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  • Data munging and data import scripting

    - by morpheous
    I need to write some scripts to carry out some tasks on my server (running Ubuntu server 8.04 TLS). The tasks are to be run periodically, so I will be running the scripts as cron jobs. I have divided the tasks into "group A" and "group B" - because (in my mind at least), they are a bit different. Task Group A import data from a file and possibly reformat it - by reformatting, I mean doing things like santizing the data, possibly normalizing it and or running calculations on 'columns' of the data Import the munged data into a database. For now, I am mostly using mySQL for the vast majority of imports - although some files will be imported into a sqlLite database. Note: The files will be mostly text files, although some of the files are in a binary format (my own proprietary format, written by a C++ application I developed). Task Group B Extract data from the database Perform calculations on the data and either insert or update tables in the database. My coding experience is is primarily as a C/C++ developer, although I have been using PHP as well for the last 2 years or so. I am from a windows background so I am still finding my feet in the linux environment. My question is this - I need to write scripts to perform the tasks I described above. Although I suppose I could write a few C++ applications to be used in the shell scripts, I think it may be better to write them in a scripting language (maybe this is a flawed assumption?). My thinking is that it would be easier to modify thins in a script - no need to rebuild etc for changes to functionality. Additionally, C++ data munging in C++ tends to involve more lines of code than "natural" scripting languages such as Perl, Python etc. Assuming that the majority of people on here agree that scripting is the way to go, herein lies my dilema. Which scripting language to use to perform the tasks above (giving my background). My gut instinct tells me that Perl (shudder) would be the most obvious choice for performing all of the above tasks. BUT (and that is a big BUT). The mere mention of Perl makes my toes curl, as I had a very, very bag experience with it a while back. The syntax seems quite unnatural to me - despite how many times I have tried to learn it - so if possible, I would really like to give it a miss. PHP (which I already know), also am not sure is a good candidate for scripting on the CLI (I have not seen many examples on how to do this etc - so I may be wrong). The last thing I must mention is that IF I have to learn a new language in order to do this, I cannot afford (time constraint) to spend more than a day, in learning the key commands/features required in order to do this (I can always learn the details of the language later, once I have actually deployed the scripts). So, which scripting language would you recommend (PHP, Python, Perl, [insert your favorite here]) - and most importantly WHY?. Or, should I just stick to writing little C++ applications that I call in a shell script?. Lastly, if you have suggested a scripting language, can you please show with a FEW lines (Perl mongers - I'm looking in your direction [nothing to cryptic!] ;) ) how I can use the language you suggested to do what I want to do. Hopefully, the lines you present will convince me that it can be done easily and elegantly in the language you suggested.

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  • Write to file depending on minSdkVersion - android

    - by Simon Rosenqvist
    Hi, I have written a filewriter for my android application. It is to function on a Galaxy Tab, so my minSdkVersion has to be at least 4, so it will fill the screen. I originally started out with SdkVersion = 2 and at that point my filewriter worked perfectly. Changing the SdkVersion to 4 introduced the problem. My filewriter doesn't work anymore! The application runs fine, but a file doesn't get created. My .java file looks like this: public class HelloAndroid extends Activity { /** Called when the activity is first created. */ @Override public void onCreate(Bundle savedInstanceState) { super.onCreate(savedInstanceState); TextView tv = new TextView(this); tv.setText("Hello, Android"); setContentView(R.layout.main); //definerer en knap kaldet button1 og sætter en listener på denne. Button button1 = (Button)findViewById(R.id.btnClickMe); button1.setOnClickListener(btnListener); //definerer en knap kaldet button2 og sætter en listener på denne. Button button2 = (Button)findViewById(R.id.btnClickMe2); button2.setOnClickListener(btnListener2); } //en variabel af typen 'long' deklæres og kaldes tid1. public long time1; private OnClickListener btnListener = new OnClickListener() { public void onClick(View v) { //Når der klikkes på button1 gemmes et tal i variablen tid1. time1 = System.currentTimeMillis(); } }; //en variabel af typen 'long' deklæres og kaldes tid2. public long time2; // en variabel af typen 'string' deklæres og kaldes tid: public String string1 = "time:"; private OnClickListener btnListener2 = new OnClickListener() { public void onClick(View v) { //Når der klikkes på button2 gemmes et tal i variablen tid2. time2 = System.currentTimeMillis(); // Herefter oprettes en fil kaldet "file.txt". try{ File file = new File(Environment.getExternalStorageDirectory(), "file.txt"); file.createNewFile(); BufferedWriter writer = new BufferedWriter(new FileWriter(file,true)); //string1 og tid2-tid1 skrives til filen. tid2-tid1 giver den tid der går fra der er trykket på den ene knap til den anden i millisekunder. writer.write(string1 + "\t" + (time2-time1)); writer.newLine(); writer.flush(); writer.close(); } catch (IOException e) { e.printStackTrace(); } } }; } And my manifest.xml looks like this: <?xml version="1.0" encoding="utf-8"?> <application android:icon="@drawable/icon" android:label="@string/app_name"> <activity android:name=".HelloAndroid" android:label="@string/app_name"> <intent-filter> <action android:name="android.intent.action.MAIN" /> <category android:name="android.intent.category.LAUNCHER" /> </intent-filter> </activity> </application> Why does my filewriter not work with minSdkVersion 2? Do i have to make a new filewriter? or what to do? Sorry for the messy code, i'm quite new to programming :)

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  • How best to modernize the 2002-era J2EE app?

    - by user331465
    I have this friend.... I have this friend who works on a java ee application (j2ee) application started in the early 2000's. Currently they add a feature here and there, but have a large codebase. Over the years the team has shrunk by 70%. [Yes, the "i have this friend is". It's me, attempting to humorously inject teenage high-school counselor shame into the mix] Java, Vintage 2002 The application uses EJB 2.1, struts 1.x, DAO's etc with straight jdbc calls (mixture of stored procedures and prepared statements). No ORM. For caching they use a mixture of OpenSymphony OSCache and a home-grown cache layer. Over the last few years, they have spent effort to modernize the UI using ajax techniques and libraries. This largely involves javascript libaries (jquery, yui, etc). Client Side On the client side, the lack of upgrade path from struts1 to struts2 discouraged them from migrating to struts2. Other web frameworks became popular (wicket, spring , jsf). Struts2 was not the "clear winner". Migrating all the existing UI from Struts1 to Struts2/wicket/etc did not seem to present much marginal benefit at a very high cost. They did not want to have a patchwork of technologies-du-jour (subsystem X in Struts2, subsystem Y in Wicket, etc.) so developer write new features using Struts 1. Server Side On the server side, they looked into moving to ejb 3, but never had a big impetus. The developers are all comfortable with ejb-jar.xml, EJBHome, EJBRemote, that "ejb 2.1 as is" represented the path of least resistance. One big complaint about the ejb environment: programmers still pretend "ejb server runs in separate jvm than servlet engine". No app server (jboss/weblogic) has ever enforced this separation. The team has never deployed the ejb server on a separate box then the app server. The ear file contains multiple copies of the same jar file; one for the 'web layer' (foo.war/WEB-INF/lib) and one for the server side (foo.ear/). The app server only loads one jar. The duplications makes for ambiguity. Caching As for caching, they use several cache implementations: OpenSymphony cache and a homegrown cache. Jgroups provides clustering support Now What? The question: The team currently has spare cycles to to invest in modernizing the application? Where would the smart investor spend them? The main criteria: 1) productivity gains. Specifically reducing the time to develope new subsystems features and reduced maintenance. 2) performance/scalability. They do not care about fashion or techno-du-jour street cred. What do you all recommend? On the persistence side Switch everything (or new development only) to JPA/JPA2? Straight hibernate? Wait for Java EE 6? On the client/web-framework side: Migrate (some or all) to struts2? wicket? jsf/jsf2? As for caching: terracotta? ehcache? coherence? stick with what they have? how best to take advantage of the huge heap sizes that the 64-bit jvms offer? Thanks in advance.

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  • Data adapter not filling my dataset

    - by Doug Ancil
    I have the following code: Imports System.Data.SqlClient Public Class Main Protected WithEvents DataGridView1 As DataGridView Dim instForm2 As New Exceptions Private Sub Button1_Click_1(ByVal sender As System.Object, ByVal e As System.EventArgs) Handles startpayrollButton.Click Dim ssql As String = "select MAX(payrolldate) AS [payrolldate], " & _ "dateadd(dd, ((datediff(dd, '17530107', MAX(payrolldate))/7)*7)+7, '17530107') AS [Sunday]" & _ "from dbo.payroll" & _ " where payrollran = 'no'" Dim oCmd As System.Data.SqlClient.SqlCommand Dim oDr As System.Data.SqlClient.SqlDataReader oCmd = New System.Data.SqlClient.SqlCommand Try With oCmd .Connection = New System.Data.SqlClient.SqlConnection("Initial Catalog=mdr;Data Source=xxxxx;uid=xxxxx;password=xxxxx") .Connection.Open() .CommandType = CommandType.Text .CommandText = ssql oDr = .ExecuteReader() End With If oDr.Read Then payperiodstartdate = oDr.GetDateTime(1) payperiodenddate = payperiodstartdate.AddSeconds(604799) Dim ButtonDialogResult As DialogResult ButtonDialogResult = MessageBox.Show(" The Next Payroll Start Date is: " & payperiodstartdate.ToString() & System.Environment.NewLine & " Through End Date: " & payperiodenddate.ToString()) If ButtonDialogResult = Windows.Forms.DialogResult.OK Then exceptionsButton.Enabled = True startpayrollButton.Enabled = False End If End If oDr.Close() oCmd.Connection.Close() Catch ex As Exception MessageBox.Show(ex.Message) oCmd.Connection.Close() End Try End Sub Private Sub Button2_Click(ByVal sender As System.Object, ByVal e As System.EventArgs) Handles exceptionsButton.Click Dim connection As System.Data.SqlClient.SqlConnection Dim adapter As System.Data.SqlClient.SqlDataAdapter = New System.Data.SqlClient.SqlDataAdapter Dim connectionString As String = "Initial Catalog=mdr;Data Source=xxxxx;uid=xxxxx;password=xxxxx" Dim ds As New DataSet Dim _sql As String = "SELECT [Exceptions].Employeenumber,[Exceptions].exceptiondate, [Exceptions].starttime, [exceptions].endtime, [Exceptions].code, datediff(minute, starttime, endtime) as duration INTO scratchpad3" & _ " FROM Employees INNER JOIN Exceptions ON [Exceptions].EmployeeNumber = [Exceptions].Employeenumber" & _ " where [Exceptions].exceptiondate between @payperiodstartdate and @payperiodenddate" & _ " GROUP BY [Exceptions].Employeenumber, [Exceptions].Exceptiondate, [Exceptions].starttime, [exceptions].endtime," & _ " [Exceptions].code, [Exceptions].exceptiondate" connection = New SqlConnection(connectionString) connection.Open() Dim _CMD As SqlCommand = New SqlCommand(_sql, connection) _CMD.Parameters.AddWithValue("@payperiodstartdate", payperiodstartdate) _CMD.Parameters.AddWithValue("@payperiodenddate", payperiodenddate) adapter.SelectCommand = _CMD Try adapter.Fill(ds) If ds Is Nothing OrElse ds.Tables.Count = 0 OrElse ds.Tables(0).Rows.Count = 0 Then 'it's empty MessageBox.Show("There was no data for this time period. Press Ok to continue", "No Data") connection.Close() Exceptions.saveButton.Enabled = False Exceptions.Hide() Else connection.Close() End If Catch ex As Exception MessageBox.Show(ex.ToString) connection.Close() End Try Exceptions.Show() End Sub Private Sub payrollButton_Click(ByVal sender As System.Object, ByVal e As System.EventArgs) Handles payrollButton.Click Payrollfinal.Show() End Sub End Class and when I run my program and press this button Private Sub Button2_Click(ByVal sender As System.Object, ByVal e As System.EventArgs) Handles exceptionsButton.Click I have my date range within a time that I know that my dataset should produce a result, but when I put a line break in my code here: adapter.Fill(ds) and look at it in debug, I show a table value of 0. If I run the same query that I have to produce these results in sql analyser, I see 1 result. Can someone see why my query on my form produces a different result than the sql analyser does? Also here is my schema for my two tables: Exceptions employeenumber varchar no 50 yes no no SQL_Latin1_General_CP1_CI_AS exceptiondate datetime no 8 yes (n/a) (n/a) NULL starttime datetime no 8 yes (n/a) (n/a) NULL endtime datetime no 8 yes (n/a) (n/a) NULL duration varchar no 50 yes no no SQL_Latin1_General_CP1_CI_AS code varchar no 50 yes no no SQL_Latin1_General_CP1_CI_AS approvedby varchar no 50 yes no no SQL_Latin1_General_CP1_CI_AS approved varchar no 50 yes no no SQL_Latin1_General_CP1_CI_AS time timestamp no 8 yes (n/a) (n/a) NULL employees employeenumber varchar no 50 no no no SQL_Latin1_General_CP1_CI_AS name varchar no 50 no no no SQL_Latin1_General_CP1_CI_AS initials varchar no 50 no no no SQL_Latin1_General_CP1_CI_AS loginname1 varchar no 50 yes no no SQL_Latin1_General_CP1_CI_AS

