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  • Impersonation on Windows 2000 to Windows XP Leaves Connections Open

    - by Tallek
    I'm running on a Windows 2000 Pro SP4 box (off domain) and trying to impersonate a local user on a Windows XP box (on domain). I'm using code very similar to the WindowsImpersonationContextFacade in the question posted here: http://stackoverflow.com/questions/879704/how-can-i-temporarily-impersonate-a-user-to-open-a-file. I am using impersonation to remotely start and stop windows services as well as access network shares (for some automated integration tests). To get this working, i had to use LOGON32_PROVIDER_DEFAULT and LOGON32_LOGON_NEW_CREDENTIALS when calling LogonUser. Everything worked beautifully ( Windows XP on domain to Windows XP on domain, Windows XP on domain to Windows Server 2003 off domain, and even Windows XP on domain to Windows 2000 off domain). The one issue was running on Windows 2000 Pro SP4 off the domain and trying to impersonate a local user on a Windows XP box running on the domain. To get the Windows 2000 piece working, i had to use LOGON32_PROVIDER_WINNT50 and LOGON32_LOGON_NEW_CREDENTIALS when calling LogonUser. This seemed to get me 95% of the way there, i could now impersonate the local user on the XP box and start/stop services as well as access a network share using the impersonated credentials. I'm running in to one problem though, calling Undo impersonation and closing the token handle seems to leave the connection to the remote box open. After about 10 or so impersonation calls, further impersonation attempts will fail with an error saying something about too many connections are currently open. If i look at the Computer Management - System Tools - Shared Folders - Sessions on my remote Windows XP box, i can see about 10 sessions open to the Windows 2000 box. I can manually close these (i think they may eventually close themselves, but not very quickly) and then impersonation begins working again few more times. This open session issue doesn't seem to be a problem in any of my other test scenarios, just when running locally on a Windows 2000 box. Any ideas? Edit 1: After some more testing and trying out many different things, this seems to be an issue with open sessions not being reused. On Windows 2000 only, every call to LogonUser to get a token and then using that token to impersonate seems to result in a new session being created. I'm guessing Windows XP & Windows Server 2003 are reusing open sessions since i don't seem to be having any issues with them. If I call LogonUser once, then cache the token, I seem to be able to make as many calls to impersonate as I need using the cached token without running in to the "too many connections" issue. This seems like an ugly work around though since i can't call CloseHandle() on my token every time i perform impersonation. Anybody have any thoughts or ideas, or am i stuck with this ugly hack? Thanks

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  • Assignment operator that calls a constructor is broken

    - by Delan Azabani
    I've implemented some of the changes suggested in this question, and (thanks very much) it works quite well, however... in the process I've seemed to break the post-declaration assignment operator. With the following code: #include <cstdio> #include "ucpp" main() { ustring a = "test"; ustring b = "ing"; ustring c = "- -"; ustring d = "cafe\xcc\x81"; printf("%s\n", (a + b + c[1] + d).encode()); } I get a nice "testing cafe´" message. However, if I modify the code slightly so that the const char * conversion is done separately, post-declaration: #include <cstdio> #include "ucpp" main() { ustring a = "test"; ustring b = "ing"; ustring c = "- -"; ustring d; d = "cafe\xcc\x81"; printf("%s\n", (a + b + c[1] + d).encode()); } the ustring named d becomes blank, and all that is output is "testing ". My new code has three constructors, one void (which is probably the one being incorrectly used, and is used in the operator+ function), one that takes a const ustring &, and one that takes a const char *. The following is my new library code: #include <cstdlib> #include <cstring> class ustring { int * values; long len; public: long length() { return len; } ustring() { len = 0; values = (int *) malloc(0); } ustring(const ustring &input) { len = input.len; values = (int *) malloc(sizeof(int) * len); for (long i = 0; i < len; i++) values[i] = input.values[i]; } ustring operator=(ustring input) { ustring result(input); return result; } ustring(const char * input) { values = (int *) malloc(0); long s = 0; // s = number of parsed chars int a, b, c, d, contNeed = 0, cont = 0; for (long i = 0; input[i]; i++) if (input[i] < 0x80) { // ASCII, direct copy (00-7f) values = (int *) realloc(values, sizeof(int) * ++s); values[s - 1] = input[i]; } else if (input[i] < 0xc0) { // this is a continuation (80-bf) if (cont == contNeed) { // no need for continuation, use U+fffd values = (int *) realloc(values, sizeof(int) * ++s); values[s - 1] = 0xfffd; } cont = cont + 1; values[s - 1] = values[s - 1] | ((input[i] & 0x3f) << ((contNeed - cont) * 6)); if (cont == contNeed) cont = contNeed = 0; } else if (input[i] < 0xc2) { // invalid byte, use U+fffd (c0-c1) values = (int *) realloc(values, sizeof(int) * ++s); values[s - 1] = 0xfffd; } else if (input[i] < 0xe0) { // start of 2-byte sequence (c2-df) contNeed = 1; values = (int *) realloc(values, sizeof(int) * ++s); values[s - 1] = (input[i] & 0x1f) << 6; } else if (input[i] < 0xf0) { // start of 3-byte sequence (e0-ef) contNeed = 2; values = (int *) realloc(values, sizeof(int) * ++s); values[s - 1] = (input[i] & 0x0f) << 12; } else if (input[i] < 0xf5) { // start of 4-byte sequence (f0-f4) contNeed = 3; values = (int *) realloc(values, sizeof(int) * ++s); values[s - 1] = (input[i] & 0x07) << 18; } else { // restricted or invalid (f5-ff) values = (int *) realloc(values, sizeof(int) * ++s); values[s - 1] = 0xfffd; } len = s; } ustring operator=(const char * input) { ustring result(input); return result; } ustring operator+(ustring input) { ustring result; result.len = len + input.len; result.values = (int *) malloc(sizeof(int) * result.len); for (long i = 0; i < len; i++) result.values[i] = values[i]; for (long i = 0; i < input.len; i++) result.values[i + len] = input.values[i]; return result; } ustring operator[](long index) { ustring result; result.len = 1; result.values = (int *) malloc(sizeof(int)); result.values[0] = values[index]; return result; } char * encode() { char * r = (char *) malloc(0); long s = 0; for (long i = 0; i < len; i++) { if (values[i] < 0x80) r = (char *) realloc(r, s + 1), r[s + 0] = char(values[i]), s += 1; else if (values[i] < 0x800) r = (char *) realloc(r, s + 2), r[s + 0] = char(values[i] >> 6 | 0x60), r[s + 1] = char(values[i] & 0x3f | 0x80), s += 2; else if (values[i] < 0x10000) r = (char *) realloc(r, s + 3), r[s + 0] = char(values[i] >> 12 | 0xe0), r[s + 1] = char(values[i] >> 6 & 0x3f | 0x80), r[s + 2] = char(values[i] & 0x3f | 0x80), s += 3; else r = (char *) realloc(r, s + 4), r[s + 0] = char(values[i] >> 18 | 0xf0), r[s + 1] = char(values[i] >> 12 & 0x3f | 0x80), r[s + 2] = char(values[i] >> 6 & 0x3f | 0x80), r[s + 3] = char(values[i] & 0x3f | 0x80), s += 4; } return r; } };

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  • Why won't jqGrid populate initially in Chrome

    - by Maxm007
    Hi, I've got a web page with a jqGrid that uses am xmlreader to populate itself with data that is spat out by a RoR service. The page loads fine in firefox and safari. In Chrome however I get a blank grid. Only when I change the sort order by clicking on the columns does it populate. <html> <head> <title>LocalFx</title> <link href="/stylesheets/main.css?1271423251" media="screen" rel="stylesheet" type="text/css" /> <link href="/stylesheets/redmond/jquery-ui-1.8.custom.css?1271404544" media="screen" rel="stylesheet" type="text/css" /> <link href="/stylesheets/ui.jqgrid.css?1265561560" media="screen" rel="stylesheet" type="text/css" /> <script src="/javascripts/jquery-1.3.2.min.js?1259426008" type="text/javascript"></script> <script src="/javascripts/i18n/grid.locale-en.js?1266140090" type="text/javascript"></script> <script src="/javascripts/jquery.jqGrid.min.js?1271437772" type="text/javascript"></script> <script type="text/javascript"> jQuery().ready(function() { jQuery("#list").jqGrid({ xmlReader: { root:"contracts", row:"contract", repeatitems:false, id:"id" }, jsonReader: { repeatitems:false, root:"contracts" }, datatype: 'xml', url:'http://localhost:3000/contracts/index/all.xml', mtype: 'GET', colNames:['User','B/S', 'Currency', 'Amount', 'Rate'], colModel :[ {name:'user', index:'username', width:100 , xmlmap:'user>username'} , {name:'side', index:'side', width:100 , xmlmap:'side'} , {name:'currency', index:'ccy', width:100 , xmlmap:'currency>ccy'} , {name:'amount', index:'amount', width:100 , xmlmap:'amount'}, {name:'rate', index:'rate', width:100 , xmlmap:'exchange-rate>rate'} ], pager: jQuery('#pager'), caption: 'Contracts', sortname: 'side', sortorder: "asc", viewrecords:true, rowNum:10, rowList:[10,20,30] }); $("#list").trigger("reloadGrid") }); </script> </head> <body> <table id="list" align="center" class="scroll"></table> <div id="pager" class="scroll" style="text-align:center;"></div> </body> </html> This is the xml: <contracts type="array"> <contract> <amount type="float">1000.0</amount> <created-at type="datetime">2010-04-16T13:59:40Z</created-at> <currency-id type="integer">488525179</currency-id> <id type="integer">18277852</id> <side>BUY</side> <updated-at type="datetime">2010-04-16T13:59:40Z</updated-at> <user-id type="integer">830138774</user-id> <exchange-rate> <contract-id type="integer">18277852</contract-id> <created-at type="datetime">2010-04-16T13:59:40Z</created-at> <denccy-id type="integer">890731696</denccy-id> <id type="integer">419011264</id> <numccy-id type="integer">488525179</numccy-id> <rate type="float">1.3</rate> <updated-at type="datetime">2010-04-16T13:59:40Z</updated-at> </exchange-rate> <user> <created-at type="datetime">2010-04-16T13:59:40Z</created-at> <id type="integer">830138774</id> <updated-at type="datetime">2010-04-16T13:59:40Z</updated-at> <username>John Doe</username> </user> <currency> <ccy>EUR</ccy> <created-at type="datetime">2010-04-16T13:59:40Z</created-at> <id type="integer">488525179</id> <updated-at type="datetime">2010-04-16T13:59:40Z</updated-at> </currency> </contract> <contract> <amount type="float">500.0</amount> <created-at type="datetime">2010-04-16T13:59:40Z</created-at> <currency-id type="integer">890731696</currency-id> <id type="integer">716237132</id> <side>SELL</side> <updated-at type="datetime">2010-04-16T13:59:40Z</updated-at> <user-id type="integer">830138774</user-id> <exchange-rate> <contract-id type="integer">716237132</contract-id> <created-at type="datetime">2010-04-16T13:59:40Z</created-at> <denccy-id type="integer">890731696</denccy-id> <id type="integer">861902380</id> <numccy-id type="integer">488525179</numccy-id> <rate type="float">1.3</rate> <updated-at type="datetime">2010-04-16T13:59:40Z</updated-at> </exchange-rate> <user> <created-at type="datetime">2010-04-16T13:59:40Z</created-at> <id type="integer">830138774</id> <updated-at type="datetime">2010-04-16T13:59:40Z</updated-at> <username>John Doe</username> </user> <currency> <ccy>GBP</ccy> <created-at type="datetime">2010-04-16T13:59:40Z</created-at> <id type="integer">890731696</id> <updated-at type="datetime">2010-04-16T13:59:40Z</updated-at> </currency> </contract> </contracts>

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  • How and why do I set up a C# build machine?

