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• ExplorerCanvas and JQuery

  - by PhubarBaz

  I am working on a Javascript app (CloudGraph) that uses HTML5 canvas and JQuery. I'm using ExplorerCanvas to support canvas in IE. I recently came across an interesting problem. What I was trying to do is restore the user's settings when the page is loaded. I read some information from a cookie that I set the last time the user accessed the application. One of these settings is the size of the canvas. I decided that the best place to do this would be when the document is ready using JQuery $(document).ready(). This worked fine in browsers that natively support the canvas element. But in IE it kept getting errors the first time I would hit the page. It seemed that the excanvas element wasn't initialized yet because I was getting null reference and unknown properties errors. If I refreshed the page the errors went away but the resized canvas wasn't drawing on the entire area of the canvas. It was like the clipping rectangle was still set to the default canvas size. I found that the canvas element when using excanvas has a div child element which is where the actual drawing takes place. When I changed the width and height of the canvas element in document.ready it didn't change the width and height of the child div. Initially my solution was to also change the div element when changing the canvas element and that worked. But then I realized that having to refresh the page every time I started the app in IE really sucked. That wouldn't be acceptable for users. Since it seemed like the canvas wasn't getting initialized before I was trying to use it I decided to try to initialize my app at a different time. I figured the next best place was in the onload event. Sure enough, moving my initialization to onload fixed all of the problems. So, it looks like the canvas shouldn't be manipulated until the onload event when using ExplorerCanvas. There might be ways to do it when the document is ready. I found some posts on initializing excanvas manually, but for me waiting until onload worked just fine.

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• Should we persist with an employee still writing bad code after many years?

  - by user94986

  I've been assigned the task of managing developers for a well-established company. They have a single developer who specialises in all their C++ coding (since forever), but the quality of the work is abysmal. Code reviews and testing have revealed many problems, one of the worst being memory leaks. The developer has never tested his code for leaks, and I discovered that the applications could leak many MBs with only a minute of use. User's were reporting huge slowdowns, and his take was, "it's nothing to do with me - if they quit and restart, it's all good again." I've given him tools to detect and trace the leaks, and sat down with him for many hours to demonstrate how the tools are used, where the problems occur, and what to do to fix them. We're 6 months down the track, and I assigned him to write a new module. I reviewed it before it was integrated into our larger code base, and was dismayed to discover the same bad coding as before. The part that I find incomprehensible is that some of the coding is worse than amateurish. For example, he wanted a class (Foo) that could populate an object of another class (Bar). He decided that Foo would hold a reference to Bar, e.g.: class Foo { public: Foo(Bar& bar) : m_bar(bar) {} private: Bar& m_bar; }; But (for other reasons) he also needed a default constructor for Foo and, rather than question his initial design, he wrote this gem: Foo::Foo() : m_bar(*(new Bar)) {} So every time the default constructor is called, a Bar is leaked. To make matters worse, Foo allocates memory from the heap for 2 other objects, but he didn't write a destructor or copy constructor. So every allocation of Foo actually leaks 3 different objects, and you can imagine what happened when a Foo was copied. And - it only gets better - he repeated the same pattern on three other classes, so it isn't a one-off slip. The whole concept is wrong on so many levels. I would feel more understanding if this came from a total novice. But this guy has been doing this for many years and has had very focussed training and advice over the past few months. I realise he has been working without mentoring or peer reviews most of that time, but I'm beginning to feel he can't change. So my question is, would you persist with someone who is writing such obviously bad code?

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• Flattening System.Web.UI ControlCollection

  - by evovision

  Hi,   Sometimes one may need to get a list of child controls inside specific container and don't care about the underlying hierarchy.   The result is beautifully achieved using this extension method:   using System;using System.Collections.Generic;using System.Linq;using System.Web;using System.Web.UI;    public static class ControlCollectionExtensionMethods    {        public static IEnumerable<Control> FlattenedList(this ControlCollection controls)        {            foreach (Control ctrl in controls)            {                  // return parent control                   yield return ctrl;                              // and dive into child collection                   foreach (Control child in ctrl.Controls.FlattenedList())                         yield return child;            }        }    }   P.S.: don't forget about namespaces when using it in your code, if above class is wrapped into namespace, for example: Sample, the source code file with calling code must explicitly reference it: using Sample;

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• Getting my younger brother started on programming

  - by SmartLemon

  My younger brother is 13 years old, I started programming when I started to develop Android applications when I was 15, last year my brother gained an interest in it and he would always pestering me about letting him make something himself, so I wrote him a few tutorials and he built himself a small application that had a few buttons that did something, I think you put in your dob and it would tell you what day you were born on, he took a couple of days building up to his final application, maybe even a week, learning everything he needed. Since then he hasn't really done much more because I have been engulfed in work and such where I have my own programming problems to sort out. I told him that when he was my age (I am 17) that he should be better then me, he was a bit sceptical about this however. I dont think he has as much logical reasoning as I would think he needs to solve more complex problems, but shouldnt that just develop over time as it did with me? He has been pestering me for the past week or something to write him more tutorials, but I didn't have time. All I had with me was a playlist I had downloaded from the new boston from youtube for C++, it's about 73 videos. He is currently about 20-30 videos in, he has come to ask me a few questions about it and thats it. Should I have really properly started him with C++? Should I stop him now and start him again on python or ruby? I know that C++ shouldn't really be a beginners language, especially for someone who is only 13, by the time this question is answered will probably be up to learning about inheritance or something. Some people may see this as not a real question, but it is, and should be used as a reference for others. I want to know, should I start him on a different language whch is more easy? What language then? And would it be better for me to teach him myself (I would make time) or just continue him with the new boston? There are a few more questions throughout this question but these are the main ones. Part of the question people seem to be neglecting is me asking whether I should change what language he is learning to another, or since he is already pretty far through the tutorials should I just leave him with C++ and he can learn the other languages freely by himself?

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• How can I improve the battery life under 12.04 on my Inspiron 14z? [duplicate]

  - by cfogelberg

  This question already has an answer here: Tips to extend battery life for laptops and notebooks 24 answers How do I improve the battery life of my Inspiron 14z under Ubuntu 12.04? This laptop gets 4-5 hours of battery life using Windows (e.g. here). I've removed Windows, installed Ubuntu 12.04 and the initial battery life was only 2 hours. With some tweaks (described below) it's still only ~2.5 hours. For reference, the laptop is the latest model of the 14z: i5-3337U processor 32GB MSATA, 500GB HDD (5400rpm) AMD Radeon HD7570M graphics card I have put ext4 partitions on both the SSD and the HDD, and have mounted / to the SSD and /home to the HDD. I also put a 24gb linux swap partition at the start of the HDD, though I figure this won't be used all that much (the laptop has 8gb of RAM). After googling around and reading Ask Ubuntu and other sites extensively, I have done the following steps, and they have improved the battery life ~30 minutes (exact improvement not clear, but battery life is still nowhere near 4-5 hours). Installed Jupiter (and set Performance to "Power Saving") Installed laptop-mode-tools cat /proc/sys/vm/laptop_mode now outputs 5 (previously it output 0) But it's not clear that this will help: AskUbuntu question Turned down the brightness of my screen from full to 1/3 Other things I have heard about but have not tried for fear of frying the laptop or my linux install: Add "pcie_aspm=force" at the end of the line with "quiet splash" in /boot/grub/grub.cfg Enable ALPM, but it may already be enabled in 12.04? Enable i915 framebuffer compression Use a propietary driver for the graphics card? Turn off the graphics card? (what would happen if I relied on the internal Intel bridge?) Use TLP? Spin down the HDD more aggressively (howto, but I think laptop-mode-tools does this already) The only other thing I've noticed is that plastic just above the F5, F6 and F7 keys gets really hot. According to Jupiter my CPU temperature is only 69 celsius and the System Monitor shows CPU load at 7% so I don't think it's the CPU. Maybe it's the graphics card? Also, I've set up MongoDB and LAMP on the machine as well. When I run powertop MongoDB is high in the list, but I'm not sure if that's relevant to battery life because I'm not actually doing anything with MongoDB most of the time. Edit - Additional info as requested $ lspci -nnk | grep -iEA3 "(graphics|vga)" 00:02.0 VGA compatible controller [0300]: Intel Corporation Ivy Bridge Graphics Controller [8086:0166] (rev 09) Subsystem: Dell Device [1028:057f] Kernel driver in use: i915 Kernel modules: i915 -- 02:00.0 VGA compatible controller [0300]: Advanced Micro Devices [AMD] nee ATI Thames [Radeon 7500M/7600M Series] [1002:6841] Subsystem: Dell Device [1028:057f] Kernel driver in use: radeon Kernel modules: radeon

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• Recommended learning path?