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  • Parse filename, insert to SQL

    - by jakesankey
    Thanks to Code Poet, I am now working off of this code to parse all .txt files in a directory and store them in a database. I need a bit more help though... The file names are R303717COMP_148A2075_20100520.txt (the middle section is unique per file). I would like to add something to code so that it can parse out the R303717COMP and put that in the left column of the database such as: (this is not the only R number we have) R303717COMP data data data R303717COMP data data data R303717COMP data data data etc Lastly, I would like to have it store each full file name into another table that gets checked so that it doesn't get processed twice.. Any Help is appreciated. using System; using System.Data; using System.Data.SQLite; using System.IO; namespace CSVImport { internal class Program { private static void Main(string[] args) { using (SQLiteConnection con = new SQLiteConnection("data source=data.db3")) { if (!File.Exists("data.db3")) { con.Open(); using (SQLiteCommand cmd = con.CreateCommand()) { cmd.CommandText = @" CREATE TABLE [Import] ( [RowId] integer PRIMARY KEY AUTOINCREMENT NOT NULL, [FeatType] varchar, [FeatName] varchar, [Value] varchar, [Actual] decimal, [Nominal] decimal, [Dev] decimal, [TolMin] decimal, [TolPlus] decimal, [OutOfTol] decimal, [Comment] nvarchar);"; cmd.ExecuteNonQuery(); } con.Close(); } con.Open(); using (SQLiteCommand insertCommand = con.CreateCommand()) { insertCommand.CommandText = @" INSERT INTO Import (FeatType, FeatName, Value, Actual, Nominal, Dev, TolMin, TolPlus, OutOfTol, Comment) VALUES (@FeatType, @FeatName, @Value, @Actual, @Nominal, @Dev, @TolMin, @TolPlus, @OutOfTol, @Comment);"; insertCommand.Parameters.Add(new SQLiteParameter("@FeatType", DbType.String)); insertCommand.Parameters.Add(new SQLiteParameter("@FeatName", DbType.String)); insertCommand.Parameters.Add(new SQLiteParameter("@Value", DbType.String)); insertCommand.Parameters.Add(new SQLiteParameter("@Actual", DbType.Decimal)); insertCommand.Parameters.Add(new SQLiteParameter("@Nominal", DbType.Decimal)); insertCommand.Parameters.Add(new SQLiteParameter("@Dev", DbType.Decimal)); insertCommand.Parameters.Add(new SQLiteParameter("@TolMin", DbType.Decimal)); insertCommand.Parameters.Add(new SQLiteParameter("@TolPlus", DbType.Decimal)); insertCommand.Parameters.Add(new SQLiteParameter("@OutOfTol", DbType.Decimal)); insertCommand.Parameters.Add(new SQLiteParameter("@Comment", DbType.String)); string[] files = Directory.GetFiles(Environment.CurrentDirectory, "TextFile*.*"); foreach (string file in files) { string[] lines = File.ReadAllLines(file); bool parse = false; foreach (string tmpLine in lines) { string line = tmpLine.Trim(); if (!parse && line.StartsWith("Feat. Type,")) { parse = true; continue; } if (!parse || string.IsNullOrEmpty(line)) { continue; } foreach (SQLiteParameter parameter in insertCommand.Parameters) { parameter.Value = null; } string[] values = line.Split(new[] {','}); for (int i = 0; i < values.Length - 1; i++) { SQLiteParameter param = insertCommand.Parameters[i]; if (param.DbType == DbType.Decimal) { decimal value; param.Value = decimal.TryParse(values[i], out value) ? value : 0; } else { param.Value = values[i]; } } insertCommand.ExecuteNonQuery(); } } } con.Close(); } } } }

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  • Migrate Spring JPA DAO unit testing to google app engine

    - by twingocerise
    I'm trying to put together a simple environment where I can get Spring, Maven, JPA, Google App Engine and DAO unit testing working happily all together. The goal is to be able to run a simple DAO unit test creating an entity and then load it again with a simple find to check it's been created properly - all of this from my maven build. My dao is making use of the JPA entity manager (query(), persist(), etc.) I've got it working no problem with hsqldb and a datasource, etc. but I'm struggling to get it working with appengine. My questions are: 1) I'm using an entity manager, injecting my persistence unit as followed. Is it OK? Is there any need for a datasource or something special? I thought not but correct me if I'm wrong. applicationContext.xml <bean id='entityManagerFactory' class='org.springframework.orm.jpa.LocalContainerEntityManagerFactoryBean'> <property name="persistenceUnitName" value="transactions-optional" /> </bean> Persistence.xml <persistence-unit name="transactions-optional"> <provider>org.datanucleus.store.appengine.jpa.DatastorePersistenceProvider</provider> <properties> <property name="datanucleus.NontransactionalRead" value="true"/> <property name="datanucleus.NontransactionalWrite" value="true"/> <property name="datanucleus.ConnectionURL" value="appengine"/> </properties> </persistence-unit> 2) what are the dependencies I need to add to my pom file to be able to run the unit test making use of the entityManager? What about versions ? I found loads of things about appengine-api-labs/stubs/testing but none them got it working i.e. I'm getting jdo dependency missing while I'm using JPA... I also get loads of conflicts when I try to add some jars (datanucleus and stuff). So far I'm trying appengine-api-1.0-sdk v1.7.0 - ASM-all v3.3 - datanucleus core/api-jpa/enhancer v3.1.0 - datanucleus-appengine v2.0.1.1 and all the gae testing jars v1.7.0 3) Is there anything I need to add to my surefire plugin (test runner) to make sure it picks up all the dependencies? I'm getting an exhausting ClassNotFound on DatastorePersistenceProvider while it is in my classpath (I checked the jars and the mvn dependency:tree) I had a look at this but it doesn't seem to be working at all: http://www.vertigrated.com/blog/2011/02/working-maven-3-google-app-engine-plugin-with-gwt-support/ 4) Do I need to use any sot of localhelper to test my DAOs? Ideally I'd want to test my dao layer "as is" with the entity manager... what's your opinion ? Has anyone managed to run a unit test using JPA on google app engine ? 5) Do I need to set up any sort of gae.home somewhere in my pom file? Would anyone make use of it (a plugin or something) ? 6) Is the gwt-maven plugin any helpful if I don't use gwt - I'm writing a simple webservice making use of appengine, not a GWT app... Any help would be much appreciated as I've been struggling for 2 days now... Cheers, V.

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  • Using Maven to Deploy to Weblogic Clusters

    - by Mark Sailes
    org.codehaus.mojo weblogic-maven-plugin 2.9.1 We're currently using the weblogic maven plugin successfully to deploy to our local WebLogic 9.2 instances. When we try to deploy to a remote environment we have a problem. We use a two machine cluster, with the admin server and managed server on one machine, and another managed server on a seperate machine. When your plugin uploads the application to the admin server, it doesn't copy it to the second managed server on the seperate machine. This then causes the second managed server a problem, as it cannot find the application in the location where the admin server saved it on its own machine. Config below <configuration> <adminServerHostName>${weblogic.adminServerHostName}</adminServerHostName> <adminServerPort>${weblogic.adminServerPort}</adminServerPort> <adminServerProtocol>${weblogic.adminServerProtocol}</adminServerProtocol> <userId>${weblogic.userId}</userId> <password>${weblogic.password}</password> <upload>${weblogic.upload}</upload> <remote>${weblogic.remote}</remote> <verbose>${weblogic.verbose}</verbose> <debug>${weblogic.debug}</debug> <stage>${weblogic.stage}</stage> <targetNames>${weblogic.targetNames}</targetNames> <exploded>${weblogic.exploded}</exploded> </configuration> <profile> <id>localhost</id> <properties> <weblogic.adminServerHostName>localhost</weblogic.adminServerHostName> <weblogic.adminServerPort>7001</weblogic.adminServerPort> <weblogic.adminServerProtocol>t3</weblogic.adminServerProtocol> <weblogic.userId>weblogic</weblogic.userId> <weblogic.password>weblogic</weblogic.password> <weblogic.upload>false</weblogic.upload> <weblogic.remote>false</weblogic.remote> <weblogic.verbose>true</weblogic.verbose> <weblogic.debug>true</weblogic.debug> <weblogic.stage>false</weblogic.stage> <weblogic.targetNames>AdminServer</weblogic.targetNames> <weblogic.exploded>false</weblogic.exploded> </properties> </profile> <profile> <id>dev</id> <properties> <weblogic.adminServerHostName>******</weblogic.adminServerHostName> <weblogic.adminServerPort>9141</weblogic.adminServerPort> <weblogic.adminServerProtocol>t3</weblogic.adminServerProtocol> <weblogic.userId>******</weblogic.userId> <weblogic.password>******</weblogic.password> <weblogic.upload>true</weblogic.upload> <weblogic.remote>true</weblogic.remote> <weblogic.verbose>true</weblogic.verbose> <weblogic.debug>true</weblogic.debug> <weblogic.stage>true</weblogic.stage> <weblogic.targetNames>dev_cluster01</weblogic.targetNames> <weblogic.exploded>false</weblogic.exploded> </properties> </profile>

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  • What version-control system is most trivial to set up and use for toy projects?

    - by Norman Ramsey
    I teach the third required intro course in a CS department. One of my homework assignments asks students to speed up code they have written for a previous assignment. Factor-of-ten speedups are routine; factors of 100 or 1000 are not unheard of. (For a factor of 1000 speedup you have to have made rookie mistakes with malloc().) Programs are improved by a sequence is small changes. I ask students to record and describe each change and the resulting improvement. While you're improving a program it is also possible to break it. Wouldn't it be nice to back out? You can see where I'm going with this: my students would benefit enormously from version control. But there are some caveats: Our computing environment is locked down. Anything that depends on a central repository is suspect. Our students are incredibly overloaded. Not just classes but jobs, sports, music, you name it. For them to use a new tool it has to be incredibly easy and have obvious benefits. Our students do most work in pairs. Getting bits back and forth between accounts is problematic. Could this problem also be solved by distributed version control? Complexity is the enemy. I know setting up a CVS repository is too baffling---I myself still have trouble because I only do it once a year. I'm told SVN is even harder. Here are my comments on existing systems: I think central version control (CVS or SVN) is ruled out because our students don't have the administrative privileges needed to make a repository that they can share with one other student. (We are stuck with Unix file permissions.) Also, setup on CVS or SVN is too hard. darcs is way easy to set up, but it's not obvious how you share things. darcs send (to send patches by email) seems promising but it's not clear how to set it up. The introductory documentation for git is not for beginners. Like CVS setup, it's something I myself have trouble with. I'm soliciting suggestions for what source-control to use with beginning students. I suspect we can find resources to put a thin veneer over an existing system and to simplify existing documentation. We probably don't have resources to write new documentation. So, what's really easy to setup, commit, revert, and share changes with a partner but does not have to be easy to merge or to work at scale? A key constraint is that programming pairs have to be able to share work with each other and only each other, and pairs change every week. Our infrastructure is Linux, Solaris, and Windows with a netapp filer. I doubt my IT staff wants to create a Unix group for each pair of students. Is there an easier solution I've overlooked? (Thanks for the accepted answer, which beats the others on account of its excellent reference to Git Magic as well as the helpful comments.)

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  • Java, LDAP: Make it not ignore blank passwords?