    - by mmr
    Hi all, I'm working with a small (4 person) development team on a C# project. I've proposed setting up a build machine which will do nightly builds and tests of the project, because I understand that this is a Good Thing. Trouble is, we don't have a whole lot of budget here, so I have to justify the expense to the powers that be. So I want to know: What kind of tools/licenses will I need? Right now, we use Visual Studio and Smart Assembly to build, and Perforce for source control. Will I need something else, or is there an equivalent of a cron job for running automated scripts? What, exactly, will this get me, other than an indication of a broken build? Should I set up test projects in this solution (sln file) that will be run by these scripts, so I can have particular functions tested? We have, at the moment, two such tests, because we haven't had the time (or frankly, the experience) to make good unit tests. What kind of hardware will I need for this? Once a build has been finished and tested, is it a common practice to put that build up on an ftp site or have some other way for internal access? The idea is that this machine makes the build, and we all go to it, but can make debug builds if we have to. How often should we make this kind of build? How is space managed? If we make nightly builds, should we keep around all the old builds, or start to ditch them after about a week or so? Is there anything else I'm not seeing here? I realize that this is a very large topic, and I'm just starting out. I couldn't find a duplicate of this question here, and if there's a book out there I should just get, please let me know. EDIT: I finally got it to work! Hudson is completely fantastic, and FxCop is showing that some features we thought were implemented were actually incomplete. We also had to change the installer type from Old-And-Busted vdproj to New Hotness WiX. Basically, for those who are paying attention, if you can run your build from the command line, then you can put it into hudson. Making the build run from the command line via MSBuild is a useful exercise in itself, because it forces your tools to be current.

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  • What is a proper way to store site-level global variables in a SharePoint site?

    - by ccomet
    One thing that has driven me nuts about SharePoint2007 is the apparent inability to have defineable settings that apply specifically to a site or site collection itself, and not the content. I mean, you have some pre-defined settings like the Site Logo, the Site Name, and various other things, but there doesn't appear to be anywhere to add new kinds of settings. The application I am working on needs to be able to create multiple kinds of "project site collections" that all follow a basic template, but have certain additional settings that apply specifically to that site collection and that one alone. In addition to the standard site name we also need to define the Project Number, the Project Name, and the Client Name. And given the requests of some of our clients, we also reach a point where we have to have configurable settings that alter how some of the workflows work, like whether files are marked with Letters or Numbers. Our current solution, which I'm hesitant about, has been to store an XML file on the SharePoint server. This file contains one node for each site collection, identified by the URL of the root site. Inside the node are all of the elements that need to be defined for that site collection. When we need them, we have to access the XML file (which will always require SPSecurity.RunWithElevatedPrivileges to access files right on the server) every time to load it and retrieve the data. There are a lot of automated processes which will have to do this, and I'm hesitant about the stability of this method when we reach hundreds of sites with thousands of files running tens of thousands of workflows, all wanting to access this file. Maybe they're unfounded worries, but I'd rather worry than risk everything breaking in a couple years. I was looking into the SPWeb object and found the AllProperties hashtable. It looks like just the kind of thing which might work, but I don't know how safe it is to be modifying this. I read through both MSDN and the WSS SDK but found nothing that clarified on adding completely new properties into AllProperties. Is it safe to use AllProperties for this kind of thing? Or is there yet another feature that I am missing, which could handle the concept of global variables at the site collection or site scope?

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  • SQL/Schema comparison and upgrade

    - by Workshop Alex
    I have a simple situation. A large organisation is using several different versions of some (desktop) application and each version has it's own database structure. There are about 200 offices and each office will have it's own version, which can be one of 7 different ones. The company wants to upgrade all applications to the latest versions, which will be version 8. The problem is that they don't have a separate database for each version. Nor do they have a separate database for each office. They have one single database which is handled by a dedicated server, thus keeping things like management and backups easier. Every office has it's own database schema and within the schema there's the whole database structure for their specific application version. As a result, I'm dealing with 200 different schema's which need to be upgraded, each with 7 possible versions. Fortunately, every schema knows the proper version so checking the version isn't difficult. But my problem is that I need to create upgrade scripts which can upgrade from version 1 to version 2 to version 3 to etc... Basically, all schema's need to be bumped up one version until they're all version 8. Writing the code that will do this is no problem. the challenge is how to create the upgrade script from one version to the other? Preferably with some automated tool. I've examined RedGate's SQL Compare and Altova's DatabaseSpy but they're not practical. Altova is way too slow. RedGate requires too much processing afterwards, since the generated SQL Script still has a few errors and it refers to the schema name. Furthermore, the code needs to become part of a stored procedure and the code generated by RedGate doesn't really fit inside a single procedure. (Plus, it's doing too much transaction-handling, while I need everything within a single transaction. I have been considering using another SQL Comparison tool but it seems to me that my case is just too different from what standard tools can deliver. So I'm going to write my own comparison tool. To do this, I'll be using ADOX with Delphi to read the catalogues for every schema version in the database, then use this to write the SQL Statements that will need to upgrade these schema's to their next version. (Comparing 1 with 2, 2 with 3, 3 with 4, etc.) I'm not unfamiliar with generating SQL-Script-Generators so I don't expect too many problems. And I'll only be upgrading the table structures, not any of the other database objects. So, does anyone have some good tips and tricks to apply when doing this kind of comparisons? Things to be aware of? Practical tips to increase speed?

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  • java web application best practices

    - by Bruce
    Hi all I'm trying to figure out the optimum way to develop and release a fairly simple web application, and I'm running into several problems. I'll outline the decisions I've made, because somewhere I've clearly gone off the rails.. Hugely grateful for any help! I have what I think is a fairly simple web application. It contains a couple of jsps that reference a couple of java beans, and the usual static html, js, css and images. Decision 1) I wanted to have a clear and clean release procedure, such that I could develop on my local machine and then release reliably to a production machine. I therefore made the decision to package the application into a war file (including all the static resources), to minimize the separate bits and pieces I would need to release. So far so good? Decision 2) I wanted things on my local machine to be as similar as possible to the production environment. So in my html, for example, I may have a reference to a static file such as http://static.foo.com/file . To keep this code working seamlessly on dev and prod, I decided to put static.foo.com in my /etc/hosts when developing locally, so that all the urls work correctly without changing anything. Decision 3) I decided to use eclipse and maven to give me a best practice environment for administering and building my project. So I have a nice tight set up now, except that: Every time I want to change anything in development, like one line in an html file, I have to rebuild the entire project and then wait for tomcat to load the war before I can see if it's what I wanted. So my questions are: 1) Is there a way to connect up eclipse and tomcat so that I don't have to rebuild the war each time? ie tomcat is looking straight at my actual workspace to serve up the static files? 2)I think I'm maybe making things harder by using /etc/hosts to reflect production urls - is there a better way that doesn't involve manually changing over urls (relative urls are fine of course, but where you have many subdomains, say one for static files and one for dynamic, you have to write out the full path, surely?) 3) Is this really best practice?? How do people set things up so that they balance the requirement for an automated, all-encompassing build process on the one hand, and the speed and flexibility to be able to develop javascript and html and css quickly, as quickly as if one just pointed apache at the directory and developed live? What do people find works? Many thanks!

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  • How to handle environment-specific application configuration organization-wide?

    - by Stuart Lange
    Problem Your organization has many separate applications, some of which interact with each other (to form "systems"). You need to deploy these applications to separate environments to facilitate staged testing (for example, DEV, QA, UAT, PROD). A given application needs to be configured slightly differently in each environment (each environment has a separate database, for example). You want this re-configuration to be handled by some sort of automated mechanism so that your release managers don't have to manually configure each application every time it is deployed to a different environment. Desired Features I would like to design an organization-wide configuration solution with the following properties (ideally): Supports "one click" deployments (only the environment needs to be specified, and no manual re-configuration during/after deployment should be necessary). There should be a single "system of record" where a shared environment-dependent property is specified (such as a database connection string that is shared by many applications). Supports re-configuration of deployed applications (in the event that an environment-specific property needs to change), ideally without requiring a re-deployment of the application. Allows an application to be run on the same machine, but in different environments (run a PROD instance and a DEV instance simultaneously). Possible Solutions I see two basic directions in which a solution could go: Make all applications "environment aware". You would pass the environment name (DEV, QA, etc) at the command line to the app, and then the app is "smart" enough to figure out the environment-specific configuration values at run-time. The app could fetch the values from flat files deployed along with the app, or from a central configuration service. Applications are not "smart" as they are in #1, and simply fetch configuration by property name from config files deployed with the app. The values of these properties are injected into the config files at deploy-time by the install program/script. That install script takes the environment name and fetches all relevant configuration values from a central configuration service. Question How would/have you achieved a configuration solution that solves these problems and supports these desired features? Am I on target with the two possible solutions? Do you have a preference between those solutions? Also, please feel free to tell me that I'm thinking about the problem all wrong. Any feedback would be greatly appreciated.

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  • Committed JDO writes do not apply on local GAE HRD, or possibly reused transaction

    - by eeeeaaii
    I'm using JDO 2.3 on app engine. I was using the Master/Slave datastore for local testing and recently switched over to using the HRD datastore for local testing, and parts of my app are breaking (which is to be expected). One part of the app that's breaking is where it sends a lot of writes quickly - that is because of the 1-second limit thing, it's failing with a concurrent modification exception. Okay, so that's also to be expected, so I have the browser retry the writes again later when they fail (maybe not the best hack but I'm just trying to get it working quickly). But a weird thing is happening. Some of the writes which should be succeeding (the ones that DON'T get the concurrent modification exception) are also failing, even though the commit phase completes and the request returns my success code. I can see from the log that the retried requests are working okay, but these other requests that seem to have committed on the first try are, I guess, never "applied." But from what I read about the Apply phase, writing again to that same entity should force the apply... but it doesn't. Code follows. Some things to note: I am attempting to use automatic JDO caching. So this is where JDO uses memcache under the covers. This doesn't actually work unless you wrap everything in a transaction. all the requests are doing is reading a string out of an entity, modifying part of the string, and saving that string back to the entity. If these requests weren't in transactions, you'd of course have the "dirty read" problem. But with transactions, isolation is supposed to be at the level of "serializable" so I don't see what's happening here. the entity being modified is a root entity (not in a group) I have cross-group transactions enabled Another weird thing is happening. If the concurrent modification thing happens, and I subsequently edit more than 5 more entities (this is the max for cross-group transactions), then nothing happens right away, but when I stop and restart the server I get "IllegalArgumentException: operating on too many entity groups in a single transaction". Could it be possible that the PMF is returning the same PersistenceManager every time, or the PM is reusing the same transaction every time? I don't see how I could possibly get the above error otherwise. The code inside the transaction just edits one root entity. I can't think of any other way that GAE would give me the "too many entity groups" error. The relevant code (this is a simplified version) PersistenceManager pm = PMF.getManager(); Transaction tx = pm.currentTransaction(); String responsetext = ""; try { tx.begin(); // I have extra calls to "makePersistent" because I found that relying // on pm.close didn't always write the objects to cache, maybe that // was only a DataNucleus 1.x issue though Key userkey = obtainUserKeyFromCookie(); User u = pm.getObjectById(User.class, userkey); pm.makePersistent(u); // to make sure it gets cached for next time Key mapkey = obtainMapKeyFromQueryString(); // this is NOT a java.util.Map, just FYI Map currentmap = pm.getObjectById(Map.class, mapkey); Text mapData = currentmap.getMapData(); // mapData is JSON stored in the entity Text newMapData = parseModifyAndReturn(mapData); // transform the map currentmap.setMapData(newMapData); // mutate the Map object pm.makePersistent(currentmap); // make sure to persist so there is a cache hit tx.commit(); responsetext = "OK"; } catch (JDOCanRetryException jdoe) { // log jdoe responsetext = "RETRY"; } catch (Exception e) { // log e responsetext = "ERROR"; } finally { if (tx.isActive()) { tx.rollback(); } pm.close(); } resp.getWriter().println(responsetext); EDIT: so I have verified that it fails after exactly 5 transactions. Here's what I do: I create a Foo (root entity), do a bunch of concurrent operations on that Foo, and some fail and get retried, and some commit but don't apply (as described above). Then, I start creating more Foos, and do a few operations on those new Foos. If I only create four Foos, stopping and restarting app engine does NOT give me the IllegalArgumentException. However if I create five Foos (which is the limit for cross-group transactions), then when I stop and restart app engine, I do get the exception. So it seems that somehow these new Foos I am creating are counting toward the limit of 5 max entities per transaction, even though they are supposed to be handled by separate transactions. It's as if a transaction is still open and is being reused by the servlet when it handles the new requests for the 2nd through 5th Foos. EDIT2: it looks like the IllegalArgument thing is independent of the other bug. In other words, it always happens when I create five Foos, even if I don't get the concurrent modification exception. I don't know if it's a symptom of the same problem or if it's unrelated. EDIT3: I found out what was causing the (unrelated) IllegalArgumentException, it was a dumb mistake on my part. But the other issue is still happening. EDIT4: added pseudocode for the datastore access EDIT5: I am pretty sure I know why this is happening, but I will still award the bounty to anyone who can confirm it. Basically, I think the problem is that transactions are not really implemented in the local version of the datastore. References: https://groups.google.com/forum/?fromgroups=#!topic/google-appengine-java/gVMS1dFSpcU https://groups.google.com/forum/?fromgroups=#!topic/google-appengine-java/deGasFdIO-M https://groups.google.com/forum/?hl=en&fromgroups=#!msg/google-appengine-java/4YuNb6TVD6I/gSttMmHYwo0J Because transactions are not implemented, rollback is essentially a no-op. Therefore, I get a dirty read when two transactions try to modify the record at the same time. In other words, A reads the data and B reads the data at the same time. A attempts to modify the data, and B attempts to modify a different part of the data. A writes to the datastore, then B writes, obliterating A's changes. Then B is "rolled back" by app engine, but since rollbacks are a no-op when running on the local datastore, B's changes stay, and A's do not. Meanwhile, since B is the thread that threw the exception, the client retries B, but does not retry A (since A was supposedly the transaction that succeeded).