  - by stairmast0r

  First, my current standing: I know C++ at an.. advanced beginner level? I've gone through a book, I know the syntax well enough, I know a fair amount of standard library functions, and I've programmed some simple console stuff with it. I'd probably be able to program more with it if I knew how to structure a program, but I just can't seem to wrap my head around the whole concept of structuring something remotely complex. I've messed around with Java for a day or two, and the syntax was extremely easy to get the hang of, except that I didn't really know any functions. I'm plenty willing to learn, and to work hard to do so, but I don't really know where to go from here. Now, at the risk of sounding cliche, what I'd like to become is someone like the great three of id; Carmack, Romero, and Abrash. To be considered a genius. I believe anything can be learned, and nothing mentally limits anyone except lack of desire to learn. But I don't know how to learn this. They learned by doing, and making do with what resources they had. On the other hand, I have access to almost any books I want, access to the internet, and access to a more than capable computer and software. Should I learn more languages? Assembly? LISP? BASIC? Haskell? Should I dive straight into advanced topics like OpenGL? Or should I wait until I feel I've come closer to mastering the simpler things, like console programs, first? Should I follow tutorials? Should I follow books? Should I just dive into writing something and follow a reference manual as I go? What order should I do all this in? How should I do it? I want to completely master this; to be considered a genius. The most perfect career I can imagine is to start the next id. I have the drive to do it, I just don't know where to begin...

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• Silverlight 4 Twitter Client &ndash; Part 5

  - by Max

  In this post, we will look at implementing the following in our twitter client - displaying the profile picture in the home page after logging in. So to accomplish this, we will need to do the following steps. 1) First of all, let us create another field in our Global variable static class to hold the profile image url. So just create a static string field in that. public static string profileImage { get; set; } 2) In the Login.xaml.cs file, before the line where we redirect to the Home page if the login was successful. Create another WebClient request to fetch details from this url - .xml">http://api.twitter.com/1/users/show/<TwitterUsername>.xml I am interested only in the “profile_image_url” node in this xml string returned. You can choose what ever field you need – the reference to the Twitter API Documentation is here. 3) This particular url does not require any authentication so no need of network credentials for this and also the useDefaultCredentials will be true. So this code would look like. WebRequest.RegisterPrefix("http://", System.Net.Browser.WebRequestCreator.ClientHttp); WebClient myService = new WebClient(); myService.AllowReadStreamBuffering = myService.UseDefaultCredentials = true; myService.DownloadStringCompleted += new DownloadStringCompletedEventHandler(TimelineRequestCompleted_User);) myService.DownloadStringAsync(new Uri("http://api.twitter.com/1/users/show/" + TwitterUsername.Text + ".xml")); 4) To parse the xml file returned by this and get the value of the “profile_image_url” node, the LINQ code will look something like XDocument xdoc; xdoc = XDocument.Parse(e.Result); var answer = (from status in xdoc.Descendants("user") select status.Element("profile_image_url").Value); GlobalVariable.profileImage = answer.First().ToString(); 5) After this, we can then redirect to Home page completing our login formalities. 6) In the home page, add a image tag and give it a name, something like <Image Name="image1" Stretch="Fill" Margin="29,29,0,0" Height="73" VerticalAlignment="Top" HorizontalAlignment="Left" Width="73" /> 7) In the OnNavigated to event in home page, we can set the image source url as below image1.Source = new BitmapImage(new Uri(GlobalVariable.profileImage, UriKind.Absolute)); Hope this helps. If you have any queries / suggestions, please feel free to comment below or contact me. Technorati Tags: Silverlight,LINQ,Twitter API,Twitter

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• Application workflow

  - by manseuk

  I am in the planning process for a new application, the application will be written in PHP (using the Symfony 2 framework) but I'm not sure how relevant that is. The application will be browser based, although there will eventually be API access for other systems to interact with the data stored within the application, again probably not relavent at this point. The application manages SIM cards for lots of different providers - each SIM card belongs to a single provider but a single customer might have many SIM cards across many providers. The application allows the user to perform actions against the SIM card - for example Activate it, Barr it, Check on its status etc Some of the providers provide an API for doing this - so a single access point with multiple methods eg activateSIM, getStatus, barrSIM etc. The method names differ for each provider and some providers offer methods for extra functions that others don't. Some providers don't have APIs but do offer these methods by sending emails with attachments - the attachments are normally a CSV file that contains the SIM reference and action required - the email is processed by the provider and replied to once the action has been complete. To give you an example - the front end of my application will provide a customer with a list of SIM cards they own and give them access to the actions that are provided by the provider of each specific SIM card - some methods may require extra data which will either be stored in the backend or collected from the user frontend. Once the user has selected their action and added any required data I will handle the process in the backend and provide either instant feedback, in the case of the providers with APIs, or start the process off by sending an email and waiting for its reply before processing it and updating the backend so that next time the user checks the SIM card its status is correct (ie updated by a backend process). My reason for creating this question is because I'm stuck !! I'm confused about how to approach the actual workflow logic. I was thinking about creating a Provider Interface with the most common methods getStatus, activateSIM and barrSIM and then implementing that interface for each provider. So class Provider1 implements Provider - Then use a Factory to create the required class depending on user selected SIM card and invoking the method selected. This would work fine if all providers offered the same methods but they don't - there are a subset which are common but some providers offer extra methods - how can I implement that flexibly ? How can I deal with the processes where the workflow is different - ie some methods require and API call and value returned and some require an email to be sent and the next stage of the process doesn't start until the email reply is recieved ... Please help ! (I hope this is a readable question and that this is the correct place to be asking) Update I guess what I'm trying to avoid is a big if or switch / case statement - some design pattern that gives me a flexible approach to implementing this kind of fluid workflow .. anyone ?

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• [EF + Oracle] Inserting Data (Sequences) (2/2)

  - by JTorrecilla

  Prologue In the previous chapter we have see how to create DB records with EF, now we are going to Some Questions about Oracle.   ORACLE One characteristic from SQL Server that differs from Oracle is “Identity”. To all that has not worked with SQL Server, this property, that applies to Integer Columns, lets indicate that there is auto increment columns, by that way it will be filled automatically, without writing it on the insert statement. In EF with SQL Server, the properties whose match with Identity columns, will be filled after invoking SaveChanges method. In Oracle DB, there is no Identity Property, but there is something similar. Sequences Sequences are DB objects, that allow to create auto increment, but there are not related directly to a Table. The syntax is as follows: name, min value, max value and begin value. 1: CREATE SEQUENCE nombre_secuencia 2: INCREMENT BY numero_incremento 3: START WITH numero_por_el_que_empezara 4: MAXVALUE valor_maximo | NOMAXVALUE 5: MINVALUE valor_minimo | NOMINVALUE 6: CYCLE | NOCYCLE 7: ORDER | NOORDER 8:    How to get sequence value? To obtain the next val from the sequence: 1: SELECT nb_secuencia.Nextval 2: From Dual Due to there is no direct way to indicate that a column is related to a sequence, there is several ways to imitate the behavior: Use a Trigger (DB), Use Stored Procedures or Functions(…) or my particularly option. EF model, only, imports Table Objects, Stored Procedures or Functions, but not sequences. By that, I decide to create my own extension Method to invoke Next Val from a sequence: 1: public static class EFSequence 2: { 3: public static int GetNextValue(this ObjectContext contexto, string SequenceName) 4: { 5: string Connection = ConfigurationManager.ConnectionStrings["JTorrecillaEntities2"].ConnectionString; 6: Connection=Connection.Substring(Connection.IndexOf(@"connection string=")+19); 7: Connection = Connection.Remove(Connection.Length - 1, 1); 8: using (IDbConnection con = new Oracle.DataAccess.Client.OracleConnection(Connection)) 9: { 10: using (IDbCommand cmd = con.CreateCommand()) 11: { 12: con.Open(); 13: cmd.CommandText = String.Format("Select {0}.nextval from DUAL", SequenceName); 14: return Convert.ToInt32(cmd.ExecuteScalar()); 15: } 16: } 17:  18: } 19: } This Object Context’s extension method are going to invoke a query with the Sequence indicated by parameter. It takes the connection strings from the App settings, removing the meta data, that was created by VS when generated the EF model. And then, returns the next value from the Sequence. The next value of a Sequence is unique, by that, when some concurrent users are going to create records in the DB using the sequence will not get duplicates. This is my own implementation, I know that it could be several ways to do and better ways. If I find any other way, I promise to post it. To use the example is needed to add a reference to the Oracle (ODP.NET) dll.

