_after_swapgs+0x8a/0x8f Almost all traces has __mutex_lock_slowpath as top-level. Only some has different trace: : [284847.737386] INFO: task apache2:12472 blocked for more than 120 seconds. : [284847.777551] "echo 0 > /proc/sys/kernel/hung_task_timeout_secs" disables this message. : [284847.824881] apache2 D ffff8101bc6b7ab0 0 12472 14358 : [284847.824886] ffff8101b9cc1c50 0000000000000086 ffffffffa0131e0a 0000000000000002 : [284847.824889] ffff8102e7454300 ffff810324c6cad0 ffff8102e7454588 0000000000000000 : [284847.824893] 0000000000000001 0000000000000296 0000000000000003 ffff8101b9cc1c58 : [284847.824896] Call Trace: : [284847.828403] [<ffffffffa0131e0a>] :ext3:__ext3_journal_dirty_metadata+0x1e/0x46 : [284847.828412] [<ffffffff80429b0d>] __mutex_lock_slowpath+0x64/0x9b : [284847.828418] [<ffffffff80429972>] mutex_lock+0xa/0xb : [284847.828421] [<ffffffff802a20b9>] do_lookup+0x82/0x1c1 : [284847.828427] [<ffffffff802a4271>] __link_path_walk+0x87a/0xd19 : [284847.828428] [<ffffffff80271296>] find_lock_page+0x1f/0x8a : [284847.828428] [<ffffffff80273182>] filemap_fault+0x1c2/0x33c : [284847.828428] [<ffffffff802a4756>] path_walk+0x46/0x8b : [284847.828428] [<ffffffff802a4a82>] do_path_lookup+0x158/0x1cf : [284847.828428] [<ffffffff802a3879>] getname+0x140/0x1a7 : [284847.828428] [<ffffffff802a53f1>] __user_walk_fd+0x37/0x4c : [284847.828428] [<ffffffff8029e381>] vfs_lstat_fd+0x18/0x47 : [284847.828428] [<ffffffff8029e3c9>] sys_newlstat+0x19/0x31 : [284847.828428] [<ffffffff8020beda>] system_call_after_swapgs+0x8a/0x8f kernel: [1912668.466347] INFO: task apache2:17984 blocked for more than 120 seconds. [1912668.507035] "echo 0 > /proc/sys/kernel/hung_task_timeout_secs" disables this message. : [1912668.555165] apache2 D ffff8101c5637ba0 0 17984 17282 : [1912668.596752] ffff810166a7dd30 0000000000000086 0000000000000000 ffff810166a7dcd8 : [1912668.643341] ffff8101c563c880 ffff81024505f000 0000000000000002 ffff810166a7dd68 : [1912668.699566] 0000000000000086 00000000000cb1a0 0000000000000000 ffff81017f344d60 : [1912668.744773] Call Trace: : [1912668.761754] [<ffffffff8022a3ed>] pick_next_task_fair+0x6e/0x7a : [1912668.829311] [<ffffffff802be0e2>] bio_alloc_bioset+0x89/0xd9 : [1912668.861930] [<ffffffff8024ac3a>] getnstimeofday+0x39/0x98 : [1912668.897005] [<ffffffff802710f6>] sync_page+0x0/0x41 : [1912668.927868] [<ffffffff80429487>] io_schedule+0x5c/0x9e : [1912668.960286] [<ffffffff80271132>] sync_page+0x3c/0x41 : [1912668.991756] [<ffffffff804295fa>] __wait_on_bit_lock+0x36/0x66 : [1912669.031757] [<ffffffff802710e3>] __lock_page+0x5e/0x64 : [1912669.064191] [<ffffffff802461d3>] wake_bit_function+0x0/0x23 : [1912669.100100] [<ffffffff80281bc5>] handle_mm_fault+0x5e4/0x8de : [1912669.134531] [<ffffffff802461a5>] autoremove_wake_function+0x0/0x2e : [1912669.174623] [<ffffffff802aa108>] fcntl_setlk+0x1cf/0x291 : [1912669.210623] [<ffffffff802461a5>] autoremove_wake_function+0x0/0x2e : [1912669.246923] [<ffffffff802a677f>] sys_fcntl+0x280/0x2f7 After googling for "mutex_lock_slowpath" I can only find the Kernel mailing list discussions that this issue was introduced in some commit. Wthout reference to verison. Discussions as recent as Jan 25, 2011. The Kernel I am using is form Debian Lenny, year ago. What should I do? Is this bug even fixed in kernel? if it's such obvious bug why it happens so rarely? Should I download latest kernel from kernel.org and upgrade? Should I use Debian backports to install new "Approved" kernel? Am I missing something? What to do?

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• How do access a secure website within a sharepoint webpart?

