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JQuery Dynamic Element - In DOM but unable to bind

- by Grant80

Hi All, I'm new to using JQuery so bear with me. I had implmented some code based on a js file that I found online which enables a series of div tags within a nested structure on my page to step through and show each one individually on the page. This all works great when I define the div tags as static entries in the masterpage. I should add that this is being implemented in a SharePoint master page. Ultimately though, with a static collection of div tags ideally containing an image with some descriptive text, and a hyperlink its not very flexible. Roll on my changes to make this a little more configurable. I have implemented some additional code that will read from a SharePoint list via an ajax call to the lists web service. For each entry in the list I am building a div tag that contains the information required dynamically. For testing, I am only pulling the title through at present. I have used the following code: $('#beltDiv').append(divHTML) to append the divs in the loop that are created to my nested structure on the page. I figured that this would cause the fade code to work as expected but I was wrong. It doesn't do anything at all. When check the source on the page, the div tags are not shown. They are however available in the DOM model when viewed through the IE developer toolbar. The issue (I think) looks to be that the initiation of the featureFade code is not working due to the div tags being unavailable. Is there a way to address this? The code used is shown below: <script type="text/javascript"> $(document).ready(function() { var soapEnv = "<soapenv:Envelope xmlns:soapenv='http://schemas.xmlsoap.org/soap/envelope/'> \ <soapenv:Body> \ <GetListItems xmlns='http://schemas.microsoft.com/sharepoint/soap/'> \ <listName>Carousel Items</listName> \ <viewFields> \ <ViewFields> \ <FieldRef Name='Title' /> \ </ViewFields> \ </viewFields> \ </GetListItems> \ </soapenv:Body> \ </soapenv:Envelope>"; $.ajax({ url: "_vti_bin/lists.asmx", type: "POST", dataType: "xml", data: soapEnv, complete: processResult, contentType: "text/xml; charset=\"utf-8\"" }); }); function processResult(xData, status) { $(xData.responseXML).find("z\\:row").each(function() { var divHTML = "<div id=\"divPanel_" + $(this).attr("ows_Title") + "\" class=\"panel\" style=\"background:url('http://devSP2010/sites/SPSOPS/Style Library/SharePointOps/Images/01.jpg') no-repeat; width:650px; height:55px;\"><div><div class=\"content\"><div><P><A style=\"COLOR: #cc0000\" href=\"www.google.com\">" + $(this).attr("ows_Title") + "</A></P><P> </P><P> </P><P> </P><P> </P></div></div></div></div>"; $("#beltDiv").append(divHTML); }); } featureFade.setup({ galleryid: 'headlines', beltclass: 'belt', panelclass: 'panel', autostep: { enable: true, moveby: 1, pause: 10000 }, panelbehavior: { speed: 1000, wraparound: true }, stepImgIDs: ["ftOne", "ftTwo", "ftThree", "ftFour","ftFive"], defaultButtons: { itemOn: "Style Library/SharePointOps/Images/dotOn.png", itemOff: "Style Library/SharePointOps/Images/dotOff.png" } }); The section where the div tags are dynamically appended is shown below. I've commented out the static div tags that work as expected. The only change is that these are implmented by the JQuery logic: <div class="homeFeature" style="display:inline-block"> <div id="headlines" class="headlines"> <div id="beltDiv" class="belt">  </div> </div> I'm stumped as to why it's not recognising the dynamically added elements in the DOM. Any help would be greatly appreciated on this. I'm happy to provide any further information on this. Thanks in advance, Grant Further to the answer recieved: I have modified the function call: function processResult(xData, status) { $(xData.responseXML).find("z\\:row").each( function() { /*alert($(this).attr("ows_ImagePath"));*/ var divHTML = "<div id=\"divPanel_" + $(this).attr("ows_Title") + "\" class=\"panel\" style=\"background:url('http://devSP2010/sites/SPSOPS/Style Library/SharePointOps/Images/ClydePort01big.jpg') no-repeat; width:650px; height:55px;\"><div><div class=\"content\"><div><P><A style=\"COLOR: #cc0000\" href=\"www.google.com\">" + $(this).attr("ows_Title") + "</A></P><P> </P><P> </P><P> </P><P> </P></div></div></div></div>"; $("#beltDiv").append(divHTML); } ); featureFade.setup( { galleryid: 'headlines', beltclass: 'belt', panelclass: 'panel', autostep: { enable: true, moveby: 1, pause: 10000 }, panelbehavior: { speed: 1000, wraparound: true }, stepImgIDs: ["ftOne", "ftTwo", "ftThree", "ftFour","ftFive"], defaultButtons: { itemOn: "Style Library/SharePointOps/Images/dotOn.png", itemOff: "Style Library/SharePointOps/Images/dotOff.png" } } ); }

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Fastest way to move records from a oracle DB into MS sql server after processing

- by user347748

Hi.. Ok this is the scenario...I have a table in Oracle that acts like a queue... A VB.net program reads the queue and calls a stored proc in MS SQL Server that processes and then inserts the message into another SQL server table and then deletes the record from the oracle table. We use a datareader to read the records from Oracle and then call the stored proc for each of the records. The program seems to be a little slow. The stored procedure itself isnt slow. The SP by itself when called in a loop can process about 2000 records in 20 seconds. BUt when called from the .Net program, the execution time is about 5 records per second. I have seen that most of the time consumed is in calling the stored procedure and waiting for it to return. Is there a better way of doing this? Here is a snippet of the actual code Function StartDataXfer() As Boolean Dim status As Boolean = False Try SqlConn.Open() OraConn.Open() c.ErrorLog(Now.ToString & "--Going to Get the messages from oracle", 1) If GetMsgsFromOracle() Then c.ErrorLog(Now.ToString & "--Got messages from oracle", 1) If ProcessMessages() Then c.ErrorLog(Now.ToString & "--Finished Processing all messages in the queue", 0) status = True Else c.ErrorLog(Now.ToString & "--Failed to Process all messages in the queue", 0) status = False End If Else status = True End If StartDataXfer = status Catch ex As Exception Finally SqlConn.Close() OraConn.Close() End Try End Function Private Function GetMsgsFromOracle() As Boolean Try OraDataAdapter = New OleDb.OleDbDataAdapter OraDataTable = New System.Data.DataTable OraSelCmd = New OleDb.OleDbCommand GetMsgsFromOracle = False With OraSelCmd .CommandType = CommandType.Text .Connection = OraConn .CommandText = GetMsgSql End With OraDataAdapter.SelectCommand = OraSelCmd OraDataAdapter.Fill(OraDataTable) If OraDataTable.Rows.Count > 0 Then GetMsgsFromOracle = True End If Catch ex As Exception GetMsgsFromOracle = False End Try End Function Private Function ProcessMessages() As Boolean Try ProcessMessages = False PrepareSQLInsert() PrepOraDel() i = 0 Dim Method As Integer Dim OraDataRow As DataRow c.ErrorLog(Now.ToString & "--Going to call message sending procedure", 2) For Each OraDataRow In OraDataTable.Rows With OraDataRow Method = GetMethod(.Item(0)) SQLInsCmd.Parameters("RelLifeTime").Value = c.RelLifetime SQLInsCmd.Parameters("Param1").Value = Nothing SQLInsCmd.Parameters("ID").Value = GenerateTransactionID() ' Nothing SQLInsCmd.Parameters("UID").Value = Nothing SQLInsCmd.Parameters("Param").Value = Nothing SQLInsCmd.Parameters("Credit").Value = 0 SQLInsCmd.ExecuteNonQuery() 'check the return value If SQLInsCmd.Parameters("ReturnValue").Value = 1 And SQLInsCmd.Parameters("OutPutParam").Value = 0 Then 'success 'delete the input record from the source table once it is logged c.ErrorLog(Now.ToString & "--Moved record successfully", 2) OraDataAdapter.DeleteCommand.Parameters("P(0)").Value = OraDataRow.Item(6) OraDataAdapter.DeleteCommand.ExecuteNonQuery() c.ErrorLog(Now.ToString & "--Deleted record successfully", 2) OraDataAdapter.Update(OraDataTable) c.ErrorLog(Now.ToString & "--Committed record successfully", 2) i = i + 1 Else 'failure c.ErrorLog(Now.ToString & "--Failed to exec: " & c.DestIns & "Status: " & SQLInsCmd.Parameters("OutPutParam").Value & " and TrackId: " & SQLInsCmd.Parameters("TrackID").Value.ToString, 0) End If If File.Exists("stop.txt") Then c.ErrorLog(Now.ToString & "--Stop File Found", 1) 'ProcessMessages = True 'Exit Function Exit For End If End With Next OraDataAdapter.Update(OraDataTable) c.ErrorLog(Now.ToString & "--Updated Oracle Table", 1) c.ErrorLog(Now.ToString & "--Moved " & i & " records from Oracle to SQL Table", 1) ProcessMessages = True Catch ex As Exception ProcessMessages = False c.ErrorLog(Now.ToString & "--MoveMsgsToSQL: " & ex.Message, 0) Finally OraDataTable.Clear() OraDataTable.Dispose() OraDataAdapter.Dispose() OraDelCmd.Dispose() OraDelCmd = Nothing OraSelCmd = Nothing OraDataTable = Nothing OraDataAdapter = Nothing End Try End Function Public Function GenerateTransactionID() As Int64 Dim SeqNo As Int64 Dim qry As String Dim SqlTransCmd As New OleDb.OleDbCommand qry = " select seqno from StoreSeqNo" SqlTransCmd.CommandType = CommandType.Text SqlTransCmd.Connection = SqlConn SqlTransCmd.CommandText = qry SeqNo = SqlTransCmd.ExecuteScalar If SeqNo > 2147483647 Then qry = "update StoreSeqNo set seqno=1" SqlTransCmd.CommandText = qry SqlTransCmd.ExecuteNonQuery() GenerateTransactionID = 1 Else qry = "update StoreSeqNo set seqno=" & SeqNo + 1 SqlTransCmd.CommandText = qry SqlTransCmd.ExecuteNonQuery() GenerateTransactionID = SeqNo End If End Function Private Function PrepareSQLInsert() As Boolean 'function to prepare the insert statement for the insert into the SQL stmt using 'the sql procedure SMSProcessAndDispatch Try Dim dr As DataRow SQLInsCmd = New OleDb.OleDbCommand With SQLInsCmd .CommandType = CommandType.StoredProcedure .Connection = SqlConn .CommandText = SQLInsProc .Parameters.Add("ReturnValue", OleDb.OleDbType.Integer) .Parameters("ReturnValue").Direction = ParameterDirection.ReturnValue .Parameters.Add("OutPutParam", OleDb.OleDbType.Integer) .Parameters("OutPutParam").Direction = ParameterDirection.Output .Parameters.Add("TrackID", OleDb.OleDbType.VarChar, 70) .Parameters.Add("RelLifeTime", OleDb.OleDbType.TinyInt) .Parameters("RelLifeTime").Direction = ParameterDirection.Input .Parameters.Add("Param1", OleDb.OleDbType.VarChar, 160) .Parameters("Param1").Direction = ParameterDirection.Input .Parameters.Add("TransID", OleDb.OleDbType.VarChar, 70) .Parameters("TransID").Direction = ParameterDirection.Input .Parameters.Add("UID", OleDb.OleDbType.VarChar, 20) .Parameters("UID").Direction = ParameterDirection.Input .Parameters.Add("Param", OleDb.OleDbType.VarChar, 160) .Parameters("Param").Direction = ParameterDirection.Input .Parameters.Add("CheckCredit", OleDb.OleDbType.Integer) .Parameters("CheckCredit").Direction = ParameterDirection.Input .Prepare() End With Catch ex As Exception c.ErrorLog(Now.ToString & "--PrepareSQLInsert: " & ex.Message) End Try End Function Private Function PrepOraDel() As Boolean OraDelCmd = New OleDb.OleDbCommand Try PrepOraDel = False With OraDelCmd .CommandType = CommandType.Text .Connection = OraConn .CommandText = DelSrcSQL .Parameters.Add("P(0)", OleDb.OleDbType.VarChar, 160) 'RowID .Parameters("P(0)").Direction = ParameterDirection.Input .Prepare() End With OraDataAdapter.DeleteCommand = OraDelCmd PrepOraDel = True Catch ex As Exception PrepOraDel = False End Try End Function WHat i would like to know is, if there is anyway to speed up this program? Any ideas/suggestions would be highly appreciated... Regardss, Chetan

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Why does this C++ code result in a segmentation fault?

- by user69514

I keep getting a segmentation fault when the readAuthor() method is called. Does anybody know why this happens? I am supposed to use dynamic arrays, I know this would be so easy if I was using static array. #include <iostream> #include <string> #include <cstring> #include <cstdlib> using namespace std; /** declare arrays **/ int* isbnArr = new int[25]; char* authorArr = new char[25]; char* publisherArr = new char[25]; char* titleArr = new char[25]; int* editionArr = new int[25]; int* yearArr = new int[25]; int* pagesArr = new int[25]; float* retailPriceArr = new float[25]; float* discountedPriceArr = new float[25]; int* stockArr = new int[25]; /** function prototypes **/ int readIsbn(); char* readAuthor(); char* readPublisher(); char* readTitle(); int readEdition(); int readYear(); int readPages(); float readMsrp(); float readDiscountedPrice(); int readStockAmount(); void readonebook(int* isbn, char* author, char* title, char* publisher, int* edition, int* year, int* pages, float* msrp, float* discounted, int* inventory); int main() { bool stop = false; //flag when to stop loop int ind = 0; //index for current book while( !stop ){ cout << "Add book: press A: "; cout << "another thing here "; char choice; cin >> choice; if( choice == 'a' || choice == 'A' ){ readonebook(&isbnArr[ind], &authorArr[ind], &titleArr[ind], &publisherArr[ind], &editionArr[ind], &yearArr[ind], &pagesArr[ind], &retailPriceArr[ind], &discountedPriceArr[ind], &stockArr[ind]); test(&authorArr[ind]); ind++; } } return 0; } /** define functions **/ int readIsbn(){ int isbn; cout << "ISBN: "; cin >> isbn; return isbn; } char* readAuthor(){ char* author; cout << "Author: "; cin >> author; return author; } char* readPublisher(){ char* publisher = NULL; cout << "Publisher: "; cin >> publisher; return publisher; } char* readTitle(){ char* title = NULL; cout << "Title: "; cin >> title; return title; } int readEdition(){ int edition; cout << "Edition: "; cin >> edition; return edition; } int readYear(){ int year; cout << "Year: "; cin >> year; return year; } int readPages(){ int pages; cout << "Pages: "; cin >> pages; return pages; } float readMsrp(){ float price; cout << "Retail Price: "; cin >> price; return price; } float readDiscountedPrice(){ float price; cout << "Discounted Price: "; cin >> price; return price; } int readStockAmount(){ int amount; cout << "Stock Amount: "; cin >> amount; return amount; } void readonebook(int* isbn, char* author, char* title, char* publisher, int* edition, int* year, int* pages, float* msrp, float* discounted, int* inventory){ *isbn = readIsbn(); author = readAuthor(); title = readTitle(); publisher = readPublisher(); *edition = readEdition(); *year = readYear(); *pages = readPages(); *msrp = readMsrp(); *discounted = readDiscountedPrice(); *inventory = readStockAmount(); }

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more radio problems * sighs * been at this for one week now