    - by Steve
    I'm maintaining some legacy Java LDAP code. I know next to nothing about LDAP. The program below basically just sends the userid and password to the LDAP server, receives notification back if the credentials are good. If so, it prints out the LDAP attributes received from the LDAP server, if not it prints out an exception. All works well if a bad password is given. An "invalid credentials" exception gets thrown. However, if a blank password is sent to the LDAP Server, authentication will still happen, LDAP attributes will still be returned. Is this unhappy situation due to the LDAP server allowing blank passwords, or does the code below need to be adjusted such a blank password will get fed to the LDAP server in such a way so it will get rejected? I do have data validation in place. I took it off in a testing environment to solve another issue and noticed this problem. I would prefer not to have this problem underneath the data validation. Thanks much in advance for any information import javax.naming.*; import javax.naming.directory.*; import java.util.*; import java.sql.*; public class LDAPTEST { public static void main(String args[]) { String lcf = "com.sun.jndi.ldap.LdapCtxFactory"; String ldapurl = "ldaps://ldap-cit.smew.acme.com:636/o=acme.com"; String loginid = "George.Jetson"; String password = ""; DirContext ctx = null; Hashtable env = new Hashtable(); Attributes attr = null; Attributes resultsAttrs = null; SearchResult result = null; NamingEnumeration results = null; int iResults = 0; int iAttributes = 0; env.put(Context.INITIAL_CONTEXT_FACTORY, lcf); env.put(Context.PROVIDER_URL, ldapurl); env.put(Context.SECURITY_PROTOCOL, "ssl"); env.put(Context.SECURITY_AUTHENTICATION, "simple"); env.put(Context.SECURITY_PRINCIPAL, "uid=" + loginid + ",ou=People,o=acme.com"); env.put(Context.SECURITY_CREDENTIALS, password); try { ctx = new InitialDirContext(env); attr = new BasicAttributes(true); attr.put(new BasicAttribute("uid",loginid)); results = ctx.search("ou=People",attr); while (results.hasMore()) { result = (SearchResult)results.next(); resultsAttrs = result.getAttributes(); for (NamingEnumeration enumAttributes = resultsAttrs.getAll(); enumAttributes.hasMore();) { Attribute a = (Attribute)enumAttributes.next(); System.out.println("attribute: " + a.getID() + " : " + a.get().toString()); iAttributes++; }// end for loop iResults++; }// end while loop System.out.println("Records == " + iResults + " Attributes: " + iAttributes); }// end try catch (Exception e) { e.printStackTrace(); } }// end function main() }// end class LDAPTEST

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  • Dot Net Nuke module works in "Edit" mode but not for "View": cache problem?

    - by Godeke
    I have a DNN task that simply runs some Javascript to compute a price based on a few input fields. This module works fine on our production site, but we had a company do a skin for us to improve the look of the site and the module fails under this new system. (DNN 05.06.00 (459) although it was 5.5 prior... I updated in a futile hope that it was a bug in the old revision.) What is incredibly odd about this is that the module works fine when I'm logged in to DNN and using the Edit mode as an administrator. In this case the small snippet of JavaScript loads fine and filling the fields results in a price. On the other hand it I click "View" (or more importantly, if I'm not logged in at all) the page loads a cached copy. Even odder, I have found the cache files in \Portals\2\Cache\Pages are generated and then only the cached data is being used. When the cached copy is loaded, the JavaScript doesn't appear (it is normally created via a Page.ClientScript.RegisterClientScriptBlock(). Additionally, the button which posts the data to the server doesn't execute any of the server side code (confirmed with a debugger) but instead just reloads the cached copy. If I manually delete the files in \Portals\2\Cache\Pages then everything works properly, but I have to do so after every page load: failing to do so simply loads the page as it was last generated repeatedly. Resetting the application (either via the UI or editing web.config) doesn't change this and clearing the cache from the Host Settings page doesn't actually clear these cached pages. I'm guessing that Edit mode bypasses the cache in some way, but I have gone as far as turning off all caching on the site (which is horrible for performance) and the cached version is still loaded. Has anyone seen anything like this? Shouldn't clearing the cache clear the files (I'm using the File provider for caching)? Shouldn't even a cached page go back to the server if the user posts back? EDIT: I should point out that permissions don't appear to be a problem on the cache directory... other pages cached output are deleted from this folder, just this page has this issue. EDIT 2: Clarifying some settings and conditions which I didn't provide. First, this module works fine in production under DNN 5.6.0. In our test environment with the consulting company's changes it fails (the changes are skin and page layout only in theory: the module source itself verifies as unchanged). All cache settings and the like have been verified the same between the two and we only resorted to setting the module cache to 0 and -1 (and disabling the test site's cache entirely) when we couldn't find another cause for the problem. I have watched the cache work correctly on many other pages in test: there is something about this page that is causing the problem. We have punted and are creating an installable skin based on the consultant's work as I suspect they have somehow corrupted the DNN install (database side I think).

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  • clock and date showing on a live site but not on localhost

    - by grumpypanda
    I've got clock.swf and date.swf working fine on a live site, now I am using the same code to set up a local develop environment. Everything is working well except the clock.swf and date.swf stopped working on localhost. Two same yellow errors "You need to update your Flash plugin. Click here if you want to continue." but of course my Flash player is up to date since the live site is working fine. I'll post the code below which I think has caused the error. I've been searching online for the last couple of hours but no luck, anyone has got into an issue like this before? What can be the possible cause? Any help is appreciated. This is on the index.php, I can post more code here if needed. <?php embed_flash("swf/clock.swf", CLOCK_WIDTH, CLOCK_HEIGHT, "8", '', "flashcontent");?> <?php embed_flash("swf/date.swf", DATE_WIDTH, DATE_HEIGHT, "8", '', "flashcontent_date");?> configure.php define('CLOCK_WIDTH', '450'); define('CLOCK_HEIGHT', ''); define('DATE_WIDTH', '440'); define('DATE_HEIGHT', ''); flash_function.php <?php function embed_flash($name, $w, $h, $version, $bgcolor, $id) { $cacheBuster = rand(); $padTop = $h/3; ?> <style> a.noflash:link, a.noflash:visited, a.noflash:active {color: #1860C2; text-decoration: none; background:#FFFFFF;} a.noflash:hover {color:#000; text-decoration:none; background:#EEEEEE;} .message { width: <?=$w;?>px; font-size:12px; font-weight:normal; margin-bottom: 10px; padding: 5px; color: #EEE; background: orange;"} </style> <div id="<?=$id; ?>" align="center"> <noscript> <div class="message"> Please enable <a href="https://www.google.com/support/adsense/bin/answer.py?answer=12654" target="_blank" class="noflash">&nbsp;JavaScript&nbsp;</a> to view this page properly. </div> </noscript> <div class="message"> You need to update your Flash plugin. Click <a href="http://www.adobe.com/shockwave/download/download.cgi?P1_Prod_Version=ShockwaveFlash&promoid=BIOW" target="_blank" class="noflash">&nbsp;here&nbsp;</a> if you want to continue. </div> </div> <script type="text/javascript"> // <![CDATA[ var so = new SWFObject("<?=$name;?>", "", "<?=$w;?>", "<?=$h;?>", "<?=$version;?>", "<?=$bgcolor;?>"); so.addParam("quality", "high"); so.addParam("allowScriptAccess", "sameDomain"); so.addParam("scale", "showall"); so.addParam("loop", "false"); so.addParam("wmode", "transparent"); so.write("<?=$id;?>"); // ]]> </script>

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  • Accessing Layout Items from inside Widget AppWidgetProvider

    - by cam4mav
    I am starting to go insane trying to figure this out. It seems like it should be very easy, I'm starting to wonder if it's possible. What I am trying to do is create a home screen widget, that only contains an ImageButton. When it is pressed, the idea is to change some setting (like the wi-fi toggle) and then change the Buttons image. I have the ImageButton declared like this in my main.xml <ImageButton android:id="@+id/buttonOne" android:src="@drawable/button_normal_ringer" android:layout_width="wrap_content" android:layout_height="wrap_content" android:layout_gravity="center" /> my AppWidgetProvider class, named ButtonWidget * note that the RemoteViews class is a locally stored variable. this allowed me to get access to the RViews layout elements... or so I thought. @Override public void onUpdate(Context context, AppWidgetManager appWidgetManager, int[] appWidgetIds) { remoteViews = new RemoteViews(context.getPackageName(), R.layout.main); Intent active = new Intent(context, ButtonWidget.class); active.setAction(VIBRATE_UPDATE); active.putExtra("msg","TESTING"); PendingIntent actionPendingIntent = PendingIntent.getBroadcast(context, 0, active, 0); remoteViews.setOnClickPendingIntent(R.id.buttonOne, actionPendingIntent); appWidgetManager.updateAppWidget(appWidgetIds, remoteViews); } @Override public void onReceive(Context context, Intent intent) { // v1.5 fix that doesn't call onDelete Action final String action = intent.getAction(); Log.d("onReceive",action); if (AppWidgetManager.ACTION_APPWIDGET_DELETED.equals(action)) { final int appWidgetId = intent.getExtras().getInt( AppWidgetManager.EXTRA_APPWIDGET_ID, AppWidgetManager.INVALID_APPWIDGET_ID); if (appWidgetId != AppWidgetManager.INVALID_APPWIDGET_ID) { this.onDeleted(context, new int[] { appWidgetId }); } } else { // check, if our Action was called if (intent.getAction().equals(VIBRATE_UPDATE)) { String msg = "null"; try { msg = intent.getStringExtra("msg"); } catch (NullPointerException e) { Log.e("Error", "msg = null"); } Log.d("onReceive",msg); if(remoteViews != null){ Log.d("onReceive",""+remoteViews.getLayoutId()); remoteViews.setImageViewResource(R.id.buttonOne, R.drawable.button_pressed_ringer); Log.d("onReceive", "tried to switch"); } else{ Log.d("F!", "--naughty language used here!!!--"); } } super.onReceive(context, intent); } } so, I've been testing this and the onReceive method works great, I'm able to send notifications and all sorts of stuff (removed from code for ease of reading) the one thing I can't do is change any properties of the view elements. To try and fix this, I made RemoteViews a local and static private variable. Using log's I was able to see that When multiple instances of the app are on screen, they all refer to the one instance of RemoteViews. perfect for what I'm trying to do The trouble is in trying to change the image of the ImageButton. I can do this from within the onUpdate method using this. remoteViews.setImageViewResource(R.id.buttonOne, R.drawable.button_pressed_ringer); that doesn't do me any good though once the widget is created. For some reason, even though its inside the same class, being inside the onReceive method makes that line not work. That line used to throw a Null pointer as a matter of fact, until I changed the variable to static. now it passes the null test, refers to the same layoutId as it did at the start, reads the line, but it does nothing. Its like the code isn't even there, just keeps chugging along. SO...... Is there any way to modify layout elements from within a widget after the widget has been created!? I want to do this based on the environment, not with a configuration activity launch. I've been looking at various questions and this seems to be an issue that really hasn't been solved, such as link text and link text oh and for anyone who finds this and wants a good starting tutorial for widgets, this is easy to follow (though a bit old, it gets you comfortable with widgets) .pdf link text hopefully someone can help here. I kinda have the feeling that this is illegal and there is a different way to go about this. I would LOVE to be told another approach!!!! Thanks

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  • How do I detect a file write error in C?

    - by rich
    I have an embedded environment where a user might insert or remove a USB flash drive. I would like to know if the drive has been removed, or if there is some other problem when I try to write to the drive. However, Linux just saves the information in its buffers and returns with no indicated error. The computer I'm using comes with a 2.4.26 kernel and libc 2.3.2. I'm mounting the drive this way: i = mount(MEMORY_DEV_PATH, MEMORY_MNT_PATH, "vfat", MS_SYNCHRONOUS, NULL); That works: 50:/root # mount /dev/scsi/host0/bus0/target0/lun0/part1 on /mem type vfat (rw,sync) 50:/root # Later, I try to copy a file to it: int ifile, ofile; ifile = open("/tmp/tmpmidi.mid", O_RDONLY); if (ifile < 0) { perror("open in"); break; } ofile = open(current_file_name.c_str(), O_WRONLY | O_SYNC); if (ofile < 0) { perror("open out"); break; } #define BUFSZ 256 char buffer[BUFSZ]; while (1) { i = read(ifile, buffer, BUFSZ); if (i < 0) { perror("read"); break; } j = write(ofile, buffer, i); if (j < 0) { perror("write"); break; } if (i != j) { perror("Sizes wrong"); break; } if (i < BUFSZ) { printf("Copy is finished, I hope\n"); close(ifile); close(ofile); break; } } If this snippet of code is executed with a write-protected USB memory, the result is Copy is finished, I hope amid a flurry of error messages from the kernel on the console. I believe the same thing would happen if I simply removed the USB drive (without unmounting it). I have also fiddled with devfs. I figured out how to get it to automatically mount the drive, (with the REGISTER event) but it never seems to trigger the UNREGISTER when I pull out the memory. How can I determine in my program whether I have successfully created a file? Update 4 July: It was a silly oversight of me not to check the result from close(). Unfortunately, the file can be closed without error. So that didn't help. What about fsync()? That sounds like a good idea, but that didn't catch the error either. There might be some interesting information in /sys if I had such a thing. I believe that didn't get added until 2.6.?. The comment(s) about the quality of my flash drive are probably justified. It's one of the earlier ones. In fact, write protect switches seem to be extremely rare these days. I think I have to use the overkill option: Create a file, unmount & remount the drive, and check to see if the file is there. If that doesn't solve my problem, then something is really messed up! Note to myself: Make sure the file you try to create isn't already there! By the way, this does happen to be a C++ program. You can tell by the .c_str() which I had intended to edit out for simplicity.

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  • WCF/REST Get image into picturebox?