  - by Bill

  How do access a secure website within a sharepoint webpart? The following code works fine as a console application but if you run it in a webpart, you will get a access violation WebRequest request = WebRequest.Create("https://somesecuresite.com"); WebResponse firstResponse = null; try { firstResponse = request.GetResponse(); } catch (WebException ex) { writer.WriteLine("Error: " + ex.ToString()); return; } if you access a non secure site, it also works. Any ideas? Error: System.Net.WebException: The underlying connection was closed: An unexpected error occurred on a receive. --- System.AccessViolationException: Attempted to read or write protected memory. This is often an indication that other memory is corrupt. at System.Net.UnsafeNclNativeMethods.NativePKI.CertVerifyCertificateChainPolicy(IntPtr policy, SafeFreeCertChain chainContext, ChainPolicyParameter& cpp, ChainPolicyStatus& ps) at System.Net.PolicyWrapper.VerifyChainPolicy(SafeFreeCertChain chainContext, ChainPolicyParameter& cpp) at System.Net.Security.SecureChannel.VerifyRemoteCertificate(RemoteCertValidationCallback remoteCertValidationCallback) at System.Net.Security.SslState.CompleteHandshake() at System.Net.Security.SslState.CheckCompletionBeforeNextReceive(ProtocolToken message, AsyncProtocolRequest asyncRequest) at System.Net.Security.SslState.StartSendBlob(Byte[] incoming, Int32 count, AsyncProtocolRequest asyncRequest) at System.Net.Security.SslState.ProcessReceivedBlob(Byte[] buffer, Int32 count, AsyncProtocolRequest asyncRequest) at System.Net.Security.SslState.StartReadFrame(Byte[] buffer, Int32 readBytes, AsyncProtocolRequest asyncRequest) at System.Net.Security.SslState.StartReceiveBlob(Byte[] buffer, AsyncProtocolRequest asyncRequest) at System.Net.Security.SslState.CheckCompletionBeforeNextReceive(ProtocolToken message, AsyncProtocolRequest asyncRequest) at System.Net.Security.SslState.StartSendBlob(Byte[] incoming, Int32 count, AsyncProtocolRequest asyncRequest) at System.Net.Security.SslState.ProcessReceivedBlob(Byte[] buffer, Int32 count, AsyncProtocolRequest asyncRequest) at System.Net.Security.SslState.StartReadFrame(Byte[] buffer, Int32 readBytes, AsyncProtocolRequest asyncRequest) at System.Net.Security.SslState.StartReceiveBlob(Byte[] buffer, AsyncProtocolRequest asyncRequest) at System.Net.Security.SslState.CheckCompletionBeforeNextReceive(ProtocolToken message, AsyncProtocolRequest asyncRequest) at System.Net.Security.SslState.StartSendBlob(Byte[] incoming, Int32 count, AsyncProtocolRequest asyncRequest) at System.Net.Security.SslState.ProcessReceivedBlob(Byte[] buffer, Int32 count, AsyncProtocolRequest asyncRequest) at System.Net.Security.SslState.StartReadFrame(Byte[] buffer, Int32 readBytes, AsyncProtocolRequest asyncRequest) at System.Net.Security.SslState.StartReceiveBlob(Byte[] buffer, AsyncProtocolRequest asyncRequest) at System.Net.Security.SslState.CheckCompletionBeforeNextReceive(ProtocolToken message, AsyncProtocolRequest asyncRequest) at System.Net.Security.SslState.StartSendBlob(Byte[] incoming, Int32 count, AsyncProtocolRequest asyncRequest) at System.Net.Security.SslState.ForceAuthentication(Boolean receiveFirst, Byte[] buffer, AsyncProtocolRequest asyncRequest) at System.Net.Security.SslState.ProcessAuthentication(LazyAsyncResult lazyResult) at System.Net.TlsStream.CallProcessAuthentication(Object state) at System.Threading.ExecutionContext.runTryCode(Object userData) at System.Runtime.CompilerServices.RuntimeHelpers.ExecuteCodeWithGuaranteedCleanup(TryCode code, CleanupCode backoutCode, Object userData) at System.Threading.ExecutionContext.RunInternal(ExecutionContext executionContext, ContextCallback callback, Object state) at System.Threading.ExecutionContext.Run(ExecutionContext executionContext, ContextCallback callback, Object state) at System.Net.TlsStream.ProcessAuthentication(LazyAsyncResult result) at System.Net.TlsStream.Write(Byte[] buffer, Int32 offset, Int32 size) at System.Net.PooledStream.Write(Byte[] buffer, Int32 offset, Int32 size) at System.Net.ConnectStream.WriteHeaders(Boolean async) --- End of inner exception stack trace --- at System.Net.HttpWebRequest.GetResponse()

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• Is READ UNCOMMITTED / NOLOCK safe in this situation?

  - by Ben Challenor

  I know that snapshot isolation would fix this problem, but I'm wondering if NOLOCK is safe in this specific case so that I can avoid the overhead. I have a table that looks something like this: drop table Data create table Data ( Id BIGINT NOT NULL, Date BIGINT NOT NULL, Value BIGINT, constraint Cx primary key (Date, Id) ) create nonclustered index Ix on Data (Id, Date) There are no updates to the table, ever. Deletes can occur but they should never contend with the SELECT because they affect the other, older end of the table. Inserts are regular and page splits to the (Id, Date) index are extremely common. I have a deadlock situation between a standard INSERT and a SELECT that looks like this: select top 1 Date, Value from Data where Id = @p0 order by Date desc because the INSERT acquires a lock on Cx (Date, Id; Value) and then Ix (Id, Date), but the SELECT acquires a lock on Ix (Id, Date) and then Cx (Date, Id; Value). This is because the SELECT first seeks on Ix and then joins to a seek on Cx. Swapping the clustered and non-clustered index would break this cycle, but it is not an acceptable solution because it would introduce cycles with other (more complex) SELECTs. If I add NOLOCK to the SELECT, can it go wrong in this case? Can it return: More than one row, even though I asked for TOP 1? No rows, even though one exists and has been committed? Worst of all, a row that doesn't satisfy the WHERE clause? I've done a lot of reading about this online, but the only reproductions of over- or under-count anomalies I've seen (one, two) involve a scan. This involves only seeks. Jeff Atwood has a post about using NOLOCK that generated a good discussion. I was particularly interested in a comment by Rick Townsend: Secondly, if you read dirty data, the risk you run is of reading the entirely wrong row. For example, if your select reads an index to find your row, then the update changes the location of the rows (e.g.: due to a page split or an update to the clustered index), when your select goes to read the actual data row, it's either no longer there, or a different row altogether! Is this possible with inserts only, and no updates? If so, then I guess even my seeks on an insert-only table could be dangerous. Update: I'm trying to figure out how snapshot isolation works. It seems to be row-based, where transactions read the table (with no shared lock!), find the row they are interested in, and then see if they need to get an old version of the row from the version store in tempdb. But in my case, no row will have more than one version, so the version store seems rather pointless. And if the row was found with no shared lock, how is it different to just using NOLOCK?

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• Linq to sql C# updating reference Tables