- by jeansymolanza

hi guys really i don't know what im doing wrong... i am trying to process multiple radios with mysql. the first step has involved me echoing the results on the process page. each time i select any radio option it simply displays the first row result. this is that i see after trying to submit the form: Notifications Thank you. The notifications have been updated successfully. statusid: 14 notc2: 1 Return this is the code for the form: <div style="padding: 15px;"> <span class="loginfail" style="font-size:24px; font-weight: bold">Notifications</span><p> <?php include("progress_insertcomment.php"); ?> <?php // Make a MySQL Connection mysql_select_db("speedycm_data") or die(mysql_error()); $query_comment = "select * from tbl_alert order by id desc limit 1"; $comment = mysql_query($query_comment, $speedycms) or die(mysql_error()); $row_comment = mysql_fetch_assoc($comment); $totalRows_comment = mysql_num_rows($comment); ?>  <form method="post" action="<?php echo $_SERVER['PHP_SELF']; ?>"> <span id="sprytextarea1"> <textarea id='comment' name="comment" style="height: 75px; width:330px;"><?php echo $row_comment['comment']; ?></textarea> </span> <p> <button type="submit">Add</button> <input type="hidden" name="notc" value="1"/> </form>  <form method="post" action="<?php echo $_SERVER['PHP_SELF']; ?>"> <table border="0" cellspacing="2" cellpadding="2"> <?php if ( $row_comment == 0 ) { echo "<span style='font-size: 11px;'>No current alerts.</span>"; } else { // Get all the data from the "example" table $result = mysql_query("SELECT * FROM tbl_alert ORDER BY id DESC") or die(mysql_error()); while($rows=mysql_fetch_array($result)){ ?> <tr> <td> <?php echo "<div class='bubble'><div class='pimped'> <blockquote>" . $rows['comment'] . " </blockquote></div> <cite><strong>" . $rows['user'] . "</strong> @ " . $rows['date'] . "</cite> <span style='font-size: 10px;'> <p> <a href='editalert.php?id=". $rows['id'] ."' class='form' >Edit</a> • <a href='deletealert.php?id=". $rows['id'] ."' class='form'>Delete</a> </span> </div> "; ?> </td> <td valign="top" align="center"><div style="padding-left: 30px;"><span style="font-size: 10px;">Completed?</span> <p class="field switch">  <?php $status = $rows['status']; $id = $rows['id']; ?> <input type="radio" name="selstatus[<?php echo $id; ?>]" value="no" <?php if($status == 'yes') {echo 'checked';} else {echo '';} ?>/> <input type="radio" name="selstatus[<?php echo $id; ?>]" value="yes" <?php if($status == 'yes') {echo 'checked';} else {echo '';} ?>/> <input type="hidden" name="statusid" value="<?php echo $id; ?>"/> <label for="radio1" class="cb-enable <?php if($status == 'yes') {echo 'selected';} else {echo '';} ?>"><span>Yes</span></label> <label for="radio2" class="cb-disable <?php if($status == 'no') {echo 'selected';} else {echo '';} ?>"><span>No</span></label> </p> </div></td> </tr> <tr> <td></td> <?php } } ?> <td align="center"><div style="padding-left: 30px;"> <button type="submit">Update</button> <input type="hidden" name="notc2" value="1"/> </div></td> </tr> </table> </form> </div> this is the code for the processing page: <?php // 6) update notifications if (array_key_exists('notc2',$_POST)) { $update = $_POST['selstatus']; echo "<p style='font-size: 12px;'>Thank you. The notifications have been updated successfully.<p>"; foreach($_POST as $key => $value){ echo $key . ': ' . $value . '<br>'; } echo "<p><span style='font-size: 12px;'> <a onClick=\"history.go(-1)\" class='form'>Return</a></p> <p></span> "; exit; }; ?> how do i resolve this problem? please any help would be appreciated. i would like to display the id and value of the selected checkbox on the processing page before inserting it into the table. thanking you in advance. GOD BLESS

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OpenGL texture shifted somewhat to the left when applied to a quad

- by user308226

I'm a bit new to OpenGL and I've been having a problem with using textures. The texture seems to load fine, but when I run the program, the texture displays shifted a couple pixels to the left, with the section cut off by the shift appearing on the right side. I don't know if the problem here is in the my TGA loader or if it's the way I'm applying the texture to the quad. Here is the loader: #include "texture.h" #include <iostream> GLubyte uncompressedheader[12] = {0,0, 2,0,0,0,0,0,0,0,0,0}; GLubyte compressedheader[12] = {0,0,10,0,0,0,0,0,0,0,0,0}; TGA::TGA() { } //Private loading function called by LoadTGA. Loads uncompressed TGA files //Returns: TRUE on success, FALSE on failure bool TGA::LoadCompressedTGA(char *filename,ifstream &texturestream) { return false; } bool TGA::LoadUncompressedTGA(char *filename,ifstream &texturestream) { cout << "G position status:" << texturestream.tellg() << endl; texturestream.read((char*)header, sizeof(header)); //read 6 bytes into the file to get the tga header width = (GLuint)header[1] * 256 + (GLuint)header[0]; //read and calculate width and save height = (GLuint)header[3] * 256 + (GLuint)header[2]; //read and calculate height and save bpp = (GLuint)header[4]; //read bpp and save cout << bpp << endl; if((width <= 0) || (height <= 0) || ((bpp != 24) && (bpp !=32))) //check to make sure the height, width, and bpp are valid { return false; } if(bpp == 24) { type = GL_RGB; } else { type = GL_RGBA; } imagesize = ((bpp/8) * width * height); //determine size in bytes of the image cout << imagesize << endl; imagedata = new GLubyte[imagesize]; //allocate memory for our imagedata variable texturestream.read((char*)imagedata,imagesize); //read according the the size of the image and save into imagedata for(GLuint cswap = 0; cswap < (GLuint)imagesize; cswap += (bpp/8)) //loop through and reverse the tga's BGR format to RGB { imagedata[cswap] ^= imagedata[cswap+2] ^= //1st Byte XOR 3rd Byte XOR 1st Byte XOR 3rd Byte imagedata[cswap] ^= imagedata[cswap+2]; } texturestream.close(); //close ifstream because we're done with it cout << "image loaded" << endl; glGenTextures(1, &texID); // Generate OpenGL texture IDs glBindTexture(GL_TEXTURE_2D, texID); // Bind Our Texture glTexParameterf(GL_TEXTURE_2D, GL_TEXTURE_MIN_FILTER, GL_LINEAR); // Linear Filtered glTexParameterf(GL_TEXTURE_2D, GL_TEXTURE_MAG_FILTER, GL_LINEAR); glTexImage2D(GL_TEXTURE_2D, 0, type, width, height, 0, type, GL_UNSIGNED_BYTE, imagedata); delete imagedata; return true; } //Public loading function for TGA images. Opens TGA file and determines //its type, if any, then loads it and calls the appropriate function. //Returns: TRUE on success, FALSE on failure bool TGA::loadTGA(char *filename) { cout << width << endl; ifstream texturestream; texturestream.open(filename,ios::binary); texturestream.read((char*)header,sizeof(header)); //read 6 bytes into the file, its the header. //if it matches the uncompressed header's first 6 bytes, load it as uncompressed LoadUncompressedTGA(filename,texturestream); return true; } GLubyte* TGA::getImageData() { return imagedata; } GLuint& TGA::getTexID() { return texID; } And here's the quad: void Square::show() { glEnable(GL_TEXTURE_2D); glBindTexture(GL_TEXTURE_2D, texture.texID); //Move to offset glTranslatef( x, y, 0 ); //Start quad glBegin( GL_QUADS ); //Set color to white glColor4f( 1.0, 1.0, 1.0, 1.0 ); //Draw square glTexCoord2f(0.0f, 0.0f); glVertex3f( 0, 0, 0 ); glTexCoord2f(1.0f, 0.0f); glVertex3f( SQUARE_WIDTH, 0, 0 ); glTexCoord2f(1.0f, 1.0f); glVertex3f( SQUARE_WIDTH, SQUARE_HEIGHT, 0 ); glTexCoord2f(0.0f, 1.0f); glVertex3f( 0, SQUARE_HEIGHT, 0 ); //End quad glEnd(); //Reset glLoadIdentity(); }

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Unwanted character being added to string in C

- by Church

I have a program that gives you shipping addresses from an input file. However at the beginning of one of the strings, order.add_one, a number is being added to the beginning of the string, that number is equivalent to the variable "choice" every time. Why is it doing this? #include <stdio.h> #include <math.h> #include <string.h> //structure typedef struct {char cust_name[25]; char cust_id[3]; char add_one[30]; char add_two[30]; char bike; char risky; int number_ordered; char cust_information[500]; }ORDER; ORDER order; int main(void){ fflush(stdin); system ( "clear" ); //initialize variables float price; float m = 359.95; float s = 279.95; //while loop, runs until user declares they no longer wish to input orders while (1==1){ printf("Options: \nEnter Customer information manually : 1 \nSearch Customer by ID(input.txt reader) : 2 \n"); int option = 0; scanf(" %d", &option); if (option == 1){ //Print and scan statements printf("Enter Customer Information\n"); printf("Customer Name: "); scanf(" %[^\n]s", &order.cust_name); printf("\nEnter Address Line One: "); scanf(" %[^\n]s", &order.add_one); printf("\nEnter Addres Line Two: "); scanf(" %[^\n]s", &order.add_two); printf("\nHow Many Bicycles Are Ordered: "); scanf(" %d", &order.number_ordered); printf("\nWhat Type Of Bike Is Ordered\n M Mountain Bike \n S Street Bike"); printf("\nChoose One (M or S): "); scanf(" %c", &order.bike); printf("\nIs The Customer Risky (Y/N): "); scanf(" %c", &order.risky); system ( "clear" ); } if (option == 2){ FILE *fpt; fpt = fopen("input.txt", "r"); if (fpt==NULL){ printf("Text file did not open\n"); return 1; } printf("Enter Customer ID: "); scanf("%s", &order.cust_id); char choice; choice = order.cust_id[0]; char x[3]; int w, u, y, z; char a[10], b[10], c[10], d[10], e[20], f[10], g[10], i[1], j[1]; int h; printf("%s value of c", c); if (choice >='1'){ while ((w = fgetc(fpt)) != '\n' ){ } } if (choice >='2'){ while ((u = fgetc(fpt)) != '\n' ){ } } if (choice >='3'){ while ((y = fgetc(fpt)) != '\n' ){ } } if (choice >= '4'){ while ((z = fgetc(fpt)) != '\n' ){ } } printf("\n"); fscanf(fpt, "%s", x); fscanf(fpt, "%s", a); printf("%s", a); strcat(order.cust_name, a); fscanf(fpt, " %s", b); printf(" %s", b); strcat(order.cust_name, " "); strcat(order.cust_name, b); fscanf(fpt, "%s", c); printf(" %s", c); strcat(order.add_one, "\0"); strcat(order.add_one, c); fscanf(fpt, "%s", d); printf(" %s", d); strcat(order.add_one, " "); strcat(order.add_one, d); fscanf(fpt, "%s", e); printf(" %s", e); strcat(order.add_two, e); fscanf(fpt, "%s", f); printf(" %s", f); strcat(order.add_two, " "); strcat(order.add_two, f); fscanf(fpt, "%s", g); printf(" %s", g); strcat(order.add_two, " "); strcat(order.add_two, g); strcat(order.add_two, "\0"); fscanf(fpt, "%d", &h); printf(" %d", h); order.number_ordered = h; fscanf(fpt, "%s", i); printf(" %s", i); order.bike = i[0]; fscanf(fpt, "%s", j); printf(" %s", j); order.risky = j[0]; fclose(fpt); printf("%s %s %s %d %c %c", order.cust_name, order.add_one, order.add_two, order.number_ordered, order.bike, order.risky); }

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interactive login on perl cgi script- updated question [closed]

- by raindrop18

I am new for perl-cgi script. and my objective is to create user/pass interactive script to log and get information from multiple device at once. instead of add the user/pass credential on the script itself. since i am new if some one show me how to write the interactive part of the script. thanks much!!! here is the current code usr/local/bin/perl -wl use CGI ':standard'; use Net::Telnet::Cisco; # ### set the error fields to nulls $selerror = ""; # ### Input from the screen - make sure data was input if (param() and param('Switches') ne "" and (param('Mac') ne "" or param('Interface Description') ne "" or param('VLAN') ne "" )) { # ### Put the input devices into an array. @devices = param('Switches'); # ### format the header data print header(); print start_html(-title=>"ShowSwitches",-BGCOLOR=>'aqua'); print "\n",h1("<CENTER>Show Switches</CENTER>"); print "\n",hr(),"\n"; # ### Go thru the device array. foreach(@devices) { $error_msg = ""; $TAC_login_error = ""; $open_error = ""; $retry_open_error = ""; $prompt_error = ""; $password_error = ""; # ### Take input host and use to send to Telnet $host = $_; $session = Net::Telnet::Cisco->new(Errmode => 'return', Timeout => 30); # ### Connect to the host $session->open(Host =>"$host", Timeout => 15); $open_error = $session->errmsg; # ### Login with TACACS if host can be connected to if ($open_error eq "") { $session->login('USER', 'PASS'); $TAC_login_error = $session->errmsg; # ### Login with TACACS failed - try standard login if ($TAC_login_error ne "") { # ### Connect to host $session->open(Host =>"$host", Timeout => 15); $retry_open_error = $session->errmsg; # ### Wait for password prompt - multiple matches - devices may have different device prompts. if ($retry_open_error eq "") { $session->waitfor(Match => '/Password:.*$/', Match => '/Enter password:.*$/', Timeout => 20); $prompt_error = $session->errmsg; if ($prompt_error eq "") { # ### Input password $session->print('getmeout'); $password_error = $session->errmsg; $session->waitfor('/.*>$/'); $password_error = $session->errmsg; } } } } # ### No errors, then issue "show commands". if ($open_error eq "" and ($TAC_login_error eq "" or $retry_open_error eq "") and $prompt_error eq "" and $password_error eq "") { # ### Show Mac if (param('Mac')) { $cmd = 'sh mac'; @output = $session->cmd("$cmd"); $show_error = ""; $show_error = $session->errmsg; print "\n",h2($host . ' - ' . $cmd); if ($show_error ne "") { $error_msg = 'Error for show mac - ' . $show_error; print b($error_msg),(br); print hr(),"\n"; $error_msg = ''; } else { print pre(@output); print hr(),"\n"; } } # ### Show Interface Description if (param('Interface Description')) { $cmd = 'sh interface description'; @output = $session->cmd("$cmd"); $show_error = ""; $show_error = $session->errmsg; print "\n",h2($host . ' - ' . $cmd); if ($show_error ne "") { $error_msg = 'Error for show mac - ' . $show_error; print b($error_msg),(br); print hr(),"\n"; $error_msg = ''; } else { print pre(@output); print hr(),"\n"; } } # ### Show VLAN if (param('VLAN')) { $cmd = 'sh vlan'; @output = $session->cmd("$cmd"); $show_error = ""; $show_error = $session->errmsg; print "\n",h2($host . ' - ' . $cmd); if ($show_error ne "") { $error_msg = 'Error for show vlan - ' . $show_error; print b($error_msg),(br); print hr(),"\n"; $error_msg = ''; } else { print pre(@output); print hr(),"\n"; } } } elsif ($TAC_login_error ne "" and $password_error ne "") { $error_msg = "Error - $host " . $TAC_login_error . ' - possible incorrect TACACS or standard password parameters on device.'; } elsif ($open_error ne "") { $error_msg = "Error - $host " . $open_error . ' - cannot connect to host - is it down??'; } elsif ($prompt_error ne "") { $error_msg = "Error - $host " . $prompt_error . ' - password prompt not recognized - invalid TACACS (or user) password.'; } elsif ($password_error ne "") { $error_msg = "Error - $host " . $password_error . ' - possible incorrect user/password parameters on device.'; } if ($error_msg ne "" ) { print b($error_msg),(br); print hr(),"\n"; } print hr(),"\n"; print end_html(),"\n"; } } else { # ### No Show command was selected. if (param('Submit') and param('Mac') eq "" and param('Interface Description') eq "" and param('VLAN') eq "" ) { $selerror = 'No Show Displays were selected. Try again please!!'; } elsif # ### No switch was selected. (param('Submit') and param('Switches') eq "") { $selerror = 'No devices were selected. Try again please!!'; } # ### This formats the initial Show Web page. print header(-Pragma=>'no-cache'), start_html(-title=>"Show Displays",-BGCOLOR=>'aqua'), h1("<CENTER>Show Switches</CENTER>"),hr(), start_form(), b("Select Show Commands:"), br(), br(), checkbox(-name=>'Mac'), checkbox(-name=>'Interface Description'), checkbox(-name=>'VLAN'), br(),br(),hr(),br(), b("Select One or More Devices:"), br(), br(), scrolling_list (-name => 'Switches', -default=> "NONE", -values => ['cs6a', 'cs7a', 'cs7b', 'cs8b', 'cs9a', 'c9b', 'csa' ], -multiple => 'true', -size => 7, ), p(submit('Submit'),reset('Reset')), b($selerror), end_form(),hr(), end_html(); } #