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• Sql Table Refactoring Challenge

  Ive been working a bit on cleaning up a large table to make it more efficient.  I pretty much know what I need to do at this point, but I figured Id offer up a challenge for my readers, to see if they can catch everything I have as well as to see if Ive missed anything.  So to that end, I give you my table: CREATE TABLE [dbo].[lq_ActivityLog]( [ID] [bigint] IDENTITY(1,1) NOT NULL, [PlacementID] [int] NOT NULL, [CreativeID] [int] NOT NULL, [PublisherID] [int] NOT NULL, [CountryCode] [nvarchar](10) NOT NULL, [RequestedZoneID] [int] NOT NULL, [AboveFold] [int] NOT NULL, [Period] [datetime] NOT NULL, [Clicks] [int] NOT NULL, [Impressions] [int] NOT NULL, CONSTRAINT [PK_lq_ActivityLog2] PRIMARY KEY CLUSTERED ( [Period] ASC, [PlacementID] ASC, [CreativeID] ASC, [PublisherID] ASC, [RequestedZoneID] ASC, [AboveFold] ASC, [CountryCode] ASC)WITH (PAD_INDEX = OFF, STATISTICS_NORECOMPUTE = OFF, IGNORE_DUP_KEY = OFF, ALLOW_ROW_LOCKS = ON, ALLOW_PAGE_LOCKS = ON) ON [PRIMARY]) ON [PRIMARY] And now some assumptions and additional information: The table has 200,000,000 rows currently PlacementID ranges from 1 to 5000 and should support at least 50,000 CreativeID ranges from 1 to 5000 and should support at least 50,000 PublisherID ranges from 1 to 500 and should support at least 50,000 CountryCode is a 2-character ISO standard (e.g. US) and there is a country table with an integer ID already.  There are < 300 rows. RequestedZoneID ranges from 1 to 100 and should support at least 50,000 AboveFold has values of 1, 0, or 1 only. Period is a date (no time). Clicks range from 0 to 5000. Impressions range from 0 to 5000000. The table is currently write-mostly.  Its primary purpose is to log advertising activity as quickly as possible.  Nothing in the rest of the system reads from it except for batch jobs that pull the data into summary tables. Heres the current information on the database tables size: Design Goals This table has been in use for about 5 years and has performed very well during that time.  The only complaints we have are that it is quite large and also there are occasionally timeouts for queries that reference it, particularly when batch jobs are pulling data from it.  Any changes should be made with an eye toward keeping write performance optimal  while trying to reduce space and improve read performance / eliminate timeouts during read operations. Refactor There are, I suggest to you, some glaringly obvious optimizations that can be made to this table.  And Im sure there are some ninja tweaks known to SQL gurus that would be a big help as well.  Ill post my own suggested changes in a follow-up post for now feel free to comment with your suggestions. Did you know that DotNetSlackers also publishes .net articles written by top known .net Authors? We already have over 80 articles in several categories including Silverlight. Take a look: here.

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• ??????WLST

  - by Masa Sasaki

  WebLogic Server?????????????WebLogic Server????????7?23?????????38?WebLogic Server???@??????????????WLST?(?????? ??????????? ?? ??)??????????????WLST(WebLogic Scripting Tool)?WebLogic Server???????????????????????????????????WLST?????TIPS??????????????????????WLST????????????????????????????????????????????·????????????(?????? Fusion Middleware?????? ??? ??) WLST?? WebLogic Scripting Tool???WebLogic ??????????????????????????·?????????? WebLogic Server 9.x???????????????????Java??????·?????????Jython??????????? WLST???????????????????????????Jython??????WebLogic Server??????????????(????)????????WLST?Jython??????????????????????????? WLST???????? WLST?????????????????????????WebLogic????????????????????????????????·????????????????·??????(???????????)???????? (WLST????????????????????????????????????????????) WLST??????Java Management Extensions (JMX)???????????????JMX??????????????????????????????????Bean (MBean)??????????????????????? WLST???????? WLST??????WebLogic Server?????????????????·??????????????????????????????????????????????????????????????????????????????????????????????????????????WLST??????WebLogic Server??????????JMX????????????????????????????????WebLogic???????????????WLST????????????????????? ???????????·???????????? ????? ???????????·???·??????WLST?????????????????????????????????????????????????????????????????????????????????????????????????????WLST??????????????????????????????????????????? ?????·??? ???.py????????·????(?????filename.py)?WLST?????????????????????????????WLST???????????????????????????·?????Jython??????????????????????? ?????? ?????????Java?????WLST??????????????????????WLST????????????????????????????????·????????WLST???????????????????????????? WLST?MBean WebLogic Server?????(?????????)?JMX(Java Management Extension)?????????????JMX???????????????Bean(MBean)?????? MBean?????? MBean????????????????????????????????MBean???????????????????????????????????????????????????????????????????? WLST????????? ? ?????? ? ???????????(MBean???)??? ? ?????????? - ??????? - ?????????? - ?? - ?????? WLST???? ??·??·?????WLST???????????????????????????????????????????WLST??????????????? TIPS?MBean???? TIPS?????????????????WLST????????????????????????????????????????????(SSL)?????????????????????????????????SSL??????????????????????????????????????????????MBean????????WLST ls?????WLST find?????JRockit Mission Control?config.xml???????????????????WebLogic Server MBean Reference????????????????? ??? WLST??????????????????????????????????????????????????????????????????ThreadPoolRuntimeMBean??????JMS?????????WLST??????????????? ??????????????????????????? WLST ????????? $WL_HOME/samples/server/examples/src/examples/wlst/online WLST????????????? $WL_HOME/common/templates/scripts/wlst ???????????????????????????????????????????????????????????????????WLST???????????????????·??????????WLST????????????????????? ?????? WebLogic Server??? WebLogic Server?????????WebLogic Server?????! WebLogic Server??????(???????????) WebLogic Server???????? WebLogic Server??????

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• git | error: Unable to append to .git/logs/refs/remotes/origin/master: Permission denied [SOLVED]

  - by Corbin Tarrant

  I am having a strange issue that I can't seem to resolve. Here is what happend: I had some log files in a github repository that I didn't want there. I found this script that removes files completely from git history like so: #!/bin/bash set -o errexit # Author: David Underhill # Script to permanently delete files/folders from your git repository. To use # it, cd to your repository's root and then run the script with a list of paths # you want to delete, e.g., git-delete-history path1 path2 if [ $# -eq 0 ]; then exit 0are still fi # make sure we're at the root of git repo if [ ! -d .git ]; then echo "Error: must run this script from the root of a git repository" exit 1 fi # remove all paths passed as arguments from the history of the repo files=$@ git filter-branch --index-filter "git rm -rf --cached --ignore-unmatch $files" HEAD # remove the temporary history git-filter-branch otherwise leaves behind for a long time rm -rf .git/refs/original/ && git reflog expire --all && git gc --aggressive --prune I, of course, made a backup first and then tried it. It seemed to work fine. I then did a git push -f and was greeted with the following messages: error: Unable to append to .git/logs/refs/remotes/origin/master: Permission denied error: Cannot update the ref 'refs/remotes/origin/master'. Everything seems to have pushed fine though, because the files seem to be gone from the GitHub repository, if I try and push again I get the same thing: error: Unable to append to .git/logs/refs/remotes/origin/master: Permission denied error: Cannot update the ref 'refs/remotes/origin/master'. Everything up-to-date EDIT $ sudo chgrp {user} .git/logs/refs/remotes/origin/master $ sudo chown {user} .git/logs/refs/remotes/origin/master $ git push Everything up-to-date Thanks! EDIT Uh Oh. Problem. I've been working on this project all night and just went to commit my changes: error: Unable to append to .git/logs/refs/heads/master: Permission denied fatal: cannot update HEAD ref So I: sudo chown {user} .git/logs/refs/heads/master sudo chgrp {user} .git/logs/refs/heads/master I try the commit again and I get: error: Unable to append to .git/logs/HEAD: Permission denied fatal: cannot update HEAD ref So I: sudo chown {user} .git/logs/HEAD sudo chgrp {user} .git/logs/HEAD And then I try the commit again: 16 files changed, 499 insertions(+), 284 deletions(-) create mode 100644 logs/DBerrors.xsl delete mode 100644 logs/emptyPHPerrors.php create mode 100644 logs/trimXMLerrors.php rewrite public/codeCore/Classes/php/DatabaseConnection.php (77%) create mode 100644 public/codeSite/php/init.php $ git push Counting objects: 49, done. Delta compression using up to 2 threads. Compressing objects: 100% (27/27), done. Writing objects: 100% (27/27), 7.72 KiB, done. Total 27 (delta 15), reused 0 (delta 0) To [email protected]:IAmCorbin/MooKit.git 59da24e..68b6397 master -> master Hooray. I jump on http://GitHub.com and check out the repository, and my latest commit is no where to be found. ::scratch head:: So I push again: Everything up-to-date Umm...it doesn't look like it. I've never had this issue before, could this be a problem with github? or did I mess something up with my git project? EDIT Nevermind, I did a simple: git push origin master and it pushed fine.