  - by Laurence Burke

  ok reclarification I am adding a new address and I know the structure as AddressID = PK and all other entities are non nullable. Now on insert of a new row the addrID Pk is autogened and I am wondering if I would have to get that to create a new row in the referencing table or does that automatically get generated also. also I want to be able to repopulate the dropdownlist that lists the current employee's addresses with the newly created address. static uint _curEmpID; protected void btnAdd_Click(object sender, EventArgs e) { if (txtZip.Text != "" && txtAdd1.Text != "" && txtCity.Text != "") { TestDataClassDataContext dc = new TestDataClassDataContext(); Address addr = new Address() { AddressLine1 = txtAdd1.Text, AddressLine2 = txtAdd2.Text, City = txtCity.Text, PostalCode = txtZip.Text, StateProvinceID = Convert.ToInt32(ddlState.SelectedValue) }; dc.Addresses.InsertOnSubmit(addr); lblSuccess.Visible = true; lblErrMsg.Visible = false; dc.SubmitChanges(); // // TODO: add reference from new address to CurEmp Table // SetAddrList(); } else { lblErrMsg.Text = "Invalid Input"; lblErrMsg.Visible = true; } } protected void ddlAddList_SelectedIndexChanged(object sender, EventArgs e) { lblErrMsg.Visible = false; lblSuccess.Visible = false; TestDataClassDataContext dc = new TestDataClassDataContext(); dc.ObjectTrackingEnabled = false; if (ddlAddList.SelectedValue != "-1") { var addr = (from a in dc.Addresses where a.AddressID == Convert.ToInt32(ddlAddList.SelectedValue) select a).FirstOrDefault(); txtAdd1.Text = addr.AddressLine1; txtAdd2.Text = addr.AddressLine2; txtCity.Text = addr.City; txtZip.Text = addr.PostalCode; ddlState.SelectedValue = addr.StateProvinceID.ToString(); btnSubmit.Visible = true; btnAdd.Visible = false; } else { txtAdd1.Text = ""; txtAdd2.Text = ""; txtCity.Text = ""; txtZip.Text = ""; btnAdd.Visible = true; btnSubmit.Visible = false; } } protected void SetAddrList() { TestDataClassDataContext dc = new TestDataClassDataContext(); dc.ObjectTrackingEnabled = false; var addList = from addr in dc.Addresses from eaddr in dc.EmployeeAddresses where eaddr.EmployeeID == _curEmpID && addr.AddressID == eaddr.AddressID select new { AddValue = addr.AddressID, AddText = addr.AddressID, }; ddlAddList.DataSource = addList; ddlAddList.DataValueField = "AddValue"; ddlAddList.DataTextField = "AddText"; ddlAddList.DataBind(); ddlAddList.Items.Add(new ListItem("<Add Address>", "-1")); } OK I am hoping that I did not include too much code. I would really appreciate any other comments about I could otherwise improve this code in any other ways also.

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• When should I use indexed arrays of OpenGL vertices?

  - by Tartley

  I'm trying to get a clear idea of when I should be using indexed arrays of OpenGL vertices, drawn with gl[Multi]DrawElements and the like, versus when I should simply use contiguous arrays of vertices, drawn with gl[Multi]DrawArrays. (Update: The consensus in the replies I got is that one should always be using indexed vertices.) I have gone back and forth on this issue several times, so I'm going to outline my current understanding, in the hopes someone can either tell me I'm now finally more or less correct, or else point out where my remaining misunderstandings are. Specifically, I have three conclusions, in bold. Please correct them if they are wrong. One simple case is if my geometry consists of meshes to form curved surfaces. In this case, the vertices in the middle of the mesh will have identical attributes (position, normal, color, texture coord, etc) for every triangle which uses the vertex. This leads me to conclude that: 1. For geometry with few seams, indexed arrays are a big win. Follow rule 1 always, except: For geometry that is very 'blocky', in which every edge represents a seam, the benefit of indexed arrays is less obvious. To take a simple cube as an example, although each vertex is used in three different faces, we can't share vertices between them, because for a single vertex, the surface normals (and possible other things, like color and texture co-ord) will differ on each face. Hence we need to explicitly introduce redundant vertex positions into our array, so that the same position can be used several times with different normals, etc. This means that indexed arrays are of less use. e.g. When rendering a single face of a cube: 0 1 o---o |\ | | \ | | \| o---o 3 2 (this can be considered in isolation, because the seams between this face and all adjacent faces mean than none of these vertices can be shared between faces) if rendering using GL_TRIANGLE_FAN (or _STRIP), then each face of the cube can be rendered thus: verts = [v0, v1, v2, v3] colors = [c0, c0, c0, c0] normal = [n0, n0, n0, n0] Adding indices does not allow us to simplify this. From this I conclude that: 2. When rendering geometry which is all seams or mostly seams, when using GL_TRIANGLE_STRIP or _FAN, then I should never use indexed arrays, and should instead always use gl[Multi]DrawArrays. (Update: Replies indicate that this conclusion is wrong. Even though indices don't allow us to reduce the size of the arrays here, they should still be used because of other performance benefits, as discussed in the comments) The only exception to rule 2 is: When using GL_TRIANGLES (instead of strips or fans), then half of the vertices can still be re-used twice, with identical normals and colors, etc, because each cube face is rendered as two separate triangles. Again, for the same single cube face: 0 1 o---o |\ | | \ | | \| o---o 3 2 Without indices, using GL_TRIANGLES, the arrays would be something like: verts = [v0, v1, v2, v2, v3, v0] normals = [n0, n0, n0, n0, n0, n0] colors = [c0, c0, c0, c0, c0, c0] Since a vertex and a normal are often 3 floats each, and a color is often 3 bytes, that gives, for each cube face, about: verts = 6 * 3 floats = 18 floats normals = 6 * 3 floats = 18 floats colors = 6 * 3 bytes = 18 bytes = 36 floats and 18 bytes per cube face. (I understand the number of bytes might change if different types are used, the exact figures are just for illustration.) With indices, we can simplify this a little, giving: verts = [v0, v1, v2, v3] (4 * 3 = 12 floats) normals = [n0, n0, n0, n0] (4 * 3 = 12 floats) colors = [c0, c0, c0, c0] (4 * 3 = 12 bytes) indices = [0, 1, 2, 2, 3, 0] (6 shorts) = 24 floats + 12 bytes, and maybe 6 shorts, per cube face. See how in the latter case, vertices 0 and 2 are used twice, but only represented once in each of the verts, normals and colors arrays. This sounds like a small win for using indices, even in the extreme case of every single geometry edge being a seam. This leads me to conclude that: 3. When using GL_TRIANGLES, one should always use indexed arrays, even for geometry which is all seams. Please correct my conclusions in bold if they are wrong.

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• SQL efficiency argument, add a column or solvable by query?

  - by theTurk

  I am a recent college graduate and a new hire for software development. Things have been a little slow lately so I was given a db task. My db skills are limited to pet projects with Rails and Django. So, I was a little surprised with my latest task. I have been asked by my manager to subclass Person with a 'Parent' table and add a reference to their custodian in the Person table. This is to facilitate going from Parent to Form when the custodian, not the Parent, is the FormContact. Here is a simplified, mock structure of a sql-db I am working with. I would have drawn the relationship tables if I had access to Visio. We have a table 'Person' and we have a table 'Form'. There is a table, 'FormContact', that relates a Person to a Form, not all Persons are related to a Form. There is a relationship table for Person to Person relationships (Employer, Parent, etc.) I've asked, "Why this couldn't be handled by a query?" Response, Inefficient. (Really!?!) So, I ask, "Why not have a reference to the Form? That would be more efficient since you wouldn't be querying the FormContacts table with the reference from child/custodian." Response, this would essentially make the Parent is a FormContact. (Fair enough.) I went ahead an wrote a query to get from non-FormContact Parent to Form, and tested on the production server. The response time was instantaneous. *SOME_VALUE* is the Parent's fk ID. SELECT FormID FROM FormContact WHERE FormContact.ContactID IN (SELECT SourceContactID FROM ContactRelationship WHERE (ContactRelationship.RelatedContactID = *SOME_VALUE*) AND (ContactRelationship.Relationship = 'Parent')); If I am right, "This is an unnecessary change." What should I do, defend my position or should I concede to the managers request? If I am wrong. What is my error? Is there a better solution than the manager's?