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Redrawing content of UIWebView

- by btate

I have a bunch of webviews with static html content that I'm putting in a scroll view as pages. That works fine, but having 20 something full screen subviews of the scroll view is causing some lag. I solved that by only placing 5 at a time in there. The current view, and the two next and two previous. The problem now is that any web view that is not a subview of the scroll view is not drawing correctly based on the device orientation. The frame of the webview is printing out correct, but the actual content is drawing in portrait mode. So there is essentially a strip of blank space to the right of the content. How do I go about re rendering the content without reloading the page? Here's the relevant code: - (void)didRotateFromInterfaceOrientation:(UIInterfaceOrientation) fromInterfaceOrientation{ [self resizeSubViews]; } - (void) setupWebViews{ if(_webViews == nil) _webViews = [[NSMutableArray alloc] init]; // This is where the navigation would come into play as far as loading up available web views [_webViews removeAllObjects]; // for loop here to create web views for (int i = 0; i < 20; i++) { //CDCWebViewController *webView = [[[MyInternalWebView alloc] initWithFrame:_webViewWrapper.frame] retain]; MyInternalWebView *page = [[[MyInternalWebView alloc] init] retain]; [page loadRequest:[NSURLRequest requestWithURL:_url]]; [page setCdcIWVdelegate:self]; [page setTestIndex:i]; [page setPageIndex:i]; [_webViews addObject:page]; [self loadScrollViewWithPage:i]; } [self clearUnusedWebViews:_pageControl.currentPage]; } - (void) setupWebViewWrapper{ // a page is the width of the scroll view _webViewWrapper.pagingEnabled = YES; _pageControl = [[[UIPageControl alloc] init] retain]; _webViewWrapper.showsHorizontalScrollIndicator = NO; _webViewWrapper.showsVerticalScrollIndicator = NO; _webViewWrapper.scrollsToTop = NO; _webViewWrapper.delegate = self; _pageControl.numberOfPages = [_webViews count]; _pageControl.currentPage = 0; } - (void) resizeSubViews{ // The frame is set in IB _webViewWrapper.contentSize = CGSizeMake(_webViewWrapper.frame.size.width * [_webViews count], _webViewWrapper.frame.size.height); // Move the content offset. _webViewWrapper.contentOffset = CGPointMake(_webViewWrapper.frame.size.width * _pageControl.currentPage, _webViewWrapper.contentOffset.y); for (MyInternalWebView *subview in _webViewWrapper.subviews) { // Reset the frame height and width here? CGRect frame = _webViewWrapper.frame; frame.origin.x = frame.size.width * subview.pageIndex; frame.origin.y = 0; [subview setFrame:frame]; } } //***************************************************** //* //* ScrollView Functions //* //***************************************************** - (void)loadScrollViewWithPage:(int)page { // Make sure we're not out of bounds if (page < 0) return; if (page >= [_webViews count]) return; MyInternalWebView *webView = [_webViews objectAtIndex:page]; // Add the preloaded webview to the scrollview if it's not there already if (nil == [webView superview]) { CGRect frame = _webViewWrapper.frame; //NSLog(@"width = %f", frame.size.width); frame.origin.x = frame.size.width * page; frame.origin.y = 0; //NSLog(@"setting frame for page %d %@", page, NSStringFromCGRect(frame)); [webView setFrame:frame]; [_webViewWrapper addSubview:[_webViews objectAtIndex:page]]; // Now that the new one is loaded, clear what doesn't need to be here [self clearUnusedWebViews:page]; } } - (void) clearUnusedWebViews: (NSInteger) page{ for (int i = 0; i < [_webViews count]; i++) { if ((page - i) <= 2 && i - page <= 2) { continue; } [[_webViews objectAtIndex:i] removeFromSuperview]; } } - (void)scrollViewDidScroll:(UIScrollView *)sender { // Switch the indicator when more than 50% of the previous/next page is visible CGFloat pageWidth = _webViewWrapper.frame.size.width; NSInteger page = floor((_webViewWrapper.contentOffset.x - pageWidth / 2) / pageWidth) + 1; _pageControl.currentPage = page; // load the visible page and the page on either side of it (to avoid flashes when the user starts scrolling) [self loadScrollViewWithPage:page - 2]; [self loadScrollViewWithPage:page - 1]; [self loadScrollViewWithPage:page]; [self loadScrollViewWithPage:page + 1]; [self loadScrollViewWithPage:page + 2]; }

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Does this language feature already exist?

- by Pindatjuh

I'm currently developing a new language for programming in a continuous environment (compare it to electrical engineering), and I've got some ideas on a certain language construction. Let me explain the feature by explanation and then by definition: x = a U b; Where x is a variable and a and b are other variables (or static values). This works like a union between a and b; no duplicates and no specific order. with(x) { // regular 'with' usage; using the global interpretation of "x" x = 5; // will replace the original definition of "x = a U b;" } with(x = a) { // this code block is executed when the "x" variable // has the "a" variable assigned. All references in // this code-block to "x" are references to "a". So saying: x = 5; // would only change the variable "a". If the variable "a" // later on changes, x still equals to 5, in this fashion: // 'x = a U b U 5;' // '[currentscope] = 5;' // thus, 'a = 5;' } with(x = b) { // same but with "b" } with(x != a) { // here the "x" variable refers to any variable // but "a"; thus saying x = 5; // is equal to the rewriting of // 'x = a U b U 5;' // 'b = 5;' (since it was the scope of this block) } with(x = (a U b)) { // guaranteed that "x" is 'a U b'; interacting with "x" // will interact with both "a" and "b". x = 5; // makes both "a" and "b" equal to 5; also the "x" variable // is updated to contain: // 'x = a U b U 5;' // '[currentscope] = 5;' // 'a U b = 5;' // and thus: 'a = 5; b = 5;'. } // etc. In the above, all code-blocks are executed, but the "scope" changes in each block how x is interpreted. In the first block, x is guaranteed to be a: thus interacting with x inside that block will interact on a. The second and the third code-block are only equal in this situation (because not a: then there only remains b). The last block guarantees that x is at least a or b. Further more; U is not the "bitwise or operator", but I've called it the "and/or"-operator. Its definition is: "U" = "and" U "or" (On my blog, http://cplang.wordpress.com/2009/12/19/binop-and-or/, there is more (mathematical) background information on this operator. It's loosely based on sets. Using different syntax, changed it in this question.) Update: more examples. print = "Hello world!" U "How are you?"; // this will print // both values, but the // order doesn't matter. // 'userkey' is a variable containing a key. with(userkey = "a") { print = userkey; // will only print "a". } with(userkey = ("shift" U "a")) { // pressed both "shift" and the "a" key. print = userkey; // will "print" shift and "a", even // if the user also pressed "ctrl": // the interpretation of "userkey" is changed, // such that it only contains the matched cases. } with((userkey = "shift") U (userkey = "a")) { // same as if-statement above this one, showing the distributivity. } x = 5 U 6 U 7; y = x + x; // will be: // y = (5 U 6 U 7) + (5 U 6 U 7) // = 10 U 11 U 12 U 13 U 14 somewantedkey = "ctrl" U "alt" U "space" with(userkey = somewantedkey) { // must match all elements of "somewantedkey" // (distributed the Boolean equals operated) // thus only executed when all the defined keys are pressed } with(somewantedkey = userkey) { // matches only one of the provided "somewantedkey" // thus when only "space" is pressed, this block is executed. } Update2: more examples and some more context. with(x = (a U b)) { // this } // can be written as with((x = a) U (x = b)) { // this: changing the variable like x = 5; // will be rewritten as: // a = 5 and b = 5 } Some background information: I'm building a language which is "time-independent", like Java is "platform-independant". Everything stated in the language is "as is", and is continuously actively executed. This means; the programmer does not know in which order (unless explicitly stated using constructions) elements are, nor when statements are executed. The language is completely separated from the "time"-concept, i.e. it's continuously executed: with(a < 5) { a++; } // this is a loop-structure; // how and when it's executed isn't known however. with(a) { // everytime the "a" variable changes, this code-block is executed. b = 4; with(b < 3) { // runs only three times. } with(b > 0) { b = b - 1; // runs four times } } Update 3: After pondering on the type of this language feature; it closely resemblances Netbeans Platform's Lookup, where each "with"-statement a synchronized agent is, working on it's specific "filter" of objects. Instead of type-based, this is variable-based (fundamentally quite the same; just a different way of identifiying objects). I greatly thank all of you for providing me with very insightful information and links/hints to great topics I can research. Thanks. I do not know if this construction already exists, so that's my question: does this language feature already exist?

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Couldn't match expected type - Haskell Code

- by wvyar

I'm trying to learn Haskell, but the small bit of sample code I tried to write is running into a fairly large amount of "Couldn't match expected type" errors. Can anyone give me some guidance as to what I'm doing wrong/how I should go about this? These are the errors, but I'm not really sure how I should be writing my code. toDoSchedulerSimple.hs:6:14: Couldn't match expected type `[t0]' with actual type `IO String' In the return type of a call of `readFile' In a stmt of a 'do' block: f <- readFile inFile In the expression: do { f <- readFile inFile; lines f } toDoSchedulerSimple.hs:27:9: Couldn't match expected type `[a0]' with actual type `IO ()' In the return type of a call of `putStr' In a stmt of a 'do' block: putStr "Enter task name: " In the expression: do { putStr "Enter task name: "; task <- getLine; return inFileArray : task } toDoSchedulerSimple.hs:34:9: Couldn't match expected type `IO ()' with actual type `[a0]' In a stmt of a 'do' block: putStrLn "Your task is: " ++ (inFileArray !! i) In the expression: do { i <- randomRIO (0, (length inFileArray - 1)); putStrLn "Your task is: " ++ (inFileArray !! i) } In an equation for `getTask': getTask inFileArray = do { i <- randomRIO (0, (length inFileArray - 1)); putStrLn "Your task is: " ++ (inFileArray !! i) } toDoSchedulerSimple.hs:41:9: Couldn't match expected type `[a0]' with actual type `IO ()' In the return type of a call of `putStr' In a stmt of a 'do' block: putStr "Enter the task you would like to end: " In the expression: do { putStr "Enter the task you would like to end: "; task <- getLine; filter (endTaskCheck task) inFileArray } toDoSchedulerSimple.hs:60:53: Couldn't match expected type `IO ()' with actual type `[String] -> IO ()' In a stmt of a 'do' block: schedulerSimpleMain In the expression: do { (getTask inFileArray); schedulerSimpleMain } In a case alternative: "get-task" -> do { (getTask inFileArray); schedulerSimpleMain } This is the code itself. I think it's fairly straightforward, but the idea is to run a loop, take input, and perform actions based off of it by calling other functions. import System.Random (randomRIO) import Data.List (lines) initializeFile :: [char] -> [String] initializeFile inFile = do f <- readFile inFile let parsedFile = lines f return parsedFile displayHelp :: IO() displayHelp = do putStrLn "Welcome to To Do Scheduler Simple, written in Haskell." putStrLn "Here are some commands you might find useful:" putStrLn " 'help' : Display this menu." putStrLn " 'quit' : Exit the program." putStrLn " 'new-task' : Create a new task." putStrLn " 'get-task' : Randomly select a task." putStrLn " 'end-task' : Mark a task as finished." putStrLn " 'view-tasks' : View all of your tasks." quit :: IO() quit = do putStrLn "We're very sad to see you go...:(" putStrLn "Come back soon!" createTask :: [String] -> [String] createTask inFileArray = do putStr "Enter task name: " task <- getLine return inFileArray:task getTask :: [String] -> IO() getTask inFileArray = do i <- randomRIO (0, (length inFileArray - 1)) putStrLn "Your task is: " ++ (inFileArray !! i) endTaskCheck :: String -> String -> Bool endTaskCheck str1 str2 = str1 /= str2 endTask :: [String] -> [String] endTask inFileArray = do putStr "Enter the task you would like to end: " task <- getLine return filter (endTaskCheck task) inFileArray viewTasks :: [String] -> IO() viewTasks inFileArray = case inFileArray of [] -> do putStrLn "\nEnd of tasks." _ -> do putStrLn (head inFileArray) viewTasks (tail inFileArray) schedulerSimpleMain :: [String] -> IO() schedulerSimpleMain inFileArray = do putStr "SchedulerSimple> " input <- getLine case input of "help" -> displayHelp "quit" -> quit "new-task" -> schedulerSimpleMain (createTask inFileArray) "get-task" -> do (getTask inFileArray); schedulerSimpleMain "end-task" -> schedulerSimpleMain (endTask inFileArray) "view-tasks" -> do (viewTasks inFileArray); schedulerSimpleMain _ -> do putStrLn "Invalid input."; schedulerSimpleMain main :: IO() main = do putStr "What is the name of the schedule? " sName <- getLine schedulerSimpleMain (initializeFile sName) Thanks, and apologies if this isn't the correct place to be asking such a question.

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How to get data from dynamically created EditText views and insert it into an array?