    - by Garrith
    So I have wcf rest service which succesfuly runs from a console app, if I navigate to: http://localhost:8000/Service/picture/300/400 my image is displayed note the 300/400 sets the width and height of the image within the body of the html page. The code looks like this: namespace WcfServiceLibrary1 { [ServiceContract] public interface IReceiveData { [OperationContract] [WebInvoke(Method = "GET", BodyStyle = WebMessageBodyStyle.Wrapped, ResponseFormat = WebMessageFormat.Xml, UriTemplate = "picture/{width}/{height}")] Stream GetImage(string width, string height); } public class RawDataService : IReceiveData { public Stream GetImage(string width, string height) { int w, h; if (!Int32.TryParse(width, out w)) { w = 640; } // Handle error if (!Int32.TryParse(height, out h)) { h = 400; } Bitmap bitmap = new Bitmap(w, h); for (int i = 0; i < bitmap.Width; i++) { for (int j = 0; j < bitmap.Height; j++) { bitmap.SetPixel(i, j, (Math.Abs(i - j) < 2) ? Color.Blue : Color.Yellow); } } MemoryStream ms = new MemoryStream(); bitmap.Save(ms, System.Drawing.Imaging.ImageFormat.Jpeg); ms.Position = 0; WebOperationContext.Current.OutgoingResponse.ContentType = "image/jpeg"; return ms; } } } What I want to do now is use a client application "my windows form app" and add that image into a picturebox. Im abit stuck as to how this can be achieved as I would like the width and height of the image from my wcf rest service to be set by the width and height of the picturebox. I have tryed this but on two of the lines have errors and im not even sure if it will work as the code for my wcf rest service seperates width and height with a "/" if you notice in the url. string uri = "http://localhost:8080/Service/picture"; private void button1_Click(object sender, EventArgs e) { StringBuilder sb = new StringBuilder(); sb.AppendLine("<picture>"); sb.AppendLine("<width>" + pictureBox1.Image.Width + "</width>"); // the url looks like this http://localhost:8080/Service/picture/300/400 when accessing the image so I am trying to set this here sb.AppendLine("<height>" + pictureBox1.Image.Height + "</height>"); sb.AppendLine("</picture>"); string picture = sb.ToString(); byte[] getimage = Encoding.UTF8.GetBytes(picture); // not sure this is right HttpWebRequest req = WebRequest.Create(uri); //cant convert webrequest to httpwebrequest req.Method = "GET"; req.ContentType = "image/jpg"; req.ContentLength = getimage.Length; MemoryStream reqStrm = req.GetRequestStream(); //cant convert IO stream to IO Memory stream reqStrm.Write(getimage, 0, getimage.Length); reqStrm.Close(); HttpWebResponse resp = req.GetResponse(); // cant convert web respone to httpwebresponse MessageBox.Show(resp.StatusDescription); pictureBox1.Image = Image.FromStream(reqStrm); reqStrm.Close(); resp.Close(); } So just wondering if some one could help me out with this futile attempt at adding a variable image size from my rest service to a picture box on button click. This is the host app aswell: namespace ConsoleApplication1 { class Program { static void Main(string[] args) { string baseAddress = "http://" + Environment.MachineName + ":8000/Service"; ServiceHost host = new ServiceHost(typeof(RawDataService), new Uri(baseAddress)); host.AddServiceEndpoint(typeof(IReceiveData), new WebHttpBinding(), "").Behaviors.Add(new WebHttpBehavior()); host.Open(); Console.WriteLine("Host opened"); Console.ReadLine();

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  • Same source, multiple targets with different resources (Visual Studio .Net 2008)

    - by Mike Bell
    A set of software products differ only by their resource strings, binary resources, and by the strings / graphics / product keys used by their Visual Studio Setup projects. What is the best way to create, organize, and maintain them? i.e. All the products essentially consist of the same core functionality customized by graphics, strings, and other resource data to form each product. Imagine you are creating a set of products like "Excel for Bankers", Excel for Gardeners", "Excel for CEOs", etc. Each product has the the same functionality, but differs in name, graphics, help files, included templates etc. The environment in which these are being built is: vanilla Windows.Forms / Visual Studio 2008 / C# / .Net. The ideal solution would be easy to maintain. e.g. If I introduce a new string / new resource projects I haven't added the resource to should fail at compile time, not run time. (And subsequent localization of the products should also be feasible). Hopefully I've missed the blindingly-obvious and easy way of doing all this. What is it? ============ Clarification(s) ================ By "product" I mean the package of software that gets installed by the installer and sold to the end user. Currently I have one solution, consisting of multiple projects, (including a Setup project), which builds a set of assemblies and create a single installer. What I need to produce are multiple products/installers, all with similar functionality, which are built from the same set of assemblies but differ in the set of resources used by one of the assemblies. What's the best way of doing this? ------------ The 95% Solution ----------------- Based upon Daminen_the_unbeliever's answer, a resource file per configuration can be achieved as follows: Create a class library project ("Satellite"). Delete the default .cs file and add a folder ("Default") Create a resource file in the folder "MyResources" Properties - set CustomToolNamespace to something appropriate (e.g. "XXX") Make sure the access modifier for the resources is "Public". Add the resources. Edit the source code. Refer to the resources in your code as XXX.MyResources.ResourceName) Create Configurations for each product variant ("ConfigN") For each product variant, create a folder ("VariantN") Copy and Paste the MyResources file into each VariantN folder Unload the "Satellite" project, and edit the .csproj file For each "VariantN/MyResources" <Compile> or <EmbeddedResource> tag, add a Condition="'$(Configuration)' == 'ConfigN'" attribute. Save, Reload the .csproj, and you're done... This creates a per-configuration resource file, which can (presumably) be further localized. Compile error messages are produced for any configuration that where a a resource is missing. The resource files can be localized using the standard method (create a second resources file (MyResources.fr.resx) and edit .csproj as before). The reason this is a 95% solution is that resources used to initialize forms (e.g. Form Titles, button texts) can't be easily handled in the same manner - the easiest approach seems to be to overwrite these with values from the satellite assembly.

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  • Please clarify how create/update happens against child entities of an aggregate root

    - by christian
    After much reading and thinking as I begin to get my head wrapped around DDD, I am a bit confused about the best practices for dealing with complex hierarchies under an aggregate root. I think this is a FAQ but after reading countless examples and discussions, no one is quite talking about the issue I'm seeing. If I am aligned with the DDD thinking, entities below the aggregate root should be immutable. This is the crux of my trouble, so if that isn't correct, that is why I'm lost. Here is a fabricated example...hope it holds enough water to discuss. Consider an automobile insurance policy (I'm not in insurance, but this matches the language I hear when on the phone w/ my insurance company). Policy is clearly an entity. Within the policy, let's say we have Auto. Auto, for the sake of this example, only exists within a policy (maybe you could transfer an Auto to another policy, so this is potential for an aggregate as well, which changes Policy...but assume it simpler than that for now). Since an Auto cannot exist without a Policy, I think it should be an Entity but not a root. So Policy in this case is an aggregate root. Now, to create a Policy, let's assume it has to have at least one auto. This is where I get frustrated. Assume Auto is fairly complex, including many fields and maybe a child for where it is garaged (a Location). If I understand correctly, a "create Policy" constructor/factory would have to take as input an Auto or be restricted via a builder to not be created without this Auto. And the Auto's creation, since it is an entity, can't be done beforehand (because it is immutable? maybe this is just an incorrect interpretation). So you don't get to say new Auto and then setX, setY, add(Z). If Auto is more than somewhat trivial, you end up having to build a huge hierarchy of builders and such to try to manage creating an Auto within the context of the Policy. One more twist to this is later, after the Policy is created and one wishes to add another Auto...or update an existing Auto. Clearly, the Policy controls this...fine...but Policy.addAuto() won't quite fly because one can't just pass in a new Auto (right!?). Examples say things like Policy.addAuto(VIN, make, model, etc.) but are all so simple that that looks reasonable. But if this factory method approach falls apart with too many parameters (the entire Auto interface, conceivably) I need a solution. From that point in my thinking, I'm realizing that having a transient reference to an entity is OK. So, maybe it is fine to have a entity created outside of its parent within the aggregate in a transient environment, so maybe it is OK to say something like: auto = AutoFactory.createAuto(); auto.setX auto.setY or if sticking to immutability, AutoBuilder.new().setX().setY().build() and then have it get sorted out when you say Policy.addAuto(auto) This insurance example gets more interesting if you add Events, such as an Accident with its PolicyReports or RepairEstimates...some value objects but most entities that are all really meaningless outside the policy...at least for my simple example. The lifecycle of Policy with its growing hierarchy over time seems the fundamental picture I must draw before really starting to dig in...and it is more the factory concept or how the child entities get built/attached to an aggregate root that I haven't seen a solid example of. I think I'm close. Hope this is clear and not just a repeat FAQ that has answers all over the place.

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  • Which of CouchDB or MongoDB suits my needs?

    - by vonconrad
    Where I work, we use Ruby on Rails to create both backend and frontend applications. Usually, these applications interact with the same MySQL database. It works great for a majority of our data, but we have one situation which I would like to move to a NoSQL environment. We have clients, and our clients have what we call "inventories"--one or more of them. An inventory can have many thousands of items. This is currently done through two relational database tables, inventories and inventory_items. The problems start when two different inventories have different parameters: # Inventory item from inventory 1, televisions { inventory_id: 1 sku: 12345 name: Samsung LCD 40 inches model: 582903-4 brand: Samsung screen_size: 40 type: LCD price: 999.95 } # Inventory item from inventory 2, accomodation { inventory_id: 2 sku: 48cab23fa name: New York Hilton accomodation_type: hotel star_rating: 5 price_per_night: 395 } Since we obviously can't use brand or star_rating as the column name in inventory_items, our solution so far has been to use generic column names such as text_a, text_b, float_a, int_a, etc, and introduce a third table, inventory_schemas. The tables now look like this: # Inventory schema for inventory 1, televisions { inventory_id: 1 int_a: sku text_a: name text_b: model text_c: brand int_b: screen_size text_d: type float_a: price } # Inventory item from inventory 1, televisions { inventory_id: 1 int_a: 12345 text_a: Samsung LCD 40 inches text_b: 582903-4 text_c: Samsung int_a: 40 text_d: LCD float_a: 999.95 } This has worked well... up to a point. It's clunky, it's unintuitive and it lacks scalability. We have to devote resources to set up inventory schemas. Using separate tables is not an option. Enter NoSQL. With it, we could let each and every item have their own parameters and still store them together. From the research I've done, it certainly seems like a great alterative for this situation. Specifically, I've looked at CouchDB and MongoDB. Both look great. However, there are a few other bits and pieces we need to be able to do with our inventory: We need to be able to select items from only one (or several) inventories. We need to be able to filter items based on its parameters (eg. get all items from inventory 2 where type is 'hotel'). We need to be able to group items based on parameters (eg. get the lowest price from items in inventory 1 where brand is 'Samsung'). We need to (potentially) be able to retrieve thousands of items at a time. We need to be able to access the data from multiple applications; both backend (to process data) and frontend (to display data). Rapid bulk insertion is desired, though not required. Based on the structure, and the requirements, are either CouchDB or MongoDB suitable for us? If so, which one will be the best fit? Thanks for reading, and thanks in advance for answers. EDIT: One of the reasons I like CouchDB is that it would be possible for us in the frontend application to request data via JavaScript directly from the server after page load, and display the results without having to use any backend code whatsoever. This would lead to better page load and less server strain, as the fetching/processing of the data would be done client-side.

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  • How to copy bytes from buffer into the managed struct?

    - by Chupo_cro
    I have a problem with getting the code to work in a managed environment (VS2008 C++/CLI Win Forms App). The problem is I cannot declare the unmanaged struct (is that even possible?) inside the managed code, so I've declared a managed struct but now I have a problem how to copy bytes from buffer into that struct. Here is the pure C++ code that obviously works as expected: typedef struct GPS_point { float point_unknown_1; float latitude; float longitude; float altitude; // x10000 float time; int point_unknown_2; int speed; // x100 int manually_logged_point; // flag (1 --> point logged manually) } track_point; int offset = 0; int filesize = 256; // simulates filesize int point_num = 10; // simulates number of records int main () { char *buffer_dyn = new char[filesize]; // allocate RAM // here, the file would have been read into the buffer buffer_dyn[0xa8] = 0x1e; // simulates the speed data (1e 00 00 00) buffer_dyn[0xa9] = 0x00; buffer_dyn[0xaa] = 0x00; buffer_dyn[0xab] = 0x00; offset = 0x90; // if the data with this offset is transfered trom buffer // to struct, int speed is alligned with the buffer at the // offset of 0xa8 track_point *points = new track_point[point_num]; points[0].speed = 0xff; // (debug) it should change into 0x1e memcpy(&points[0],buffer_dyn+offset,32); cout << "offset: " << offset << "\r\n"; //cout << "speed: " << points[0].speed << "\r\n"; printf ("speed : 0x%x\r\n",points[0].speed); printf("byte at offset 0xa8: 0x%x\r\n",(unsigned char)buffer_dyn[0xa8]); // should be 0x1e delete[] buffer_dyn; // release RAM delete[] points; /* What I need is to rewrite the lines 29 and 31 to work in the managed code (VS2008 Win Forms C++/CLI) What should I have after: array<track_point^>^ points = gcnew array<track_point^>(point_num); so I can copy 32 bytes from buffer_dyn to the managed struct declared as typedef ref struct GPS_point { float point_unknown_1; float latitude; float longitude; float altitude; // x10000 float time; int point_unknown_2; int speed; // x100 int manually_logged_point; // flag (1 --> point logged manually) } track_point; */ return 0; } Here is the paste to codepad.org so it can be seen the code is OK. What I need is to rewrite these two lines: track_point *points = new track_point[point_num]; memcpy(&points[0],buffer_dyn+offset,32); to something that will work in a managed application. I wrote: array<track_point^>^ points = gcnew array<track_point^>(point_num); and now trying to reproduce the described copying of the data from buffer over the struct, but haven't any idea how it should be done. Alternatively, if there is a way to use an unmanaged struct in the same way shown in my code, then I would like to avoid working with managed struct.