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• Using label tags in a validation summary error list?

  - by patridge

  I was thinking about making use of <label> tags in my validation error summary on a failed form submit and I can't figure out if it is going to get me in trouble down the line. Can anyone think of a good reason to avoid this approach? Usability, functionality, design, or other issues are all helpful. I really like the idea of clicking a line item in the error list and being jumped to the offending input element, especially in a mobile HTML scenario where vertical orientation is more common and scrolling is a pain. So far the only problem I can find is that labels don't navigate for radio buttons or checkboxes without individual IDs (Clicking a label for a single ID-tagged radio/checkbox element alters its selection). It doesn't make it any worse than no label, though. Here is a stripped down HTML test sample of this idea (CSS omitted for simplicity). <div class="validation-errors"> <p>There was a problem saving your form.</p> <ul> <li><label for="select1">Select 1 is invalid.</label></li> <li><label for="text1">Text 1 is invalid.</label></li> <li><label for="textarea1">TextArea 1 is invalid.</label></li> <li><label for="radio1">Radio 1 is invalid.</label></li> <li><label for="checkbox1">Checkbox 1 is invalid.</label></li> </ul> </div> <form action="/somewhere"> <fieldset><legend>Some Form</legend> <ol> <li><label for="select1">select1</label> <select id="select1" name="select1"> <option value="value1">Value 1</option> <option value="value2">Value 2</option> <option selected="selected" value="value3">Value 3</option> </select></li> <li><label for="text1">text1</label> <input id="text1" name="text1" type="text" value="sometext" /></li> <li><label for="textarea1">textarea1</label> <textarea id="textarea1" name="textarea1" rows="5" cols="25">sometext</textarea></li> <li><ul> <li><label><input type="radio" name="radio1" value="value1" />Value 1</label></li> <li><label><input type="radio" name="radio1" value="value2" checked="checked" />Value 2</label></li> <li><label><input type="radio" name="radio1" value="value3" />Value 3</label></li> </ul></li> <li><ul> <li><label><input type="checkbox" name="checkbox1" value="value1" checked="checked" />Value 1</label></li> <li><label><input type="checkbox" name="checkbox1" value="value2" />Value 2</label></li> <li><label><input type="checkbox" name="checkbox1" value="value3" checked="checked" />Value 3</label></li> </ul></li> <li><input type="submit" value="Save &amp; Continue" /></li> </ol> </fieldset> </form> The only thing I have added to make the click-capable behavior more obvious is to add a CSS rule for the labels. .validation-errors label { text-decoration: underline; cursor: pointer; }

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• can you simlify and generalize this useful jQuery function?

  - by user199368

  Hi, I'm doing an eshop with goods displayed as "tiles" in grid as usual. I just want to use various sizes of tiles and make sure (via jQuery) there are no free spaces. In basic situation, I have a 960px wrapper and want to use 240x180px (class .grid_4) tiles and 480x360px (class .grid_8) tiles. See image (imagine no margins/paddings there): Problems without jQuery: - when the CMS provides the big tile as 6th, there would be a free space under the 5th one - when the CMS provides the big tile as 7th, there would be a free space under 5th and 6th - when the CMS provides the big tile as 8th, it would shift to next line, leaving position no.8 free My solution so far looks like this: $(".grid_8").each(function(){ //console.log("BIG on position "+($(this).index()+1)+" which is "+(($(this).index()+1)%2?"ODD":"EVEN")); switch (($(this).index()+1)%4) { case 1: // nothing needed //console.log("case 1"); break; case 2: //need to shift one position and wrap into 240px div //console.log("case 2"); $(this).insertAfter($(this).next()); //swaps this with next $(this).prevAll(":nth(0), :nth(1)").wrapAll("<div class=\"grid_4\" />"); break; case 3: //need to shift two positions and wrap into 480px div //console.log("case 3"); $(this).prevAll(":nth(0), :nth(1)").wrapAll("<div class=\"grid_4\" />"); //wraps previous two - forcing them into column $(this).nextAll(":nth(0), :nth(1)").wrapAll("<div class=\"grid_4\" />"); //wraps next two - forcing them into column $(this).insertAfter($(this).next()); //moves behind the second column break; case 0: //need to shift one position //console.log("case 4"); $(this).insertAfter($(this).next()); //console.log("shifted to next line"); break; } }); It should be obvious from the comments how it works - generally always makes sure that the big tile is on odd position (count of preceding small tiles is even) by shifting one position back if needed. Also small tiles to the left from the big one need to be wrapped in another div so that they appear in column rather than row. Now finally the questions: how to generalize the function so that I can use even more tile dimensions like 720x360 (3x2), 480x540 (2x3), etc.? is there a way to simplify the function? I need to make sure that big tile counts as a multiple of small tiles when checking the actual position. Because using index() on the tile on position 12 (last tile in 3rd row) would now return 7 (position 8) because tiles on positions 5 and 9 are wrapped together in one culumn and the big tile is also just a single div, but spans 2x2 positions. any clean way to ensure this? Thank you very much for any hints. Feel free to reuse the code, I think it can be useful. Josef

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• Adding rows with linq trouble with reference table