- by Snwspeckle

So basically what I need my program to do at this point is that when I click the submit button, I need to loop through each dynamic row of the ListView and grab the value in the EditText view and then insert that into an array which I will do further calculations after. Here is my code right now. package com.hello_world; import java.util.ArrayList; import com.hello_world.ByteInputActivity.MyAdapter.ViewHolder; import android.app.Activity; import android.content.Context; import android.os.Bundle; import android.util.Log; import android.view.KeyEvent; import android.view.LayoutInflater; import android.view.View; import android.view.View.OnFocusChangeListener; import android.view.View.OnKeyListener; import android.view.ViewGroup; import android.widget.BaseAdapter; import android.widget.Button; import android.widget.EditText; import android.widget.ListView; import android.widget.TextView; public class ByteInputActivity extends Activity { private ListView myList; private MyAdapter myAdapter; private Integer resQuestions; private Integer indexVal = 0; private View caption; ViewHolder holder; ArrayList<Integer> intArrayList = new ArrayList<Integer>(); @Override protected void onCreate(Bundle savedInstanceState) { super.onCreate(savedInstanceState); setContentView(R.layout.fieldlist); //Gets number of questions from MainActivity Bundle extras = getIntent().getExtras(); if(extras !=null) { resQuestions = extras.getInt("index"); } myList = (ListView) findViewById(R.id.FieldList); myList.setItemsCanFocus(true); myAdapter = new MyAdapter(); myList.setAdapter(myAdapter); Button submit = (Button) findViewById(R.id.btn_New); submit.setOnClickListener(new View.OnClickListener() { @Override public void onClick(View v) { for (int i = 0; i < myList.getCount() ; i++) { View vListSortOrder; vListSortOrder = myList.getChildAt(i); String temp = holder.caption.getText().toString(); Log.e("VALUES", "" +temp); } } }); } public class MyAdapter extends BaseAdapter { private LayoutInflater mInflater; public ArrayList myItems = new ArrayList(); public MyAdapter() { mInflater = (LayoutInflater) getSystemService(Context.LAYOUT_INFLATER_SERVICE); for (int i = 0; i < resQuestions; i++) { ListItem listItem = new ListItem(); listItem.caption = "Index " + i; listItem.indexText = "Index " + i; myItems.add(listItem); indexVal += 1; } notifyDataSetChanged(); } public int getCount() { return myItems.size(); } public Object getItem(int position) { return position; } public long getItemId(int position) { return position; } public View getView(int position, View convertView, ViewGroup parent) { if (convertView == null) { holder = new ViewHolder(); convertView = mInflater.inflate(R.layout.item, null); holder.indexText = (TextView) convertView .findViewById(R.id.textView1); holder.caption = (EditText) convertView .findViewById(R.id.ItemCaption); convertView.setTag(holder); } else { holder = (ViewHolder) convertView.getTag(); } //Fill EditText with the value you have in data source holder.caption.setText(""); holder.caption.setId(position); holder.indexText.setText("Index " + position); holder.indexText.setId(position); //we need to update adapter once we finish with editing holder.caption.setOnFocusChangeListener(new OnFocusChangeListener() { public void onFocusChange(View v, boolean hasFocus) { if (!hasFocus){ final int position = v.getId(); final EditText Caption = (EditText) v; myItems.set(position, Caption.getText().toString()); } } }); return convertView; } class ViewHolder { EditText caption; TextView indexText; } class ListItem { String caption; String indexText; } } }

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Fastest way to move records from an Oracle database into SQL Server

- by user347748

Ok this is the scenario... I have a table in Oracle that acts like a queue... A VB.net program reads the queue and calls a stored proc in SQL Server that processes and then inserts the message into another SQL Server table and then deletes the record from the oracle table. We use a DataReader to read the records from Oracle and then call the stored proc for each of the records. The program seems to be a little slow. The stored procedure itself isn't slow. The SP by itself when called in a loop can process about 2000 records in 20 seconds. But when called from the .Net program, the execution time is about 5 records per second. I have seen that most of the time consumed is in calling the stored procedure and waiting for it to return. Is there a better way of doing this? Here is a snippet of the actual code Function StartDataXfer() As Boolean Dim status As Boolean = False Try SqlConn.Open() OraConn.Open() c.ErrorLog(Now.ToString & "--Going to Get the messages from oracle", 1) If GetMsgsFromOracle() Then c.ErrorLog(Now.ToString & "--Got messages from oracle", 1) If ProcessMessages() Then c.ErrorLog(Now.ToString & "--Finished Processing all messages in the queue", 0) status = True Else c.ErrorLog(Now.ToString & "--Failed to Process all messages in the queue", 0) status = False End If Else status = True End If StartDataXfer = status Catch ex As Exception Finally SqlConn.Close() OraConn.Close() End Try End Function Private Function GetMsgsFromOracle() As Boolean Try OraDataAdapter = New OleDb.OleDbDataAdapter OraDataTable = New System.Data.DataTable OraSelCmd = New OleDb.OleDbCommand GetMsgsFromOracle = False With OraSelCmd .CommandType = CommandType.Text .Connection = OraConn .CommandText = GetMsgSql End With OraDataAdapter.SelectCommand = OraSelCmd OraDataAdapter.Fill(OraDataTable) If OraDataTable.Rows.Count > 0 Then GetMsgsFromOracle = True End If Catch ex As Exception GetMsgsFromOracle = False End Try End Function Private Function ProcessMessages() As Boolean Try ProcessMessages = False PrepareSQLInsert() PrepOraDel() i = 0 Dim Method As Integer Dim OraDataRow As DataRow c.ErrorLog(Now.ToString & "--Going to call message sending procedure", 2) For Each OraDataRow In OraDataTable.Rows With OraDataRow Method = GetMethod(.Item(0)) SQLInsCmd.Parameters("RelLifeTime").Value = c.RelLifetime SQLInsCmd.Parameters("Param1").Value = Nothing SQLInsCmd.Parameters("ID").Value = GenerateTransactionID() ' Nothing SQLInsCmd.Parameters("UID").Value = Nothing SQLInsCmd.Parameters("Param").Value = Nothing SQLInsCmd.Parameters("Credit").Value = 0 SQLInsCmd.ExecuteNonQuery() 'check the return value If SQLInsCmd.Parameters("ReturnValue").Value = 1 And SQLInsCmd.Parameters("OutPutParam").Value = 0 Then 'success 'delete the input record from the source table once it is logged c.ErrorLog(Now.ToString & "--Moved record successfully", 2) OraDataAdapter.DeleteCommand.Parameters("P(0)").Value = OraDataRow.Item(6) OraDataAdapter.DeleteCommand.ExecuteNonQuery() c.ErrorLog(Now.ToString & "--Deleted record successfully", 2) OraDataAdapter.Update(OraDataTable) c.ErrorLog(Now.ToString & "--Committed record successfully", 2) i = i + 1 Else 'failure c.ErrorLog(Now.ToString & "--Failed to exec: " & c.DestIns & "Status: " & SQLInsCmd.Parameters("OutPutParam").Value & " and TrackId: " & SQLInsCmd.Parameters("TrackID").Value.ToString, 0) End If If File.Exists("stop.txt") Then c.ErrorLog(Now.ToString & "--Stop File Found", 1) 'ProcessMessages = True 'Exit Function Exit For End If End With Next OraDataAdapter.Update(OraDataTable) c.ErrorLog(Now.ToString & "--Updated Oracle Table", 1) c.ErrorLog(Now.ToString & "--Moved " & i & " records from Oracle to SQL Table", 1) ProcessMessages = True Catch ex As Exception ProcessMessages = False c.ErrorLog(Now.ToString & "--MoveMsgsToSQL: " & ex.Message, 0) Finally OraDataTable.Clear() OraDataTable.Dispose() OraDataAdapter.Dispose() OraDelCmd.Dispose() OraDelCmd = Nothing OraSelCmd = Nothing OraDataTable = Nothing OraDataAdapter = Nothing End Try End Function Public Function GenerateTransactionID() As Int64 Dim SeqNo As Int64 Dim qry As String Dim SqlTransCmd As New OleDb.OleDbCommand qry = " select seqno from StoreSeqNo" SqlTransCmd.CommandType = CommandType.Text SqlTransCmd.Connection = SqlConn SqlTransCmd.CommandText = qry SeqNo = SqlTransCmd.ExecuteScalar If SeqNo > 2147483647 Then qry = "update StoreSeqNo set seqno=1" SqlTransCmd.CommandText = qry SqlTransCmd.ExecuteNonQuery() GenerateTransactionID = 1 Else qry = "update StoreSeqNo set seqno=" & SeqNo + 1 SqlTransCmd.CommandText = qry SqlTransCmd.ExecuteNonQuery() GenerateTransactionID = SeqNo End If End Function Private Function PrepareSQLInsert() As Boolean 'function to prepare the insert statement for the insert into the SQL stmt using 'the sql procedure SMSProcessAndDispatch Try Dim dr As DataRow SQLInsCmd = New OleDb.OleDbCommand With SQLInsCmd .CommandType = CommandType.StoredProcedure .Connection = SqlConn .CommandText = SQLInsProc .Parameters.Add("ReturnValue", OleDb.OleDbType.Integer) .Parameters("ReturnValue").Direction = ParameterDirection.ReturnValue .Parameters.Add("OutPutParam", OleDb.OleDbType.Integer) .Parameters("OutPutParam").Direction = ParameterDirection.Output .Parameters.Add("TrackID", OleDb.OleDbType.VarChar, 70) .Parameters.Add("RelLifeTime", OleDb.OleDbType.TinyInt) .Parameters("RelLifeTime").Direction = ParameterDirection.Input .Parameters.Add("Param1", OleDb.OleDbType.VarChar, 160) .Parameters("Param1").Direction = ParameterDirection.Input .Parameters.Add("TransID", OleDb.OleDbType.VarChar, 70) .Parameters("TransID").Direction = ParameterDirection.Input .Parameters.Add("UID", OleDb.OleDbType.VarChar, 20) .Parameters("UID").Direction = ParameterDirection.Input .Parameters.Add("Param", OleDb.OleDbType.VarChar, 160) .Parameters("Param").Direction = ParameterDirection.Input .Parameters.Add("CheckCredit", OleDb.OleDbType.Integer) .Parameters("CheckCredit").Direction = ParameterDirection.Input .Prepare() End With Catch ex As Exception c.ErrorLog(Now.ToString & "--PrepareSQLInsert: " & ex.Message) End Try End Function Private Function PrepOraDel() As Boolean OraDelCmd = New OleDb.OleDbCommand Try PrepOraDel = False With OraDelCmd .CommandType = CommandType.Text .Connection = OraConn .CommandText = DelSrcSQL .Parameters.Add("P(0)", OleDb.OleDbType.VarChar, 160) 'RowID .Parameters("P(0)").Direction = ParameterDirection.Input .Prepare() End With OraDataAdapter.DeleteCommand = OraDelCmd PrepOraDel = True Catch ex As Exception PrepOraDel = False End Try End Function WHat i would like to know is, if there is anyway to speed up this program? Any ideas/suggestions would be highly appreciated... Regardss, Chetan

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Building my first Javascript Application (jQuery), struggling on something

- by Jason Wells

I'd really appreciate recommendations on the most efficient way to approach this. I'm building a simple javascript application which displays a list of records and allows the user to edit a record by clicking an "Edit" link in the records row. The user also can click the "Add" link to pop open a dialog allowing them to add a new record. Here's a working prototype of this: http://jsfiddle.net/FfRcG/ You'll note if you click "Edit" a dialog pops up with some canned values. And, if you click "Add", a dialog pops up with empty values. I need help on how to approach two problems I believe we need to pass our index to our edit dialog and reference the values within the JSON, but I am unsure how to pass the index when the user clicks edit. It bothers me that the Edit and Add div contents are so similiar (Edit just pre populates the values). I feel like there is a more efficient way of doing this but am at a loss. Here is my code for reference $(document).ready( function(){ // Our JSON (This would actually be coming from an AJAX database call) people = { "COLUMNS":["DATEMODIFIED", "NAME","AGE"], "DATA":[ ["9/6/2012", "Person 1","32"], ["9/5/2012","Person 2","23"] ] } // Here we loop over our JSON and build our HTML (Will refactor to use templating eventually) members = people.DATA; var newcontent = '<table width=50%><tr><td>date</td><td>name</td><td>age</td><td></td></tr>'; for(var i=0;i<members.length;i++) { newcontent+= '<tr id="member'+i+'"><td>' + members[i][0] + '</td>'; newcontent+= '<td>' + members[i][1] + '</td>'; newcontent+= '<td>' + members[i][2] + '</td>'; newcontent+= '<td><a href="#" class="edit" id=edit'+i+'>Edit</a></td><td>'; } newcontent += "</table>"; $("#result").html(newcontent); // Bind a dialog to the edit link $(".edit").click( function(){ // Trigger our dialog to open $("#edit").dialog("open"); // Not sure the most efficient way to change our dialog field values $("#name").val() // ??? alert($()); return false; }); // Bind a dialog to the add link $(".edit").click( function(){ // Trigger our dialog to open $("#add").dialog("open"); return false; }); // Bind a dialog to our edit DIV $("#edit").dialog(); // Bind a dialog to our add DIV $("#add").dialog(); }); And here's the HTML <h1>People</h1> <a href="#" class="add">Add a new person</a>  <div id="result"></div>  <div id="edit"> <form> <p>Name:<br/><input type="text" id="name" value="foo"></p> <p>Age:<br/><input type="text" id="age" value="33"></p> <input type="submit" value="Save" id="submitEdit"> </form> </div>  <div id="add"> <form> <p>Name:<br/><input type="text" id="name"></p> <p>Age:<br/><input type="text" id="age"></p> <input type="submit" value="Save" id="submitEdit"> </form> </div>

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HTG Reviews the CODE Keyboard: Old School Construction Meets Modern Amenities