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  • ActionScript 3 Cant see Movieclip

    - by user3697993
    When I play my game it does not show my _Player Movieclip, but it does collide with the ground which is very confusing. So I believe the movieclip is there but not showing the texture/Sprite. I think the problem is in "function Spawn" (First Function). public class PewdyBird extends MovieClip { //Player variables public var Up_Speed:int = 25; public var speed:Number = 0; public var _grav:Number = 0.5; public var isJump:Boolean = false; public var Score:int = 0; public var Player_Live:Boolean = true; public var _Player:Player = new Player(); //Other variables //Environment variables var Floor:int = 480; var Clock:Number = 0; var Clock_restart:Number = 0; var Clock_ON:Boolean = false; var Clock_max:int = 15; var Player_Stage:Boolean = true; private var _X:int; private var _Y:int; private var hit_ground:Boolean = false; private var width_BG:int = 479; //SPAWN function Spawn(e:Event){ _Player.x = 200; _Player.y = 200; stage.addChild(_Player); } //Keyboard Input private function KeyboardListener(e:KeyboardEvent){ if(e.keyCode == Keyboard.SPACE){ Clock = Clock_restart; Clock_ON = true; isJump = true; if(isJump){ _Player.gotoAndPlay("Fly"); speed = -Up_Speed; isJump = false; } } } //Mouse Input & Spawn Listener private function MouseListener(m:MouseEvent){ if(MouseEvent.CLICK){ Clock = Clock_restart; Clock_ON = true; isJump = true; if(isJump){ _Player.gotoAndPlay("Fly"); speed = -Up_Speed; isJump = false; } } } //Rotation Fly function Rot_Fly(){ if(Clock < Clock_max){ _Player.rotation = -15; }else if(Clock >= Clock_max){ if(_Player.rotation < 90){ _Player.rotation += 10; }else if(_Player.rotation >= 90){ _Player.rotation = 90; } } } //END //Update Function function enter_frame(e:Event):void{ Rot_Fly(); //Clock if(Clock_ON){ Clock++; }else if(Clock > Clock_max){ Clock = Clock_max; } //Fall Limits if(speed >= 20){ _Player.y += 20; return; _Player.gotoAndPlay("Fall"); } //Physics speed += _grav*3; _Player.y += speed; } //Hit Ground function Hit_Ground(e:Event){ if(_Player.hitTestObject(Ground1)){ _grav = 0; speed = 0; trace("HIT GROUND"); }else if(_Player.hitTestObject(Ground2)){ _grav = 0; speed = 0; trace("HIT GROUND"); }else if(_Player.hitTestObject(Ground1) == false){ _grav = 1; }else if(_Player.hitTestObject(Ground2) == false){ _grav = 1; } } //Background Slide (Left) private function Background_Move(e:Event):void{ Background1.x -= 1.5; Background2.x -= 1.5; Ground1.x -= 4; Ground2.x -= 4; if(Background1.x < -width_BG){ Background1.x = width_BG; } else if(Background2.x < -width_BG){ Background2.x = width_BG; } else if(Ground1.x < -width_BG){ Ground1.x = width_BG; } else if(Ground2.x < -width_BG){ Ground2.x = width_BG; } } } The eventListeners are in flash it self stage.addEventListener(Event.ENTER_FRAME, enter_frame); stage.addEventListener(Event.ENTER_FRAME, Hit_Ground); stage.addEventListener(KeyboardEvent.KEY_UP, KeyboardListener); stage.addEventListener(MouseEvent.CLICK, MouseListener); stage.addEventListener(Event.ENTER_FRAME, Background_Move); stage.addEventListener(Event.ADDED_TO_STAGE, Spawn);

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  • Doesn't get the output in Java Database Connectivity

    - by Dooree
    I'm working on Java Database Connectivity through Eclipse IDE. I built a database through Ubuntu Terminal, and I need to connect and work with it. However, when I tried to run the following code, I don't get any error, but the following output is showed, anybody knows why I don't get the output from the code ? //STEP 1. Import required packages import java.sql.*; public class FirstExample { // JDBC driver name and database URL static final String JDBC_DRIVER = "com.mysql.jdbc.Driver"; static final String DB_URL = "jdbc:mysql://localhost/EMP"; // Database credentials static final String USER = "username"; static final String PASS = "password"; public static void main(String[] args) { Connection conn = null; Statement stmt = null; try{ //STEP 2: Register JDBC driver Class.forName("com.mysql.jdbc.Driver"); //STEP 3: Open a connection System.out.println("Connecting to database..."); conn = DriverManager.getConnection(DB_URL,USER,PASS); //STEP 4: Execute a query System.out.println("Creating statement..."); stmt = conn.createStatement(); String sql; sql = "SELECT id, first, last, age FROM Employees"; ResultSet rs = stmt.executeQuery(sql); //STEP 5: Extract data from result set while(rs.next()){ //Retrieve by column name int id = rs.getInt("id"); int age = rs.getInt("age"); String first = rs.getString("first"); String last = rs.getString("last"); //Display values System.out.print("ID: " + id); System.out.print(", Age: " + age); System.out.print(", First: " + first); System.out.println(", Last: " + last); } //STEP 6: Clean-up environment rs.close(); stmt.close(); conn.close(); }catch(SQLException se){ //Handle errors for JDBC se.printStackTrace(); }catch(Exception e){ //Handle errors for Class.forName e.printStackTrace(); }finally{ //finally block used to close resources try{ if(stmt!=null) stmt.close(); }catch(SQLException se2){ }// nothing we can do try{ if(conn!=null) conn.close(); }catch(SQLException se){ se.printStackTrace(); }//end finally try }//end try System.out.println("Goodbye!"); }//end main }//end FirstExample <ConnectionProperties> <PropertyCategory name="Connection/Authentication"> <Property name="user" required="No" default="" sortOrder="-2147483647" since="all"> The user to connect as </Property> <Property name="password" required="No" default="" sortOrder="-2147483646" since="all"> The password to use when connecting </Property> <Property name="socketFactory" required="No" default="com.mysql.jdbc.StandardSocketFactory" sortOrder="4" since="3.0.3"> The name of the class that the driver should use for creating socket connections to the server. This class must implement the interface 'com.mysql.jdbc.SocketFactory' and have public no-args constructor. </Property> <Property name="connectTimeout" required="No" default="0" sortOrder="9" since="3.0.1"> Timeout for socket connect (in milliseconds), with 0 being no timeout. Only works on JDK-1.4 or newer. Defaults to '0'. </Property> ...

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  • about getadrrinfo() C++?

    - by Isavel
    I'm reading this book called beej's guide to network programming and there's a part in the book were it provide a sample code which illustrate the use of getaddrinfo(); the book state that the code below "will print the IP addresses for whatever host you specify on the command line" - beej's guide to network programming. now I'm curious and want to try it out and run the code, but I guess the code was develop in UNIX environment and I'm using visual studio 2012 windows 7 OS, and most of the headers was not supported so I did a bit of research and find out that I need to include the winsock.h and ws2_32.lib for windows, for it to get working, fortunately everything compiled no errors, but when I run it using the debugger and put in 'www.google.com' as command argument I was disappointed that it did not print any ipaddress, the output that I got from the console is "getaddrinfo: E" what does the letter E mean? Do I need to configure something out of the debugger? Interestingly I left the command argument blank and the output changed to "usage: showip hostname" Any help would be appreciated. #ifdef _WIN32 #endif #include <sys/types.h> #include <winsock2.h> #include <ws2tcpip.h> #include <iostream> using namespace std; #include <stdio.h> #include <string.h> #include <sys/types.h> #include <winsock.h> #pragma comment(lib, "ws2_32.lib") int main(int argc, char *argv[]) { struct addrinfo hints, *res, *p; int status; char ipstr[INET6_ADDRSTRLEN]; if (argc != 2) { fprintf(stderr,"usage: showip hostname\n"); system("PAUSE"); return 1; } memset(&hints, 0, sizeof hints); hints.ai_family = AF_UNSPEC; // AF_INET or AF_INET6 to force version hints.ai_socktype = SOCK_STREAM; if ((status = getaddrinfo(argv[1], NULL, &hints, &res)) != 0) { fprintf(stderr, "getaddrinfo: %s\n", gai_strerror(status)); system("PAUSE"); return 2; } printf("IP addresses for %s:\n\n", argv[1]); for(p = res;p != NULL; p = p->ai_next) { void *addr; char *ipver; // get the pointer to the address itself, // different fields in IPv4 and IPv6: if (p->ai_family == AF_INET) { // IPv4 struct sockaddr_in *ipv4 = (struct sockaddr_in *)p->ai_addr; addr = &(ipv4->sin_addr); ipver = "IPv4"; } else { // IPv6 struct sockaddr_in6 *ipv6 = (struct sockaddr_in6 *)p->ai_addr; addr = &(ipv6->sin6_addr); ipver = "IPv6"; } // convert the IP to a string and print it: inet_ntop(p->ai_family, addr, ipstr, sizeof ipstr); printf(" %s: %s\n", ipver, ipstr); } freeaddrinfo(res); // free the linked list system("PAUSE"); return 0; }

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  • Can't save my picture

    - by mamii
    I want to save the image that I draw, but I always failure is reported. I have tested and tried but I can correct any errors. Therefore, I appeal to you. This store is for me as a "cancer sore". And what is the drawing application without the possibility shranjevnja? sucks: D Question: What is wrong with my code for storage? or anything else? Posts: 09-12 07:30:34.346: E / Panel (8003): IOEception 09-12 07:30:34.346: E / Panel (8003): java.io.IOException: Parent directory of file does not exist: / sdcard/anppp/2012Sep1273034.png 09-12 07:30:34.346: E / Panel (8003): at java.io.File.createNewFile (File.java: 1263) 09-12 07:30:34.346: E / Panel (8003): at aa.bb.cc.Panel.saveapp (Panel.java: 67) 09-12 07:30:34.346: E / Panel (8003): at aa.bb.cc.AndroidPaint.onOptionsItemSelected (AndroidPaint.java: 94) 09-12 07:30:34.346: E / Panel (8003): at android.app.Activity.onMenuItemSelected (Activity.java: 2170) 09-12 07:30:34.346: E / Panel (8003): at com.android.internal.policy.impl.PhoneWindow.onMenuItemSelected (PhoneWindow.java: 730) 09-12 07:30:34.346: E / Panel (8003): at com.android.internal.view.menu.MenuItemImpl.invoke (MenuItemImpl.java: 139) 09-12 07:30:34.346: E / Panel (8003): at com.android.internal.view.menu.MenuBuilder.performItemAction (MenuBuilder.java: 855) 09-12 07:30:34.346: E / Panel (8003): at com.android.internal.view.menu.ExpandedMenuView.invokeItem (ExpandedMenuView.java: 89) 09-12 07:30:34.346: E / Panel (8003): at com.android.internal.view.menu.ExpandedMenuView.onItemClick (ExpandedMenuView.java: 93) 09-12 07:30:34.346: E / Panel (8003): at android.widget.AdapterView.performItemClick (AdapterView.java: 284) 09-12 07:30:34.346: E / Panel (8003): at android.widget.ListView.performItemClick (ListView.java: 3285) 09-12 07:30:34.346: E / Panel (8003): at android.widget.AbsListView $ PerformClick.run (AbsListView.java: 1640) 09-12 07:30:34.346: E / Panel (8003): at android.os.Handler.handleCallback (Handler.java: 587) 09-12 07:30:34.346: E / Panel (8003): at android.os.Handler.dispatchMessage (Handler.java: 92) 09-12 07:30:34.346: E / Panel (8003): at android.os.Looper.loop (Looper.java: 123) 09-12 07:30:34.346: E / Panel (8003): at android.app.ActivityThread.main (ActivityThread.java: 4363) 09-12 07:30:34.346: E / Panel (8003): at java.lang.reflect.Method.invokeNative (Native Method) 09-12 07:30:34.346: E / Panel (8003): at java.lang.reflect.Method.invoke (Method.java: 521) 09-12 07:30:34.346: E / Panel (8003): at com.android.internal.os.ZygoteInit $ MethodAndArgsCaller.run (ZygoteInit.java: 860) 09-12 07:30:34.346: E / Panel (8003): at com.android.internal.os.ZygoteInit.main (ZygoteInit.java: 618) 09-12 07:30:34.346: E / Panel (8003): at dalvik.system.NativeStart.main (Native Method) There is code: private Bitmap mBitmap; private Canvas mCanvas; private Bitmap tmpBitmap; private Canvas tmpCanvas; private DrawHandler mDrawHandler; private Canvas tCanvas; private String mImagePath = Environment.getExternalStorageDirectory() + "/anppp"; private File file; public void saveapp() { Calendar currentDate = Calendar.getInstance(); SimpleDateFormat formatter= new SimpleDateFormat("yyyyMMMddHmmss"); String dateNow = formatter.format(currentDate.getTime()); file = new File(mImagePath + "/" + dateNow +".png"); FileOutputStream fos; try { file.createNewFile(); fos = new FileOutputStream(file); tmpBitmap.compress(Bitmap.CompressFormat.PNG, 100, fos); fos.close(); } catch (FileNotFoundException e) { Log.e("Panel", "FileNotFoundException", e); } catch (IOException e) { Log.e("Panel", "IOEception", e); } } That's it .. I do not know what could be wrong ;(