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  • Symfony2 Forms: is it possible to bind a form in an "unconventional way"?

    - by DonCallisto
    Imagine this scenario: in our company there is an employee that "play" around graphic,css,html and so on. Our new project will born under symfony2 so we're trying some silly - but "real" - stuff (like authentication from db, submit data from a form and persist it to db and so on..) The problem As far i know, learnt from symfony2 "book" that i found on the site (you can find it here), there is an "automated" way for creating and rendering forms: 1) Build the form up into a controller in this way $form = $this->createFormBuilder($task) ->add('task','text'), ->add('dueDate','date'), ->getForm(); return $this->render('pathToBundle:Controller:templateTwig', array('form'=>$form->createview()); 2) Into templateTwig render the template {{ form_widget(form) }} // or single rows method 3) Into a controller (the same that have a route where you can submit data), take back submitted information if($rquest->getMethod()=='POST'){ $form->bindRequest($request); /* and so on */ } Return to scenario Our graphic employee don't want to access controllers, write php and other stuff like those. So he'll write a twig template with a "unconventional" (from symfony2 point of view, but conventional from HTML point of view) method: /* into twig template */ <form action="{{ path('SestanteUserBundle_homepage') }}" method="post" name="userForm"> <div> USERNAME: <input type="text" name="user_name" value="{{ user.username}}"/> </div> <div> EMAIL: <input type="text" name="user_mail" value="{{ user.email }}"/> </div> <input type="hidden" name="user_id" value="{{ id }}" /> <input type="submit" value="modifica i dati"> </form> Now, if into the controller that handle the submission of data we do something like that public function indexAction(Request $request) { if($request->getMethod() == 'POST'){ // sono arrivato per via di un submit, quindi devo modificare i dati prima di farli vedere a video $defaultData = array('message'=>'ho visto questa cosa in esempio, ma non capisco se posso farne a meno'); $form = $this->createFormBuilder($defaultData) ->add('user_name','text') ->add('user_mail','email') ->add('user_id','integer') ->getForm(); $form->bindRequest($request); //bindo la form ad una request $data = $form->getData(); //mi aspetto un'array chiave=>valore /* .... */ We expected that $data will contain an array with key,value from the submitted form. We found that it isn't true. After googling for a while and try with other "bad" ideas, we're frozen into that. So, if you have a "graphic office" that can't handle directly php code, how can we interface from form(s) to controller(s) ? UPDATE It seems that Symfony2 use a different convention for form's field name and lookup once you've submitted that. In particular, if my form's name is addUser and a field is named userName, the field's name will be AddUser[username] so maybe it have a "dynamic" lookup method that will extract form's name, field's name, concat them and lookup for values. Is it possible?

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  • ASP.Net / MySQL : Translating content into several languages

    - by philwilks
    I have an ASP.Net website which uses a MySQL database for the back end. The website is an English e-commerce system, and we are looking at the possibility of translating it into about five other languages (French, Spanish etc). We will be getting human translators to perform the translation - we've looked at automated services but these aren't good enough. The static text on the site (e.g. headings, buttons etc) can easily be served up in multiple languages via .Net's built in localization features (resx files etc). The thing that I'm not so sure about it how best to store and retrieve the multi-language content in the database. For example, there is a products table that includes these fields... productId (int) categoryId (int) title (varchar) summary (varchar) description (text) features (text) The title, summary, description and features text would need to be available in all the different languages. Here are the two options that I've come up with... Create additional field for each language For example we could have titleEn, titleFr, titleEs etc for all the languages, and repeat this for all text columns. We would then adapt our code to use the appropriate field depending on the language selected. This feels a bit hacky, and also would lead to some very large tables. Also, if we wanted to add additional languages in the future it would be time consuming to add even more columns. Use a lookup table We could create a new table with the following format... textId | languageId | content ------------------------------- 10 | EN | Car 10 | FR | Voiture 10 | ES | Coche 11 | EN | Bike 11 | FR | Vélo We'd then adapt our products table to reference the appropriate textId for the title, summary, description and features instead of having the text stored in the product table. This seems much more elegant, but I can't think of a simple way of getting this data out of the database and onto the page without using complex SQL statements. Of course adding new languages in the future would be very simple compared to the previous option. I'd be very grateful for any suggestions about the best way to achieve this! Is there any "best practice" guidance out there? Has anyone done this before?

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  • Databinding, using formulas for unusual binding possible?

    - by Rattenmann
    Edit: added Info for WPF being used I am trying to bind a list of custom objects to a DataGrid. Straight binding seems easy enough, but i need to specify some complex formulas for some extra fields that do not directly show up in my class. Also i want to be able to EDIT the data in the Grid and get updates on related fields. Let me show you an example, because it is really hard to explain. I will simplify it to rooms with items. Each item can be red and blue. My Class looks like this: public class room { public string strRoomName { set; get; } public string strItemname { set; get; } public int intRedItem { set; get; } public int intBlueItem { set; get; } } Now if i use dataTable.ItemSource = myList; i get something like this: nr. | room | name | red | blue 1. living room, ball, 2, 1 2. sleeping room, bunny, 4, 1 3. living room, chair, 3, 2 4. kitchen, ball, 4, 7 5. garage, chair, 1, 4 Now for the complex part i need help with. I want every item to be the same number, red and blue. And because this does not hold true i want to see the "inbalance" per room AND globally like this: nr. | room | name | red | blue | missing | global red | global blue | global missing 1. living room, ball, 2, 1, 1 blue, 6, 7, 1 red 2. sleeping room, bunny, 4, 1, 3 blue, 4, 1, 3 blue 3. living room, chair, 3, 2, 1 blue, 4, 6, 2 red 4. kitchen, ball, 4, 7, 3 red, 6, 7, 1 red 5. garage, chair, 1, 4, 3 red, 4, 6, 2 red As you can see this smeels like excel formulas, i am unsure how to handle this in c# code however. You can also see i need to use data in the same row, but also get data from other rows that match one propertiy (the items name). Also if i change the blue value=1 in line 1 to value=2, i want line 1 to read like this: 1. living room, ball, 2, 2, even, 6, 8, 2 red and of corse line 4 needs to change to: 4. kitchen, ball, 4, 7, 3 red, 6, 8, 2 red As i said, this smells like excel, that's why i am really upset about myself not finding an easy solution. Surely enough c# offers some way to handle this stuff, right? Disclaimer: It is totally possible that i need a complete differend approach, pointing that out ot me is perfectly fine. Be it other ways to handle this, or a better way to structure my class. I am ok with every way to handle this as it is for learning purposes. I am simply doing programms for fun next to my college and just so happen to hit these kinda things that bug me out because i don't find a clean solution. And then i neglect my studies because i want to solve my (unreleated to studys,...) issue. Just can't stand having unsolved coding stuff around, don't judge me! ;-) And big thanks in advance if you have gotten this far in my post. It sure must be confusing with all those reds and blues. Edit: After reading trough your answers and testing my skills to implement your hints, i now have the following code as my class: public class RoomList : ObservableCollection<room> { public RoomList() : base() { Add(new room() { strRoomName = "living room", strItemname = "ball", intRedItem = 2, intBlueItem = 1 }); Add(new room() { strRoomName = "sleeping room", strItemname = "bunny", intRedItem = 4, intBlueItem = 1 }); Add(new room() { strRoomName = "living room", strItemname = "chair", intRedItem = 3, intBlueItem = 2 }); Add(new room() { strRoomName = "kitchen", strItemname = "ball", intRedItem = 4, intBlueItem = 7 }); Add(new room() { strRoomName = "garage", strItemname = "chair", intRedItem = 1, intBlueItem = 4 }); } } //rooms public class room : INotifyPropertyChanged { public string strRoomName { set; get; } public string strItemname { set; get; } public int intRedItem { get { return intRedItem; } set { intRedItem = value; NotifyPropertyChanged("intRedItem", "strMissing"); } } public int intBlueItem { get { return intBlueItem; } set { intBlueItem = value; NotifyPropertyChanged("intBlueItem", "strMissing"); } } public string strMissing { get { int missingCount = intRedItem - intBlueItem; return missingCount == 0 ? "Even" : missingCount.ToString(); } } public event PropertyChangedEventHandler PropertyChanged; public void NotifyPropertyChanged(params string[] propertyNames) { if (PropertyChanged != null) { foreach (string propertyName in propertyNames) { PropertyChanged(this, new PropertyChangedEventArgs(propertyName)); } } } } I got the "missing" field working right away, thanks alot for that tip. It really was as easy as i imagined and will be of great use for future projects. Still two (three maybe....) things i am missing tho. The above code terminates with a "System.StackOverflowException" in the setter of intRedItem and intBlueItem. I fail to see the error, that could be due to being 4:30am here, or my lack of understanding. Second issue: I followed the link to ObservableCollections as you can see from my code above. Yet i am unsure how to actually use that collection. Putting it as DataContent like suggested on that page shows a missing ressource. Adding it as a ressource like listed there crashes my VSExpress designer and leads to the programm not starting. So for now i am still using my old approach of a list like this: listRooms.Add(new room() { strRoomName = "living room", strItemname = "ball", intRedItem = 2, intBlueItem = 1 }); listRooms.Add(new room() { strRoomName = "sleeping room", strItemname = "bunny", intRedItem = 4, intBlueItem = 1 }); listRooms.Add(new room() { strRoomName = "living room", strItemname = "chair", intRedItem = 3, intBlueItem = 2 }); listRooms.Add(new room() { strRoomName = "kitchen", strItemname = "ball", intRedItem = 4, intBlueItem = 7 }); listRooms.Add(new room() { strRoomName = "garage", strItemname = "chair", intRedItem = 1, intBlueItem = 4 }); datagridRooms.ItemsSource = listRooms; And lastly: When testing before adding the notifyevents i tried to implement a proterty that looped trough the other objects, without any luck. The "missingItem" property worked so easy, yet it only tries to access "it's own" properties kind of. I need to access other objects, like "all objects that have the same room value". The idea behind this is that i am trying to calculate a value from other objects without even having those objects yet, at least in my logic. Where is the flaw in my thinking? Those 5 objects are added and created (?) one after another. So if the first tries to set it's "all red balls in my room AND all other rooms" value,.. how could it know about the balls in the kitchen, that get added as 4th object? So far so good tho, got on the right track i think. Just need some sleep first.

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  • Visual Studio not recognizing "BuildStep"

    - by AmbiguousX
    I'm trying to add an automatic post-build trigger to run NDepend after an automated team build in TFS 2010. NDepend's website provided code for integrating this capability, and so I have pasted their code into my .csproj file where they said for it to go, but I receive errors on the build. The errors refer to two of the three "BuildStep" tags I have in the code snippet. The following two snippets are giving me errors: <BuildStep TeamFoundationServerUrl="$(TeamFoundationServerUrl)" BuildUri="$(BuildUri)" Message="Running NDepend analysis"> <Output TaskParameter="Id" PropertyName="StepId" /> </BuildStep> and <BuildStep TeamFoundationServerUrl="$(TeamFoundationServerUrl)" BuildUri="$(BuildUri)" Id="$(StepId)" Status="Failed" /> However, this code snippet is NOT throwing up any problems: <BuildStep TeamFoundationServerUrl="$(TeamFoundationServerUrl)" BuildUri="$(BuildUri)" Id="$(StepId)" Status="Succeeded" /> I just don't understand why one works fine and a nearly identically laid out BuildStep tag does not. Is there something simple that I'm just overlooking? EDIT: Here is how it looks all together, if this makes a difference: <Target Name="NDepend" > <PropertyGroup> <NDPath>c:\tools\NDepend\NDepend.console.exe</NDPath> <NDProject>$(SolutionDir)MyProject.ndproj</NDProject> <NDOut>$(TargetDir)NDepend</NDOut> <NDIn>$(TargetDir)</NDIn> </PropertyGroup> <Exec Command='"$(NDPath)" "$(NDProject)" /OutDir "$(NDOut)" /InDirs "$(NDIn)"'/> </Target> <Target Name="AfterBuild"> <BuildStep TeamFoundationServerUrl="$(TeamFoundationServerUrl)" BuildUri="$(BuildUri)" Message="Running NDepend analysis"> <Output TaskParameter="Id" PropertyName="StepId" /> </BuildStep> <PropertyGroup> <NDPath>c:\tools\NDepend\NDepend.console.exe</NDPath> <NDProject>$(SolutionRoot)\Main\src\MyProject.ndproj</NDProject> <NDOut>$(BinariesRoot)\NDepend</NDOut> <NDIn>$(BinariesRoot)\Release</NDIn> </PropertyGroup> <Exec Command='$(NDPath) "$(NDProject)" /OutDir "$(NDOut)" /InDirs "$(NDIn)"'/> <BuildStep TeamFoundationServerUrl="$(TeamFoundationServerUrl)" BuildUri="$(BuildUri)" Id="$(StepId)" Status="Succeeded" /> <OnError ExecuteTargets="MarkBuildStepAsFailed" /> </Target> <Target Name="MarkBuildStepAsFailed"> <BuildStep TeamFoundationServerUrl="$(TeamFoundationServerUrl)" BuildUri="$(BuildUri)" Id="$(StepId)" Status="Failed" /> </Target> EDIT: Added a bounty because I really need to get this going for my team. Thank you in advance!