  - by Laurence Burke

  I am adding a new address and I know the structure as AddressID = PK and all other entities are non nullable. Now on insert of a new row the addrID Pk is autogened and I am wondering if I would have to get that to create a new row in the referencing table EmployeeAddress or does that automatically get generated also. also I want to be able to repopulate the dropdownlist that lists the current employee's addresses with the newly created address. static uint _curEmpID; protected void btnAdd_Click(object sender, EventArgs e) { if (txtZip.Text != "" && txtAdd1.Text != "" && txtCity.Text != "") { TestDataClassDataContext dc = new TestDataClassDataContext(); Address addr = new Address() { AddressLine1 = txtAdd1.Text, AddressLine2 = txtAdd2.Text, City = txtCity.Text, PostalCode = txtZip.Text, StateProvinceID = Convert.ToInt32(ddlState.SelectedValue) }; dc.Addresses.InsertOnSubmit(addr); lblSuccess.Visible = true; lblErrMsg.Visible = false; dc.SubmitChanges(); // // TODO: insert new row in EmployeeAddress to reference CurEmp to newly created address // SetAddrList(); } else { lblErrMsg.Text = "Invalid Input"; lblErrMsg.Visible = true; } } protected void ddlAddList_SelectedIndexChanged(object sender, EventArgs e) { lblErrMsg.Visible = false; lblSuccess.Visible = false; TestDataClassDataContext dc = new TestDataClassDataContext(); dc.ObjectTrackingEnabled = false; if (ddlAddList.SelectedValue != "-1") { var addr = (from a in dc.Addresses where a.AddressID == Convert.ToInt32(ddlAddList.SelectedValue) select a).FirstOrDefault(); txtAdd1.Text = addr.AddressLine1; txtAdd2.Text = addr.AddressLine2; txtCity.Text = addr.City; txtZip.Text = addr.PostalCode; ddlState.SelectedValue = addr.StateProvinceID.ToString(); btnSubmit.Visible = true; btnAdd.Visible = false; } else { txtAdd1.Text = ""; txtAdd2.Text = ""; txtCity.Text = ""; txtZip.Text = ""; btnAdd.Visible = true; btnSubmit.Visible = false; } } protected void SetAddrList() { TestDataClassDataContext dc = new TestDataClassDataContext(); dc.ObjectTrackingEnabled = false; var addList = from addr in dc.Addresses from eaddr in dc.EmployeeAddresses where eaddr.EmployeeID == _curEmpID && addr.AddressID == eaddr.AddressID select new { AddValue = addr.AddressID, AddText = addr.AddressID, }; ddlAddList.DataSource = addList; ddlAddList.DataValueField = "AddValue"; ddlAddList.DataTextField = "AddText"; ddlAddList.DataBind(); ddlAddList.Items.Add(new ListItem("<Add Address>", "-1")); } OK I am hoping that I did not include too much code. I would really appreciate any other comments about I could otherwise improve this code in any other ways also.

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• Multiprogramming in Django, writing to the Database

  - by Marcus Whybrow

  Introduction I have the following code which checks to see if a similar model exists in the database, and if it does not it creates the new model: class BookProfile(): # ... def save(self, *args, **kwargs): uniqueConstraint = {'book_instance': self.book_instance, 'collection': self.collection} # Test for other objects with identical values profiles = BookProfile.objects.filter(Q(**uniqueConstraint) & ~Q(pk=self.pk)) # If none are found create the object, else fail. if len(profiles) == 0: super(BookProfile, self).save(*args, **kwargs) else: raise ValidationError('A Book Profile for that book instance in that collection already exists') I first build my constraints, then search for a model with those values which I am enforcing must be unique Q(**uniqueConstraint). In addition I ensure that if the save method is updating and not inserting, that we do not find this object when looking for other similar objects ~Q(pk=self.pk). I should mention that I ham implementing soft delete (with a modified objects manager which only shows non-deleted objects) which is why I must check for myself rather then relying on unique_together errors. Problem Right thats the introduction out of the way. My problem is that when multiple identical objects are saved in quick (or as near as simultaneous) succession, sometimes both get added even though the first being added should prevent the second. I have tested the code in the shell and it succeeds every time I run it. Thus my assumption is if say we have two objects being added Object A and Object B. Object A runs its check upon save() being called. Then the process saving Object B gets some time on the processor. Object B runs that same test, but Object A has not yet been added so Object B is added to the database. Then Object A regains control of the processor, and has allready run its test, even though identical Object B is in the database, it adds it regardless. My Thoughts The reason I fear multiprogramming could be involved is that each Object A and Object is being added through an API save view, so a request to the view is made for each save, thus not a single request with multiple sequential saves on objects. It might be the case that Apache is creating a process for each request, and thus causing the problems I think I am seeing. As you would expect, the problem only occurs sometimes, which is characteristic of multiprogramming or multiprocessing errors. If this is the case, is there a way to make the test and set parts of the save() method a critical section, so that a process switch cannot happen between the test and the set?

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• deleted gen folder, eclipse isn't generating it now :(

  - by LuxuryMode

  I accidentally deleted my gen folder and now, predictably, my resources are all messed up. I just created a gen folder myself and tried to project clean - that didn't work. Tried right-clicking project and going to android tools fix project properties - didn't work. Tried unchecking build automatically...didn't work. cleaned, closed project, closed eclipse, restarted, etc, etc. Nothing is working and I keep seeing this error: gen already exists but is not a source folder. Convert to a source folder or rename it. EDIT - OK was able to generate R.java, but now I'm getting crazy stuff in the console: [2011-06-14 17:06:11 - fastapp] Conversion to Dalvik format failed with error 1 [2011-06-14 17:06:42 - fastapp] Dx trouble processing "java/awt/font/NumericShaper.class": Ill-advised or mistaken usage of a core class (java.* or javax.*) when not building a core library. This is often due to inadvertently including a core library file in your application's project, when using an IDE (such as Eclipse). If you are sure you're not intentionally defining a core class, then this is the most likely explanation of what's going on. However, you might actually be trying to define a class in a core namespace, the source of which you may have taken, for example, from a non-Android virtual machine project. This will most assuredly not work. At a minimum, it jeopardizes the compatibility of your app with future versions of the platform. It is also often of questionable legality. If you really intend to build a core library -- which is only appropriate as part of creating a full virtual machine distribution, as opposed to compiling an application -- then use the "--core-library" option to suppress this error message. If you go ahead and use "--core-library" but are in fact building an application, then be forewarned that your application will still fail to build or run, at some point. Please be prepared for angry customers who find, for example, that your application ceases to function once they upgrade their operating system. You will be to blame for this problem. If you are legitimately using some code that happens to be in a core package, then the easiest safe alternative you have is to repackage that code. That is, move the classes in question into your own package namespace. This means that they will never be in conflict with core system classes. JarJar is a tool that may help you in this endeavor. If you find that you cannot do this, then that is an indication that the path you are on will ultimately lead to pain, suffering, grief, and lamentation. [2011-06-14 17:06:42 - fastapp] Dx 1 error; aborting [2011-06-14 17:06:42 - fastapp] Conversion to Dalvik format failed with error 1 And eclipse can't resolve the import of my resources import com.me.fastapp.R;

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• Ruby on Rails check box not updating on form submission