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• email-spec destroys my rake cucumber:all

  - by Leonardo Dario Perna

  This works fine: $ rake cucumber:all Then $ script/plugin install git://github.com/bmabey/email-spec.git remote: Counting objects: 162, done. remote: Compressing objects: 100% (130/130), done. remote: Total 162 (delta 18), reused 79 (delta 13) Receiving objects: 100% (162/162), 127.65 KiB | 15 KiB/s, done. Resolving deltas: 100% (18/18), done. From git://github.com/bmabey/email-spec * branch HEAD - FETCH_HEAD And $ script/generate email_spec exists features/step_definitions create features/step_definitions/email_steps.rb And I add 'require 'email_spec/cucumber' in /feature/support/env.rb so it looks somethinng like: require File.expand_path(File.dirname(__FILE__) + '/../../config/environment') require 'cucumber/rails/world' require 'cucumber/formatter/unicode' # Comment out this line if you don't want Cucumber Unicode support require 'email_spec/cucumber' and now: rake cucumber:all gives me this error: $ rake cucumber:all --trace (in /Users/leonardodarioperna/Projects/frestyl/frestyl) ** Invoke cucumber:all (first_time) ** Invoke cucumber:ok (first_time) ** Invoke db:test:prepare (first_time) ** Invoke db:abort_if_pending_migrations (first_time) ** Invoke environment (first_time) ** Execute environment ** Execute db:abort_if_pending_migrations ** Execute db:test:prepare ** Invoke db:test:load (first_time) ** Invoke db:test:purge (first_time) ** Invoke environment ** Execute db:test:purge ** Execute db:test:load ** Invoke db:schema:load (first_time) ** Invoke environment ** Execute db:schema:load ** Execute cucumber:ok /System/Library/Frameworks/Ruby.framework/Versions/1.8/usr/bin/ruby -I "/Library/Ruby/Gems/1.8/gems/cucumber-0.4.4/lib:lib" "/Library/Ruby/Gems/1.8/gems/cucumber-0.4.4/bin/cucumber" --profile default cucumber.yml was not found. Please refer to cucumber's documentation on defining profiles in cucumber.yml. You must define a 'default' profile to use the cucumber command without any arguments. Type 'cucumber --help' for usage. rake aborted! Command failed with status (1): [/System/Library/Frameworks/Ruby.framework/...] /Library/Ruby/Gems/1.8/gems/rake-0.8.7/lib/rake.rb:995:in `sh' /Library/Ruby/Gems/1.8/gems/rake-0.8.7/lib/rake.rb:1010:in `call' /Library/Ruby/Gems/1.8/gems/rake-0.8.7/lib/rake.rb:1010:in `sh' /Library/Ruby/Gems/1.8/gems/rake-0.8.7/lib/rake.rb:1094:in `sh' /Library/Ruby/Gems/1.8/gems/rake-0.8.7/lib/rake.rb:1029:in `ruby' /Library/Ruby/Gems/1.8/gems/rake-0.8.7/lib/rake.rb:1094:in `ruby' /Library/Ruby/Gems/1.8/gems/cucumber-0.4.4/lib/cucumber/rake/task.rb:68:in `run' /Library/Ruby/Gems/1.8/gems/cucumber-0.4.4/lib/cucumber/rake/task.rb:138:in `define_task' /Library/Ruby/Gems/1.8/gems/rake-0.8.7/lib/rake.rb:636:in `call' /Library/Ruby/Gems/1.8/gems/rake-0.8.7/lib/rake.rb:636:in `execute' /Library/Ruby/Gems/1.8/gems/rake-0.8.7/lib/rake.rb:631:in `each' /Library/Ruby/Gems/1.8/gems/rake-0.8.7/lib/rake.rb:631:in `execute' /Library/Ruby/Gems/1.8/gems/rake-0.8.7/lib/rake.rb:597:in `invoke_with_call_chain' /System/Library/Frameworks/Ruby.framework/Versions/1.8/usr/lib/ruby/1.8/monitor.rb:242:in `synchronize' /Library/Ruby/Gems/1.8/gems/rake-0.8.7/lib/rake.rb:590:in `invoke_with_call_chain' /Library/Ruby/Gems/1.8/gems/rake-0.8.7/lib/rake.rb:607:in `invoke_prerequisites' /Library/Ruby/Gems/1.8/gems/rake-0.8.7/lib/rake.rb:604:in `each' /Library/Ruby/Gems/1.8/gems/rake-0.8.7/lib/rake.rb:604:in `invoke_prerequisites' /Library/Ruby/Gems/1.8/gems/rake-0.8.7/lib/rake.rb:596:in `invoke_with_call_chain' /System/Library/Frameworks/Ruby.framework/Versions/1.8/usr/lib/ruby/1.8/monitor.rb:242:in `synchronize' /Library/Ruby/Gems/1.8/gems/rake-0.8.7/lib/rake.rb:590:in `invoke_with_call_chain' /Library/Ruby/Gems/1.8/gems/rake-0.8.7/lib/rake.rb:583:in `invoke' /Library/Ruby/Gems/1.8/gems/rake-0.8.7/lib/rake.rb:2051:in `invoke_task' /Library/Ruby/Gems/1.8/gems/rake-0.8.7/lib/rake.rb:2029:in `top_level' /Library/Ruby/Gems/1.8/gems/rake-0.8.7/lib/rake.rb:2029:in `each' /Library/Ruby/Gems/1.8/gems/rake-0.8.7/lib/rake.rb:2029:in `top_level' /Library/Ruby/Gems/1.8/gems/rake-0.8.7/lib/rake.rb:2068:in `standard_exception_handling' /Library/Ruby/Gems/1.8/gems/rake-0.8.7/lib/rake.rb:2023:in `top_level' /Library/Ruby/Gems/1.8/gems/rake-0.8.7/lib/rake.rb:2001:in `run' /Library/Ruby/Gems/1.8/gems/rake-0.8.7/lib/rake.rb:2068:in `standard_exception_handling' /Library/Ruby/Gems/1.8/gems/rake-0.8.7/lib/rake.rb:1998:in `run' /Library/Ruby/Gems/1.8/gems/rake-0.8.7/bin/rake:31 /usr/bin/rake:19:in `load' /usr/bin/rake:19 WHY? but the command: $ cucumber still works Any idea?

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• Android AsyncTask testing problem with Android Test Framework