- by Jason Fitzpatrick

There’s nothing quite as satisfying as the smooth and crisp action of a well built keyboard. If you’re tired of mushy keys and cheap feeling keyboards, a well-constructed mechanical keyboard is a welcome respite from the $10 keyboard that came with your computer. Read on as we put the CODE mechanical keyboard through the paces. What is the CODE Keyboard? The CODE keyboard is a collaboration between manufacturer WASD Keyboards and Jeff Atwood of Coding Horror (the guy behind the Stack Exchange network and Discourse forum software). Atwood’s focus was incorporating the best of traditional mechanical keyboards and the best of modern keyboard usability improvements. In his own words: The world is awash in terrible, crappy, no name how-cheap-can-we-make-it keyboards. There are a few dozen better mechanical keyboard options out there. I’ve owned and used at least six different expensive mechanical keyboards, but I wasn’t satisfied with any of them, either: they didn’t have backlighting, were ugly, had terrible design, or were missing basic functions like media keys. That’s why I originally contacted Weyman Kwong of WASD Keyboards way back in early 2012. I told him that the state of keyboards was unacceptable to me as a geek, and I proposed a partnership wherein I was willing to work with him to do whatever it takes to produce a truly great mechanical keyboard. Even the ardent skeptic who questions whether Atwood has indeed created a truly great mechanical keyboard certainly can’t argue with the position he starts from: there are so many agonizingly crappy keyboards out there. Even worse, in our opinion, is that unless you’re a typist of a certain vintage there’s a good chance you’ve never actually typed on a really nice keyboard. Those that didn’t start using computers until the mid-to-late 1990s most likely have always typed on modern mushy-key keyboards and never known the joy of typing on a really responsive and crisp mechanical keyboard. Is our preference for and love of mechanical keyboards shining through here? Good. We’re not even going to try and hide it. So where does the CODE keyboard stack up in pantheon of keyboards? Read on as we walk you through the simple setup and our experience using the CODE. Setting Up the CODE Keyboard Although the setup of the CODE keyboard is essentially plug and play, there are two distinct setup steps that you likely haven’t had to perform on a previous keyboard. Both highlight the degree of care put into the keyboard and the amount of customization available. Inside the box you’ll find the keyboard, a micro USB cable, a USB-to-PS2 adapter, and a tool which you may be unfamiliar with: a key puller. We’ll return to the key puller in a moment. Unlike the majority of keyboards on the market, the cord isn’t permanently affixed to the keyboard. What does this mean for you? Aside from the obvious need to plug it in yourself, it makes it dead simple to repair your own keyboard cord if it gets attacked by a pet, mangled in a mechanism on your desk, or otherwise damaged. It also makes it easy to take advantage of the cable routing channels in on the underside of the keyboard to route your cable exactly where you want it. While we’re staring at the underside of the keyboard, check out those beefy rubber feet. By peripherals standards they’re huge (and there is six instead of the usual four). Once you plunk the keyboard down where you want it, it might as well be glued down the rubber feet work so well. After you’ve secured the cable and adjusted it to your liking, there is one more task before plug the keyboard into the computer. On the bottom left-hand side of the keyboard, you’ll find a small recess in the plastic with some dip switches inside: The dip switches are there to switch hardware functions for various operating systems, keyboard layouts, and to enable/disable function keys. By toggling the dip switches you can change the keyboard from QWERTY mode to Dvorak mode and Colemak mode, the two most popular alternative keyboard configurations. You can also use the switches to enable Mac-functionality (for Command/Option keys). One of our favorite little toggles is the SW3 dip switch: you can disable the Caps Lock key; goodbye accidentally pressing Caps when you mean to press Shift. You can review the entire dip switch configuration chart here. The quick-start for Windows users is simple: double check that all the switches are in the off position (as seen in the photo above) and then simply toggle SW6 on to enable the media and backlighting function keys (this turns the menu key on the keyboard into a function key as typically found on laptop keyboards). After adjusting the dip switches to your liking, plug the keyboard into an open USB port on your computer (or into your PS/2 port using the included adapter). Design, Layout, and Backlighting The CODE keyboard comes in two flavors, a traditional 87-key layout (no number pad) and a traditional 104-key layout (number pad on the right hand side). We identify the layout as traditional because, despite some modern trapping and sneaky shortcuts, the actual form factor of the keyboard from the shape of the keys to the spacing and position is as classic as it comes. You won’t have to learn a new keyboard layout and spend weeks conditioning yourself to a smaller than normal backspace key or a PgUp/PgDn pair in an unconventional location. Just because the keyboard is very conventional in layout, however, doesn’t mean you’ll be missing modern amenities like media-control keys. The following additional functions are hidden in the F11, F12, Pause button, and the 2×6 grid formed by the Insert and Delete rows: keyboard illumination brightness, keyboard illumination on/off, mute, and then the typical play/pause, forward/backward, stop, and volume +/- in Insert and Delete rows, respectively. While we weren’t sure what we’d think of the function-key system at first (especially after retiring a Microsoft Sidewinder keyboard with a huge and easily accessible volume knob on it), it took less than a day for us to adapt to using the Fn key, located next to the right Ctrl key, to adjust our media playback on the fly. Keyboard backlighting is a largely hit-or-miss undertaking but the CODE keyboard nails it. Not only does it have pleasant and easily adjustable through-the-keys lighting but the key switches the keys themselves are attached to are mounted to a steel plate with white paint. Enough of the light reflects off the interior cavity of the keys and then diffuses across the white plate to provide nice even illumination in between the keys. Highlighting the steel plate beneath the keys brings us to the actual construction of the keyboard. It’s rock solid. The 87-key model, the one we tested, is 2.0 pounds. The 104-key is nearly a half pound heavier at 2.42 pounds. Between the steel plate, the extra-thick PCB board beneath the steel plate, and the thick ABS plastic housing, the keyboard has very solid feel to it. Combine that heft with the previously mentioned thick rubber feet and you have a tank-like keyboard that won’t budge a millimeter during normal use. Examining The Keys This is the section of the review the hardcore typists and keyboard ninjas have been waiting for. We’ve looked at the layout of the keyboard, we’ve looked at the general construction of it, but what about the actual keys? There are a wide variety of keyboard construction techniques but the vast majority of modern keyboards use a rubber-dome construction. The key is floated in a plastic frame over a rubber membrane that has a little rubber dome for each key. The press of the physical key compresses the rubber dome downwards and a little bit of conductive material on the inside of the dome’s apex connects with the circuit board. Despite the near ubiquity of the design, many people dislike it. The principal complaint is that dome keyboards require a complete compression to register a keystroke; keyboard designers and enthusiasts refer to this as “bottoming out”. In other words, the register the “b” key, you need to completely press that key down. As such it slows you down and requires additional pressure and movement that, over the course of tens of thousands of keystrokes, adds up to a whole lot of wasted time and fatigue. The CODE keyboard features key switches manufactured by Cherry, a company that has manufactured key switches since the 1960s. Specifically the CODE features Cherry MX Clear switches. These switches feature the same classic design of the other Cherry switches (such as the MX Blue and Brown switch lineups) but they are significantly quieter (yes this is a mechanical keyboard, but no, your neighbors won’t think you’re firing off a machine gun) as they lack the audible click found in most Cherry switches. This isn’t to say that they keyboard doesn’t have a nice audible key press sound when the key is fully depressed, but that the key mechanism isn’t doesn’t create a loud click sound when triggered. One of the great features of the Cherry MX clear is a tactile “bump” that indicates the key has been compressed enough to register the stroke. For touch typists the very subtle tactile feedback is a great indicator that you can move on to the next stroke and provides a welcome speed boost. Even if you’re not trying to break any word-per-minute records, that little bump when pressing the key is satisfying. The Cherry key switches, in addition to providing a much more pleasant typing experience, are also significantly more durable than dome-style key switch. Rubber dome switch membrane keyboards are typically rated for 5-10 million contacts whereas the Cherry mechanical switches are rated for 50 million contacts. You’d have to write the next War and Peace and follow that up with A Tale of Two Cities: Zombie Edition, and then turn around and transcribe them both into a dozen different languages to even begin putting a tiny dent in the lifecycle of this keyboard. So what do the switches look like under the classicly styled keys? You can take a look yourself with the included key puller. Slide the loop between the keys and then gently beneath the key you wish to remove: Wiggle the key puller gently back and forth while exerting a gentle upward pressure to pop the key off; You can repeat the process for every key, if you ever find yourself needing to extract piles of cat hair, Cheeto dust, or other foreign objects from your keyboard. There it is, the naked switch, the source of that wonderful crisp action with the tactile bump on each keystroke. The last feature worthy of a mention is the N-key rollover functionality of the keyboard. This is a feature you simply won’t find on non-mechanical keyboards and even gaming keyboards typically only have any sort of key roller on the high-frequency keys like WASD. So what is N-key rollover and why do you care? On a typical mass-produced rubber-dome keyboard you cannot simultaneously press more than two keys as the third one doesn’t register. PS/2 keyboards allow for unlimited rollover (in other words you can’t out type the keyboard as all of your keystrokes, no matter how fast, will register); if you use the CODE keyboard with the PS/2 adapter you gain this ability. If you don’t use the PS/2 adapter and use the native USB, you still get 6-key rollover (and the CTRL, ALT, and SHIFT don’t count towards the 6) so realistically you still won’t be able to out type the computer as even the more finger twisting keyboard combos and high speed typing will still fall well within the 6-key rollover. The rollover absolutely doesn’t matter if you’re a slow hunt-and-peck typist, but if you’ve read this far into a keyboard review there’s a good chance that you’re a serious typist and that kind of quality construction and high-number key rollover is a fantastic feature. The Good, The Bad, and the Verdict We’ve put the CODE keyboard through the paces, we’ve played games with it, typed articles with it, left lengthy comments on Reddit, and otherwise used and abused it like we would any other keyboard. The Good: The construction is rock solid. In an emergency, we’re confident we could use the keyboard as a blunt weapon (and then resume using it later in the day with no ill effect on the keyboard). The Cherry switches are an absolute pleasure to type on; the Clear variety found in the CODE keyboard offer a really nice middle-ground between the gun-shot clack of a louder mechanical switch and the quietness of a lesser-quality dome keyboard without sacrificing quality. Touch typists will love the subtle tactile bump feedback. Dip switch system makes it very easy for users on different systems and with different keyboard layout needs to switch between operating system and keyboard layouts. If you’re investing a chunk of change in a keyboard it’s nice to know you can take it with you to a different operating system or “upgrade” it to a new layout if you decide to take up Dvorak-style typing. The backlighting is perfect. You can adjust it from a barely-visible glow to a blazing light-up-the-room brightness. Whatever your intesity preference, the white-coated steel backplate does a great job diffusing the light between the keys. You can easily remove the keys for cleaning (or to rearrange the letters to support a new keyboard layout). The weight of the unit combined with the extra thick rubber feet keep it planted exactly where you place it on the desk. The Bad: While you’re getting your money’s worth, the $150 price tag is a shock when compared to the $20-60 price tags you find on lower-end keyboards. People used to large dedicated media keys independent of the traditional key layout (such as the large buttons and volume controls found on many modern keyboards) might be off put by the Fn-key style media controls on the CODE. The Verdict: The keyboard is clearly and heavily influenced by the needs of serious typists. Whether you’re a programmer, transcriptionist, or just somebody that wants to leave the lengthiest article comments the Internet has ever seen, the CODE keyboard offers a rock solid typing experience. Yes, $150 isn’t pocket change, but the quality of the CODE keyboard is so high and the typing experience is so enjoyable, you’re easily getting ten times the value you’d get out of purchasing a lesser keyboard. Even compared to other mechanical keyboards on the market, like the Das Keyboard, you’re still getting more for your money as other mechanical keyboards don’t come with the lovely-to-type-on Cherry MX Clear switches, back lighting, and hardware-based operating system keyboard layout switching. If it’s in your budget to upgrade your keyboard (especially if you’ve been slogging along with a low-end rubber-dome keyboard) there’s no good reason to not pickup a CODE keyboard. Key animation courtesy of Geekhack.org user Lethal Squirrel.

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Azure Mobile Services: what files does it consist of?

- by svdoever

Azure Mobile Services is a platform that provides a small set of functionality consisting of authentication, custom data tables, custom API’s, scheduling scripts and push notifications to be used as the back-end of a mobile application or if you want, any application or web site. As described in my previous post Azure Mobile Services: lessons learned the documentation on what can be used in the custom scripts is a bit minimalistic. The list below of all files the complete Azure Mobile Services platform consists of ca shed some light on what is available in the platform. In following posts I will provide more detailed information on what we can conclude from this list of files. Below are the available files as available in the Azure Mobile Services platform. The bold files are files that describe your data model, api scripts, scheduler scripts and table scripts. Those are the files you configure/construct to provide the “configuration”/implementation of you mobile service. The files are located in a folder like C:\DWASFiles\Sites\youreservice\VirtualDirectory0\site\wwwroot. One file is missing in the list below and that is the event log file C:\DWASFiles\Sites\youreservice\VirtualDirectory0\site\LogFiles\eventlog.xml where your messages written with for example console.log() and exception catched by the system are written. NOTA BENE: the Azure Mobile Services system is a system that is under full development, new releases may change the list of files. ./app.js ./App_Data/config/datamodel.json ./App_Data/config/scripts/api/youreapi.js ./App_Data/config/scripts/api/youreapi.json ./App_Data/config/scripts/scheduler/placeholder ./App_Data/config/scripts/scheduler/youresheduler.js ./App_Data/config/scripts/shared/placeholder ./App_Data/config/scripts/table/placeholder ./App_Data/config/scripts/table/yourtable.insert.js ./App_Data/config/scripts/table/yourtable.update.js ./App_Data/config/scripts/table/yourtable.delete.js ./App_Data/config/scripts/table/yourtable.read.js ./node_modules/apn/index.js ./node_modules/apn/lib/connection.js ./node_modules/apn/lib/device.js ./node_modules/apn/lib/errors.js ./node_modules/apn/lib/feedback.js ./node_modules/apn/lib/notification.js ./node_modules/apn/lib/util.js ./node_modules/apn/node_modules/q/package.json ./node_modules/apn/node_modules/q/q.js 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New features of C# 4.0