  - by user284194

  I have an entries controller that allows users to add contact information the website. The user-submitted information isn't visible to users until the administrator checks a check box and submits the form. So basically my problem is that if I check the check box as an administrator while initially creating an entry (entries#new) the entry will be publicly visible as expected, but if a non-admin user creates an entry (the normal user view doesn't include the 'live' check box, only the admin one does) then that entry is stuck in limbo because the entries#edit view for some reason doesn't update the boolean check box value when logged in as an admin. entries#new view: <% form_for(@entry) do |f| %> <%= f.error_messages %> Name<br /> <%= f.text_field :name %> Mailing Address<br /> <%= f.text_field :address %> #... <%- if current_user -%> <%= f.label :live %><br /> <%= f.check_box :live %> <%- end -%> <%= f.submit 'Create' %> <% end %> entries#edit (only accessible by admin) view: <% form_for(@entry) do |f| %> <%= f.error_messages %> <%= f.label :name %><br /> <%= f.text_field :name %> Mailing Address<br /> <%= f.text_field :address %> <%= f.label :live %><br /> <%= f.check_box :live %> <%= f.submit 'Update' %> <% end %> Any ideas as to why an administrator can't update the :live check box from the edit view? I would greatly appreciate any suggestions. I'm new to rails. I can post more code if it's needed. Thanks for reading my question.

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• should I ever put a major version number into a C#/Java namespace?

  - by Andrew Patterson

  I am designing a set of 'service' layer objects (data objects and interface definitions) for a WCF web service (that will be consumed by third party clients i.e. not in-house, so outside my direct control). I know that I am not going to get the interface definition exactly right - and am wanting to prepare for the time when I know that I will have to introduce a breaking set of new data objects. However, the reality of the world I am in is that I will also need to run my first version simultaneously for quite a while. The first version of my service will have URL of http://host/app/v1service.svc and when the times comes by new version will live at http://host/app/v2service.svc However, when it comes to the data objects and interfaces, I am toying with putting the 'major' version of the interface number into the actual namespace of the classes. namespace Company.Product.V1 { [DataContract(Namespace = "company-product-v1")] public class Widget { [DataMember] string widgetName; } public interface IFunction { Widget GetWidgetData(int code); } } When the time comes for a fundamental change to the service, I will introduce some classes like namespace Company.Product.V2 { [DataContract(Namespace = "company-product-v2")] public class Widget { [DataMember] int widgetCode; [DataMember] int widgetExpiry; } public interface IFunction { Widget GetWidgetData(int code); } } The advantages as I see it are that I will be able to have a single set of code serving both interface versions, sharing functionality where possible. This is because I will be able to reference both interface versions as a distinct set of C# objects. Similarly, clients may use both interface versions simultaneously, perhaps using V1.Widget in some legacy code whilst new bits move on to V2.Widget. Can anyone tell why this is a stupid idea? I have a nagging feeling that this is a bit smelly.. notes: I am obviously not proposing every single new version of the service would be in a new namespace. Presumably I will do as many non-breaking interface changes as possible, but I know that I will hit a point where all the data modelling will probably need a significant rewrite. I understand assembly versioning etc but I think this question is tangential to that type of versioning. But I could be wrong.

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• How does the rsync algorithm correctly identify repeating blocks?

  - by Kai

  I'm on a personal quest to learn how the rsync algorithm works. After some reading and thinking, I've come up with a situation where I think the algorithm fails. I'm trying to figure out how this is resolved in an actual implementation. Consider this example, where A is the receiver and B is the sender. A = abcde1234512345fghij B = abcde12345fghij As you can see, the only change is that 12345 has been removed. Now, to make this example interesting, let's choose a block size of 5 bytes (chars). Hashing the values on the sender's side using the weak checksum gives the following values list. abcde|12345|fghij abcde -> 495 12345 -> 255 fghij -> 520 values = [495, 255, 520] Next we check to see if any hash values differ in A. If there's a matching block we can skip to the end of that block for the next check. If there's a non-matching block then we've found a difference. I'll step through this process. Hash the first block. Does this hash exist in the values list? abcde -> 495 (yes, so skip) Hash the second block. Does this hash exist in the values list? 12345 -> 255 (yes, so skip) Hash the third block. Does this hash exist in the values list? 12345 -> 255 (yes, so skip) Hash the fourth block. Does this hash exist in the values list? fghij -> 520 (yes, so skip) No more data, we're done. Since every hash was found in the values list, we conclude that A and B are the same. Which, in my humble opinion, isn't true. It seems to me this will happen whenever there is more than one block that share the same hash. What am I missing?

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• QValidator for hex input

  - by Evan Teran

  I have a Qt widget which should only accept a hex string as input. It is very simple to restrict the input characters to [0-9A-Fa-f], but I would like to have it display with a delimiter between "bytes" so for example if the delimiter is a space, and the user types 0011223344 I would like the line edit to display 00 11 22 33 44 Now if the user presses the backspace key 3 times, then I want it to display 00 11 22 3. I almost have what i want, so far there is only one subtle bug involving using the delete key to remove a delimiter. Does anyone have a better way to implement this validator? Here's my code so far: class HexStringValidator : public QValidator { public: HexStringValidator(QObject * parent) : QValidator(parent) {} public: virtual void fixup(QString &input) const { QString temp; int index = 0; // every 2 digits insert a space if they didn't explicitly type one Q_FOREACH(QChar ch, input) { if(std::isxdigit(ch.toAscii())) { if(index != 0 && (index & 1) == 0) { temp += ' '; } temp += ch.toUpper(); ++index; } } input = temp; } virtual State validate(QString &input, int &pos) const { if(!input.isEmpty()) { // TODO: can we detect if the char which was JUST deleted // (if any was deleted) was a space? and special case this? // as to not have the bug in this case? const int char_pos = pos - input.left(pos).count(' '); int chars = 0; fixup(input); pos = 0; while(chars != char_pos) { if(input[pos] != ' ') { ++chars; } ++pos; } // favor the right side of a space if(input[pos] == ' ') { ++pos; } } return QValidator::Acceptable; } }; For now this code is functional enough, but I'd love to have it work 100% as expected. Obviously the ideal would be the just separate the display of the hex string from the actual characters stored in the QLineEdit's internal buffer but I have no idea where to start with that and I imagine is a non-trivial undertaking. In essence, I would like to have a Validator which conforms to this regex: "[0-9A-Fa-f]( [0-9A-Fa-f])*" but I don't want the user to ever have to type a space as delimiter. Likewise, when editing what they types, the spaces should be managed implicitly.