  - by Vlad

  I have a very simple AsyncTask implementation example and have problem to test it using Android JUnit framework. It works just fine when I instantiate and execute it in normal application. However when it's executed from any of Android Testing framework classes (i.e. AndroidTestCase, ActivityUnitTestCase, ActivityInstrumentationTestCase2 etc) it behaves sarngely: - It executes doInBackground() method correctly - However it doesn't invokes any of its notification methods (onPostExecute(), onProgressUpdate(), etc) -- just silently ignores them whitout showing any errors. This is very simple AsyncTask example package kroz.andcookbook.threads.asynctask; import android.os.AsyncTask; import android.util.Log; import android.widget.ProgressBar; import android.widget.Toast; public class AsyncTaskDemo extends AsyncTask<Integer, Integer, String> { AsyncTaskDemoActivity _parentActivity; int _counter; int _maxCount; public AsyncTaskDemo(AsyncTaskDemoActivity asyncTaskDemoActivity) { _parentActivity = asyncTaskDemoActivity; } @Override protected void onPreExecute() { super.onPreExecute(); _parentActivity._progressBar.setVisibility(ProgressBar.VISIBLE); _parentActivity._progressBar.invalidate(); } @Override protected String doInBackground(Integer... params) { _maxCount = params[0]; for (_counter = 0; _counter <= _maxCount; _counter++) { try { Thread.sleep(1000); publishProgress(_counter); } catch (InterruptedException e) { // Ignore } } } @Override protected void onProgressUpdate(Integer... values) { super.onProgressUpdate(values); int progress = values[0]; String progressStr = "Counting " + progress + " out of " + _maxCount; _parentActivity._textView.setText(progressStr); _parentActivity._textView.invalidate(); } @Override protected void onPostExecute(String result) { super.onPostExecute(result); _parentActivity._progressBar.setVisibility(ProgressBar.INVISIBLE); _parentActivity._progressBar.invalidate(); } @Override protected void onCancelled() { super.onCancelled(); _parentActivity._textView.setText("Request to cancel AsyncTask"); } } This is a test case. Here AsyncTaskDemoActivity is a very simple Activity providing UI for testing AsyncTask in mode: package kroz.andcookbook.test.threads.asynctask; import java.util.concurrent.ExecutionException; import kroz.andcookbook.R; import kroz.andcookbook.threads.asynctask.AsyncTaskDemo; import kroz.andcookbook.threads.asynctask.AsyncTaskDemoActivity; import android.content.Intent; import android.test.ActivityUnitTestCase; import android.widget.Button; public class AsyncTaskDemoTest2 extends ActivityUnitTestCase<AsyncTaskDemoActivity> { AsyncTaskDemo _atask; private Intent _startIntent; public AsyncTaskDemoTest2() { super(AsyncTaskDemoActivity.class); } protected void setUp() throws Exception { super.setUp(); _startIntent = new Intent(Intent.ACTION_MAIN); } protected void tearDown() throws Exception { super.tearDown(); } public final void testExecute() { startActivity(_startIntent, null, null); Button btnStart = (Button) getActivity().findViewById(R.id.Button01); btnStart.performClick(); assertNotNull(getActivity()); } } All this code is working just fine, except the fact that AsynTask doesn't invoke it's notification methods when executed by whithin Android Testing Framework. Any ideas?

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• Understanding the memory consumption on iPhone

  - by zoul

  Hello! I am working on a 2D iPhone game using OpenGL ES and I keep hitting the 24 MB memory limit – my application keeps crashing with the error code 101. I tried real hard to find where the memory goes, but the numbers in Instruments are still much bigger than what I would expect. I ran the application with the Memory Monitor, Object Alloc, Leaks and OpenGL ES instruments. When the application gets loaded, free physical memory drops from 37 MB to 23 MB, the Object Alloc settles around 7 MB, Leaks show two or three leaks a few bytes in size, the Gart Object Size is about 5 MB and Memory Monitor says the application takes up about 14 MB of real memory. I am perplexed as where did the memory go – when I dig into the Object Allocations, most of the memory is in the textures, exactly as I would expect. But both my own texture allocation counter and the Gart Object Size agree that the textures should take up somewhere around 5 MB. I am not aware of allocating anything else that would be worth mentioning, and the Object Alloc agrees. Where does the memory go? (I would be glad to supply more details if this is not enough.) Update: I really tried to find where I could allocate so much memory, but with no results. What drives me wild is the difference between the Object Allocations (~7 MB) and real memory usage as shown by Memory Monitor (~14 MB). Even if there were huge leaks or huge chunks of memory I forget about, the should still show up in the Object Allocations, shouldn’t they? I’ve already tried the usual suspects, ie. the UIImage with its caching, but that did not help. Is there a way to track memory usage “debugger-style”, line by line, watching each statement’s impact on memory usage? What I have found so far: I really am using that much memory. It is not easy to measure the real memory consumption, but after a lot of counting I think the memory consumption is really that high. My fault. I found no easy way to measure the memory used. The Memory Monitor numbers are accurate (these are the numbers that really matter), but the Memory Monitor can’t tell you where exactly the memory goes. The Object Alloc tool is almost useless for tracking the real memory usage. When I create a texture, the allocated memory counter goes up for a while (reading the texture into the memory), then drops (passing the texture data to OpenGL, freeing). This is OK, but does not always happen – sometimes the memory usage stays high even after the texture has been passed on to OpenGL and freed from “my” memory. This means that the total amount of memory allocated as shown by the Object Alloc tool is smaller than the real total memory consumption, but bigger than the real consumption minus textures (real – textures < object alloc < real). Go figure. I misread the Programming Guide. The memory limit of 24 MB applies to textures and surfaces, not the whole application. The actual red line lies a bit further, but I could not find any hard numbers. The consensus is that 25–30 MB is the ceiling. When the system gets short on memory, it starts sending the memory warning. I have almost nothing to free, but other applications do release some memory back to the system, especially Safari (which seems to be caching the websites). When the free memory as shown in the Memory Monitor goes zero, the system starts killing. I had to bite the bullet and rewrite some parts of the code to be more efficient on memory, but I am probably still pushing it. I

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• Code Golf: Word Search Solver

  - by Maxim Z.

  Note: This is my first Code Golf challenge/question, so I might not be using the correct format below. I'm not really sure how to tag this particular question, and should this be community wiki? Thanks! This Code Golf challenge is about solving word searches! A word search, as defined by Wikipedia, is: A word search, word find, word seek, word sleuth or mystery word puzzle is a word game that is letters of a word in a grid, that usually has a rectangular or square shape. The objective of this puzzle is to find and mark all the words hidden inside the box. The words may be horizontally, vertically or diagonally. Often a list of the hidden words is provided, but more challenging puzzles may let the player figure them out. Many word search puzzles have a theme to which all the hidden words are related. The word searches for this challenge will all be rectangular grids with a list of words to find provided. The words can be written vertically, horizontally, or diagonally. Input/Output The user inputs their word search and then inputs a word to be found in their grid. These two inputs are passed to the function that you will be writing. It is up to you how you want to declare and handle these objects. Using a strategy described below or one of your own, the function finds the specific word in the search and outputs its starting coordinates (simply row number and column number) and ending coordinates. If you find two occurrences of the word, you must output both's set of coordinates. Example Input: A I Y R J J Y T A S V Q T Z E X B X G R Z P W V T B K U F O E A F L V F J J I A G B A J K R E S U R E P U S C Y R S Y K F B B Q Y T K O I K H E W G N G L W Z F R F H L O R W A R E J A O S F U E H Q V L O A Z B J F B G I F Q X E E A L W A C F W K Z E U U R Z R T N P L D F L M P H D F W H F E C G W Z B J S V O A O Y D L M S T C R B E S J U V T C S O O X P F F R J T L C V W R N W L Q U F I B L T O O S Q V K R O W G N D B C D E J Y E L W X J D F X M Word to find: codegolf Output: row 12, column 8 --> row 5, column 1 Strategies Here are a few strategies you might consider using. It is completely up to you to decide what strategy you want to use; it doesn't have to be in this list. Looking for the first letter of the word; on each occurrence, looking at the eight surrounding letters to see whether the next letter of the word is there. Same as above, except looking for a part of a word that has two of the same letter side-by-side. Counting how often each letter of the alphabet is present in the whole grid, then selecting one of the least-occurring letters from the word you have to find and searching for the letter. On each occurrence of the letter, you look at its eight surrounding letters to see whether the next and previous letters of the word is there.

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• How do I send an email with embedded images AND regular attachments in JavaMail?