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  • Find Label Control in Repeater Asp.net

    - by user2769165
    I am using repeater and I want to find the label control in my repeater. here is my code <asp:Repeater ID="friendRepeater" runat="server"> <table cellpadding="0" cellspacing="0"> <ItemTemplate> <tr style=" width:700px; height:120px;"> <td> <div style=" padding-left:180px;"> <div id="leftHandPost" style="float:left; width:120px; height:120px; border: medium solid #cdaf95; padding-top:5px;"> <div id="childLeft" style=" padding-left:5px;"> <div id="photo" style=" border: thin solid black; width:100px;height:100px;"> <asp:Image id="photoImage" runat="server" ImageUrl='<%# String.Concat("Images/", Eval("Picture")) %>' Width="100px" Height="100px" /> </div> </div><!--childLeft--> </div><!--leftHandPost--> </div> </td> <td> <div id="rightHandPost" style=" float:right; padding-right:260px;"> <div id="childRight" style="width:400px; height:120px; border: medium solid #cdaf95; padding-top:5px; padding-left:10px;"> <strong><asp:Label id="lblName" runat="server"><%# Eval("PersonName") %></asp:Label></strong><br /> <div style=" float:right; padding-right:10px;"><asp:Button runat="server" Text="Add" onClick="add" /></div><br /> <asp:Label id="lblID" runat="server"><%# Eval("PersonID") %></asp:Label><br /> <asp:Label id="lblEmail" runat="server"><%# Eval("Email") %></asp:Label> </div><!--childRight--> </div><!--rightHandPost--> </td> </tr> </ItemTemplate> <AlternatingItemTemplate> <tr style=" width:700px; height:120px;"> <td> <div style=" padding-left:180px;"> <div id="Div1" style="float:left; width:120px; height:120px; border: medium solid #cdaf95; padding-top:5px;"> <div id="Div2" style="padding-left:5px;"> <div id="Div3" style=" border: thin solid black; width:100px;height:100px;"> <asp:Image id="photoImage" runat="server" ImageUrl='<%# String.Concat("Images/", Eval("Picture")) %>' Width="100px" Height="100px" /> </div> </div><!--childLeft--> </div><!--leftHandPost--> </div> </td> <td> <div id="Div4" style=" float:right; padding-right:260px;"> <div id="Div5" style="width:400px; height:120px; border: medium solid #cdaf95; padding-top:5px; padding-left:10px;"> <strong><asp:Label id="lblName" runat="server"><%# Eval("PersonName")%></asp:Label></strong> <div style=" float:right; padding-right:10px;"><asp:Button id="btnAdd" runat="server" Text="Add" onClick="add"></asp:Button></div><br /> <br /> <asp:Label id="lblID" runat="server"><%# Eval("PersonID") %></asp:Label><br /> <asp:Label id="lblEmail" runat="server"><%# Eval("Email") %></asp:Label> </div><!--childRight--> </div><!--rightHandPost--> </td> </tr> </AlternatingItemTemplate> <FooterTemplate> </table> </FooterTemplate> </asp:Repeater> Here is the code behind for the add button. protected void add(object sender, EventArgs e) { DateTime date = DateTime.Now; System.Web.UI.WebControls.Label la = (System.Web.UI.WebControls.Label)friendRepeater.FindControl("PersonID"); String id = la.Text; try { MySqlConnection connStr = new MySqlConnection(); connStr.ConnectionString = "Server = localhost; Database = healthlivin; Uid = root; Pwd = khei92;"; String insertFriend = "INSERT INTO contactFriend(friendID, PersonID, PersonIDB, date) values (@id, @personIDA, @personIDB, @date)"; MySqlCommand cmdInsertStaff = new MySqlCommand(insertFriend, connStr); cmdInsertStaff.Parameters.AddWithValue("@id", "F000004"); cmdInsertStaff.Parameters.AddWithValue("@personIDA", "M000001"); cmdInsertStaff.Parameters.AddWithValue("@personIDB", id); cmdInsertStaff.Parameters.AddWithValue("@date", date); connStr.Open(); cmdInsertStaff.ExecuteNonQuery(); MessageBox.Show("inserted"); connStr.Close(); } catch (Exception ex) { MessageBox.Show(ex.ToString()); } } I have get the error of Object reference not set to an instance of an object. I think is because there are no value in the Label. The Find Control are not working. May I know how can fix this problem? Thank you very much

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

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

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  • Recommendations for distributed processing/distributed storage systems

    - by Eddie
    At my organization we have a processing and storage system spread across two dozen linux machines that handles over a petabyte of data. The system right now is very ad-hoc; processing automation and data management is handled by a collection of large perl programs on independent machines. I am looking at distributed processing and storage systems to make it easier to maintain, evenly distribute load and data with replication, and grow in disk space and compute power. The system needs to be able to handle millions of files, varying in size between 50 megabytes to 50 gigabytes. Once created, the files will not be appended to, only replaced completely if need be. The files need to be accessible via HTTP for customer download. Right now, processing is automated by perl scripts (that I have complete control over) which call a series of other programs (that I don't have control over because they are closed source) that essentially transforms one data set into another. No data mining happening here. Here is a quick list of things I am looking for: Reliability: These data must be accessible over HTTP about 99% of the time so I need something that does data replication across the cluster. Scalability: I want to be able to add more processing power and storage easily and rebalance the data on across the cluster. Distributed processing: Easy and automatic job scheduling and load balancing that fits with processing workflow I briefly described above. Data location awareness: Not strictly required but desirable. Since data and processing will be on the same set of nodes I would like the job scheduler to schedule jobs on or close to the node that the data is actually on to cut down on network traffic. Here is what I've looked at so far: Storage Management: GlusterFS: Looks really nice and easy to use but doesn't seem to have a way to figure out what node(s) a file actually resides on to supply as a hint to the job scheduler. GPFS: Seems like the gold standard of clustered filesystems. Meets most of my requirements except, like glusterfs, data location awareness. Ceph: Seems way to immature right now. Distributed processing: Sun Grid Engine: I have a lot of experience with this and it's relatively easy to use (once it is configured properly that is). But Oracle got its icy grip around it and it no longer seems very desirable. Both: Hadoop/HDFS: At first glance it looked like hadoop was perfect for my situation. Distributed storage and job scheduling and it was the only thing I found that would give me the data location awareness that I wanted. But I don't like the namename being a single point of failure. Also, I'm not really sure if the MapReduce paradigm fits the type of processing workflow that I have. It seems like you need to write all your software specifically for MapReduce instead of just using Hadoop as a generic job scheduler. OpenStack: I've done some reading on this but I'm having trouble deciding if it fits well with my problem or not. Does anyone have opinions or recommendations for technologies that would fit my problem well? Any suggestions or advise would be greatly appreciated. Thanks!

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  • What is wrong in my DKIM setup? I'm getting all fails

    - by djechelon
    I own a domain name I have implemented SPF and DKIM to avoid my mails being junked. I have also upgraded to DMARC in monitor mode. Since I received a few failure reports recently I wanted to investigate more. I have only one server sending outbound emails, running postfix + dkimproxy. I trust that dkimproxy has no major software bugs resulting in bad messages. I have tested ReturnPath's automated DKIM test and this is the part related to DKIM/DomainKeys DKIM Results ============ Result = failed: invalid key for signature: Syntax error in tag: \"v Domain = domain.org Selector = sel DNS Record(s) = sel._domainkey.domain.org TXT "v=1; p=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; t=s" Public Key Length = 4096 DomainKeys Results ================== Domain = domain.org Selector = sel DNS Record(s) = sel._domainkey.domain.org TXT "v=1; p=MIICIjANBgkqhkiG9w0BAQEFAAOCAg8AMIICCgKCAgEAsMMLhxzXkU+tagc44oMi7eX2BsFb8BsWeT8MRL+hxi4Lsosx7tuPm90iYgilNteyJoXuSP5SUf8B2tDAifdzYQhfhctr0hX9b6ocBCukGq5p0GHpNsCPWyFvxZsCkGqLRmkfb0c36quEAWBeQLe4Z/BwXBBiW1g96WFNb2/GRI1+9OHhligdfuo4PPuU+xiwX4GB0Ik50cJL4xTdBf7lrFwoGYa03ZkXuzKxeGE4cTk50OeIs6eqrzAfbmej4nCex2qGOUt1TWI7ZvCY7u3Gxj+XKaE7VFrQACZof+NP0k2pXPHg9saGJqZrr2i6+RoxGD0w/ibjAWij9enwqlnv2ORsZfe+FmXNOLJAhlYvhHaruubDpte1c7V3ZKDceM45ZawnVmSdLCfBrMbsqipzy8NXN5MxuANYFBkx5EDT+Ieab+zqcnf08m9bgDc4RXMYppDT1/lUy6On+nyfZEnJWiH3BUtgxS8X0uXciXbsooTmPnpkzzvvKXAE/Tv3XqL90q51geqP0EmaZI6lRTpiqoX7zFGlEBiiF7/u8oheszATks8LsNZ/boTFy0OVldbYNhxlIuRmqeXkqD6+kM5ObKtMEv3AdaeBiZmvyJTP8tCsSmPt+e954RLlz2HaDjjNnZNgsj/39U2RzZsFbVqW6uyQh36/y1X4joOiPf366GkCAwEAAQ==; t=s" The mail displays an anonymised DNS record with genuine public key. It reports an error in tag v. A few hours ago I noticed my v tag was v=DKIM1 instead of v=1 as specified in RFC. I thought it was an error made by me during the initial setup months ago and fixed to v=1, but anyway I received one DMARC success from Google. Let me explain better: I enforced DMARC a couple of days ago. On 4/16 morning I got a mail from Google telling me that DMARC fully passes, then since 4/17 I get all failures. Then I discovered the v=DKIM1 tag and replaced with v=1 without success I have not modified my DNS records before that. So, keeping in topic with the question, why does ReturnPath refuse my DKIM DNS record? Is something wrong in my DKIM implementation at DNS level? [Add] I have just tried port25.com's tester but at least DKIM passes ---------------------------------------------------------- DomainKeys check details: ---------------------------------------------------------- Result: permerror (DK_STAT_BADKEY: Unusable key, public if verifying, private if signing.) ID(s) verified: header.From=########### DNS record(s): sel._domainkey.domain.org. 1800 IN TXT ""v=1; p=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; t=s"" ---------------------------------------------------------- DKIM check details: ---------------------------------------------------------- Result: pass (matches From: #########) ID(s) verified: header.d=domain.org Canonicalized Headers: message-id:<[email protected]>'0D''0A' date:Thu,'20'18'20'Apr'20'2013'20'11:40:26'20'+0200'0D''0A' from:#############'0D''0A' mime-version:1.0'0D''0A' to:[email protected]'0D''0A' subject:Test'0D''0A' content-type:text/plain;'20'charset=ISO-8859-15;'20'format=flowed'0D''0A' content-transfer-encoding:7bit'0D''0A' dkim-signature:v=1;'20'a=rsa-sha1;'20'c=relaxed;'20'd=domain.org;'20'h='20'message-id:date:from:mime-version:to:subject:content-type'20':content-transfer-encoding;'20's=dom;'20'bh=uoq1oCgLlTqpdDX/iUbLy7J1Wi'20'c=;'20'b= Canonicalized Body: '0D''0A' DNS record(s): sel._domainkey.domain.org. 1800 IN TXT ""v=1; p=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; t=s"" Public key used for verification: sel._domainkey.domain.org (4096 bits)