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• How to know when a user has really released a key in Java?

  - by Luis Soeiro

  (Edited for clarity) I want to detect when a user presses and releases a key in Java Swing, ignoring the keyboard auto repeat feature. I also would like a pure Java approach the works on Linux, Mac OS and Windows. Requirements: When the user presses some key I want to know what key is that; When the user releases some key, I want to know what key is that; I want to ignore the system auto repeat options: I want to receive just one keypress event for each key press and just one key release event for each key release; If possible, I would use items 1 to 3 to know if the user is holding more than one key at a time (i.e, she hits 'a' and without releasing it, she hits "Enter"). The problem I'm facing in Java is that under Linux, when the user holds some key, there are many keyPress and keyRelease events being fired (because of the keyboard repeat feature). I've tried some approaches with no success: Get the last time a key event occurred - in Linux, they seem to be zero for key repeat, however, in Mac OS they are not; Consider an event only if the current keyCode is different from the last one - this way the user can't hit twice the same key in a row; Here is the basic (non working) part of code: import java.awt.event.KeyListener; public class Example implements KeyListener { public void keyTyped(KeyEvent e) { } public void keyPressed(KeyEvent e) { System.out.println("KeyPressed: "+e.getKeyCode()+", ts="+e.getWhen()); } public void keyReleased(KeyEvent e) { System.out.println("KeyReleased: "+e.getKeyCode()+", ts="+e.getWhen()); } } When a user holds a key (i.e, 'p') the system shows: KeyPressed: 80, ts=1253637271673 KeyReleased: 80, ts=1253637271923 KeyPressed: 80, ts=1253637271923 KeyReleased: 80, ts=1253637271956 KeyPressed: 80, ts=1253637271956 KeyReleased: 80, ts=1253637271990 KeyPressed: 80, ts=1253637271990 KeyReleased: 80, ts=1253637272023 KeyPressed: 80, ts=1253637272023 ... At least under Linux, the JVM keeps resending all the key events when a key is being hold. To make things more difficult, on my system (Kubuntu 9.04 Core 2 Duo) the timestamps keep changing. The JVM sends a key new release and new key press with the same timestamp. This makes it hard to know when a key is really released. Any ideas? Thanks

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• Editing a Gridview row with drop-down lists gets too wide - how can I use popup panels instead?

  - by David

  I have a series of GridViews in a Tab Panel - databound to a generic List of Business Objects. The columns in the Gridview are all similar to the following: <asp:TemplateField HeaderText="Company" SortExpression="Company.ShortName"> <ItemTemplate> <asp:Label ID="lblCompany" runat="server" Text='<%# Bind("Company.ShortName") %>'></asp:Label> </ItemTemplate> <EditItemTemplate> <asp:DropDownList ID="ddlCompany" runat="server"></asp:DropDownList> </EditItemTemplate> </asp:TemplateField> The GridView generates the "Edit" link at the beginning of the row, all the events fire ok. The problem is that the data is getting long. When in 'display mode', it's fine because the GridView control is smart enough to break some text into multiple lines (in particular Project, Title and Worker names can get pretty long). The problem come in editing mode. Drop-down lists DON'T break entries into multiple lines (for obvious reasons). Going into Edit ode on a row in the Gridview can make the Griview expand horizontally to twice the screen size (blowing through the width limits in the Master page and CSS but that's only a related problem). What I need is something like the ModalPopup - but trying to tie it to an ID in an EditItemTemplate gives me errors when the page renders (because the 'ddlXXXX' doesn't exist at the time). In addition I don't know how to dynamically populate the panel so that I can get a response from it (like the ID of the Company they selected). I'm also trying to avoid javascript and would like this to be a 'pure' aspx/code-behind solution (for simplicity's sake among others). All the examples I find are of Modal Popups with the panels pre-defined. Even if it (the popup panel) were something like a list of checkboxes, it could be databound to the SortedList I have ready to go and an OK/Cancel button combination to accept or ignore things. I'm just not sure of what goes where. I'm open to suggestions. Thanks in advance. EDIT: Final solution looks as follows: <asp:TemplateField HeaderText="Company" SortExpression="Company.ShortName"> <ItemTemplate> <asp:Label ID="lblCompany" runat="server" Text='<%# Bind("Company.ShortName") %>'></asp:Label> </ItemTemplate> <EditItemTemplate> <asp:LinkButton ID="lnkCompany" runat="server" Text='<%# Bind("Company.ShortName") %>'></asp:LinkButton> <asp:Panel ID="pnlCompany" runat="server" style="display:none"> <div> <asp:DropDownList ID="ddlCompany" runat="server" ></asp:DropDownList> <br/> <asp:ImageButton ID="btnOKCo" runat="server" ImageUrl="~/Images/greencheck.gif" OnCommand="PopupButton_Command" CommandName="SelectCO" /> <asp:ImageButton ID="btnCxlCo" runat="server" ImageUrl="~/Images/RedX.gif" /> </div> </asp:Panel> <cc1:ModalPopupExtender ID="mpeCompany" runat="server" TargetControlID="lnkCompany" PopupControlID="pnlCompany" BackgroundCssClass="modalBackground" CancelControlID="btnCxlCo" DropShadow="true" PopupDragHandleControlID="pnlCompany" /> </EditItemTemplate> </asp:TemplateField> And in the code-behind, lstIDLabor is the generic List of data lines (of which Company is one of the properties that is also a business object) that is bound to the GridView: Sub PopupButton_Command(ByVal sender As Object, ByVal e As CommandEventArgs) Dim intRow As Integer Dim intVal As Integer RestoreFromSessionVariables() Select Case e.CommandName Case "SelectCO" intRow = grdIDCostLabor.EditIndex Dim ddlCo As DropDownList = CType(grdIDCost.Rows(intRow).FindControl("ddlCompany"), DropDownList) intVal = ddlCo.SelectedValue lstIDLabor(intRow).CompanyID = intVal lstIDLabor(intRow).Company = Company.Read(intVal) Case Else ' End Select MakeSessionVariables() BindGrids() End Sub

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• Mocking a concrete class : templates and avoiding conditional compilation