  - by Chris

  Hi, I'd like to know how to build an SMTP multipart message in the correct order so that it will render correctly on the iPhone mail client (rendering correctly in GMail). I'm using Javamail to build up an email containing the following parts: A body part with content type "text/html; UTF-8" An embedded image attachment. A file attachment I am sending the mail via GMail SMTP (via SSL) and the mail is sent and rendered correctly using a GMail account, however, the mail does not render correctly on the iPhone mail client. On the iPhone mail client, the image is rendered before the "Before Image" text when it should be rendered afterwards. After the "Before Image" text there is an icon with a question mark (I assume it means it couldn't find the referenced CID). I'm not sure if this is a limitation of the iPhone mail client or a bug in my mail sending code (I strongly assume the latter). I think that perhaps the headers on my parts might by incorrect or perhaps I am providing the multiparts in the wrong order. I include the text of the received mail as output by gmail (which renders the file correc Message-ID: <[email protected]> Subject: =?UTF-8?Q?Test_from_=E3=82=AF=E3=83=AA=E3=82=B9?= MIME-Version: 1.0 Content-Type: multipart/mixed; boundary="----=_Part_0_20870565.1274154021755" ------=_Part_0_20870565.1274154021755 Content-Type: application/octet-stream Content-Transfer-Encoding: base64 Content-ID: <20100518124021763_368238_0> iVBORw0K ----- TRIMMED FOR CONCISENESS 6p1VVy4alAAAAABJRU5ErkJggg== ------=_Part_0_20870565.1274154021755 Content-Type: text/html; charset=UTF-8 Content-Transfer-Encoding: 7bit <html><head><title>Employees Favourite Foods</title> <style> body { font: normal 8pt arial; } th { font: bold 8pt arial; white-space: nowrap; } td { font: normal 8pt arial; white-space: nowrap; } </style></head><body> Before Image<br><img src="cid:20100518124021763_368238_0"> After Image<br><table border="0"> <tr> <th colspan="4">Employees Favourite Foods</th> </tr> <tr> <th align="left">Name</th><th align="left">Age</th><th align="left">Tel.No</th><th align="left">Fav.Food</th> </tr> <tr style="background-color:#e0e0e0"> <td>Chris</td><td>34</td><td>555-123-4567</td><td>Pancakes</td> </tr> </table></body></html> ------=_Part_0_20870565.1274154021755 Content-Type: text/plain; charset=us-ascii; name=textfile.txt Content-Transfer-Encoding: 7bit Content-Disposition: attachment; filename=textfile.txt This is a textfile with numbers counting from one to ten beneath this line: one two three four five six seven eight nine ten(no trailing carriage return) ------=_Part_0_20870565.1274154021755-- Even if you can't assist me with this, I would appreciate it if any members of the forum could forward me a (non-personal) mail that includes inline images (not external hyperlinked images though). I just need to find a working sample then I can move past this. Thanks, Chris.

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• Problem trying to achieve a join using the `comments` contrib in Django

  - by NiKo

  Hi, Django rookie here. I have this model, comments are managed with the django_comments contrib: class Fortune(models.Model): author = models.CharField(max_length=45, blank=False) title = models.CharField(max_length=200, blank=False) slug = models.SlugField(_('slug'), db_index=True, max_length=255, unique_for_date='pub_date') content = models.TextField(blank=False) pub_date = models.DateTimeField(_('published date'), db_index=True, default=datetime.now()) votes = models.IntegerField(default=0) comments = generic.GenericRelation( Comment, content_type_field='content_type', object_id_field='object_pk' ) I want to retrieve Fortune objects with a supplementary nb_comments value for each, counting their respectve number of comments ; I try this query: >>> Fortune.objects.annotate(nb_comments=models.Count('comments')) From the shell: >>> from django_fortunes.models import Fortune >>> from django.db.models import Count >>> Fortune.objects.annotate(nb_comments=Count('comments')) [<Fortune: My first fortune, from NiKo>, <Fortune: Another One, from Dude>, <Fortune: A funny one, from NiKo>] >>> from django.db import connection >>> connection.queries.pop() {'time': '0.000', 'sql': u'SELECT "django_fortunes_fortune"."id", "django_fortunes_fortune"."author", "django_fortunes_fortune"."title", "django_fortunes_fortune"."slug", "django_fortunes_fortune"."content", "django_fortunes_fortune"."pub_date", "django_fortunes_fortune"."votes", COUNT("django_comments"."id") AS "nb_comments" FROM "django_fortunes_fortune" LEFT OUTER JOIN "django_comments" ON ("django_fortunes_fortune"."id" = "django_comments"."object_pk") GROUP BY "django_fortunes_fortune"."id", "django_fortunes_fortune"."author", "django_fortunes_fortune"."title", "django_fortunes_fortune"."slug", "django_fortunes_fortune"."content", "django_fortunes_fortune"."pub_date", "django_fortunes_fortune"."votes" LIMIT 21'} Below is the properly formatted sql query: SELECT "django_fortunes_fortune"."id", "django_fortunes_fortune"."author", "django_fortunes_fortune"."title", "django_fortunes_fortune"."slug", "django_fortunes_fortune"."content", "django_fortunes_fortune"."pub_date", "django_fortunes_fortune"."votes", COUNT("django_comments"."id") AS "nb_comments" FROM "django_fortunes_fortune" LEFT OUTER JOIN "django_comments" ON ("django_fortunes_fortune"."id" = "django_comments"."object_pk") GROUP BY "django_fortunes_fortune"."id", "django_fortunes_fortune"."author", "django_fortunes_fortune"."title", "django_fortunes_fortune"."slug", "django_fortunes_fortune"."content", "django_fortunes_fortune"."pub_date", "django_fortunes_fortune"."votes" LIMIT 21 Can you spot the problem? Django won't LEFT JOIN the django_comments table with the content_type data (which contains a reference to the fortune one). This is the kind of query I'd like to be able to generate using the ORM: SELECT "django_fortunes_fortune"."id", "django_fortunes_fortune"."author", "django_fortunes_fortune"."title", COUNT("django_comments"."id") AS "nb_comments" FROM "django_fortunes_fortune" LEFT OUTER JOIN "django_comments" ON ("django_fortunes_fortune"."id" = "django_comments"."object_pk") LEFT OUTER JOIN "django_content_type" ON ("django_comments"."content_type_id" = "django_content_type"."id") GROUP BY "django_fortunes_fortune"."id", "django_fortunes_fortune"."author", "django_fortunes_fortune"."title", "django_fortunes_fortune"."slug", "django_fortunes_fortune"."content", "django_fortunes_fortune"."pub_date", "django_fortunes_fortune"."votes" LIMIT 21 But I don't manage to do it, so help from Django veterans would be much appreciated :) Hint: I'm using Django 1.2-DEV Thanks in advance for your help.

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• Explain the Peak and Flag Algorithm

  - by Isaac Levin

  EDIT Just was pointed that the requirements state peaks cannot be ends of Arrays. So I ran across this site http://codility.com/ Which gives you programming problems and gives you certificates if you can solve them in 2 hours. The very first question is one I have seen before, typically called the Peaks and Flags question. If you are not familiar A non-empty zero-indexed array A consisting of N integers is given. A peak is an array element which is larger than its neighbours. More precisely, it is an index P such that 0 < P < N - 1 and A[P - 1] < A[P] A[P + 1] . For example, the following array A: A[0] = 1 A[1] = 5 A[2] = 3 A[3] = 4 A[4] = 3 A[5] = 4 A[6] = 1 A[7] = 2 A[8] = 3 A[9] = 4 A[10] = 6 A[11] = 2 has exactly four peaks: elements 1, 3, 5 and 10. You are going on a trip to a range of mountains whose relative heights are represented by array A. You have to choose how many flags you should take with you. The goal is to set the maximum number of flags on the peaks, according to certain rules. Flags can only be set on peaks. What's more, if you take K flags, then the distance between any two flags should be greater than or equal to K. The distance between indices P and Q is the absolute value |P - Q|. For example, given the mountain range represented by array A, above, with N = 12, if you take: two flags, you can set them on peaks 1 and 5; three flags, you can set them on peaks 1, 5 and 10; four flags, you can set only three flags, on peaks 1, 5 and 10. You can therefore set a maximum of three flags in this case. Write a function that, given a non-empty zero-indexed array A of N integers, returns the maximum number of flags that can be set on the peaks of the array. For example, given the array above the function should return 3, as explained above. Assume that: N is an integer within the range [1..100,000]; each element of array A is an integer within the range [0..1,000,000,000]. Complexity: expected worst-case time complexity is O(N); expected worst-case space complexity is O(N), beyond input storage (not counting the storage required for input arguments). Elements of input arrays can be modified. So this makes sense, but I failed it using this code public int GetFlags(int[] A) { List<int> peakList = new List<int>(); for (int i = 0; i <= A.Length - 1; i++) { if ((A[i] > A[i + 1] && A[i] > A[i - 1])) { peakList.Add(i); } } List<int> flagList = new List<int>(); int distance = peakList.Count; flagList.Add(peakList[0]); for (int i = 1, j = 0, max = peakList.Count; i < max; i++) { if (Math.Abs(Convert.ToDecimal(peakList[j]) - Convert.ToDecimal(peakList[i])) >= distance) { flagList.Add(peakList[i]); j = i; } } return flagList.Count; } EDIT int[] A = new int[] { 7, 10, 4, 5, 7, 4, 6, 1, 4, 3, 3, 7 }; The correct answer is 3, but my application says 2 This I do not get, since there are 4 peaks (indices 1,4,6,8) and from that, you should be able to place a flag at 2 of the peaks (1 and 6) Am I missing something here? Obviously my assumption is that the beginning or end of an Array can be a peak, is this not the case? If this needs to go in Stack Exchange Programmers, I will move it, but thought dialog here would be helpful. EDIT

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• Order of parts in SMTP multipart messages