This article covers New features of C# 4.0. Article has been divided into below sections. Introduction. Dynamic Lookup. Named and Optional Arguments. Features for COM interop. Variance. Relationship with Visual Basic. Resources. Other interested readings… 22 New Features of Visual Studio 2008 for .NET Professionals 50 New Features of SQL Server 2008 IIS 7.0 New features Introduction It is now close to a year since Microsoft Visual C# 3.0 shipped as part of Visual Studio 2008. In the VS Managed Languages team we are hard at work on creating the next version of the language (with the unsurprising working title of C# 4.0), and this document is a first public description of the planned language features as we currently see them. Please be advised that all this is in early stages of production and is subject to change. Part of the reason for sharing our plans in public so early is precisely to get the kind of feedback that will cause us to improve the final product before it rolls out. Simultaneously with the publication of this whitepaper, a first public CTP (community technology preview) of Visual Studio 2010 is going out as a Virtual PC image for everyone to try. Please use it to play and experiment with the features, and let us know of any thoughts you have. We ask for your understanding and patience working with very early bits, where especially new or newly implemented features do not have the quality or stability of a final product. The aim of the CTP is not to give you a productive work environment but to give you the best possible impression of what we are working on for the next release. The CTP contains a number of walkthroughs, some of which highlight the new language features of C# 4.0. Those are excellent for getting a hands-on guided tour through the details of some common scenarios for the features. You may consider this whitepaper a companion document to these walkthroughs, complementing them with a focus on the overall language features and how they work, as opposed to the specifics of the concrete scenarios. C# 4.0 The major theme for C# 4.0 is dynamic programming. Increasingly, objects are “dynamic” in the sense that their structure and behavior is not captured by a static type, or at least not one that the compiler knows about when compiling your program. Some examples include a. objects from dynamic programming languages, such as Python or Ruby b. COM objects accessed through IDispatch c. ordinary .NET types accessed through reflection d. objects with changing structure, such as HTML DOM objects While C# remains a statically typed language, we aim to vastly improve the interaction with such objects. A secondary theme is co-evolution with Visual Basic. Going forward we will aim to maintain the individual character of each language, but at the same time important new features should be introduced in both languages at the same time. They should be differentiated more by style and feel than by feature set. The new features in C# 4.0 fall into four groups: Dynamic lookup Dynamic lookup allows you to write method, operator and indexer calls, property and field accesses, and even object invocations which bypass the C# static type checking and instead gets resolved at runtime. Named and optional parameters Parameters in C# can now be specified as optional by providing a default value for them in a member declaration. When the member is invoked, optional arguments can be omitted. Furthermore, any argument can be passed by parameter name instead of position. COM specific interop features Dynamic lookup as well as named and optional parameters both help making programming against COM less painful than today. On top of that, however, we are adding a number of other small features that further improve the interop experience. Variance It used to be that an IEnumerable<string> wasn’t an IEnumerable<object>. Now it is – C# embraces type safe “co-and contravariance” and common BCL types are updated to take advantage of that. Dynamic Lookup Dynamic lookup allows you a unified approach to invoking things dynamically. With dynamic lookup, when you have an object in your hand you do not need to worry about whether it comes from COM, IronPython, the HTML DOM or reflection; you just apply operations to it and leave it to the runtime to figure out what exactly those operations mean for that particular object. This affords you enormous flexibility, and can greatly simplify your code, but it does come with a significant drawback: Static typing is not maintained for these operations. A dynamic object is assumed at compile time to support any operation, and only at runtime will you get an error if it wasn’t so. Oftentimes this will be no loss, because the object wouldn’t have a static type anyway, in other cases it is a tradeoff between brevity and safety. In order to facilitate this tradeoff, it is a design goal of C# to allow you to opt in or opt out of dynamic behavior on every single call. The dynamic type C# 4.0 introduces a new static type called dynamic. When you have an object of type dynamic you can “do things to it” that are resolved only at runtime: dynamic d = GetDynamicObject(…); d.M(7); The C# compiler allows you to call a method with any name and any arguments on d because it is of type dynamic. At runtime the actual object that d refers to will be examined to determine what it means to “call M with an int” on it. The type dynamic can be thought of as a special version of the type object, which signals that the object can be used dynamically. It is easy to opt in or out of dynamic behavior: any object can be implicitly converted to dynamic, “suspending belief” until runtime. Conversely, there is an “assignment conversion” from dynamic to any other type, which allows implicit conversion in assignment-like constructs: dynamic d = 7; // implicit conversion int i = d; // assignment conversion Dynamic operations Not only method calls, but also field and property accesses, indexer and operator calls and even delegate invocations can be dispatched dynamically: dynamic d = GetDynamicObject(…); d.M(7); // calling methods d.f = d.P; // getting and settings fields and properties d[“one”] = d[“two”]; // getting and setting thorugh indexers int i = d + 3; // calling operators string s = d(5,7); // invoking as a delegate The role of the C# compiler here is simply to package up the necessary information about “what is being done to d”, so that the runtime can pick it up and determine what the exact meaning of it is given an actual object d. Think of it as deferring part of the compiler’s job to runtime. The result of any dynamic operation is itself of type dynamic. Runtime lookup At runtime a dynamic operation is dispatched according to the nature of its target object d: COM objects If d is a COM object, the operation is dispatched dynamically through COM IDispatch. This allows calling to COM types that don’t have a Primary Interop Assembly (PIA), and relying on COM features that don’t have a counterpart in C#, such as indexed properties and default properties. Dynamic objects If d implements the interface IDynamicObject d itself is asked to perform the operation. Thus by implementing IDynamicObject a type can completely redefine the meaning of dynamic operations. This is used intensively by dynamic languages such as IronPython and IronRuby to implement their own dynamic object models. It will also be used by APIs, e.g. by the HTML DOM to allow direct access to the object’s properties using property syntax. Plain objects Otherwise d is a standard .NET object, and the operation will be dispatched using reflection on its type and a C# “runtime binder” which implements C#’s lookup and overload resolution semantics at runtime. This is essentially a part of the C# compiler running as a runtime component to “finish the work” on dynamic operations that was deferred by the static compiler. Example Assume the following code: dynamic d1 = new Foo(); dynamic d2 = new Bar(); string s; d1.M(s, d2, 3, null); Because the receiver of the call to M is dynamic, the C# compiler does not try to resolve the meaning of the call. Instead it stashes away information for the runtime about the call. This information (often referred to as the “payload”) is essentially equivalent to: “Perform an instance method call of M with the following arguments: 1. a string 2. a dynamic 3. a literal int 3 4. a literal object null” At runtime, assume that the actual type Foo of d1 is not a COM type and does not implement IDynamicObject. In this case the C# runtime binder picks up to finish the overload resolution job based on runtime type information, proceeding as follows: 1. Reflection is used to obtain the actual runtime types of the two objects, d1 and d2, that did not have a static type (or rather had the static type dynamic). The result is Foo for d1 and Bar for d2. 2. Method lookup and overload resolution is performed on the type Foo with the call M(string,Bar,3,null) using ordinary C# semantics. 3. If the method is found it is invoked; otherwise a runtime exception is thrown. Overload resolution with dynamic arguments Even if the receiver of a method call is of a static type, overload resolution can still happen at runtime. This can happen if one or more of the arguments have the type dynamic: Foo foo = new Foo(); dynamic d = new Bar(); var result = foo.M(d); The C# runtime binder will choose between the statically known overloads of M on Foo, based on the runtime type of d, namely Bar. The result is again of type dynamic. The Dynamic Language Runtime An important component in the underlying implementation of dynamic lookup is the Dynamic Language Runtime (DLR), which is a new API in .NET 4.0. The DLR provides most of the infrastructure behind not only C# dynamic lookup but also the implementation of several dynamic programming languages on .NET, such as IronPython and IronRuby. Through this common infrastructure a high degree of interoperability is ensured, but just as importantly the DLR provides excellent caching mechanisms which serve to greatly enhance the efficiency of runtime dispatch. To the user of dynamic lookup in C#, the DLR is invisible except for the improved efficiency. However, if you want to implement your own dynamically dispatched objects, the IDynamicObject interface allows you to interoperate with the DLR and plug in your own behavior. This is a rather advanced task, which requires you to understand a good deal more about the inner workings of the DLR. For API writers, however, it can definitely be worth the trouble in order to vastly improve the usability of e.g. a library representing an inherently dynamic domain. Open issues There are a few limitations and things that might work differently than you would expect. · The DLR allows objects to be created from objects that represent classes. However, the current implementation of C# doesn’t have syntax to support this. · Dynamic lookup will not be able to find extension methods. Whether extension methods apply or not depends on the static context of the call (i.e. which using clauses occur), and this context information is not currently kept as part of the payload. · Anonymous functions (i.e. lambda expressions) cannot appear as arguments to a dynamic method call. The compiler cannot bind (i.e. “understand”) an anonymous function without knowing what type it is converted to. One consequence of these limitations is that you cannot easily use LINQ queries over dynamic objects: dynamic collection = …; var result = collection.Select(e => e + 5); If the Select method is an extension method, dynamic lookup will not find it. Even if it is an instance method, the above does not compile, because a lambda expression cannot be passed as an argument to a dynamic operation. There are no plans to address these limitations in C# 4.0. Named and Optional Arguments Named and optional parameters are really two distinct features, but are often useful together. Optional parameters allow you to omit arguments to member invocations, whereas named arguments is a way to provide an argument using the name of the corresponding parameter instead of relying on its position in the parameter list. Some APIs, most notably COM interfaces such as the Office automation APIs, are written specifically with named and optional parameters in mind. Up until now it has been very painful to call into these APIs from C#, with sometimes as many as thirty arguments having to be explicitly passed, most of which have reasonable default values and could be omitted. Even in APIs for .NET however you sometimes find yourself compelled to write many overloads of a method with different combinations of parameters, in order to provide maximum usability to the callers. Optional parameters are a useful alternative for these situations. Optional parameters A parameter is declared optional simply by providing a default value for it: public void M(int x, int y = 5, int z = 7); Here y and z are optional parameters and can be omitted in calls: M(1, 2, 3); // ordinary call of M M(1, 2); // omitting z – equivalent to M(1, 2, 7) M(1); // omitting both y and z – equivalent to M(1, 5, 7) Named and optional arguments C# 4.0 does not permit you to omit arguments between commas as in M(1,,3). This could lead to highly unreadable comma-counting code. Instead any argument can be passed by name. Thus if you want to omit only y from a call of M you can write: M(1, z: 3); // passing z by name or M(x: 1, z: 3); // passing both x and z by name or even M(z: 3, x: 1); // reversing the order of arguments All forms are equivalent, except that arguments are always evaluated in the order they appear, so in the last example the 3 is evaluated before the 1. Optional and named arguments can be used not only with methods but also with indexers and constructors. Overload resolution Named and optional arguments affect overload resolution, but the changes are relatively simple: A signature is applicable if all its parameters are either optional or have exactly one corresponding argument (by name or position) in the call which is convertible to the parameter type. Betterness rules on conversions are only applied for arguments that are explicitly given – omitted optional arguments are ignored for betterness purposes. If two signatures are equally good, one that does not omit optional parameters is preferred. M(string s, int i = 1); M(object o); M(int i, string s = “Hello”); M(int i); M(5); Given these overloads, we can see the working of the rules above. M(string,int) is not applicable because 5 doesn’t convert to string. M(int,string) is applicable because its second parameter is optional, and so, obviously are M(object) and M(int). M(int,string) and M(int) are both better than M(object) because the conversion from 5 to int is better than the conversion from 5 to object. Finally M(int) is better than M(int,string) because no optional arguments are omitted. Thus the method that gets called is M(int). Features for COM interop Dynamic lookup as well as named and optional parameters greatly improve the experience of interoperating with COM APIs such as the Office Automation APIs. In order to remove even more of the speed bumps, a couple of small COM-specific features are also added to C# 4.0. Dynamic import Many COM methods accept and return variant types, which are represented in the PIAs as object. In the vast majority of cases, a programmer calling these methods already knows the static type of a returned object from context, but explicitly has to perform a cast on the returned value to make use of that knowledge. These casts are so common that they constitute a major nuisance. In order to facilitate a smoother experience, you can now choose to import these COM APIs in such a way that variants are instead represented using the type dynamic. In other words, from your point of view, COM signatures now have occurrences of dynamic instead of object in them. This means that you can easily access members directly off a returned object, or you can assign it to a strongly typed local variable without having to cast. To illustrate, you can now say excel.Cells[1, 1].Value = "Hello"; instead of ((Excel.Range)excel.Cells[1, 1]).Value2 = "Hello"; and Excel.Range range = excel.Cells[1, 1]; instead of Excel.Range range = (Excel.Range)excel.Cells[1, 1]; Compiling without PIAs Primary Interop Assemblies are large .NET assemblies generated from COM interfaces to facilitate strongly typed interoperability. They provide great support at design time, where your experience of the interop is as good as if the types where really defined in .NET. However, at runtime these large assemblies can easily bloat your program, and also cause versioning issues because they are distributed independently of your application. The no-PIA feature allows you to continue to use PIAs at design time without having them around at runtime. Instead, the C# compiler will bake the small part of the PIA that a program actually uses directly into its assembly. At runtime the PIA does not have to be loaded. Omitting ref Because of a different programming model, many COM APIs contain a lot of reference parameters. Contrary to refs in C#, these are typically not meant to mutate a passed-in argument for the subsequent benefit of the caller, but are simply another way of passing value parameters. It therefore seems unreasonable that a C# programmer should have to create temporary variables for all such ref parameters and pass these by reference. Instead, specifically for COM methods, the C# compiler will allow you to pass arguments by value to such a method, and will automatically generate temporary variables to hold the passed-in values, subsequently discarding these when the call returns. In this way the caller sees value semantics, and will not experience any side effects, but the called method still gets a reference. Open issues A few COM interface features still are not surfaced in C#. Most notably these include indexed properties and default properties. As mentioned above these will be respected if you access COM dynamically, but statically typed C# code will still not recognize them. There are currently no plans to address these remaining speed bumps in C# 4.0. Variance An aspect of generics that often comes across as surprising is that the following is illegal: IList<string> strings = new List<string>(); IList<object> objects = strings; The second assignment is disallowed because strings does not have the same element type as objects. There is a perfectly good reason for this. If it were allowed you could write: objects[0] = 5; string s = strings[0]; Allowing an int to be inserted into a list of strings and subsequently extracted as a string. This would be a breach of type safety. However, there are certain interfaces where the above cannot occur, notably where there is no way to insert an object into the collection. Such an interface is IEnumerable<T>. If instead you say: IEnumerable<object> objects = strings; There is no way we can put the wrong kind of thing into strings through objects, because objects doesn’t have a method that takes an element in. Variance is about allowing assignments such as this in cases where it is safe. The result is that a lot of situations that were previously surprising now just work. Covariance In .NET 4.0 the IEnumerable<T> interface will be declared in the following way: public interface IEnumerable<out T> : IEnumerable { IEnumerator<T> GetEnumerator(); } public interface IEnumerator<out T> : IEnumerator { bool MoveNext(); T Current { get; } } The “out” in these declarations signifies that the T can only occur in output position in the interface – the compiler will complain otherwise. In return for this restriction, the interface becomes “covariant” in T, which means that an IEnumerable<A> is considered an IEnumerable<B> if A has a reference conversion to B. As a result, any sequence of strings is also e.g. a sequence of objects. This is useful e.g. in many LINQ methods. Using the declarations above: var result = strings.Union(objects); // succeeds with an IEnumerable<object> This would previously have been disallowed, and you would have had to to some cumbersome wrapping to get the two sequences to have the same element type. Contravariance Type parameters can also have an “in” modifier, restricting them to occur only in input positions. An example is IComparer<T>: public interface IComparer<in T> { public int Compare(T left, T right); } The somewhat baffling result is that an IComparer<object> can in fact be considered an IComparer<string>! It makes sense when you think about it: If a comparer can compare any two objects, it can certainly also compare two strings. This property is referred to as contravariance. A generic type can have both in and out modifiers on its type parameters, as is the case with the Func<…> delegate types: public delegate TResult Func<in TArg, out TResult>(TArg arg); Obviously the argument only ever comes in, and the result only ever comes out. Therefore a Func<object,string> can in fact be used as a Func<string,object>. Limitations Variant type parameters can only be declared on interfaces and delegate types, due to a restriction in the CLR. Variance only applies when there is a reference conversion between the type arguments. For instance, an IEnumerable<int> is not an IEnumerable<object> because the conversion from int to object is a boxing conversion, not a reference conversion. Also please note that the CTP does not contain the new versions of the .NET types mentioned above. In order to experiment with variance you have to declare your own variant interfaces and delegate types. COM Example Here is a larger Office automation example that shows many of the new C# features in action. using System; using System.Diagnostics; using System.Linq; using Excel = Microsoft.Office.Interop.Excel; using Word = Microsoft.Office.Interop.Word; class Program { static void Main(string[] args) { var excel = new Excel.Application(); excel.Visible = true; excel.Workbooks.Add(); // optional arguments omitted excel.Cells[1, 1].Value = "Process Name"; // no casts; Value dynamically excel.Cells[1, 2].Value = "Memory Usage"; // accessed var processes = Process.GetProcesses() .OrderByDescending(p => p.WorkingSet) .Take(10); int i = 2; foreach (var p in processes) { excel.Cells[i, 1].Value = p.ProcessName; // no casts excel.Cells[i, 2].Value = p.WorkingSet; // no casts i++; } Excel.Range range = excel.Cells[1, 1]; // no casts Excel.Chart chart = excel.ActiveWorkbook.Charts. Add(After: excel.ActiveSheet); // named and optional arguments chart.ChartWizard( Source: range.CurrentRegion, Title: "Memory Usage in " + Environment.MachineName); //named+optional chart.ChartStyle = 45; chart.CopyPicture(Excel.XlPictureAppearance.xlScreen, Excel.XlCopyPictureFormat.xlBitmap, Excel.XlPictureAppearance.xlScreen); var word = new Word.Application(); word.Visible = true; word.Documents.Add(); // optional arguments word.Selection.Paste(); } } The code is much more terse and readable than the C# 3.0 counterpart. Note especially how the Value property is accessed dynamically. This is actually an indexed property, i.e. a property that takes an argument; something which C# does not understand. However the argument is optional. Since the access is dynamic, it goes through the runtime COM binder which knows to substitute the default value and call the indexed property. Thus, dynamic COM allows you to avoid accesses to the puzzling Value2 property of Excel ranges. Relationship with Visual Basic A number of the features introduced to C# 4.0 already exist or will be introduced in some form or other in Visual Basic: · Late binding in VB is similar in many ways to dynamic lookup in C#, and can be expected to make more use of the DLR in the future, leading to further parity with C#. · Named and optional arguments have been part of Visual Basic for a long time, and the C# version of the feature is explicitly engineered with maximal VB interoperability in mind. · NoPIA and variance are both being introduced to VB and C# at the same time. VB in turn is adding a number of features that have hitherto been a mainstay of C#. As a result future versions of C# and VB will have much better feature parity, for the benefit of everyone. Resources All available resources concerning C# 4.0 can be accessed through the C# Dev Center. Specifically, this white paper and other resources can be found at the Code Gallery site. Enjoy! span.fullpost {display:none;}

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The Incremental Architect’s Napkin - #5 - Design functions for extensibility and readability