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  • Using FiddlerCore to capture HTTP Requests with .NET

    - by Rick Strahl
    Over the last few weeks I’ve been working on my Web load testing utility West Wind WebSurge. One of the key components of a load testing tool is the ability to capture URLs effectively so that you can play them back later under load. One of the options in WebSurge for capturing URLs is to use its built-in capture tool which acts as an HTTP proxy to capture any HTTP and HTTPS traffic from most Windows HTTP clients, including Web Browsers as well as standalone Windows applications and services. To make this happen, I used Eric Lawrence’s awesome FiddlerCore library, which provides most of the functionality of his desktop Fiddler application, all rolled into an easy to use library that you can plug into your own applications. FiddlerCore makes it almost too easy to capture HTTP content! For WebSurge I needed to capture all HTTP traffic in order to capture the full HTTP request – URL, headers and any content posted by the client. The result of what I ended up creating is this semi-generic capture form: In this post I’m going to demonstrate how easy it is to use FiddlerCore to build this HTTP Capture Form.  If you want to jump right in here are the links to get Telerik’s Fiddler Core and the code for the demo provided here. FiddlerCore Download FiddlerCore on NuGet Show me the Code (WebSurge Integration code from GitHub) Download the WinForms Sample Form West Wind Web Surge (example implementation in live app) Note that FiddlerCore is bound by a license for commercial usage – see license.txt in the FiddlerCore distribution for details. Integrating FiddlerCore FiddlerCore is a library that simply plugs into your application. You can download it from the Telerik site and manually add the assemblies to your project, or you can simply install the NuGet package via:       PM> Install-Package FiddlerCore The library consists of the FiddlerCore.dll as well as a couple of support libraries (CertMaker.dll and BCMakeCert.dll) that are used for installing SSL certificates. I’ll have more on SSL captures and certificate installation later in this post. But first let’s see how easy it is to use FiddlerCore to capture HTTP content by looking at how to build the above capture form. Capturing HTTP Content Once the library is installed it’s super easy to hook up Fiddler functionality. Fiddler includes a number of static class methods on the FiddlerApplication object that can be called to hook up callback events as well as actual start monitoring HTTP URLs. In the following code directly lifted from WebSurge, I configure a few filter options on Form level object, from the user inputs shown on the form by assigning it to a capture options object. In the live application these settings are persisted configuration values, but in the demo they are one time values initialized and set on the form. Once these options are set, I hook up the AfterSessionComplete event to capture every URL that passes through the proxy after the request is completed and start up the Proxy service:void Start() { if (tbIgnoreResources.Checked) CaptureConfiguration.IgnoreResources = true; else CaptureConfiguration.IgnoreResources = false; string strProcId = txtProcessId.Text; if (strProcId.Contains('-')) strProcId = strProcId.Substring(strProcId.IndexOf('-') + 1).Trim(); strProcId = strProcId.Trim(); int procId = 0; if (!string.IsNullOrEmpty(strProcId)) { if (!int.TryParse(strProcId, out procId)) procId = 0; } CaptureConfiguration.ProcessId = procId; CaptureConfiguration.CaptureDomain = txtCaptureDomain.Text; FiddlerApplication.AfterSessionComplete += FiddlerApplication_AfterSessionComplete; FiddlerApplication.Startup(8888, true, true, true); } The key lines for FiddlerCore are just the last two lines of code that include the event hookup code as well as the Startup() method call. Here I only hook up to the AfterSessionComplete event but there are a number of other events that hook various stages of the HTTP request cycle you can also hook into. Other events include BeforeRequest, BeforeResponse, RequestHeadersAvailable, ResponseHeadersAvailable and so on. In my case I want to capture the request data and I actually have several options to capture this data. AfterSessionComplete is the last event that fires in the request sequence and it’s the most common choice to capture all request and response data. I could have used several other events, but AfterSessionComplete is one place where you can look both at the request and response data, so this will be the most common place to hook into if you’re capturing content. The implementation of AfterSessionComplete is responsible for capturing all HTTP request headers and it looks something like this:private void FiddlerApplication_AfterSessionComplete(Session sess) { // Ignore HTTPS connect requests if (sess.RequestMethod == "CONNECT") return; if (CaptureConfiguration.ProcessId > 0) { if (sess.LocalProcessID != 0 && sess.LocalProcessID != CaptureConfiguration.ProcessId) return; } if (!string.IsNullOrEmpty(CaptureConfiguration.CaptureDomain)) { if (sess.hostname.ToLower() != CaptureConfiguration.CaptureDomain.Trim().ToLower()) return; } if (CaptureConfiguration.IgnoreResources) { string url = sess.fullUrl.ToLower(); var extensions = CaptureConfiguration.ExtensionFilterExclusions; foreach (var ext in extensions) { if (url.Contains(ext)) return; } var filters = CaptureConfiguration.UrlFilterExclusions; foreach (var urlFilter in filters) { if (url.Contains(urlFilter)) return; } } if (sess == null || sess.oRequest == null || sess.oRequest.headers == null) return; string headers = sess.oRequest.headers.ToString(); var reqBody = sess.GetRequestBodyAsString(); // if you wanted to capture the response //string respHeaders = session.oResponse.headers.ToString(); //var respBody = session.GetResponseBodyAsString(); // replace the HTTP line to inject full URL string firstLine = sess.RequestMethod + " " + sess.fullUrl + " " + sess.oRequest.headers.HTTPVersion; int at = headers.IndexOf("\r\n"); if (at < 0) return; headers = firstLine + "\r\n" + headers.Substring(at + 1); string output = headers + "\r\n" + (!string.IsNullOrEmpty(reqBody) ? reqBody + "\r\n" : string.Empty) + Separator + "\r\n\r\n"; BeginInvoke(new Action<string>((text) => { txtCapture.AppendText(text); UpdateButtonStatus(); }), output); } The code starts by filtering out some requests based on the CaptureOptions I set before the capture is started. These options/filters are applied when requests actually come in. This is very useful to help narrow down the requests that are captured for playback based on options the user picked. I find it useful to limit requests to a certain domain for captures, as well as filtering out some request types like static resources – images, css, scripts etc. This is of course optional, but I think it’s a common scenario and WebSurge makes good use of this feature. AfterSessionComplete like other FiddlerCore events, provides a Session object parameter which contains all the request and response details. There are oRequest and oResponse objects to hold their respective data. In my case I’m interested in the raw request headers and body only, as you can see in the commented code you can also retrieve the response headers and body. Here the code captures the request headers and body and simply appends the output to the textbox on the screen. Note that the Fiddler events are asynchronous, so in order to display the content in the UI they have to be marshaled back the UI thread with BeginInvoke, which here simply takes the generated headers and appends it to the existing textbox test on the form. As each request is processed, the headers are captured and appended to the bottom of the textbox resulting in a Session HTTP capture in the format that Web Surge internally supports, which is basically raw request headers with a customized 1st HTTP Header line that includes the full URL rather than a server relative URL. When the capture is done the user can either copy the raw HTTP session to the clipboard, or directly save it to file. This raw capture format is the same format WebSurge and also Fiddler use to import/export request data. While this code is application specific, it demonstrates the kind of logic that you can easily apply to the request capture process, which is one of the reasonsof why FiddlerCore is so powerful. You get to choose what content you want to look up as part of your own application logic and you can then decide how to capture or use that data as part of your application. The actual captured data in this case is only a string. The user can edit the data by hand or in the the case of WebSurge, save it to disk and automatically open the captured session as a new load test. Stopping the FiddlerCore Proxy Finally to stop capturing requests you simply disconnect the event handler and call the FiddlerApplication.ShutDown() method:void Stop() { FiddlerApplication.AfterSessionComplete -= FiddlerApplication_AfterSessionComplete; if (FiddlerApplication.IsStarted()) FiddlerApplication.Shutdown(); } As you can see, adding HTTP capture functionality to an application is very straight forward. FiddlerCore offers tons of features I’m not even touching on here – I suspect basic captures are the most common scenario, but a lot of different things can be done with FiddlerCore’s simple API interface. Sky’s the limit! The source code for this sample capture form (WinForms) is provided as part of this article. Adding Fiddler Certificates with FiddlerCore One of the sticking points in West Wind WebSurge has been that if you wanted to capture HTTPS/SSL traffic, you needed to have the full version of Fiddler and have HTTPS decryption enabled. Essentially you had to use Fiddler to configure HTTPS decryption and the associated installation of the Fiddler local client certificate that is used for local decryption of incoming SSL traffic. While this works just fine, requiring to have Fiddler installed and then using a separate application to configure the SSL functionality isn’t ideal. Fortunately FiddlerCore actually includes the tools to register the Fiddler Certificate directly using FiddlerCore. Why does Fiddler need a Certificate in the first Place? Fiddler and FiddlerCore are essentially HTTP proxies which means they inject themselves into the HTTP conversation by re-routing HTTP traffic to a special HTTP port (8888 by default for Fiddler) and then forward the HTTP data to the original client. Fiddler injects itself as the system proxy in using the WinInet Windows settings  which are the same settings that Internet Explorer uses and that are configured in the Windows and Internet Explorer Internet Settings dialog. Most HTTP clients running on Windows pick up and apply these system level Proxy settings before establishing new HTTP connections and that’s why most clients automatically work once Fiddler – or FiddlerCore/WebSurge are running. For plain HTTP requests this just works – Fiddler intercepts the HTTP requests on the proxy port and then forwards them to the original port (80 for HTTP and 443 for SSL typically but it could be any port). For SSL however, this is not quite as simple – Fiddler can easily act as an HTTPS/SSL client to capture inbound requests from the server, but when it forwards the request to the client it has to also act as an SSL server and provide a certificate that the client trusts. This won’t be the original certificate from the remote site, but rather a custom local certificate that effectively simulates an SSL connection between the proxy and the client. If there is no custom certificate configured for Fiddler the SSL request fails with a certificate validation error. The key for this to work is that a custom certificate has to be installed that the HTTPS client trusts on the local machine. For a much more detailed description of the process you can check out Eric Lawrence’s blog post on