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  • multi-user rvm gem install failure when called from CloudFormation::Init

    - by Peter Mounce
    I've taken an Amazon Linux AMI (based on CentOS) and installed RVM (1.10.3) to it in multi-user fashion (see {1} below). I used that to install ruby 1.9.3-p125, rubygems 1.8.17, and bundler 1.1 as the baseline requirements for most things I'm going to be using the instances for. I've captured that instance to an AMI, and am now launching it via CloudFormation, with some CloudFormation::Init commands. One of them is to use s3cmd to pull down a private gem from S3, and the next one, the one that fails, is to install that gem. It fails with an error message 2012-03-15 16:53:20,201 [ERROR] Command 20_install_gems (/usr/local/rvm/rubies/ruby-1.9.3-p125/bin/gem install ./*.gem) failed 2012-03-15 16:53:20,202 [DEBUG] Command 20_install_gems output: /usr/local/rvm/rubies/ruby-1.9.3-p125/bin/gem:12:in `require': no such file to load -- rubygems (LoadError) from /usr/local/rvm/rubies/ruby-1.9.3-p125/bin/gem:12 Now, that happens during the cfn-init execution - I assume, but haven't checked yet, that cfn-init is being run with an environment different from that of ec2-user (there are no other users on the instance). If I run gem install mygem.gem in an interactive session then that works fine. So, my question really, is what should I do to make this work for cfn-init? Have I correctly set up rvm as multi-user? I've confirmed that cfn-init is being run as the root user, with his restricted environment. How should I source the /etc/profile.d/rvm.sh into root's sessions? {1} My semi-automated rvm installation steps (run in interactive session as ec2-user): sudo bash -s stable < <(curl -s https://raw.github.com/wayneeseguin/rvm/master/binscripts/rvm-installer ) sudo gpasswd -a ec2-user rvm # iconv-devel is baked into centos' glibc sudo yum install -y autoconf automake bison bzip2 gcc-c++ git libffi-devel libtool libxml2-devel libxslt-devel libyaml-devel make openssl-devel patch readline readline-devel zlib zlib-devel source /etc/profile.d/rvm.sh rvm list known # in a new session: rvm install ruby-1.9.3-p125 rvm use 1.9.3 --default gem update --system # gems required by public_web-awareness gem install aws-sdk bundler cocaine sinatra echo -e "gem: --no-ri --no-rdoc\n" > /home/ec2-user/.gemrc # delete unnecessary documentation files rm -rf `gem env gemdir`/doc sudo -s sudo echo -e "gem: --no-ri --no-rdoc\n" > /etc/skel/.gemrc sudo echo -e "gem: --no-ri --no-rdoc\n" > /etc/gemrc # ctrl + d out of the sudo session Some environment information: [ec2-user@ip ~]$ echo $PATH /usr/local/rvm/gems/ruby-1.9.3-p125/bin:/usr/local/rvm/gems/ruby-1.9.3-p125@global/bin:/usr/local/rvm/rubies/ruby-1.9.3-p125/bin:/usr/local/rvm/bin:/usr/local/bin:/bin:/usr/bin:/usr/local/sbin:/usr/sbin:/sbin:/opt/aws/bin:/home/ec2-user/bin [ec2-user@ip ~]$ echo $GEM_HOME /usr/local/rvm/gems/ruby-1.9.3-p125 [ec2-user@ip ~]$ echo $GEM_PATH /usr/local/rvm/gems/ruby-1.9.3-p125:/usr/local/rvm/gems/ruby-1.9.3-p125@global [ec2-user@ip ~]$ echo $BUNDLE_PATH [ec2-user@ip ~]$ gem list *** LOCAL GEMS *** aws-sdk (1.3.6) bundler (1.1.0) cocaine (0.2.1) httparty (0.8.1) json (1.6.5) multi_json (1.1.0) multi_xml (0.4.1) nokogiri (1.5.1, 1.5.0) rack (1.4.1) rack-protection (1.2.0) rake (0.9.2) sinatra (1.3.2) tilt (1.3.3) uuidtools (2.1.2) yamler (0.1.0)

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  • ubuntu 10.04; kvm bridged networking not working with public ip addresses

    - by senorsmile
    I have a dedicated hosted server box with ubuntu 10.04 64 bit installed. I would like to run kvm with ubuntu 8.04 installed for some php 5.2 compatible apps(they don't work right with php 5.3, the default in ubuntu 10.04). I installed KVM as instructed at https://help.ubuntu.com/community/KVM/Installation . I installed the vm using virt-manager. I never could figure out how use virt-install or any of those automated installers. I just installed it using the disc. I set up bridged networking as per https://help.ubuntu.com/community/KVM/Networking . However, the bridged connection doesn't work. Here's my /etc/network/interfaces on the host, running ubuntu 10.04. (with specific public ip blanked) auto lo iface lo inet loopback auto eth0 iface eth0 inet manual auto br0 iface br0 inet static address xx.xx.xx.xx netmask 255.255.255.248 gateway xx.xx.xx.xa bridge_ports eth0 bridge_stp on bridge_fd 0 bridge_maxwait 10 ` Here's my /etc/network/interfaces on the guest, running ubuntu 8.04. auto lo iface lo inet loopback auto eth0 iface eth0 inet static address xx.xx.xx.xy netmask 255.255.255.248 gateway xx.xx.xx.xa The two vm's can communicate to each other. But, the guest vm can't access anyone in the real world. Here's my /etc/libvirt/qemu/store_804.xml <domain type='kvm'> <name>store_804</name> <uuid>27acfb75-4f90-a34c-9a0b-70a6927ae84c</uuid> <memory>2097152</memory> <currentMemory>2097152</currentMemory> <vcpu>2</vcpu> <os> <type arch='x86_64' machine='pc-0.12'>hvm</type> <boot dev='hd'/> </os> <features> <acpi/> <apic/> <pae/> </features> <clock offset='utc'/> <on_poweroff>destroy</on_poweroff> <on_reboot>restart</on_reboot> <on_crash>restart</on_crash> <devices> <emulator>/usr/bin/kvm</emulator> <disk type='file' device='disk'> <driver name='qemu' type='raw'/> <source file='/var/lib/libvirt/images/store_804.img'/> <target dev='hda' bus='ide'/> </disk> <disk type='block' device='cdrom'> <driver name='qemu' type='raw'/> <target dev='hdc' bus='ide'/> <readonly/> </disk> <interface type='bridge'> <mac address='52:54:00:26:0b:c6'/> <source bridge='br0'/> <model type='virtio'/> </interface> <console type='pty'> <target port='0'/> </console> <console type='pty'> <target port='0'/> </console> <input type='mouse' bus='ps2'/> <graphics type='vnc' port='-1' autoport='yes'/> <sound model='es1370'/> <video> <model type='cirrus' vram='9216' heads='1'/> </video> </devices> </domain> Any idea where I've gone wrong?

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  • How to automate org-refile for multiple todo

    - by lawlist
    I'm looking to automate org-refile so that it will find all of the matches and re-file them to a specific location (but not archive). I found a fully automated method of archiving multiple todo, and I am hopeful to find or create (with some help) something similar to this awesome function (but for a different heading / location other than archiving): https://github.com/tonyday567/jwiegley-dot-emacs/blob/master/dot-org.el (defun org-archive-done-tasks () (interactive) (save-excursion (goto-char (point-min)) (while (re-search-forward "\* \\(None\\|Someday\\) " nil t) (if (save-restriction (save-excursion (org-narrow-to-subtree) (search-forward ":LOGBOOK:" nil t))) (forward-line) (org-archive-subtree) (goto-char (line-beginning-position)))))) I also found this (written by aculich), which is a step in the right direction, but still requires repeating the function manually: http://stackoverflow.com/questions/7509463/how-to-move-a-subtree-to-another-subtree-in-org-mode-emacs ;; I also wanted a way for org-refile to refile easily to a subtree, so I wrote some code and generalized it so that it will set an arbitrary immediate target anywhere (not just in the same file). ;; Basic usage is to move somewhere in Tree B and type C-c C-x C-m to mark the target for refiling, then move to the entry in Tree A that you want to refile and type C-c C-w which will immediately refile into the target location you set in Tree B without prompting you, unless you called org-refile-immediate-target with a prefix arg C-u C-c C-x C-m. ;; Note that if you press C-c C-w in rapid succession to refile multiple entries it will preserve the order of your entries even if org-reverse-note-order is set to t, but you can turn it off to respect the setting of org-reverse-note-order with a double prefix arg C-u C-u C-c C-x C-m. (defvar org-refile-immediate nil "Refile immediately using `org-refile-immediate-target' instead of prompting.") (make-local-variable 'org-refile-immediate) (defvar org-refile-immediate-preserve-order t "If last command was also `org-refile' then preserve ordering.") (make-local-variable 'org-refile-immediate-preserve-order) (defvar org-refile-immediate-target nil) "Value uses the same format as an item in `org-refile-targets'." (make-local-variable 'org-refile-immediate-target) (defadvice org-refile (around org-immediate activate) (if (not org-refile-immediate) ad-do-it ;; if last command was `org-refile' then preserve ordering (let ((org-reverse-note-order (if (and org-refile-immediate-preserve-order (eq last-command 'org-refile)) nil org-reverse-note-order))) (ad-set-arg 2 (assoc org-refile-immediate-target (org-refile-get-targets))) (prog1 ad-do-it (setq this-command 'org-refile))))) (defadvice org-refile-cache-clear (after org-refile-history-clear activate) (setq org-refile-targets (default-value 'org-refile-targets)) (setq org-refile-immediate nil) (setq org-refile-immediate-target nil) (setq org-refile-history nil)) ;;;###autoload (defun org-refile-immediate-target (&optional arg) "Set current entry as `org-refile' target. Non-nil turns off `org-refile-immediate', otherwise `org-refile' will immediately refile without prompting for target using most recent entry in `org-refile-targets' that matches `org-refile-immediate-target' as the default." (interactive "P") (if (equal arg '(16)) (progn (setq org-refile-immediate-preserve-order (not org-refile-immediate-preserve-order)) (message "Order preserving is turned: %s" (if org-refile-immediate-preserve-order "on" "off"))) (setq org-refile-immediate (unless arg t)) (make-local-variable 'org-refile-targets) (let* ((components (org-heading-components)) (level (first components)) (heading (nth 4 components)) (string (substring-no-properties heading))) (add-to-list 'org-refile-targets (append (list (buffer-file-name)) (cons :regexp (format "^%s %s$" (make-string level ?*) string)))) (setq org-refile-immediate-target heading)))) (define-key org-mode-map "\C-c\C-x\C-m" 'org-refile-immediate-target) It sure would be helpful if aculich, or some other maven, could please create a variable similar to (setq org-archive-location "~/0.todo.org::* Archived Tasks") so users can specify the file and heading, which is already a part of the org-archive-subtree functionality. I'm doing a search and mark because I don't have the wherewithal to create something like org-archive-location for this setup. EDIT: One step closer -- almost home free . . . (defun lawlist-auto-refile () (interactive) (beginning-of-buffer) (re-search-forward "\* UNDATED") (org-refile-immediate-target) ;; cursor must be on a heading to work. (save-excursion (re-search-backward "\* UNDATED") ;; must be written in such a way so that sub-entries of * UNDATED are not searched; or else infinity loop. (while (re-search-backward "\* \\(None\\|Someday\\) " nil t) (org-refile) ) ) )

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  • Computer not POSTing "randomly"?