  - by Chris

  Hi, I'd like to know how to build an SMTP multipart message in the correct order so that it will render correctly on the iPhone mail client (rendering correctly in GMail). I'm using Javamail to build up an email containing the following parts: A body part with content type "text/html; UTF-8" An embedded image attachment. A file attachment I am sending the mail via GMail SMTP (via SSL) and the mail is sent and rendered correctly using a GMail account, however, the mail does not render correctly on the iPhone mail client. On the iPhone mail client, the image is rendered before the "Before Image" text when it should be rendered afterwards. After the "Before Image" text there is an icon with a question mark (I assume it means it couldn't find the referenced CID). I'm not sure if this is a limitation of the iPhone mail client or a bug in my mail sending code (I strongly assume the latter). I think that perhaps the headers on my parts might by incorrect or perhaps I am providing the multiparts in the wrong order. I include the text of the received mail as output by gmail (which renders the file correc Message-ID: <[email protected]> Subject: =?UTF-8?Q?Test_from_=E3=82=AF=E3=83=AA=E3=82=B9?= MIME-Version: 1.0 Content-Type: multipart/mixed; boundary="----=_Part_0_20870565.1274154021755" ------=_Part_0_20870565.1274154021755 Content-Type: application/octet-stream Content-Transfer-Encoding: base64 Content-ID: <20100518124021763_368238_0> iVBORw0K ----- TRIMMED FOR CONCISENESS 6p1VVy4alAAAAABJRU5ErkJggg== ------=_Part_0_20870565.1274154021755 Content-Type: text/html; charset=UTF-8 Content-Transfer-Encoding: 7bit <html><head><title>Employees Favourite Foods</title> <style> body { font: normal 8pt arial; } th { font: bold 8pt arial; white-space: nowrap; } td { font: normal 8pt arial; white-space: nowrap; } </style></head><body> Before Image<br><img src="cid:20100518124021763_368238_0"> After Image<br><table border="0"> <tr> <th colspan="4">Employees Favourite Foods</th> </tr> <tr> <th align="left">Name</th><th align="left">Age</th><th align="left">Tel.No</th><th align="left">Fav.Food</th> </tr> <tr style="background-color:#e0e0e0"> <td>Chris</td><td>34</td><td>555-123-4567</td><td>Pancakes</td> </tr> </table></body></html> ------=_Part_0_20870565.1274154021755 Content-Type: text/plain; charset=us-ascii; name=textfile.txt Content-Transfer-Encoding: 7bit Content-Disposition: attachment; filename=textfile.txt This is a textfile with numbers counting from one to ten beneath this line: one two three four five six seven eight nine ten(no trailing carriage return) ------=_Part_0_20870565.1274154021755-- Even if you can't assist me with this, I would appreciate it if any members of the forum could forward me a (non-personal) mail that includes inline images (not external hyperlinked images though). I just need to find a working sample then I can move past this. Thanks, Chris.

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• SQL Server 2008: Using Multiple dts Ranges to Build a Set of Dates

  - by raoulcousins

  I'm trying to build a query for a medical database that counts the number of patients that were on at least one medication from a class of medications (the medications listed below in the FAST_MEDS CTE) and had either: 1) A diagnosis of myopathy (the list of diagnoses in the FAST_DX CTE) 2) A CPK lab value above 1000 (the lab value in the FAST_LABS CTE) and this diagnosis or lab happened AFTER a patient was on a statin. The query I've included below does that under the assumption that once a patient is on a statin, they're on a statin forever. The first CTE collects the ids of patients that were on a statin along with the first date of their diagnosis, the second those with a diagnosis, and the third those with a high lab value. After this I count those that match the above criteria. What I would like to do is drop the assumption that once a patient is on a statin, they're on it for life. The table edw_dm.patient_medications has a column called start_dts and end_dts. This table has one row for each prescription written, with start_dts and end_dts denoting the start and end date of the prescription. End_dts could be null, which I'll take to assume that the patient is currently on this medication (it could be a missing record, but I can't do anything about this). If a patient is on two different statins, the start and ends dates can overlap, and there may be multiple records of the same medication for a patient, as in a record showing 3-11-2000 to 4-5-2003 and another for the same patient showing 5-6-2007 to 7-8-2009. I would like to use these two columns to build a query where I'm only counting the patients that had a lab value or diagnosis done during a time when they were already on a statin, or in the first n (say 3) months after they stopped taking a statin. I'm really not sure how to go about rewriting the first CTE to get this information and how to do the comparison after the CTEs are built. I know this is a vague question, but I'm really stumped. Any ideas? As always, thank you in advance. Here's the current query: WITH FAST_MEDS AS ( select distinct statins.mrd_pt_id, min(year(statins.order_dts)) as statin_yr from edw_dm.patient_medications as statins inner join mrd.medications as mrd on statins.mrd_med_id = mrd.mrd_med_id WHERE mrd.generic_nm in ( 'Lovastatin (9664708500)', 'lovastatin-niacin', 'Lovastatin/Niacin', 'Lovastatin', 'Simvastatin (9678583966)', 'ezetimibe-simvastatin', 'niacin-simvastatin', 'ezetimibe/Simvastatin', 'Niacin/Simvastatin', 'Simvastatin', 'Aspirin Buffered-Pravastatin', 'aspirin-pravastatin', 'Aspirin/Pravastatin', 'Pravastatin', 'amlodipine-atorvastatin', 'Amlodipine/atorvastatin', 'atorvastatin', 'fluvastatin', 'rosuvastatin' ) and YEAR(statins.order_dts) IS NOT NULL and statins.mrd_pt_id IS NOT NULL group by statins.mrd_pt_id ) select * into #meds from FAST_MEDS ; --return patients who had a diagnosis in the list and the year that --diagnosis was given with FAST_DX AS ( SELECT pd.mrd_pt_id, YEAR(pd.init_noted_dts) as init_yr FROM edw_dm.patient_diagnoses as pd inner join mrd.diagnoses as mrd on pd.mrd_dx_id = mrd.mrd_dx_id and mrd.icd9_cd in ('728.89','729.1','710.4','728.3','729.0','728.81','781.0','791.3') ) select * into #dx from FAST_DX; --return patients who had a high cpk value along with the year the cpk --value was taken with FAST_LABS AS ( SELECT pl.mrd_pt_id, YEAR(pl.order_dts) as lab_yr FROM edw_dm.patient_labs as pl inner join mrd.labs as mrd on pl.mrd_lab_id = mrd.mrd_lab_id and mrd.lab_nm = 'CK (CPK)' WHERE pl.lab_val between 1000 AND 999998 ) select * into #labs from FAST_LABS; -- count the number of patients who had a lab value or a medication -- value taken sometime AFTER their initial statin diagnosis select count(distinct p.mrd_pt_id) as ct from mrd.patient_demographics as p join #meds as m on p.mrd_pt_id = m.mrd_pt_id AND ( EXISTS ( SELECT 'A' FROM #labs l WHERE p.mrd_pt_id = l.mrd_pt_id and l.lab_yr >= m.statin_yr ) OR EXISTS( SELECT 'A' FROM #dx d WHERE p.mrd_pt_id = d.mrd_pt_id AND d.init_yr >= m.statin_yr ) )

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• Fill container with template parameters

  - by phlipsy

  I want to fill the template parameters passed to a variadic template into an array with fixed length. For that purpose I wrote the following helper function templates template<typename ForwardIterator, typename T> void fill(ForwardIterator i) { } template<typename ForwardIterator, typename T, T head, T... tail> void fill(ForwardIterator i) { *i = head; fill<ForwardIterator, T, tail...>(++i); } the following class template template<typename T, T... args> struct params_to_array; template<typename T, T last> struct params_to_array<T, last> { static const std::size_t SIZE = 1; typedef std::array<T, SIZE> array_type; static const array_type params; private: void init_params() { array_type result; fill<typename array_type::iterator, T, head, tail...>(result.begin()); return result; } }; template<typename T, T head, T... tail> struct params_to_array<T, head, tail...> { static const std::size_t SIZE = params_to_array<T, tail...>::SIZE + 1; typedef std::array<T, SIZE> array_type; static const array_type params; private: void init_params() { array_type result; fill<typename array_type::iterator, T, last>(result.begin()); return result; } }; and initialized the static constants via template<typename T, T last> const typename param_to_array<T, last>::array_type param_to_array<T, last>::params = param_to_array<T, last>::init_params(); and template<typename T, T head, T... tail> const typename param_to_array<T, head, tail...>::array_type param_to_array<T, head, tail...>::params = param_to_array<T, head, tail...>::init_params(); Now the array param_to_array<int, 1, 3, 4>::params is a std::array<int, 3> and contains the values 1, 3 and 4. I think there must be a simpler way to achieve this behavior. Any suggestions? Edit: As Noah Roberts suggested in his answer I modified my program like the following: I wrote a new struct counting the elements in a parameter list: template<typename T, T... args> struct count; template<typename T, T head, T... tail> struct count<T, head, tail...> { static const std::size_t value = count<T, tail...>::value + 1; }; template<typename T, T last> stuct count<T, last> { static const std::size_t value = 1; }; and wrote the following function template<typename T, T... args> std::array<T, count<T, args...>::value> params_to_array() { std::array<T, count<T, args...>::value> result; fill<typename std::array<T, count<T, args...>::value>::iterator, T, args...>(result.begin()); return result; } Now I get with params_to_array<int, 10, 20, 30>() a std::array<int, 3> with the content 10, 20 and 30. Any further suggestions?