Certificates. If you’re using the desktop version of Fiddler you can install a local certificate into the Windows certificate store. Fiddler proper does this from the Options menu: This operation does several things: It installs the Fiddler Root Certificate It sets trust to this Root Certificate A new client certificate is generated for each HTTPS site monitored Certificate Installation with FiddlerCore You can also provide this same functionality using FiddlerCore which includes a CertMaker class. Using CertMaker is straight forward to use and it provides an easy way to create some simple helpers that can install and uninstall a Fiddler Root certificate:public static bool InstallCertificate() { if (!CertMaker.rootCertExists()) { if (!CertMaker.createRootCert()) return false; if (!CertMaker.trustRootCert()) return false; } return true; } public static bool UninstallCertificate() { if (CertMaker.rootCertExists()) { if (!CertMaker.removeFiddlerGeneratedCerts(true)) return false; } return true; } InstallCertificate() works by first checking whether the root certificate is already installed and if it isn’t goes ahead and creates a new one. The process of creating the certificate is a two step process – first the actual certificate is created and then it’s moved into the certificate store to become trusted. I’m not sure why you’d ever split these operations up since a cert created without trust isn’t going to be of much value, but there are two distinct steps. When you trigger the trustRootCert() method, a message box will pop up on the desktop that lets you know that you’re about to trust a local private certificate. This is a security feature to ensure that you really want to trust the Fiddler root since you are essentially installing a man in the middle certificate. It’s quite safe to use this generated root certificate, because it’s been specifically generated for your machine and thus is not usable from external sources, the only way to use this certificate in a trusted way is from the local machine. IOW, unless somebody has physical access to your machine, there’s no useful way to hijack this certificate and use it for nefarious purposes (see Eric’s post for more details). Once the Root certificate has been installed, FiddlerCore/Fiddler create new certificates for each site that is connected to with HTTPS. You can end up with quite a few temporary certificates in your certificate store. To uninstall you can either use Fiddler and simply uncheck the Decrypt HTTPS traffic option followed by the remove Fiddler certificates button, or you can use FiddlerCore’s CertMaker.removeFiddlerGeneratedCerts() which removes the root cert and any of the intermediary certificates Fiddler created. Keep in mind that when you uninstall you uninstall the certificate for both FiddlerCore and Fiddler, so use UninstallCertificate() with care and realize that you might affect the Fiddler application’s operation by doing so as well. When to check for an installed Certificate Note that the check to see if the root certificate exists is pretty fast, while the actual process of installing the certificate is a relatively slow operation that even on a fast machine takes a few seconds. Further the trust operation pops up a message box so you probably don’t want to install the certificate repeatedly. Since the check for the root certificate is fast, you can easily put a call to InstallCertificate() in any capture startup code – in which case the certificate installation only triggers when a certificate is in fact not installed. Personally I like to make certificate installation explicit – just like Fiddler does, so in WebSurge I use a small drop down option on the menu to install or uninstall the SSL certificate:   This code calls the InstallCertificate and UnInstallCertificate functions respectively – the experience with this is similar to what you get in Fiddler with the extra dialog box popping up to prompt confirmation for installation of the root certificate. Once the cert is installed you can then capture SSL requests. There’s a gotcha however… Gotcha: FiddlerCore Certificates don’t stick by Default When I originally tried to use the Fiddler certificate installation I ran into an odd problem. I was able to install the certificate and immediately after installation was able to capture HTTPS requests. Then I would exit the application and come back in and try the same HTTPS capture again and it would fail due to a missing certificate. CertMaker.rootCertExists() would return false after every restart and if re-installed the certificate a new certificate would get added to the certificate store resulting in a bunch of duplicated root certificates with different keys. What the heck? CertMaker and BcMakeCert create non-sticky CertificatesI turns out that FiddlerCore by default uses different components from what the full version of Fiddler uses. Fiddler uses a Windows utility called MakeCert.exe to create the Fiddler Root certificate. FiddlerCore however installs the CertMaker.dll and BCMakeCert.dll assemblies, which use a different crypto library (Bouncy Castle) for certificate creation than MakeCert.exe which uses the Windows Crypto API. The assemblies provide support for non-windows operation for Fiddler under Mono, as well as support for some non-Windows certificate platforms like iOS and Android for decryption. The bottom line is that the FiddlerCore provided bouncy castle assemblies are not sticky by default as the certificates created with them are not cached as they are in Fiddler proper. To get certificates to ‘stick’ you have to explicitly cache the certificates in Fiddler’s internal preferences. A cache aware version of InstallCertificate looks something like this:public static bool InstallCertificate() { if (!CertMaker.rootCertExists()) { if (!CertMaker.createRootCert()) return false; if (!CertMaker.trustRootCert()) return false; App.Configuration.UrlCapture.Cert = FiddlerApplication.Prefs.GetStringPref("fiddler.certmaker.bc.cert", null); App.Configuration.UrlCapture.Key = FiddlerApplication.Prefs.GetStringPref("fiddler.certmaker.bc.key", null); } return true; } public static bool UninstallCertificate() { if (CertMaker.rootCertExists()) { if (!CertMaker.removeFiddlerGeneratedCerts(true)) return false; } App.Configuration.UrlCapture.Cert = null; App.Configuration.UrlCapture.Key = null; return true; } In this code I store the Fiddler cert and private key in an application configuration settings that’s stored with the application settings (App.Configuration.UrlCapture object). These settings automatically persist when WebSurge is shut down. The values are read out of Fiddler’s internal preferences store which is set after a new certificate has been created. Likewise I clear out the configuration settings when the certificate is uninstalled. In order for these setting to be used you have to also load the configuration settings into the Fiddler preferences *before* a call to rootCertExists() is made. I do this in the capture form’s constructor:public FiddlerCapture(StressTestForm form) { InitializeComponent(); CaptureConfiguration = App.Configuration.UrlCapture; MainForm = form; if (!string.IsNullOrEmpty(App.Configuration.UrlCapture.Cert)) { FiddlerApplication.Prefs.SetStringPref("fiddler.certmaker.bc.key", App.Configuration.UrlCapture.Key); FiddlerApplication.Prefs.SetStringPref("fiddler.certmaker.bc.cert", App.Configuration.UrlCapture.Cert); }} This is kind of a drag to do and not documented anywhere that I could find, so hopefully this will save you some grief if you want to work with the stock certificate logic that installs with FiddlerCore. MakeCert provides sticky Certificates and the same functionality as Fiddler But there’s actually an easier way. If you want to skip the above Fiddler preference configuration code in your application you can choose to distribute MakeCert.exe instead of certmaker.dll and bcmakecert.dll. When you use MakeCert.exe, the certificates settings are stored in Windows so they are available without any custom configuration inside of your application. It’s easier to integrate and as long as you run on Windows and you don’t need to support iOS or Android devices is simply easier to deal with. To integrate into your project, you can remove the reference to CertMaker.dll (and the BcMakeCert.dll assembly) from your project. Instead copy MakeCert.exe into your output folder. To make sure MakeCert.exe gets pushed out, include MakeCert.exe in your project and set the Build Action to None, and Copy to Output Directory to Copy if newer. Note that the CertMaker.dll reference in the project has been removed and on disk the files for Certmaker.dll, as well as the BCMakeCert.dll files on disk. Keep in mind that these DLLs are resources of the FiddlerCore NuGet package, so updating the package may end up pushing those files back into your project. Once MakeCert.exe is distributed FiddlerCore checks for it first before using the assemblies so as long as MakeCert.exe exists it’ll be used for certificate creation (at least on Windows). Summary FiddlerCore is a pretty sweet tool, and it’s absolutely awesome that we get to plug in most of the functionality of Fiddler right into our own applications. A few years back I tried to build this sort of functionality myself for an app and ended up giving up because it’s a big job to get HTTP right – especially if you need to support SSL. FiddlerCore now provides that functionality as a turnkey solution that can be plugged into your own apps easily. The only downside is FiddlerCore’s documentation for more advanced features like certificate installation which is pretty sketchy. While for the most part FiddlerCore’s feature set is easy to work with without any documentation, advanced features are often not intuitive to gleam by just using Intellisense or the FiddlerCore help file reference (which is not terribly useful). While Eric Lawrence is very responsive on his forum and on Twitter, there simply isn’t much useful documentation on Fiddler/FiddlerCore available online. If you run into trouble the forum is probably the first place to look and then ask a question if you can’t find the answer. The best documentation you can find is Eric’s Fiddler Book which covers a ton of functionality of Fiddler and FiddlerCore. The book is a great reference to Fiddler’s feature set as well as providing great insights into the HTTP protocol. The second half of the book that gets into the innards of HTTP is an excellent read for anybody who wants to know more about some of the more arcane aspects and special behaviors of HTTP – it’s well worth the read. While the book has tons of information in a very readable format, it’s unfortunately not a great reference as it’s hard to find things in the book and because it’s not available online you can’t electronically search for the great content in it. But it’s hard to complain about any of this given the obvious effort and love that’s gone into this awesome product for all of these years. A mighty big thanks to Eric Lawrence  for having created this useful tool that so many of us use all the time, and also to Telerik for picking up Fiddler/FiddlerCore and providing Eric the resources to support and improve this wonderful tool full time and keeping it free for all. Kudos! Resources FiddlerCore Download FiddlerCore NuGet Fiddler Capture Sample Form Fiddler Capture Form in West Wind WebSurge (GitHub) Eric Lawrence’s Fiddler Book© Rick Strahl, West Wind Technologies, 2005-2014Posted in .NET  HTTP   Tweet !function(d,s,id){var js,fjs=d.getElementsByTagName(s)[0];if(!d.getElementById(id)){js=d.createElement(s);js.id=id;js.src="//platform.twitter.com/widgets.js";fjs.parentNode.insertBefore(js,fjs);}}(document,"script","twitter-wjs"); (function() { var po = document.createElement('script'); po.type = 'text/javascript'; po.async = true; po.src = 'https://apis.google.com/js/plusone.js'; var s = document.getElementsByTagName('script')[0]; s.parentNode.insertBefore(po, s); })();