- by Ralf Westphal

Originally posted on: http://geekswithblogs.net/theArchitectsNapkin/archive/2014/08/24/the-incremental-architectrsquos-napkin---5---design-functions-for.aspx The functionality of programs is entered via Entry Points. So what we´re talking about when designing software is a bunch of functions handling the requests represented by and flowing in through those Entry Points. Designing software thus consists of at least three phases: Analyzing the requirements to find the Entry Points and their signatures Designing the functionality to be executed when those Entry Points get triggered Implementing the functionality according to the design aka coding I presume, you´re familiar with phase 1 in some way. And I guess you´re proficient in implementing functionality in some programming language. But in my experience developers in general are not experienced in going through an explicit phase 2. “Designing functionality? What´s that supposed to mean?” you might already have thought. Here´s my definition: To design functionality (or functional design for short) means thinking about… well, functions. You find a solution for what´s supposed to happen when an Entry Point gets triggered in terms of functions. A conceptual solution that is, because those functions only exist in your head (or on paper) during this phase. But you may have guess that, because it´s “design” not “coding”. And here is, what functional design is not: It´s not about logic. Logic is expressions (e.g. +, -, && etc.) and control statements (e.g. if, switch, for, while etc.). Also I consider calling external APIs as logic. It´s equally basic. It´s what code needs to do in order to deliver some functionality or quality. Logic is what´s doing that needs to be done by software. Transformations are either done through expressions or API-calls. And then there is alternative control flow depending on the result of some expression. Basically it´s just jumps in Assembler, sometimes to go forward (if, switch), sometimes to go backward (for, while, do). But calling your own function is not logic. It´s not necessary to produce any outcome. Functionality is not enhanced by adding functions (subroutine calls) to your code. Nor is quality increased by adding functions. No performance gain, no higher scalability etc. through functions. Functions are not relevant to functionality. Strange, isn´t it. What they are important for is security of investment. By introducing functions into our code we can become more productive (re-use) and can increase evolvability (higher unterstandability, easier to keep code consistent). That´s no small feat, however. Evolvable code can hardly be overestimated. That´s why to me functional design is so important. It´s at the core of software development. To sum this up: Functional design is on a level of abstraction above (!) logical design or algorithmic design. Functional design is only done until you get to a point where each function is so simple you are very confident you can easily code it. Functional design an logical design (which mostly is coding, but can also be done using pseudo code or flow charts) are complementary. Software needs both. If you start coding right away you end up in a tangled mess very quickly. Then you need back out through refactoring. Functional design on the other hand is bloodless without actual code. It´s just a theory with no experiments to prove it. But how to do functional design? An example of functional design Let´s assume a program to de-duplicate strings. The user enters a number of strings separated by commas, e.g. a, b, a, c, d, b, e, c, a. And the program is supposed to clear this list of all doubles, e.g. a, b, c, d, e. There is only one Entry Point to this program: the user triggers the de-duplication by starting the program with the string list on the command line C:\>deduplicate "a, b, a, c, d, b, e, c, a" a, b, c, d, e …or by clicking on a GUI button. This leads to the Entry Point function to get called. It´s the program´s main function in case of the batch version or a button click event handler in the GUI version. That´s the physical Entry Point so to speak. It´s inevitable. What then happens is a three step process: Transform the input data from the user into a request. Call the request handler. Transform the output of the request handler into a tangible result for the user. Or to phrase it a bit more generally: Accept input. Transform input into output. Present output. This does not mean any of these steps requires a lot of effort. Maybe it´s just one line of code to accomplish it. Nevertheless it´s a distinct step in doing the processing behind an Entry Point. Call it an aspect or a responsibility - and you will realize it most likely deserves a function of its own to satisfy the Single Responsibility Principle (SRP). Interestingly the above list of steps is already functional design. There is no logic, but nevertheless the solution is described - albeit on a higher level of abstraction than you might have done yourself. But it´s still on a meta-level. The application to the domain at hand is easy, though: Accept string list from command line De-duplicate Present de-duplicated strings on standard output And this concrete list of processing steps can easily be transformed into code:static void Main(string[] args) { var input = Accept_string_list(args); var output = Deduplicate(input); Present_deduplicated_string_list(output); } Instead of a big problem there are three much smaller problems now. If you think each of those is trivial to implement, then go for it. You can stop the functional design at this point. But maybe, just maybe, you´re not so sure how to go about with the de-duplication for example. Then just implement what´s easy right now, e.g.private static string Accept_string_list(string[] args) { return args[0]; } private static void Present_deduplicated_string_list( string[] output) { var line = string.Join(", ", output); Console.WriteLine(line); } Accept_string_list() contains logic in the form of an API-call. Present_deduplicated_string_list() contains logic in the form of an expression and an API-call. And then repeat the functional design for the remaining processing step. What´s left is the domain logic: de-duplicating a list of strings. How should that be done? Without any logic at our disposal during functional design you´re left with just functions. So which functions could make up the de-duplication? Here´s a suggestion: De-duplicate Parse the input string into a true list of strings. Register each string in a dictionary/map/set. That way duplicates get cast away. Transform the data structure into a list of unique strings. Processing step 2 obviously was the core of the solution. That´s where real creativity was needed. That´s the core of the domain. But now after this refinement the implementation of each step is easy again:private static string[] Parse_string_list(string input) { return input.Split(',') .Select(s => s.Trim()) .ToArray(); } private static Dictionary<string,object> Compile_unique_strings(string[] strings) { return strings.Aggregate( new Dictionary<string, object>(), (agg, s) => { agg[s] = null; return agg; }); } private static string[] Serialize_unique_strings( Dictionary<string,object> dict) { return dict.Keys.ToArray(); } With these three additional functions Main() now looks like this:static void Main(string[] args) { var input = Accept_string_list(args); var strings = Parse_string_list(input); var dict = Compile_unique_strings(strings); var output = Serialize_unique_strings(dict); Present_deduplicated_string_list(output); } I think that´s very understandable code: just read it from top to bottom and you know how the solution to the problem works. It´s a mirror image of the initial design: Accept string list from command line Parse the input string into a true list of strings. Register each string in a dictionary/map/set. That way duplicates get cast away. Transform the data structure into a list of unique strings. Present de-duplicated strings on standard output You can even re-generate the design by just looking at the code. Code and functional design thus are always in sync - if you follow some simple rules. But about that later. And as a bonus: all the functions making up the process are small - which means easy to understand, too. So much for an initial concrete example. Now it´s time for some theory. Because there is method to this madness ;-) The above has only scratched the surface. Introducing Flow Design Functional design starts with a given function, the Entry Point. Its goal is to describe the behavior of the program when the Entry Point is triggered using a process, not an algorithm. An algorithm consists of logic, a process on the other hand consists just of steps or stages. Each processing step transforms input into output or a side effect. Also it might access resources, e.g. a printer, a database, or just memory. Processing steps thus can rely on state of some sort. This is different from Functional Programming, where functions are supposed to not be stateful and not cause side effects.[1] In its simplest form a process can be written as a bullet point list of steps, e.g. Get data from user Output result to user Transform data Parse data Map result for output Such a compilation of steps - possibly on different levels of abstraction - often is the first artifact of functional design. It can be generated by a team in an initial design brainstorming. Next comes ordering the steps. What should happen first, what next etc.? Get data from user Parse data Transform data Map result for output Output result to user That´s great for a start into functional design. It´s better than starting to code right away on a given function using TDD. Please get me right: TDD is a valuable practice. But it can be unnecessarily hard if the scope of a functionn is too large. But how do you know beforehand without investing some thinking? And how to do this thinking in a systematic fashion? My recommendation: For any given function you´re supposed to implement first do a functional design. Then, once you´re confident you know the processing steps - which are pretty small - refine and code them using TDD. You´ll see that´s much, much easier - and leads to cleaner code right away. For more information on this approach I call “Informed TDD” read my book of the same title. Thinking before coding is smart. And writing down the solution as a bunch of functions possibly is the simplest thing you can do, I´d say. It´s more according to the KISS (Keep It Simple, Stupid) principle than returning constants or other trivial stuff TDD development often is started with. So far so good. A simple ordered list of processing steps will do to start with functional design. As shown in the above example such steps can easily be translated into functions. Moving from design to coding thus is simple. However, such a list does not scale. Processing is not always that simple to be captured in a list. And then the list is just text. Again. Like code. That means the design is lacking visuality. Textual representations need more parsing by your brain than visual representations. Plus they are limited in their “dimensionality”: text just has one dimension, it´s sequential. Alternatives and parallelism are hard to encode in text. In addition the functional design using numbered lists lacks data. It´s not visible what´s the input, output, and state of the processing steps. That´s why functional design should be done using a lightweight visual notation. No tool is necessary to draw such designs. Use pen and paper; a flipchart, a whiteboard, or even a napkin is sufficient. Visualizing processes The building block of the functional design notation is a functional unit. I mostly draw it like this: Something is done, it´s clear what goes in, it´s clear what comes out, and it´s clear what the processing step requires in terms of state or hardware. Whenever input flows into a functional unit it gets processed and output is produced and/or a side effect occurs. Flowing data is the driver of something happening. That´s why I call this approach to functional design Flow Design. It´s about data flow instead of control flow. Control flow like in algorithms is of no concern to functional design. Thinking about control flow simply is too low level. Once you start with control flow you easily get bogged down by tons of details. That´s what you want to avoid during design. Design is supposed to be quick, broad brush, abstract. It should give overview. But what about all the details? As Robert C. Martin rightly said: “Programming is abot detail”. Detail is a matter of code. Once you start coding the processing steps you designed you can worry about all the detail you want. Functional design does not eliminate all the nitty gritty. It just postpones tackling them. To me that´s also an example of the SRP. Function design has the responsibility to come up with a solution to a problem posed by a single function (Entry Point). And later coding has the responsibility to implement the solution down to the last detail (i.e. statement, API-call). TDD unfortunately mixes both responsibilities. It´s just coding - and thereby trying to find detailed implementations (green phase) plus getting the design right (refactoring). To me that´s one reason why TDD has failed to deliver on its promise for many developers. Using functional units as building blocks of functional design processes can be depicted very easily. Here´s the initial process for the example problem: For each processing step draw a functional unit and label it. Choose a verb or an “action phrase” as a label, not a noun. Functional design is about activities, not state or structure. Then make the output of an upstream step the input of a downstream step. Finally think about the data that should flow between the functional units. Write the data above the arrows connecting the functional units in the direction of the data flow. Enclose the data description in brackets. That way you can clearly see if all flows have already been specified. Empty brackets mean “no data is flowing”, but nevertheless a signal is sent. A name like “list” or “strings” in brackets describes the data content. Use lower case labels for that purpose. A name starting with an upper case letter like “String” or “Customer” on the other hand signifies a data type. If you like, you also can combine descriptions with data types by separating them with a colon, e.g. (list:string) or (strings:string[]). But these are just suggestions from my practice with Flow Design. You can do it differently, if you like. Just be sure to be consistent. Flows wired-up in this manner I call one-dimensional (1D). Each functional unit just has one input and/or one output. A functional unit without an output is possible. It´s like a black hole sucking up input without producing any output. Instead it produces side effects. A functional unit without an input, though, does make much sense. When should it start to work? What´s the trigger? That´s why in the above process even the first processing step has an input. If you like, view such 1D-flows as pipelines. Data is flowing through them from left to right. But as you can see, it´s not always the same data. It get´s transformed along its passage: (args) becomes a (list) which is turned into (strings). The Principle of Mutual Oblivion A very characteristic trait of flows put together from function units is: no functional units knows another one. They are all completely independent of each other. Functional units don´t know where their input is coming from (or even when it´s gonna arrive). They just specify a range of values they can process. And they promise a certain behavior upon input arriving. Also they don´t know where their output is going. They just produce it in their own time independent of other functional units. That means at least conceptually all functional units work in parallel. Functional units don´t know their “deployment context”. They now nothing about the overall flow they are place in. They are just consuming input from some upstream, and producing output for some downstream. That makes functional units very easy to test. At least as long as they don´t depend on state or resources. I call this the Principle of Mutual Oblivion (PoMO). Functional units are oblivious of others as well as an overall context/purpose. They are just parts of a whole focused on a single responsibility. How the whole is built, how a larger goal is achieved, is of no concern to the single functional units. By building software in such a manner, functional design interestingly follows nature. Nature´s building blocks for organisms also follow the PoMO. The cells forming your body do not know each other. Take a nerve cell “controlling” a muscle cell for example:[2] The nerve cell does not know anything about muscle cells, let alone the specific muscel cell it is “attached to”. Likewise the muscle cell does not know anything about nerve cells, let a lone a specific nerve cell “attached to” it. Saying “the nerve cell is controlling the muscle cell” thus only makes sense when viewing both from the outside. “Control” is a concept of the whole, not of its parts. Control is created by wiring-up parts in a certain way. Both cells are mutually oblivious. Both just follow a contract. One produces Acetylcholine (ACh) as output, the other consumes ACh as input. Where the ACh is going, where it´s coming from neither cell cares about. Million years of evolution have led to this kind of division of labor. And million years of evolution have produced organism designs (DNA) which lead to the production of these different cell types (and many others) and also to their co-location. The result: the overall behavior of an organism. How and why this happened in nature is a mystery. For our software, though, it´s clear: functional and quality requirements needs to be fulfilled. So we as developers have to become “intelligent designers” of “software cells” which we put together to form a “software organism” which responds in satisfying ways to triggers from it´s environment. My bet is: If nature gets complex organisms working by following the PoMO, who are we to not apply this recipe for success to our much simpler “machines”? So my rule is: Wherever there is functionality to be delivered, because there is a clear Entry Point into software, design the functionality like nature would do it. Build it from mutually oblivious functional units. That´s what Flow Design is about. In that way it´s even universal, I´d say. Its notation can also be applied to biology: Never mind labeling the functional units with nouns. That´s ok in Flow Design. You´ll do that occassionally for functional units on a higher level of abstraction or when their purpose is close to hardware. Getting a cockroach to roam your bedroom takes 1,000,000 nerve cells (neurons). Getting the de-duplication program to do its job just takes 5 “software cells” (functional units). Both, though, follow the same basic principle. Translating functional units into code Moving from functional design to code is no rocket science. In fact it´s straightforward. There are two simple rules: Translate an input port to a function. Translate an output port either to a return statement in that function or to a function pointer visible to that function. The simplest translation of a functional unit is a function. That´s what you saw in the above example. Functions are mutually oblivious. That why Functional Programming likes them so much. It makes them composable. Which is the reason, nature works according to the PoMO. Let´s be clear about one thing: There is no dependency injection in nature. For all of an organism´s complexity no DI container is used. Behavior is the result of smooth cooperation between mutually oblivious building blocks. Functions will often be the adequate translation for the functional units in your designs. But not always. Take for example the case, where a processing step should not always produce an output. Maybe the purpose is to filter input. Here the functional unit consumes words and produces words. But it does not pass along every word flowing in. Some words are swallowed. Think of a spell checker. It probably should not check acronyms for correctness. There are too many of them. Or words with no more than two letters. Such words are called “stop words”. In the above picture the optionality of the output is signified by the astrisk outside the brackets. It means: Any number of (word) data items can flow from the functional unit for each input data item. It might be none or one or even more. This I call a stream of data. Such behavior cannot be translated into a function where output is generated with return. Because a function always needs to return a value. So the output port is translated into a function pointer or continuation which gets passed to the subroutine when called:[3]void filter_stop_words( string word, Action<string> onNoStopWord) { if (...check if not a stop word...) onNoStopWord(word); } If you want to be nitpicky you might call such a function pointer parameter an injection. And technically you´re right. Conceptually, though, it´s not an injection. Because the subroutine is not functionally dependent on the continuation. Firstly continuations are procedures, i.e. subroutines without a return type. Remember: Flow Design is about unidirectional data flow. Secondly the name of the formal parameter is chosen in a way as to not assume anything about downstream processing steps. onNoStopWord describes a situation (or event) within the functional unit only. Translating output ports into function pointers helps keeping functional units mutually oblivious in cases where output is optional or produced asynchronically. Either pass the function pointer to the function upon call. Or make it global by putting it on the encompassing class. Then it´s called an event. In C# that´s even an explicit feature.class Filter { public void filter_stop_words( string word) { if (...check if not a stop word...) onNoStopWord(word); } public event Action<string> onNoStopWord; } When to use a continuation and when to use an event dependens on how a functional unit is used in flows and how it´s packed together with others into classes. You´ll see examples further down the Flow Design road. Another example of 1D functional design Let´s see Flow Design once more in action using the visual notation. How about the famous word wrap kata? Robert C. Martin has posted a much cited solution including an extensive reasoning behind his TDD approach. So maybe you want to compare it to Flow Design. The function signature given is:string WordWrap(string text, int maxLineLength) {...} That´s not an Entry Point since we don´t see an application with an environment and users. Nevertheless it´s a function which is supposed to provide a certain functionality. The text passed in has to be reformatted. The input is a single line of arbitrary length consisting of words separated by spaces. The output should consist of one or more lines of a maximum length specified. If a word is longer than a the maximum line length it can be split in multiple parts each fitting in a line. Flow Design Let´s start by brainstorming the process to accomplish the feat of reformatting the text. What´s needed? Words need to be assembled into lines Words need to be extracted from the input text The resulting lines need to be assembled into the output text Words too long to fit in a line need to be split Does sound about right? I guess so. And it shows a kind of priority. Long words are a special case. So maybe there is a hint for an incremental design here. First let´s tackle “average words” (words not longer than a line). Here´s the Flow Design for this increment: The the first three bullet points turned into functional units with explicit data added. As the signature requires a text is transformed into another text. See the input of the first functional unit and the output of the last functional unit. In between no text flows, but words and lines. That´s good to see because thereby the domain is clearly represented in the design. The requirements are talking about words and lines and here they are. But note the asterisk! It´s not outside the brackets but inside. That means it´s not a stream of words or lines, but lists or sequences. For each text a sequence of words is output. For each sequence of words a sequence of lines is produced. The asterisk is used to abstract from the concrete implementation. Like with streams. Whether the list of words gets implemented as an array or an IEnumerable is not important during design. It´s an implementation detail. Does any processing step require further refinement? I don´t think so. They all look pretty “atomic” to me. And if not… I can always backtrack and refine a process step using functional design later once I´ve gained more insight into a sub-problem. Implementation The implementation is straightforward as you can imagine. The processing steps can all be translated into functions. Each can be tested easily and separately. Each has a focused responsibility. And the process flow becomes just a sequence of function calls: Easy to understand. It clearly states how word wrapping works - on a high level of abstraction. And it´s easy to evolve as you´ll see. Flow Design - Increment 2 So far only texts consisting of “average words” are wrapped correctly. Words not fitting in a line will result in lines too long. Wrapping long words is a feature of the requested functionality. Whether it´s there or not makes a difference to the user. To quickly get feedback I decided to first implement a solution without this feature. But now it´s time to add it to deliver the full scope. Fortunately Flow Design automatically leads to code following the Open Closed Principle (OCP). It´s easy to extend it - instead of changing well tested code. How´s that possible? Flow Design allows for extension of functionality by inserting functional units into the flow. That way existing functional units need not be changed. The data flow arrow between functional units is a natural extension point. No need to resort to the Strategy Pattern. No need to think ahead where extions might need to be made in the future. I just “phase in” the remaining processing step: Since neither Extract words nor Reformat know of their environment neither needs to be touched due to the “detour”. The new processing step accepts the output of the existing upstream step and produces data compatible with the existing downstream step. Implementation - Increment 2 A trivial implementation checking the assumption if this works does not do anything to split long words. The input is just passed on: Note how clean WordWrap() stays. The solution is easy to understand. A developer looking at this code sometime in the future, when a new feature needs to be build in, quickly sees how long words are dealt with. Compare this to Robert C. Martin´s solution:[4] How does this solution handle long words? Long words are not even part of the domain language present in the code. At least I need considerable time to understand the approach. Admittedly the Flow Design solution with the full implementation of long word splitting is longer than Robert C. Martin´s. At least it seems. Because his solution does not cover all the “word wrap situations” the Flow Design solution handles. Some lines would need to be added to be on par, I guess. But even then… Is a difference in LOC that important as long as it´s in the same ball park? I value understandability and openness for extension higher than saving on the last line of code. Simplicity is not just less code, it´s also clarity in design. But don´t take my word for it. Try Flow Design on larger problems and compare for yourself. What´s the easier, more straightforward way to clean code? And keep in mind: You ain´t seen all yet ;-) There´s more to Flow Design than described in this chapter. In closing I hope I was able to give you a impression of functional design that makes you hungry for more. To me it´s an inevitable step in software development. Jumping from requirements to code does not scale. And it leads to dirty code all to quickly. Some thought should be invested first. Where there is a clear Entry Point visible, it´s functionality should be designed using data flows. Because with data flows abstraction is possible. For more background on why that´s necessary read my blog article here. For now let me point out to you - if you haven´t already noticed - that Flow Design is a general purpose declarative language. It´s “programming by intention” (Shalloway et al.). Just write down how you think the solution should work on a high level of abstraction. This breaks down a large problem in smaller problems. And by following the PoMO the solutions to those smaller problems are independent of each other. So they are easy to test. Or you could even think about getting them implemented in parallel by different team members. Flow Design not only increases evolvability, but also helps becoming more productive. All team members can participate in functional design. This goes beyon collective code ownership. We´re talking collective design/architecture ownership. Because with Flow Design there is a common visual language to talk about functional design - which is the foundation for all other design activities. PS: If you like what you read, consider getting my ebook “The Incremental Architekt´s Napkin”. It´s where I compile all the articles in this series for easier reading. I like the strictness of Function Programming - but I also find it quite hard to live by. And it certainly is not what millions of programmers are used to. Also to me it seems, the real world is full of state and side effects. So why give them such a bad image? That´s why functional design takes a more pragmatic approach. State and side effects are ok for processing steps - but be sure to follow the SRP. Don´t put too much of it into a single processing step. ? Image taken from www.physioweb.org ? My code samples are written in C#. C# sports typed function pointers called delegates. Action is such a function pointer type matching functions with signature void someName(T t). Other languages provide similar ways to work with functions as first class citizens - even Java now in version 8. I trust you find a way to map this detail of my translation to your favorite programming language. I know it works for Java, C++, Ruby, JavaScript, Python, Go. And if you´re using a Functional Programming language it´s of course a no brainer. ? Taken from his blog post “The Craftsman 62, The Dark Path”. ?