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• Using the login Details via Application

  - by ramin ss

  I have a CURL(in C++) to send my user and pass to remauth.php file so i think i do something wrong on remuth.php ( because i am basic in php and my program can not run because the auth not passed.) I use login via Application. my CURL: bool Auth_PerformSessionLogin(const char* username, const char* password) { curl_global_init(CURL_GLOBAL_ALL); CURL* curl = curl_easy_init(); if (curl) { char url[255]; _snprintf(url, sizeof(url), "http://%s/remauth.php", "SITEADDRESS.com"); char buf[8192] = {0}; char postBuf[8192]; _snprintf(postBuf, sizeof(postBuf), "%s&&%s", username, password); curl_easy_setopt(curl, CURLOPT_URL, url); curl_easy_setopt(curl, CURLOPT_WRITEFUNCTION, AuthDataReceived); curl_easy_setopt(curl, CURLOPT_WRITEDATA, (void*)&buf); curl_easy_setopt(curl, CURLOPT_USERAGENT, "IW4M"); curl_easy_setopt(curl, CURLOPT_FAILONERROR, true); curl_easy_setopt(curl, CURLOPT_POST, 1); curl_easy_setopt(curl, CURLOPT_POSTFIELDS, postBuf); curl_easy_setopt(curl, CURLOPT_POSTFIELDSIZE, -1); curl_easy_setopt(curl, CURLOPT_SSL_VERIFYPEER, false); CURLcode code = curl_easy_perform(curl); curl_easy_cleanup(curl); curl_global_cleanup(); if (code == CURLE_OK) { return Auth_ParseResultBuffer(buf); } else { Auth_Error(va("Could not reach the SITEADDRESS.comt server. Error code from CURL: %x.", code)); } return false; } curl_global_cleanup(); return false; } and my remauth.php: <?php ob_start(); $host=""; // Host name $dbusername=""; // Mysql username $dbpassword=""; // Mysql password $db_name=""; // Database name $tbl_name=""; // Table name // Connect to server and select databse. mysql_connect("$host", "$dbusername", "$dbpassword") or die(mysql_error()); mysql_select_db("$db_name") or die(mysql_error()); // Define $username and $password //$username=$username; //$password=md5($_POST['password']); //$password=$password; $username=$_POST['username']; $password=$_POST['password']; //$post_item[]='action='.$_POST['submit']; // To protect MySQL injection (more detail about MySQL injection) $username = stripslashes($username); $password = stripslashes($password); $username = mysql_real_escape_string($username); $password = mysql_real_escape_string($password); $sql="SELECT * FROM $tbl_name WHERE username='$username'"; $result=mysql_query($sql); // Mysql_num_row is counting table row $count=mysql_num_rows($result); // If result matched $username and $password, table row must be 1 row if($count==1){ $row = mysql_fetch_assoc($result); if (md5(md5($row['salt']).md5($password)) == $row['password']){ session_register("username"); session_register("password"); echo "#"; return true; } else { echo "o"; return false; } } else{ echo "o"; return false; } ob_end_flush(); ?> ///////////////////////////////////

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• SCCM 2012 - Windows 8 WSUS

  - by Owen

  We're using SCCM 2012 to deploy Windows Updates on our domain, and our Windows 8 clients have started failing with error 80240438 when they try to update. Windows 7 clients update fine, but Windows 8 clients refuse to do anything. I've done a search online and it seems to only reference Windows InTune. Has anyone seen any similar behavior on Windows 8 machines? If we don't get that error, we're getting 80244021 which seems to indicate that the server can't be found.... but they can resolve it just fine and our exceptions are defined on the proxy too. A bit stuck here! 2012-11-22 14:45:28:935 476 998 Agent ********* 2012-11-22 14:45:28:935 476 998 Agent ** END ** Agent: Finding updates [CallerId = AutomaticUpdates] 2012-11-22 14:45:28:935 476 998 Agent ************* 2012-11-22 14:45:28:935 476 998 Agent WARNING: WU client failed Searching for update with error 0x80240438 2012-11-22 14:45:28:935 476 c74 AU >>## RESUMED ## AU: Search for updates [CallId = {EAECB947-48AC-43BE-8F98-C44727E4A131} ServiceId = {3DA21691-E39D-4DA6-8A4B-B43877BCB1B7}] 2012-11-22 14:45:28:935 476 c74 AU # WARNING: Search callback failed, result = 0x80240438 2012-11-22 14:45:28:935 476 c74 AU ######### 2012-11-22 14:45:28:935 476 c74 AU ## END ## AU: Search for updates [CallId = {EAECB947-48AC-43BE-8F98-C44727E4A131} ServiceId = {3DA21691-E39D-4DA6-8A4B-B43877BCB1B7}] 2012-11-22 14:45:28:935 476 c74 AU ############# 2012-11-22 14:45:28:935 476 c74 AU All AU searches complete. 2012-11-22 14:45:28:935 476 c74 AU # WARNING: Failed to find updates with error code 80240438 2012-11-22 14:45:28:935 476 c74 AU AU setting next detection timeout to 2012-11-22 04:12:23 2012-11-22 14:45:33:936 476 c9c Report REPORT EVENT: {EE35CD79-FD2A-472D-BFC9-0420F5D60C04} 2012-11-22 14:45:28:935+1300 1 148 [AGENT_DETECTION_FAILED] 101 {00000000-0000-0000-0000-000000000000} 0 80240438 AutomaticUpdates Failure Software Synchronization Windows Update Client failed to detect with error 0x80240438. 2012-11-22 14:45:33:938 476 c9c Report CWERReporter::HandleEvents - WER report upload completed with status 0x8 2012-11-22 14:45:33:938 476 c9c Report WER Report sent: 7.8.9200.16420 0x80240438 00000000-0000-0000-0000-000000000000 Scan 101 Managed 2012-11-22 14:45:33:938 476 c9c Report CWERReporter finishing event handling. (00000000) Thanks in advance

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• Watchguard Firewall WebBlocker Regular Expression for Multiple Domains?

  - by Eric

  I'm pretty sure this is really a regex question, so you can skip to REGEX QUESTION if you want to skip the background. Our primary firewall is a Watchguard X750e running Fireware XTM v11.2. We're using webblocker to block most of the categories, and I'm allowing needed sites as they arise. Some sites are simple to add as exceptions, like Pandora radio. That one is just a pattern matched exception with "padnora.com/" as the pattern. All traffic from anywhere on pandora.com is allowed. I'm running into trouble on more sophisticated domains that reference content off of their base domains. We'll take GrooveShark as a sample. If you go to http://grooveshark.com/ and view page source, you'll see hrefs referring to gs-cdn.net as well as grooveshar.com. So adding a WebBlocker exception to grooveshark.com/ is not effective, and I have to add a second rule allowing gs-cdn.net/ as well. I see that the WebBlocker allows regex rules, so what I'd like to do in situations like this is create a single regex rule that allows traffic to all the needed domains. REGEX QUESTION: I'd like to try a regex that matches grooveshark.com/ and gs-cdn.net/. If anybody can help me write that regex, I'd appreciate it. Here is what is in the WatchGuard documentation from that section: Regular expression Regular expression matches use a Perl-compatible regular expression to make a match. For example, .[onc][eor][gtm] matches .org, .net, .com, or any other three-letter combination of one letter from each bracket, in order. Be sure to drop the leading “http://” Supports wild cards used in shell script. For example, the expression “(www)?.watchguard.[com|org|net]” will match URL paths including www.watchguard.com, www.watchguard.net, and www.watchguard.org. Thanks all!

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