    - by smoth190
    I have a custom built PC that is exhibiting some...odd... behavior, something I've never seen before. It was working fine one day, and the next day, it wouldn't start. Seeing as I wanted an upgrade anyway, I purchased a new motherboard that was compatible with all my parts. While replacing the motherboard, I accidentally damaged the CPU. Well, I wanted a new one anyway... so I got a new one. Seeing as I was replacing a ton of parts already, I bought a new PSU because the old one was super loud. When I slapped it all together, it starts up, lights, fans, drives, they all start. But I get no display from the monitor. No beeps, which I believe means it doesn't POST. I figured it was the RAM, because after removing the sound card and graphics card, there was nothing else that I hadn't replaced. When I remove both sticks of RAM, I get a continuous beeping, and according to the mobo handbook, means no RAM. So I think the mobo is functional, or atleast partly. I bought new RAM, but it still didn't work. I tried 3 monitors, with both VGA and DIV. So it's probably not the monitor, either. Now, let me get to the random part. Every 20 or so boots (I should also mention, for about 3 out of 5 boots I have to unplug the PC because it won't powerdown via the button), it will POST and I'll get display. Then, after about 2 or 3 resets, it won't work again. This confuses me so much, because even when I change nothing, it will/will not work. My thought is that maybe it has something to do with the RAM not clearing or something. I also reset the CMOS battery, incase that had anything to do with it, but no eval. I found some weird suggestion online about holding the power button for 30 seconds while it was unplugged. That did nothing, but I didn't expect it would... I've replaced just about the entire computer, and all the parts are compatible. Done about everything I can think of, but nothing has worked. Hopefully someone can help me here. And as I side note: When I do get my computer to boot, it says my hardware has changed and I have to re-activate windows. But it says I have to call Microsoft to do it. So I get this fancy automated voice that asks me to enter in a code into windows, then it asks me "How many computers have you activated with this copy of Windows?". Well, I had it on my computer before I replaced everything, so I said 1. Then he yelled at me for violating my 1-use license. I dunno what's going on there, do I have to re-purchase Windows 7? And they wonder why people pirate software... That's just a bonus question, though. Specs: 8GB of DDR2 RAM (Corsair) AMD CPU (I don't know what GHz or model because I can't find the box... (I think its 4 cores of 2.8Ghz) ASRock A785GM-LE Motherboard

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  • WiX 3 Tutorial: Generating file/directory fragments with Heat.exe

    - by Mladen Prajdic
    In previous posts I’ve shown you our SuperForm test application solution structure and how the main wxs and wxi include file look like. In this post I’ll show you how to automate inclusion of files to install into your build process. For our SuperForm application we have a single exe to install. But in the real world we have 10s or 100s of different files from dll’s to resource files like pictures. It all depends on what kind of application you’re building. Writing a directory structure for so many files by hand is out of the question. What we need is an automated way to create this structure. Enter Heat.exe. Heat is a command line utility to harvest a file, directory, Visual Studio project, IIS website or performance counters. You might ask what harvesting means? Harvesting is converting a source (file, directory, …) into a component structure saved in a WiX fragment (a wxs) file. There are 2 options you can use: Create a static wxs fragment with Heat and include it in your project. The pro of this is that you can add or remove components by hand. The con is that you have to do the pro part by hand. Automation always beats manual labor. Run heat command line utility in a pre-build event of your WiX project. I prefer this way. By always recreating the whole fragment you don’t have to worry about missing any new files you add. The con of this is that you’ll include files that you otherwise might not want to. There is no perfect solution so pick one and deal with it. I prefer using the second way. A neat way of overcoming the con of the second option is to have a post-build event on your main application project (SuperForm.MainApp in our case) to copy the files needed to be installed in a special location and have the Heat.exe read them from there. I haven’t set this up for this tutorial and I’m simply including all files from the default SuperForm.MainApp \bin directory. Remember how we created a System Environment variable called SuperFormFilesDir? This is where we’ll use it for the first time. The command line text that you have to put into the pre-build event of your WiX project looks like this: "$(WIX)bin\heat.exe" dir "$(SuperFormFilesDir)" -cg SuperFormFiles -gg -scom -sreg -sfrag -srd -dr INSTALLLOCATION -var env.SuperFormFilesDir -out "$(ProjectDir)Fragments\FilesFragment.wxs" After you install WiX you’ll get the WIX environment variable. In the pre/post-build events environment variables are referenced like this: $(WIX). By using this you don’t have to think about the installation path of the WiX. Remember: for 32 bit applications Program files folder is named differently between 32 and 64 bit systems. $(ProjectDir) is obviously the path to your project and is a Visual Studio built in variable. You can view all Heat.exe options by running it without parameters but I’ll explain some that stick out the most. dir "$(SuperFormFilesDir)": tell Heat to harvest the whole directory at the set location. That is the location we’ve set in our System Environment variable. –cg SuperFormFiles: the name of the Component group that will be created. This name is included in out Feature tag as is seen in the previous post. -dr INSTALLLOCATION: the directory reference this fragment will fall under. You can see the top level directory structure in the previous post. -var env.SuperFormFilesDir: the name of the variable that will replace the SourceDir text that would otherwise appear in the fragment file. -out "$(ProjectDir)Fragments\FilesFragment.wxs": the full path and name under which the fragment file will be saved. If you have source control you have to include the FilesFragment.wxs into your project but remove its source control binding. The auto generated FilesFragment.wxs for our test app looks like this: <?xml version="1.0" encoding="utf-8"?><Wix xmlns="http://schemas.microsoft.com/wix/2006/wi"> <Fragment> <ComponentGroup Id="SuperFormFiles"> <ComponentRef Id="cmp5BB40DB822CAA7C5295227894A07502E" /> <ComponentRef Id="cmpCFD331F5E0E471FC42A1334A1098E144" /> <ComponentRef Id="cmp4614DD03D8974B7C1FC39E7B82F19574" /> <ComponentRef Id="cmpDF166522884E2454382277128BD866EC" /> </ComponentGroup> </Fragment> <Fragment> <DirectoryRef Id="INSTALLLOCATION"> <Component Id="cmp5BB40DB822CAA7C5295227894A07502E" Guid="{117E3352-2F0C-4E19-AD96-03D354751B8D}"> <File Id="filDCA561ABF8964292B6BC0D0726E8EFAD" KeyPath="yes" Source="$(env.SuperFormFilesDir)\SuperForm.MainApp.exe" /> </Component> <Component Id="cmpCFD331F5E0E471FC42A1334A1098E144" Guid="{369A2347-97DD-45CA-A4D1-62BB706EA329}"> <File Id="filA9BE65B2AB60F3CE41105364EDE33D27" KeyPath="yes" Source="$(env.SuperFormFilesDir)\SuperForm.MainApp.pdb" /> </Component> <Component Id="cmp4614DD03D8974B7C1FC39E7B82F19574" Guid="{3443EBE2-168F-4380-BC41-26D71A0DB1C7}"> <File Id="fil5102E75B91F3DAFA6F70DA57F4C126ED" KeyPath="yes" Source="$(env.SuperFormFilesDir)\SuperForm.MainApp.vshost.exe" /> </Component> <Component Id="cmpDF166522884E2454382277128BD866EC" Guid="{0C0F3D18-56EB-41FE-B0BD-FD2C131572DB}"> <File Id="filF7CA5083B4997E1DEC435554423E675C" KeyPath="yes" Source="$(env.SuperFormFilesDir)\SuperForm.MainApp.vshost.exe.manifest" /> </Component> </DirectoryRef> </Fragment></Wix> The $(env.SuperFormFilesDir) will be replaced at build time with the directory where the files to be installed are located. There is nothing too complicated about this. In the end it turns out that this sort of automation is great! There are a few other ways that Heat.exe can compose the wxs file but this is the one I prefer. It just seems the clearest. Play with its options to see what can it do. It’s one awesome little tool.   WiX 3 tutorial by Mladen Prajdic navigation WiX 3 Tutorial: Solution/Project structure and Dev resources WiX 3 Tutorial: Understanding main wxs and wxi file WiX 3 Tutorial: Generating file/directory fragments with Heat.exe

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  • This task is currently locked by a running workflow and cannot be edited. Limitation to both Nintex and SPD workflow

    - by ybbest
    Note, this post is from Nintex Forum here. These limitations apply to both SharePoint designer Workflow and Nintex Workflow as Nintex using the SharePoint workflow engine. The common cause that I experience is that ‘parent’ workflow is generating more than one task at once. This is common as you can have multiple approvers for certain approval process. You could also have workflow running when the task is created, one of the common scenario is you would like to set a custom column value in your approval task. For me this is huge limitation, as Nintex lover I really hope Nintex could solve this problem with Microsoft going forward. Introduction “This task is currently locked by a running workflow and cannot be edited” is a common message that is seen when an error occurs while the SharePoint workflow engine is processing a task item associated with a workflow. When a workflow processes a task normally, the following sequence of events is expected to occur: 1.       The process begins. 2.       The workflow places a ‘lock’ on the task so nothing else can change the values while the workflow is processing. 3.       The workflow processes the task. 4.       The lock is released when the task processing is finished. When the message is encountered, it usually indicates that an error occurred between step 2 and 4. As a result, the lock is never released. Therefore, the ‘task locked’ message is not an error itself, rather a symptom of another error – the ‘task locked’ message does not indicate what went wrong. In most cases, once this message is encountered, the workflow cannot be made to continue and must be terminated and started again. The following is a guide that can help troubleshoot the cause of these messages.  Some initial observations to narrow down the potential causes are: Is the error consistent or intermittent? When the error is consistent, it will happen every time the workflow is run. When it is intermittent, it may happen regularly, but not every time. Does the error occur the first time the user tries to respond to a task, or do they respond and notice the workflow does not continue, and when they respond again the error occurs? If the message is present when the user first responds to the task, the issue would have occurred when the task was created. Otherwise, it would have occurred when the user attempted to respond to the task. Causes Modifying the task list A cause of this error appearing consistently the first time a user tries to respond to a task is a modification to the default task list schema. For example, changing the ‘Assigned to’ field in a task list to be a multiple selection will cause the behaviour. Deleting the workflow task then restoring it from the Recycle bin If you start a workflow, delete the workflow task then restore it from the Recycle Bin in SharePoint, the workflow will fail with the ‘task locked’ error.  This is confirmed behaviour whether using a SharePoint Designer or a Nintex workflow.  You will need to terminate the workflow and start it again. Parallel simultaneous responses A cause of this error appearing inconsistently is multiple users responding to tasks in parallel at the same time. In this scenario, one task will complete correctly and the other will not process. When the user tries again, the ‘task locked’ message will display. Nintex included a workaround for this issue in build 11000. In build 11000 and later, one of the users will receive a message on the task form when they attempt to respond, stating that they need to try again in a few moments. Additional processing on the task A cause of this error appearing consistently and inconsistently is having an additional system running on the items in the task list. Some examples include: a workflow running on the task list, an event receiver running on the task list or another automated process querying and updating workflow tasks. Note: This Microsoft help article (http://office.microsoft.com/en-us/sharepointdesigner/HA102376561033.aspx#5) explains creating a workflow that runs on the task list to update a field on the task. Our experience shows that this causes the ‘Task Locked’ issues when the ‘parent’ workflow is generating more than one task at once. Isolated system error If the error is a rare event, or a ‘one off’ event, then an isolated system error may have occurred. For example, if there is a database connectivity issue while the workflow is processing the task response, the task will lock. In this case, the user will respond to a task but the workflow will not continue. When they respond again, the ‘task locked’ message will display. In this case, there will be an error in the SharePoint ULS Logs at the time that the user originally responded. Temporary delay while workflow processes If the workflow is taking a long time to process after a user submits a task, they may notice and try to respond to the task again. They will see the task locked error, but after a number of attempts (or after waiting some time) the task response page eventually indicates the task has been responded to. In this case, nothing actually went wrong, and the error message gives an accurate indication of what is happening – the workflow temporarily locked the task while it was processing. This scenario may occur in a very large workflow, or after the SharePoint application pool has just started. Modifying the task via a web service with an invalid url If the Nintex Workflow web service is used to respond to or delegate a task, the site context part of the url must be a valid alternative access mapping url. For example, if you access the web service via the IP address of the SharePoint server, and the IP address is not a valid AAM, the task can become locked. The workflow has become stuck without any apparent errors This behaviour can occur as a result of a bug in the SharePoint 2010 workflow engine.  If you do not have the August 2010 Cumulative Update (or later) for SharePoint, and your workflow uses delays, “Flexi-task”, State machine”, “Task Reminder” actions or variables, you could be affected. Check the SharePoint 2010 Updates site here: http://technet.microsoft.com/en-us/sharepoint/ff800847.  The October CU is recommended http://support.microsoft.com/kb/2553031.   The fix is described as “Consider the following scenario. You add a Delay activity to a workflow. Then, you set the duration for the Delay activity. You deploy the workflow in SharePoint Foundation 2010. In this scenario, the workflow is not resumed after the duration of the Delay activity”. If you find this is occurring in your environment, install the October CU, terminate all the running workflows affected and run them afresh. Investigative steps The first step to isolate the issue is to create a new task list on the site and configure the workflow to use it.  Any customizations that were made to the original task list should not be made to the new task list. If the new task list eliminates the issue, then the cause can be attributed to the original task list or a change that was made to it. To change the task list that the workflow uses: In Workflow Designer select Settings -> Startup Options Then configure the task list as required If any of the scenarios above do not help, check the SharePoint logs for any messages with a category of ‘Workflow Infrastructure’. Conclusion The information in this article has been gathered from observations and investigations by Nintex. The sources of these issues are the underlying SharePoint workflow engine. This article will be updated if further causes are discovered. From <http://connect.nintex.com/forums/thread/6503.aspx>