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  • Upgrading from TFS 2010 RC to TFS 2010 RTM done

    - by Martin Hinshelwood
    Today is the big day, with the Launch of Visual Studio 2010 already done in Asia, and rolling around the world towards us, we are getting ready for the RTM (Released). We have had TFS 2010 in Production for nearly 6 months and have had only minimal problems. Update 12th April 2010  – Added Scott Hanselman’s tweet about the MSDN download release time. SSW was the first company in the world outside of Microsoft to deploy Visual Studio 2010 Team Foundation Server to production, not once, but twice. I am hoping to make it 3 in a row, but with all the hype around the new version, and with it being a production release and not just a go-live, I think there will be a lot of competition. Developers: MSDN will be updated with #vs2010 downloads and details at 10am PST *today*! @shanselman - Scott Hanselman Same as before, we need to Uninstall 2010 RC and install 2010 RTM. The installer will take care of all the complexity of actually upgrading any schema changes. If you are upgrading from TFS 2008 to TFS2010 you can follow our Rules To Better TFS 2010 Migration and read my post on our successes.   We run TFS 2010 in a Hyper-V virtual environment, so we have the advantage of running a snapshot as well as taking a DB backup. Done - Snapshot the hyper-v server Microsoft does not support taking a snapshot of a running server, for very good reason, and Brian Harry wrote a post after my last upgrade with the reason why you should never snapshot a running server. Done - Uninstall Visual Studio Team Explorer 2010 RC You will need to uninstall all of the Visual Studio 2010 RC client bits that you have on the server. Done - Uninstall TFS 2010 RC Done - Install TFS 2010 RTM Done - Configure TFS 2010 RTM Pick the Upgrade option and point it at your existing “tfs_Configuration” database to load all of the existing settings Done - Upgrade the SharePoint Extensions Upgrade Build Servers (Pending) Test the server The back out plan, and you should always have one, is to restore the snapshot. Upgrading to Team Foundation Server 2010 – Done The first thing you need to do is off the TFS server and then log into the Hyper-v server and create a snapshot. Figure: Make sure you turn the server off and delete all old snapshots before you take a new one I noticed that the snapshot that was taken before the Beta 2 to RC upgrade was still there. You should really delete old snapshots before you create a new one, but in this case the SysAdmin (who is currently tucked up in bed) asked me not to. I guess he is worried about a developer messing up his server Turn your server on and wait for it to boot in anticipation of all the nice shiny RTM’ness that is coming next. The upgrade procedure for TFS2010 is to uninstal the old version and install the new one. Figure: Remove Visual Studio 2010 Team Foundation Server RC from the system.   Figure: Most of the heavy lifting is done by the Uninstaller, but make sure you have removed any of the client bits first. Specifically Visual Studio 2010 or Team Explorer 2010.  Once the uninstall is complete, this took around 5 minutes for me, you can begin the install of the RTM. Running the 64 bit OS will allow the application to use more than 2GB RAM, which while not common may be of use in heavy load situations. Figure: It is always recommended to install the 64bit version of a server application where possible. I do not think it is likely, with SharePoint 2010 and Exchange 2010  and even Windows Server 2008 R2 being 64 bit only, I do not think there will be another release of a server app that is 32bit. You then need to choose what it is you want to install. This depends on how you are running TFS and on how many servers. In our case we run TFS and the Team Foundation Build Service (controller only) on out TFS server along with Analysis services and Reporting Services. But our SharePoint server lives elsewhere. Figure: This always confuses people, but in reality it makes sense. Don’t install what you do not need. Every extra you install has an impact of performance. If you are integrating with SharePoint you will need to run this install on every Front end server in your farm and don’t forget to upgrade your Build servers and proxy servers later. Figure: Selecting only Team Foundation Server (TFS) and Team Foundation Build Services (TFBS)   It is worth noting that if you have a lot of builds kicking off, and hence a lot of get operations against your TFS server, you can use a proxy server to cache the source control on another server in between your TFS server and your build servers. Figure: Installing Microsoft .NET Framework 4 takes the most time. Figure: Now run Windows Update, and SSW Diagnostic to make sure all your bits and bobs are up to date. Note: SSW Diagnostic will check your Power Tools, Add-on’s, Check in Policies and other bits as well. Configure Team Foundation Server 2010 – Done Now you can configure the server. If you have no key you will need to pick “Install a Trial Licence”, but it is only £500, or free with a MSDN subscription. Anyway, if you pick Trial you get 90 days to get your key. Figure: You can pick trial and add your key later using the TFS Server Admin. Here is where the real choices happen. We are doing an Upgrade from a previous version, so I will pick Upgrade the same as all you folks that are using the RC or TFS 2008. Figure: The upgrade wizard takes your existing 2010 or 2008 databases and upgraded them to the release.   Once you have entered your database server name you can click “List available databases” and it will show what it can upgrade. Figure: Select your database from the list and at this point, make sure you have a valid backup. At this point you have not made ANY changes to the databases. At this point the configuration wizard will load configuration from your existing database if you have one. If you are upgrading TFS 2008 refer to Rules To Better TFS 2010 Migration. Mostly during the wizard the default values will suffice, but depending on the configuration you want you can pick different options. Figure: Set the application tier account and Authentication method to use. We use NTLM to keep things simple as we host our TFS server externally for our remote developers.  Figure: Setting your TFS server URL’s to be the remote URL’s allows the reports to be accessed without using VPN. Very handy for those remote developers. Figure: Detected the existing Warehouse no problem. Figure: Again we love green ticks. It gives us a warm fuzzy feeling. Figure: The username for connecting to Reporting services should be a domain account (if you are on a domain that is). Figure: Setup the SharePoint integration to connect to your external SharePoint server. You can take the option to connect later.   You then need to run all of your readiness checks. These check can save your life! it will check all of the settings that you have entered as well as checking all the external services are configures and running properly. There are two reasons that TFS 2010 is so easy and painless to install where previous version were not. Microsoft changes the install to two steps, Install and configuration. The second reason is that they have pulled out all of the stops in making the install run all the checks necessary to make sure that once you start the install that it will complete. if you find any errors I recommend that you report them on http://connect.microsoft.com so everyone can benefit from your misery.   Figure: Now we have everything setup the configuration wizard can do its work.  Figure: Took a while on the “Web site” stage for some point, but zipped though after that.  Figure: last wee bit. TFS Needs to do a little tinkering with the data to complete the upgrade. Figure: All upgraded. I am not worried about the yellow triangle as SharePoint was being a little silly Exception Message: TF254021: The account name or password that you specified is not valid. (type TfsAdminException) Exception Stack Trace:    at Microsoft.TeamFoundation.Management.Controls.WizardCommon.AccountSelectionControl.TestLogon(String connectionString)    at System.ComponentModel.BackgroundWorker.WorkerThreadStart(Object argument) [Info   @16:10:16.307] Benign exception caught as part of verify: Exception Message: TF255329: The following site could not be accessed: http://projects.ssw.com.au/. The server that you specified did not return the expected response. Either you have not installed the Team Foundation Server Extensions for SharePoint Products on this server, or a firewall is blocking access to the specified site or the SharePoint Central Administration site. For more information, see the Microsoft Web site (http://go.microsoft.com/fwlink/?LinkId=161206). (type TeamFoundationServerException) Exception Stack Trace:    at Microsoft.TeamFoundation.Client.SharePoint.WssUtilities.VerifyTeamFoundationSharePointExtensions(ICredentials credentials, Uri url)    at Microsoft.TeamFoundation.Admin.VerifySharePointSitesUrl.Verify() Inner Exception Details: Exception Message: TF249064: The following Web service returned an response that is not valid: http://projects.ssw.com.au/_vti_bin/TeamFoundationIntegrationService.asmx. This Web service is used for the Team Foundation Server Extensions for SharePoint Products. Either the extensions are not installed, the request resulted in HTML being returned, or there is a problem with the URL. Verify that the following URL points to a valid SharePoint Web application and that the application is available: http://projects.ssw.com.au. If the URL is correct and the Web application is operating normally, verify that a firewall is not blocking access to the Web application. (type TeamFoundationServerInvalidResponseException) Exception Data Dictionary: ResponseStatusCode = InternalServerError I’ll look at SharePoint after, probably the SharePoint box just needs a restart or a kick If there is a problem with SharePoint it will come out in testing, But I will definatly be passing this on to Microsoft.   Upgrading the SharePoint connector to TFS 2010 You will need to upgrade the Extensions for SharePoint Products and Technologies on all of your SharePoint farm front end servers. To do this uninstall  the TFS 2010 RC from it in the same way as the server, and then install just the RTM Extensions. Figure: Only install the SharePoint Extensions on your SharePoint front end servers. TFS 2010 supports both SharePoint 2007 and SharePoint 2010.   Figure: When you configure SharePoint it uploads all of the solutions and templates. Figure: Everything is uploaded Successfully. Figure: TFS even remembered the settings from the previous installation, fantastic.   Upgrading the Team Foundation Build Servers to TFS 2010 Just like on the SharePoint servers you will need to upgrade the Build Server to the RTM. Just uninstall TFS 2010 RC and then install only the Team Foundation Build Services component. Unlike on the SharePoint server you will probably have some version of Visual Studio installed. You will need to remove this as well. (Coming Soon) Connecting Visual Studio 2010 / 2008 / 2005 and Eclipse to TFS2010 If you have developers still on Visual Studio 2005 or 2008 you will need do download the respective compatibility pack: Visual Studio Team System 2005 Service Pack 1 Forward Compatibility Update for Team Foundation Server 2010 Visual Studio Team System 2008 Service Pack 1 Forward Compatibility Update for Team Foundation Server 2010 If you are using Eclipse you can download the new Team Explorer Everywhere install for connecting to TFS. Get your developers to check that you have the latest version of your applications with SSW Diagnostic which will check for Service Packs and hot fixes to Visual Studio as well.   Technorati Tags: TFS,TFS2010,TFS 2010,Upgrade

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  • Parallelism in .NET – Part 11, Divide and Conquer via Parallel.Invoke

    - by Reed
    Many algorithms are easily written to work via recursion.  For example, most data-oriented tasks where a tree of data must be processed are much more easily handled by starting at the root, and recursively “walking” the tree.  Some algorithms work this way on flat data structures, such as arrays, as well.  This is a form of divide and conquer: an algorithm design which is based around breaking up a set of work recursively, “dividing” the total work in each recursive step, and “conquering” the work when the remaining work is small enough to be solved easily. Recursive algorithms, especially ones based on a form of divide and conquer, are often a very good candidate for parallelization. This is apparent from a common sense standpoint.  Since we’re dividing up the total work in the algorithm, we have an obvious, built-in partitioning scheme.  Once partitioned, the data can be worked upon independently, so there is good, clean isolation of data. Implementing this type of algorithm is fairly simple.  The Parallel class in .NET 4 includes a method suited for this type of operation: Parallel.Invoke.  This method works by taking any number of delegates defined as an Action, and operating them all in parallel.  The method returns when every delegate has completed: Parallel.Invoke( () => { Console.WriteLine("Action 1 executing in thread {0}", Thread.CurrentThread.ManagedThreadId); }, () => { Console.WriteLine("Action 2 executing in thread {0}", Thread.CurrentThread.ManagedThreadId); }, () => { Console.WriteLine("Action 3 executing in thread {0}", Thread.CurrentThread.ManagedThreadId); } ); .csharpcode, .csharpcode pre { font-size: small; color: black; font-family: consolas, "Courier New", courier, monospace; background-color: #ffffff; /*white-space: pre;*/ } .csharpcode pre { margin: 0em; } .csharpcode .rem { color: #008000; } .csharpcode .kwrd { color: #0000ff; } .csharpcode .str { color: #006080; } .csharpcode .op { color: #0000c0; } .csharpcode .preproc { color: #cc6633; } .csharpcode .asp { background-color: #ffff00; } .csharpcode .html { color: #800000; } .csharpcode .attr { color: #ff0000; } .csharpcode .alt { background-color: #f4f4f4; width: 100%; margin: 0em; } .csharpcode .lnum { color: #606060; } Running this simple example demonstrates the ease of using this method.  For example, on my system, I get three separate thread IDs when running the above code.  By allowing any number of delegates to be executed directly, concurrently, the Parallel.Invoke method provides us an easy way to parallelize any algorithm based on divide and conquer.  We can divide our work in each step, and execute each task in parallel, recursively. For example, suppose we wanted to implement our own quicksort routine.  The quicksort algorithm can be designed based on divide and conquer.  In each iteration, we pick a pivot point, and use that to partition the total array.  We swap the elements around the pivot, then recursively sort the lists on each side of the pivot.  For example, let’s look at this simple, sequential implementation of quicksort: public static void QuickSort<T>(T[] array) where T : IComparable<T> { QuickSortInternal(array, 0, array.Length - 1); } private static void QuickSortInternal<T>(T[] array, int left, int right) where T : IComparable<T> { if (left >= right) { return; } SwapElements(array, left, (left + right) / 2); int last = left; for (int current = left + 1; current <= right; ++current) { if (array[current].CompareTo(array[left]) < 0) { ++last; SwapElements(array, last, current); } } SwapElements(array, left, last); QuickSortInternal(array, left, last - 1); QuickSortInternal(array, last + 1, right); } static void SwapElements<T>(T[] array, int i, int j) { T temp = array[i]; array[i] = array[j]; array[j] = temp; } Here, we implement the quicksort algorithm in a very common, divide and conquer approach.  Running this against the built-in Array.Sort routine shows that we get the exact same answers (although the framework’s sort routine is slightly faster).  On my system, for example, I can use framework’s sort to sort ten million random doubles in about 7.3s, and this implementation takes about 9.3s on average. Looking at this routine, though, there is a clear opportunity to parallelize.  At the end of QuickSortInternal, we recursively call into QuickSortInternal with each partition of the array after the pivot is chosen.  This can be rewritten to use Parallel.Invoke by simply changing it to: // Code above is unchanged... SwapElements(array, left, last); Parallel.Invoke( () => QuickSortInternal(array, left, last - 1), () => QuickSortInternal(array, last + 1, right) ); } This routine will now run in parallel.  When executing, we now see the CPU usage across all cores spike while it executes.  However, there is a significant problem here – by parallelizing this routine, we took it from an execution time of 9.3s to an execution time of approximately 14 seconds!  We’re using more resources as seen in the CPU usage, but the overall result is a dramatic slowdown in overall processing time. This occurs because parallelization adds overhead.  Each time we split this array, we spawn two new tasks to parallelize this algorithm!  This is far, far too many tasks for our cores to operate upon at a single time.  In effect, we’re “over-parallelizing” this routine.  This is a common problem when working with divide and conquer algorithms, and leads to an important observation: When parallelizing a recursive routine, take special care not to add more tasks than necessary to fully utilize your system. This can be done with a few different approaches, in this case.  Typically, the way to handle this is to stop parallelizing the routine at a certain point, and revert back to the serial approach.  Since the first few recursions will all still be parallelized, our “deeper” recursive tasks will be running in parallel, and can take full advantage of the machine.  This also dramatically reduces the overhead added by parallelizing, since we’re only adding overhead for the first few recursive calls.  There are two basic approaches we can take here.  The first approach would be to look at the total work size, and if it’s smaller than a specific threshold, revert to our serial implementation.  In this case, we could just check right-left, and if it’s under a threshold, call the methods directly instead of using Parallel.Invoke. The second approach is to track how “deep” in the “tree” we are currently at, and if we are below some number of levels, stop parallelizing.  This approach is a more general-purpose approach, since it works on routines which parse trees as well as routines working off of a single array, but may not work as well if a poor partitioning strategy is chosen or the tree is not balanced evenly. This can be written very easily.  If we pass a maxDepth parameter into our internal routine, we can restrict the amount of times we parallelize by changing the recursive call to: // Code above is unchanged... SwapElements(array, left, last); if (maxDepth < 1) { QuickSortInternal(array, left, last - 1, maxDepth); QuickSortInternal(array, last + 1, right, maxDepth); } else { --maxDepth; Parallel.Invoke( () => QuickSortInternal(array, left, last - 1, maxDepth), () => QuickSortInternal(array, last + 1, right, maxDepth)); } We no longer allow this to parallelize indefinitely – only to a specific depth, at which time we revert to a serial implementation.  By starting the routine with a maxDepth equal to Environment.ProcessorCount, we can restrict the total amount of parallel operations significantly, but still provide adequate work for each processing core. With this final change, my timings are much better.  On average, I get the following timings: Framework via Array.Sort: 7.3 seconds Serial Quicksort Implementation: 9.3 seconds Naive Parallel Implementation: 14 seconds Parallel Implementation Restricting Depth: 4.7 seconds Finally, we are now faster than the framework’s Array.Sort implementation.

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