  - by AshirusNW

  I'm trying to testing a concrete object with this sort of structure. class Database { public: Database(Server server) : server_(server) {} int Query(const char* expression) { server_.Connect(); return server_.ExecuteQuery(); } private: Server server_; }; i.e. it has no virtual functions, let alone a well-defined interface. I want to a fake database which calls mock services for testing. Even worse, I want the same code to be either built against the real version or the fake so that the same testing code can both: Test the real Database implementation - for integration tests Test the fake implementation, which calls mock services To solve this, I'm using a templated fake, like this: #ifndef INTEGRATION_TESTS class FakeDatabase { public: FakeDatabase() : realDb_(mockServer_) {} int Query(const char* expression) { MOCK_EXPECT_CALL(mockServer_, Query, 3); return realDb_.Query(); } private: // in non-INTEGRATION_TESTS builds, Server is a mock Server with // extra testing methods that allows mocking Server mockServer_; Database realDb_; }; #endif template <class T> class TestDatabaseContainer { public: int Query(const char* expression) { int result = database_.Query(expression); std::cout << "LOG: " << result << endl; return result; } private: T database_; }; Edit: Note the fake Database must call the real Database (but with a mock Server). Now to switch between them I'm planning the following test framework: class DatabaseTests { public: #ifdef INTEGRATION_TESTS typedef TestDatabaseContainer<Database> TestDatabase ; #else typedef TestDatabaseContainer<FakeDatabase> TestDatabase ; #endif TestDatabase& GetDb() { return _testDatabase; } private: TestDatabase _testDatabase; }; class QueryTestCase : public DatabaseTests { public: void TestStep1() { ASSERT(GetDb().Query(static_cast<const char *>("")) == 3); return; } }; I'm not a big fan of that compile-time switching between the real and the fake. So, my question is: Whether there's a better way of switching between Database and FakeDatabase? For instance, is it possible to do it at runtime in a clean fashion? I like to avoid #ifdefs. Also, if anyone has a better way of making a fake class that mimics a concrete class, I'd appreciate it. I don't want to have templated code all over the actual test code (QueryTestCase class). Feel free to critique the code style itself, too. You can see a compiled version of this code on codepad.

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• Spring.NET & Immediacy CMS (or how to inject to server side controls without using PageHandlerFactor

  - by Simon Rice

  Is there any way to inject dependencies into an Immediacy CMS control using Spring.NET, ideally without having to use to ContextRegistry when initialising the control? Update, with my own answer The issue here is that Immediacy already has a handler defined in web.config that deals with all aspx pages, & so it's not possible add an entry for Spring.NET's PageHandlerFactory in web.config as per a normal webforms app. That rules out making the control implement ISupportsWebDependencyInjection. Furthermore, most of Immediacy's generated pages are aspx pages that don't physically exist on the drive. I have changed the title of the question to reflect this. What I have done to get Dependency Injection working is: Add the usual entries to web.config for Spring.NET as outlined in the documentation, except for the adding the entry to the <httpHandlers> section. In this case I've got my object definitions in Spring.config. Create the following abstract base class that will deal with all of the Dependency Injection work: DIControl.cs public abstract class DIControl : ImmediacyControl { protected virtual string DIName { get { return this.GetType().Name; } } protected override void OnInit(EventArgs e) { if (ContextRegistry.GetContext().GetObject(DIName, this.GetType()) != null) ContextRegistry.GetContext().ConfigureObject(this, DIName); base.OnInit(e); } } For non-immediacy controls, you can make this server side control inherit from Control or whatever subclass of that you like. For any control with which you wish to use with Spring.NET's Inversion of Control container, define it to inherit from DIControl & add the relelvant entry to Spring.config, for example: SampleControl.cs public class SampleControl : DIControl, INamingContainer { public string Text { get; set; } protected string InjectedText { get; set; } public SampleControl() : base() { Text = "Hello world"; } protected override void RenderContents(HtmlTextWriter output) { output.Write(string.Format("{0} {1}", Text, InjectedText)); } } Spring.config <objects xmlns="http://www.springframework.net"> <object id="SampleControl" type="MyProject.SampleControl, MyAssembly"> <property name="InjectedText" value="from Spring.NET" /> </object> </objects> You can optionally override DIName if you wish to name your entry in Spring.config differently from the name of your class. Provided everything's done correctly, you will have the control writing out "Hello world from Spring.NET!" when used in a page. This solution uses Spring.NET's ContextRegistry from within the control, but I would be surprised if there's no way around that for Immediacy at least since the page objects themselves aren't accessible. However, can this be improved at all from a Spring.NET perspective? Is there maybe an Immediacy plugin that already does this that I'm completely unaware of? Or is there an approach that does this in a more elegant way? I'm open to suggestions.

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• Consolidating coding styles: Funcs, private method, single method classes

  - by jdoig

  Hi all, We currently have 3 devs with, some, conflicting styles and I'm looking for a way to bring peace to the kingdom... The Coders: Foo 1: Likes to use Func's & Action's inside public methods. He uses actions to alias off lengthy method calls and Func's to perform simple tasks that can be expressed in 1 or 2 lines and will be used frequently through out the code Pros: The main body of his code is succinct and very readable, often with only one or 2 public methods per class and rarely any private methods. Cons: The start of methods contain blocks of lambda rich code that other developers don't enjoy reading; and, on occasion, can contain higher order functions that other dev's REALLY don't like reading. Foo 2: Likes to create a private method for (almost) everything the public method will have to do . Pros: Public methods remain small and readable (to all developers). Cons: Private methods are numerous. With private methods that call into other private methods, that call into... etc, etc. Making code hard to navigate. Foo 3: Likes to create a public class with a, single, public method for every, non-trivial, task that needs performing, then dependency inject them into other objects. Pros: Easily testable, easy to understand (one object, one responsibility). Cons: project gets littered by classes, opening multiple class files to understand what code does makes navigation awkward. It would be great to take the best of all these techniques... Foo-1 Has really nice, readable (almost dsl-like) code... for the most part, except for all the Action and Func lambda shenanigans bulked together at the start of a method. Foo-3 Has highly testable and extensible code that just feels a bit "belt-&-braces" for some solutions and has some code-navigation niggles (constantly hitting F12 in VS and opening 5 other .cs files to find out what a single method does). And Foo-2... Well I'm not sure I like anything about the one-huge .cs file with 2 public methods and 12 private ones, except for the fact it's easier for juniors to dig into. I admit I grossly over-simplified the explanations of those coding styles; but if any one knows of any patterns, practices or diplomatic-manoeuvres that can help unite our three developers (without just telling any of them to just "stop it!") that would be great. From a feasibility standpoint : Foo-1's style meets with the most resistance due to some developers finding lambda and/or Func's hard to read. Foo-2's style meets with a less resistance as it's just so easy to fall into. Foo-3's style requires the most forward thinking and is difficult to enforce when time is short. Any ideas on some coding styles or conventions that can make this work?

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