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  • Upgrading Windows 8 boot to VHD to Windows 8.1&ndash;Step by step guide

    - by Liam Westley
    Originally posted on: http://geekswithblogs.net/twickers/archive/2013/10/19/upgrading-windows-8-boot-to-vhd-to-windows-8.1ndashstep-by.aspxBoot to VHD – dual booting Windows 7 and Windows 8 became easy When Windows 8 arrived, quite a few people decided that they would still dual boot their machines, and instead of mucking about with resizing disk partitions to free up space for Windows 8 they decided to use the boot from VHD feature to create a huge hard disc image into which Windows 8 could be installed.  Scott Hanselman wrote this installation guide, while I myself used the installation guide from Ed Bott of ZD net fame. Boot to VHD is a great solution, it achieves a dual boot, can be backed up easily and had virtually no effect on the original Windows 7 partition. As a developer who has dual booted Windows operating systems for years, hacking boot.ini files, the boot to VHD was a much easier solution. Upgrade to Windows 8.1 – ah, you can’t do that on a virtual disk installation (boot to VHD) Last week the final version of Windows 8.1 arrived, and I went into the Windows Store to upgrade.  Luckily I’m on a fast download service, and use an SSD, because once the upgrade was downloaded and prepared Windows informed that This PC can’t run Windows 8.1, and provided the reason, You can’t install Windows on a virtual drive.  You can see an image of the message and discussion that sparked my search for a solution in this Microsoft Technet forum post. I was determined not to have to resize partitions yet again and fiddle with VHD to disk utilities and back again, and in the end I did succeed in upgrading to a Windows 8.1 boot to VHD partition.  It takes quite a bit of effort though … tldr; Simple steps of how you upgrade Boot into Windows 7 – make a copy of your Windows 8 VHD, to become Windows 8.1 Enable Hyper-V in your Windows 8 (the original boot to VHD partition) Create a new virtual machine, attaching the copy of your Windows 8 VHD Start the virtual machine, upgrade it via the Windows Store to Windows 8.1 Shutdown the virtual machine Boot into Windows 7 – use the bcedit tool to create a new Windows 8.1 boot to VHD option (pointing at the copy) Boot into the new Windows 8.1 option Reactivate Windows 8.1 (it will have become deactivated by running under Hyper-V) Remove the original Windows 8 VHD, and in Windows 7 use bcedit to remove it from the boot menu Things you’ll need A system that can run Hyper-V under Windows 8 (Intel i5, i7 class CPU) Enough space to have your original Windows 8 boot to VHD and a copy at the same time An ISO or DVD for Windows 8 to create a bootable Windows 8 partition Step by step guide Boot to your base o/s, the real one, Windows 7. Make a copy of the Windows 8 VHD file that you use to boot Windows 8 (via boot from VHD) – I copied it from a folder on C: called VHD-Win8 to VHD-Win8.1 on my N: drive. Reboot your system into Windows 8, and enable Hyper-V if not already present (this may require reboot) Use the Hyper-V manager , create a new Hyper-V machine, using half your system memory, and use the option to attach an existing VHD on the main IDE controller – this will be the new copy you made in Step 2. Start the virtual machine, use Connect to view it, and you’ll probably discover it cannot boot as there is no boot record If this is the case, go to Hyper-V manager, edit the Settings for the virtual machine to attach an ISO of a Windows 8 DVD to the second IDE controller. Start the virtual machine, use Connect to view it, and it should now attempt a fresh installation of Windows 8.  You should select Advanced Options and choose Repair - this will make VHD bootable When the setup reboots your virtual machine, turn off the virtual machine, and remove the ISO of the Windows 8 DVD from the virtual machine settings. Start virtual machine, use Connect to view it.  You will see the devices to be re-discovered (including your quad CPU becoming single CPU).  Eventually you should see the Windows Login screen. You may notice that your desktop background (Win+D) will have turned black as your Windows installation has become deactivate due to the hardware changes between your real PC and Hyper-V. Fortunately becoming deactivated, does not stop you using the Windows Store, where you can select the update to Windows 8.1. You can now watch the progress joy of the Windows 8 update; downloading, preparing to update, checking compatibility, gathering info, preparing to restart, and finally, confirm restart - remember that you are restarting your virtual machine sitting on the copy of the VHD, not the Windows 8 boot to VHD you are currently using to run Hyper-V (confused yet?) After the reboot you get the real upgrade messages; setting up x%, xx%, (quite slow) After a while, Getting ready Applying PC Settings x%, xx% (really slow) Updating your system (fast) Setting up a few more things x%, (quite slow) Getting ready, again Accept license terms Express settings Confirmed previous password Next, I had to set up a Microsoft account – which is possibly now required, and not optional Using the Microsoft account required a 2 factor authorization, via text message, a 7 digit code for me Finalising settings Blank screen, HI .. We're setting up things for you (similar to original Windows 8 install) 'You can get new apps from the Store', below which is ’Installing your apps’ - I had Windows Media Center which is counts as an app from the Store ‘Taking care of a few things’, below which is ‘Installing your apps’ ‘Taking care of a few things’, below ‘Don't turn off your PC’ ‘Getting your apps ready’, below ‘Don't turn off your PC’ ‘Almost ready’, below ‘Don't turn off your PC’ … finally, we get the Windows 8.1 start menu, and a quick Win+D to check the desktop confirmed all the application icons I expected, pinned items on the taskbar, and one app moaning about a missing drive At this point the upgrade is complete – you can shutdown the virtual machine Reboot from the original Windows 8 and return to Windows 7 to configure booting to the Windows 8.1 copy of the VHD In an administrator command prompt do following use the bcdedit tool (from an MSDN blog about configuring VHD to boot in Windows 7) Type bcedit to list the current boot options, so you can copy the GUID (complete with brackets/braces) for the original Windows 8 boot to VHD Create a new menu option, copy of the Windows 8 option; bcdedit /copy {originalguid} /d "Windows 8.1" Point the new Windows 8.1 option to the copy of the VHD; bcdedit /set {newguid} device vhd=[D:]\Image.vhd Point the new Windows 8.1 option to the copy of the VHD; bcdedit /set {newguid} osdevice vhd=[D:]\Image.vhd Set autodetection of the HAL (may already be set); bcdedit /set {newguid} detecthal on Reboot from Windows 7 and select the new option 'Windows 8.1' on the boot menu, and you’ll have some messages to look at, as your hardware is redetected (as you are back from 1 CPU to 4 CPUs) ‘Getting devices ready, blank then %xx, with occasional blank screen, for the graphics driver, (fast-ish) Getting Ready message (fast) You will have to suffer one final reboots, choose 'Windows 8.1' and you can now login to a lovely Windows 8.1 start screen running on non virtualized hardware via boot to VHD After checking everything is running fine, you can now choose to Activate Windows, which for me was a toll free phone call to the automated system where you type in lots of numbers to be given a whole bunch of new activation codes. Once you’re happy with your new Windows 8.1 boot to VHD, and no longer need the Windows 8 boot to VHD, feel free to delete the old one.  I do believe once you upgrade, you are no longer licensed to use it anyway. There, that was simple wasn’t it? Looking at the huge list of steps it took to perform this upgrade, you may wonder whether I think this is worth it.  Well, I think it is worth booting to VHD.  It makes backups a snap (go to Windows 7, copy the VHD, you backed up the o/s) and helps with disk management – want to move the o/s, you can move the VHD and repoint the boot menu to the new location. The downside is that Microsoft has complete neglected to support boot to VHD as an upgradable option.  Quite a poor decision in my opinion, and if you read twitter and the forums quite a few people agree with that view.  It’s a shame this got missed in the work on creating the upgrade packages for Windows 8.1.

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  • Red Gate Coder interviews: Alex Davies

    - by Michael Williamson
    Alex Davies has been a software engineer at Red Gate since graduating from university, and is currently busy working on .NET Demon. We talked about tackling parallel programming with his actors framework, a scientific approach to debugging, and how JavaScript is going to affect the programming languages we use in years to come. So, if we start at the start, how did you get started in programming? When I was seven or eight, I was given a BBC Micro for Christmas. I had asked for a Game Boy, but my dad thought it would be better to give me a proper computer. For a year or so, I only played games on it, but then I found the user guide for writing programs in it. I gradually started doing more stuff on it and found it fun. I liked creating. As I went into senior school I continued to write stuff on there, trying to write games that weren’t very good. I got a real computer when I was fourteen and found ways to write BASIC on it. Visual Basic to start with, and then something more interesting than that. How did you learn to program? Was there someone helping you out? Absolutely not! I learnt out of a book, or by experimenting. I remember the first time I found a loop, I was like “Oh my God! I don’t have to write out the same line over and over and over again any more. It’s amazing!” When did you think this might be something that you actually wanted to do as a career? For a long time, I thought it wasn’t something that you would do as a career, because it was too much fun to be a career. I thought I’d do chemistry at university and some kind of career based on chemical engineering. And then I went to a careers fair at school when I was seventeen or eighteen, and it just didn’t interest me whatsoever. I thought “I could be a programmer, and there’s loads of money there, and I’m good at it, and it’s fun”, but also that I shouldn’t spoil my hobby. Now I don’t really program in my spare time any more, which is a bit of a shame, but I program all the rest of the time, so I can live with it. Do you think you learnt much about programming at university? Yes, definitely! I went into university knowing how to make computers do anything I wanted them to do. However, I didn’t have the language to talk about algorithms, so the algorithms course in my first year was massively important. Learning other language paradigms like functional programming was really good for breadth of understanding. Functional programming influences normal programming through design rather than actually using it all the time. I draw inspiration from it to write imperative programs which I think is actually becoming really fashionable now, but I’ve been doing it for ages. I did it first! There were also some courses on really odd programming languages, a bit of Prolog, a little bit of C. Having a little bit of each of those is something that I would have never done on my own, so it was important. And then there are knowledge-based courses which are about not programming itself but things that have been programmed like TCP. Those are really important for examples for how to approach things. Did you do any internships while you were at university? Yeah, I spent both of my summers at the same company. I thought I could code well before I went there. Looking back at the crap that I produced, it was only surpassed in its crappiness by all of the other code already in that company. I’m so much better at writing nice code now than I used to be back then. Was there just not a culture of looking after your code? There was, they just didn’t hire people for their abilities in that area. They hired people for raw IQ. The first indicator of it going wrong was that they didn’t have any computer scientists, which is a bit odd in a programming company. But even beyond that they didn’t have people who learnt architecture from anyone else. Most of them had started straight out of university, so never really had experience or mentors to learn from. There wasn’t the experience to draw from to teach each other. In the second half of my second internship, I was being given tasks like looking at new technologies and teaching people stuff. Interns shouldn’t be teaching people how to do their jobs! All interns are going to have little nuggets of things that you don’t know about, but they shouldn’t consistently be the ones who know the most. It’s not a good environment to learn. I was going to ask how you found working with people who were more experienced than you… When I reached Red Gate, I found some people who were more experienced programmers than me, and that was difficult. I’ve been coding since I was tiny. At university there were people who were cleverer than me, but there weren’t very many who were more experienced programmers than me. During my internship, I didn’t find anyone who I classed as being a noticeably more experienced programmer than me. So, it was a shock to the system to have valid criticisms rather than just formatting criticisms. However, Red Gate’s not so big on the actual code review, at least it wasn’t when I started. We did an entire product release and then somebody looked over all of the UI of that product which I’d written and say what they didn’t like. By that point, it was way too late and I’d disagree with them. Do you think the lack of code reviews was a bad thing? I think if there’s going to be any oversight of new people, then it should be continuous rather than chunky. For me I don’t mind too much, I could go out and get oversight if I wanted it, and in those situations I felt comfortable without it. If I was managing the new person, then maybe I’d be keener on oversight and then the right way to do it is continuously and in very, very small chunks. Have you had any significant projects you’ve worked on outside of a job? When I was a teenager I wrote all sorts of stuff. I used to write games, I derived how to do isomorphic projections myself once. I didn’t know what the word was so I couldn’t Google for it, so I worked it out myself. It was horrifically complicated. But it sort of tailed off when I started at university, and is now basically zero. If I do side-projects now, they tend to be work-related side projects like my actors framework, NAct, which I started in a down tools week. Could you explain a little more about NAct? It is a little C# framework for writing parallel code more easily. Parallel programming is difficult when you need to write to shared data. Sometimes parallel programming is easy because you don’t need to write to shared data. When you do need to access shared data, you could just have your threads pile in and do their work, but then you would screw up the data because the threads would trample on each other’s toes. You could lock, but locks are really dangerous if you’re using more than one of them. You get interactions like deadlocks, and that’s just nasty. Actors instead allows you to say this piece of data belongs to this thread of execution, and nobody else can read it. If you want to read it, then ask that thread of execution for a piece of it by sending a message, and it will send the data back by a message. And that avoids deadlocks as long as you follow some obvious rules about not making your actors sit around waiting for other actors to do something. There are lots of ways to write actors, NAct allows you to do it as if it was method calls on other objects, which means you get all the strong type-safety that C# programmers like. Do you think that this is suitable for the majority of parallel programming, or do you think it’s only suitable for specific cases? It’s suitable for most difficult parallel programming. If you’ve just got a hundred web requests which are all independent of each other, then I wouldn’t bother because it’s easier to just spin them up in separate threads and they can proceed independently of each other. But where you’ve got difficult parallel programming, where you’ve got multiple threads accessing multiple bits of data in multiple ways at different times, then actors is at least as good as all other ways, and is, I reckon, easier to think about. When you’re using actors, you presumably still have to write your code in a different way from you would otherwise using single-threaded code. You can’t use actors with any methods that have return types, because you’re not allowed to call into another actor and wait for it. If you want to get a piece of data out of another actor, then you’ve got to use tasks so that you can use “async” and “await” to await asynchronously for it. But other than that, you can still stick things in classes so it’s not too different really. Rather than having thousands of objects with mutable state, you can use component-orientated design, where there are only a few mutable classes which each have a small number of instances. Then there can be thousands of immutable objects. If you tend to do that anyway, then actors isn’t much of a jump. If I’ve already built my system without any parallelism, how hard is it to add actors to exploit all eight cores on my desktop? Usually pretty easy. If you can identify even one boundary where things look like messages and you have components where some objects live on one side and these other objects live on the other side, then you can have a granddaddy object on one side be an actor and it will parallelise as it goes across that boundary. Not too difficult. If we do get 1000-core desktop PCs, do you think actors will scale up? It’s hard. There are always in the order of twenty to fifty actors in my whole program because I tend to write each component as actors, and I tend to have one instance of each component. So this won’t scale to a thousand cores. What you can do is write data structures out of actors. I use dictionaries all over the place, and if you need a dictionary that is going to be accessed concurrently, then you could build one of those out of actors in no time. You can use queuing to marshal requests between different slices of the dictionary which are living on different threads. So it’s like a distributed hash table but all of the chunks of it are on the same machine. That means that each of these thousand processors has cached one small piece of the dictionary. I reckon it wouldn’t be too big a leap to start doing proper parallelism. Do you think it helps if actors get baked into the language, similarly to Erlang? Erlang is excellent in that it has thread-local garbage collection. C# doesn’t, so there’s a limit to how well C# actors can possibly scale because there’s a single garbage collected heap shared between all of them. When you do a global garbage collection, you’ve got to stop all of the actors, which is seriously expensive, whereas in Erlang garbage collections happen per-actor, so they’re insanely cheap. However, Erlang deviated from all the sensible language design that people have used recently and has just come up with crazy stuff. You can definitely retrofit thread-local garbage collection to .NET, and then it’s quite well-suited to support actors, even if it’s not baked into the language. Speaking of language design, do you have a favourite programming language? I’ll choose a language which I’ve never written before. I like the idea of Scala. It sounds like C#, only with some of the niggles gone. I enjoy writing static types. It means you don’t have to writing tests so much. When you say it doesn’t have some of the niggles? C# doesn’t allow the use of a property as a method group. It doesn’t have Scala case classes, or sum types, where you can do a switch statement and the compiler checks that you’ve checked all the cases, which is really useful in functional-style programming. Pattern-matching, in other words. That’s actually the major niggle. C# is pretty good, and I’m quite happy with C#. And what about going even further with the type system to remove the need for tests to something like Haskell? Or is that a step too far? I’m quite a pragmatist, I don’t think I could deal with trying to write big systems in languages with too few other users, especially when learning how to structure things. I just don’t know anyone who can teach me, and the Internet won’t teach me. That’s the main reason I wouldn’t use it. If I turned up at a company that writes big systems in Haskell, I would have no objection to that, but I wouldn’t instigate it. What about things in C#? For instance, there’s contracts in C#, so you can try to statically verify a bit more about your code. Do you think that’s useful, or just not worthwhile? I’ve not really tried it. My hunch is that it needs to be built into the language and be quite mathematical for it to work in real life, and that doesn’t seem to have ended up true for C# contracts. I don’t think anyone who’s tried them thinks they’re any good. I might be wrong. On a slightly different note, how do you like to debug code? I think I’m quite an odd debugger. I use guesswork extremely rarely, especially if something seems quite difficult to debug. I’ve been bitten spending hours and hours on guesswork and not being scientific about debugging in the past, so now I’m scientific to a fault. What I want is to see the bug happening in the debugger, to step through the bug happening. To watch the program going from a valid state to an invalid state. When there’s a bug and I can’t work out why it’s happening, I try to find some piece of evidence which places the bug in one section of the code. From that experiment, I binary chop on the possible causes of the bug. I suppose that means binary chopping on places in the code, or binary chopping on a stage through a processing cycle. Basically, I’m very stupid about how I debug. I won’t make any guesses, I won’t use any intuition, I will only identify the experiment that’s going to binary chop most effectively and repeat rather than trying to guess anything. I suppose it’s quite top-down. Is most of the time then spent in the debugger? Absolutely, if at all possible I will never debug using print statements or logs. I don’t really hold much stock in outputting logs. If there’s any bug which can be reproduced locally, I’d rather do it in the debugger than outputting logs. And with SmartAssembly error reporting, there’s not a lot that can’t be either observed in an error report and just fixed, or reproduced locally. And in those other situations, maybe I’ll use logs. But I hate using logs. You stare at the log, trying to guess what’s going on, and that’s exactly what I don’t like doing. You have to just look at it and see does this look right or wrong. We’ve covered how you get to grip with bugs. How do you get to grips with an entire codebase? I watch it in the debugger. I find little bugs and then try to fix them, and mostly do it by watching them in the debugger and gradually getting an understanding of how the code works using my process of binary chopping. I have to do a lot of reading and watching code to choose where my slicing-in-half experiment is going to be. The last time I did it was SmartAssembly. The old code was a complete mess, but at least it did things top to bottom. There wasn’t too much of some of the big abstractions where flow of control goes all over the place, into a base class and back again. Code’s really hard to understand when that happens. So I like to choose a little bug and try to fix it, and choose a bigger bug and try to fix it. Definitely learn by doing. I want to always have an aim so that I get a little achievement after every few hours of debugging. Once I’ve learnt the codebase I might be able to fix all the bugs in an hour, but I’d rather be using them as an aim while I’m learning the codebase. If I was a maintainer of a codebase, what should I do to make it as easy as possible for you to understand? Keep distinct concepts in different places. And name your stuff so that it’s obvious which concepts live there. You shouldn’t have some variable that gets set miles up the top of somewhere, and then is read miles down to choose some later behaviour. I’m talking from a very much SmartAssembly point of view because the old SmartAssembly codebase had tons and tons of these things, where it would read some property of the code and then deal with it later. Just thousands of variables in scope. Loads of things to think about. If you can keep concepts separate, then it aids me in my process of fixing bugs one at a time, because each bug is going to more or less be understandable in the one place where it is. And what about tests? Do you think they help at all? I’ve never had the opportunity to learn a codebase which has had tests, I don’t know what it’s like! What about when you’re actually developing? How useful do you find tests in finding bugs or regressions? Finding regressions, absolutely. Running bits of code that would be quite hard to run otherwise, definitely. It doesn’t happen very often that a test finds a bug in the first place. I don’t really buy nebulous promises like tests being a good way to think about the spec of the code. My thinking goes something like “This code works at the moment, great, ship it! Ah, there’s a way that this code doesn’t work. Okay, write a test, demonstrate that it doesn’t work, fix it, use the test to demonstrate that it’s now fixed, and keep the test for future regressions.” The most valuable tests are for bugs that have actually happened at some point, because bugs that have actually happened at some point, despite the fact that you think you’ve fixed them, are way more likely to appear again than new bugs are. Does that mean that when you write your code the first time, there are no tests? Often. The chance of there being a bug in a new feature is relatively unaffected by whether I’ve written a test for that new feature because I’m not good enough at writing tests to think of bugs that I would have written into the code. So not writing regression tests for all of your code hasn’t affected you too badly? There are different kinds of features. Some of them just always work, and are just not flaky, they just continue working whatever you throw at them. Maybe because the type-checker is particularly effective around them. Writing tests for those features which just tend to always work is a waste of time. And because it’s a waste of time I’ll tend to wait until a feature has demonstrated its flakiness by having bugs in it before I start trying to test it. You can get a feel for whether it’s going to be flaky code as you’re writing it. I try to write it to make it not flaky, but there are some things that are just inherently flaky. And very occasionally, I’ll think “this is going to be flaky” as I’m writing, and then maybe do a test, but not most of the time. How do you think your programming style has changed over time? I’ve got clearer about what the right way of doing things is. I used to flip-flop a lot between different ideas. Five years ago I came up with some really good ideas and some really terrible ideas. All of them seemed great when I thought of them, but they were quite diverse ideas, whereas now I have a smaller set of reliable ideas that are actually good for structuring code. So my code is probably more similar to itself than it used to be back in the day, when I was trying stuff out. I’ve got more disciplined about encapsulation, I think. There are operational things like I use actors more now than I used to, and that forces me to use immutability more than I used to. The first code that I wrote in Red Gate was the memory profiler UI, and that was an actor, I just didn’t know the name of it at the time. I don’t really use object-orientation. By object-orientation, I mean having n objects of the same type which are mutable. I want a constant number of objects that are mutable, and they should be different types. I stick stuff in dictionaries and then have one thing that owns the dictionary and puts stuff in and out of it. That’s definitely a pattern that I’ve seen recently. I think maybe I’m doing functional programming. Possibly. It’s plausible. If you had to summarise the essence of programming in a pithy sentence, how would you do it? Programming is the form of art that, without losing any of the beauty of architecture or fine art, allows you to produce things that people love and you make money from. So you think it’s an art rather than a science? It’s a little bit of engineering, a smidgeon of maths, but it’s not science. Like architecture, programming is on that boundary between art and engineering. If you want to do it really nicely, it’s mostly art. You can get away with doing architecture and programming entirely by having a good engineering mind, but you’re not going to produce anything nice. You’re not going to have joy doing it if you’re an engineering mind. Architects who are just engineering minds are not going to enjoy their job. I suppose engineering is the foundation on which you build the art. Exactly. How do you think programming is going to change over the next ten years? There will be an unfortunate shift towards dynamically-typed languages, because of JavaScript. JavaScript has an unfair advantage. JavaScript’s unfair advantage will cause more people to be exposed to dynamically-typed languages, which means other dynamically-typed languages crop up and the best features go into dynamically-typed languages. Then people conflate the good features with the fact that it’s dynamically-typed, and more investment goes into dynamically-typed languages. They end up better, so people use them. What about the idea of compiling other languages, possibly statically-typed, to JavaScript? It’s a reasonable idea. I would like to do it, but I don’t think enough people in the world are going to do it to make it pick up. The hordes of beginners are the lifeblood of a language community. They are what makes there be good tools and what makes there be vibrant community websites. And any particular thing which is the same as JavaScript only with extra stuff added to it, although it might be technically great, is not going to have the hordes of beginners. JavaScript is always to be quickest and easiest way for a beginner to start programming in the browser. And dynamically-typed languages are great for beginners. Compilers are pretty scary and beginners don’t write big code. And having your errors come up in the same place, whether they’re statically checkable errors or not, is quite nice for a beginner. If someone asked me to teach them some programming, I’d teach them JavaScript. If dynamically-typed languages are great for beginners, when do you think the benefits of static typing start to kick in? The value of having a statically typed program is in the tools that rely on the static types to produce a smooth IDE experience rather than actually telling me my compile errors. And only once you’re experienced enough a programmer that having a really smooth IDE experience makes a blind bit of difference, does static typing make a blind bit of difference. So it’s not really about size of codebase. If I go and write up a tiny program, I’m still going to get value out of writing it in C# using ReSharper because I’m experienced with C# and ReSharper enough to be able to write code five times faster if I have that help. Any other visions of the future? Nobody’s going to use actors. Because everyone’s going to be running on single-core VMs connected over network-ready protocols like JSON over HTTP. So, parallelism within one operating system is going to die. But until then, you should use actors. More Red Gater Coder interviews

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