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An Introduction to Meteor

- by Stephen.Walther

The goal of this blog post is to give you a brief introduction to Meteor which is a framework for building Single Page Apps. In this blog entry, I provide a walkthrough of building a simple Movie database app. What is special about Meteor? Meteor has two jaw-dropping features: Live HTML – If you make any changes to the HTML, CSS, JavaScript, or data on the server then every client shows the changes automatically without a browser refresh. For example, if you change the background color of a page to yellow then every open browser will show the new yellow background color without a refresh. Or, if you add a new movie to a collection of movies, then every open browser will display the new movie automatically. With Live HTML, users no longer need a refresh button. Changes to an application happen everywhere automatically without any effort. The Meteor framework handles all of the messy details of keeping all of the clients in sync with the server for you. Latency Compensation – When you modify data on the client, these modifications appear as if they happened on the server without any delay. For example, if you create a new movie then the movie appears instantly. However, that is all an illusion. In the background, Meteor updates the database with the new movie. If, for whatever reason, the movie cannot be added to the database then Meteor removes the movie from the client automatically. Latency compensation is extremely important for creating a responsive web application. You want the user to be able to make instant modifications in the browser and the framework to handle the details of updating the database without slowing down the user. Installing Meteor Meteor is licensed under the open-source MIT license and you can start building production apps with the framework right now. Be warned that Meteor is still in the “early preview” stage. It has not reached a 1.0 release. According to the Meteor FAQ, Meteor will reach version 1.0 in “More than a month, less than a year.” Don’t be scared away by that. You should be aware that, unlike most open source projects, Meteor has financial backing. The Meteor project received an $11.2 million round of financing from Andreessen Horowitz. So, it would be a good bet that this project will reach the 1.0 mark. And, if it doesn’t, the framework as it exists right now is still very powerful. Meteor runs on top of Node.js. You write Meteor apps by writing JavaScript which runs both on the client and on the server. You can build Meteor apps on Windows, Mac, or Linux (Although the support for Windows is still officially unofficial). If you want to install Meteor on Windows then download the MSI from the following URL: http://win.meteor.com/ If you want to install Meteor on Mac/Linux then run the following CURL command from your terminal: curl https://install.meteor.com | /bin/sh Meteor will install all of its dependencies automatically including Node.js. However, I recommend that you install Node.js before installing Meteor by installing Node.js from the following address: http://nodejs.org/ If you let Meteor install Node.js then Meteor won’t install NPM which is the standard package manager for Node.js. If you install Node.js and then you install Meteor then you get NPM automatically. Creating a New Meteor App To get a sense of how Meteor works, I am going to walk through the steps required to create a simple Movie database app. Our app will display a list of movies and contain a form for creating a new movie. The first thing that we need to do is create our new Meteor app. Open a command prompt/terminal window and execute the following command: Meteor create MovieApp After you execute this command, you should see something like the following: Follow the instructions: execute cd MovieApp to change to your MovieApp directory, and run the meteor command. Executing the meteor command starts Meteor on port 3000. Open up your favorite web browser and navigate to http://localhost:3000 and you should see the default Meteor Hello World page: Open up your favorite development environment to see what the Meteor app looks like. Open the MovieApp folder which we just created. Here’s what the MovieApp looks like in Visual Studio 2012: Notice that our MovieApp contains three files named MovieApp.css, MovieApp.html, and MovieApp.js. In other words, it contains a Cascading Style Sheet file, an HTML file, and a JavaScript file. Just for fun, let’s see how the Live HTML feature works. Open up multiple browsers and point each browser at http://localhost:3000. Now, open the MovieApp.html page and modify the text “Hello World!” to “Hello Cruel World!” and save the change. The text in all of the browsers should update automatically without a browser refresh. Pretty amazing, right? Controlling Where JavaScript Executes You write a Meteor app using JavaScript. Some of the JavaScript executes on the client (the browser) and some of the JavaScript executes on the server and some of the JavaScript executes in both places. For a super simple app, you can use the Meteor.isServer and Meteor.isClient properties to control where your JavaScript code executes. For example, the following JavaScript contains a section of code which executes on the server and a section of code which executes in the browser: if (Meteor.isClient) { console.log("Hello Browser!"); } if (Meteor.isServer) { console.log("Hello Server!"); } console.log("Hello Browser and Server!"); When you run the app, the message “Hello Browser!” is written to the browser JavaScript console. The message “Hello Server!” is written to the command/terminal window where you ran Meteor. Finally, the message “Hello Browser and Server!” is execute on both the browser and server and the message appears in both places. For simple apps, using Meteor.isClient and Meteor.isServer to control where JavaScript executes is fine. For more complex apps, you should create separate folders for your server and client code. Here are the folders which you can use in a Meteor app: · client – This folder contains any JavaScript which executes only on the client. · server – This folder contains any JavaScript which executes only on the server. · common – This folder contains any JavaScript code which executes on both the client and server. · lib – This folder contains any JavaScript files which you want to execute before any other JavaScript files. · public – This folder contains static application assets such as images. For the Movie App, we need the client, server, and common folders. Delete the existing MovieApp.js, MovieApp.html, and MovieApp.css files. We will create new files in the right locations later in this walkthrough. Combining HTML, CSS, and JavaScript Files Meteor combines all of your JavaScript files, and all of your Cascading Style Sheet files, and all of your HTML files automatically. If you want to create one humongous JavaScript file which contains all of the code for your app then that is your business. However, if you want to build a more maintainable application, then you should break your JavaScript files into many separate JavaScript files and let Meteor combine them for you. Meteor also combines all of your HTML files into a single file. HTML files are allowed to have the following top-level elements: <head> — All <head> files are combined into a single <head> and served with the initial page load. <body> — All <body> files are combined into a single <body> and served with the initial page load. <template> — All <template> files are compiled into JavaScript templates. Because you are creating a single page app, a Meteor app typically will contain a single HTML file for the <head> and <body> content. However, a Meteor app typically will contain several template files. In other words, all of the interesting stuff happens within the <template> files. Displaying a List of Movies Let me start building the Movie App by displaying a list of movies. In order to display a list of movies, we need to create the following four files: · client\movies.html – Contains the HTML for the <head> and <body> of the page for the Movie app. · client\moviesTemplate.html – Contains the HTML template for displaying the list of movies. · client\movies.js – Contains the JavaScript for supplying data to the moviesTemplate. · server\movies.js – Contains the JavaScript for seeding the database with movies. After you create these files, your folder structure should looks like this: Here’s what the client\movies.html file looks like: <head> <title>My Movie App</title> </head> <body> <h1>Movies</h1> {{> moviesTemplate }} </body> Notice that it contains <head> and <body> top-level elements. The <body> element includes the moviesTemplate with the syntax {{> moviesTemplate }}. The moviesTemplate is defined in the client/moviesTemplate.html file: <template name="moviesTemplate"> <ul> {{#each movies}} <li> {{title}} </li> {{/each}} </ul> </template> By default, Meteor uses the Handlebars templating library. In the moviesTemplate above, Handlebars is used to loop through each of the movies using {{#each}}…{{/each}} and display the title for each movie using {{title}}. The client\movies.js JavaScript file is used to bind the moviesTemplate to the Movies collection on the client. Here’s what this JavaScript file looks like: // Declare client Movies collection Movies = new Meteor.Collection("movies"); // Bind moviesTemplate to Movies collection Template.moviesTemplate.movies = function () { return Movies.find(); }; The Movies collection is a client-side proxy for the server-side Movies database collection. Whenever you want to interact with the collection of Movies stored in the database, you use the Movies collection instead of communicating back to the server. The moviesTemplate is bound to the Movies collection by assigning a function to the Template.moviesTemplate.movies property. The function simply returns all of the movies from the Movies collection. The final file which we need is the server-side server\movies.js file: // Declare server Movies collection Movies = new Meteor.Collection("movies"); // Seed the movie database with a few movies Meteor.startup(function () { if (Movies.find().count() == 0) { Movies.insert({ title: "Star Wars", director: "Lucas" }); Movies.insert({ title: "Memento", director: "Nolan" }); Movies.insert({ title: "King Kong", director: "Jackson" }); } }); The server\movies.js file does two things. First, it declares the server-side Meteor Movies collection. When you declare a server-side Meteor collection, a collection is created in the MongoDB database associated with your Meteor app automatically (Meteor uses MongoDB as its database automatically). Second, the server\movies.js file seeds the Movies collection (MongoDB collection) with three movies. Seeding the database gives us some movies to look at when we open the Movies app in a browser. Creating New Movies Let me modify the Movies Database App so that we can add new movies to the database of movies. First, I need to create a new template file – named client\movieForm.html – which contains an HTML form for creating a new movie: <template name="movieForm"> <fieldset> <legend>Add New Movie</legend> <form> <div> <label> Title: <input id="title" /> </label> </div> <div> <label> Director: <input id="director" /> </label> </div> <div> <input type="submit" value="Add Movie" /> </div> </form> </fieldset> </template> In order for the new form to show up, I need to modify the client\movies.html file to include the movieForm.html template. Notice that I added {{> movieForm }} to the client\movies.html file: <head> <title>My Movie App</title> </head> <body> <h1>Movies</h1> {{> moviesTemplate }} {{> movieForm }} </body> After I make these modifications, our Movie app will display the form: The next step is to handle the submit event for the movie form. Below, I’ve modified the client\movies.js file so that it contains a handler for the submit event raised when you submit the form contained in the movieForm.html template: // Declare client Movies collection Movies = new Meteor.Collection("movies"); // Bind moviesTemplate to Movies collection Template.moviesTemplate.movies = function () { return Movies.find(); }; // Handle movieForm events Template.movieForm.events = { 'submit': function (e, tmpl) { // Don't postback e.preventDefault(); // create the new movie var newMovie = { title: tmpl.find("#title").value, director: tmpl.find("#director").value }; // add the movie to the db Movies.insert(newMovie); } }; The Template.movieForm.events property contains an event map which maps event names to handlers. In this case, I am mapping the form submit event to an anonymous function which handles the event. In the event handler, I am first preventing a postback by calling e.preventDefault(). This is a single page app, no postbacks are allowed! Next, I am grabbing the new movie from the HTML form. I’m taking advantage of the template find() method to retrieve the form field values. Finally, I am calling Movies.insert() to insert the new movie into the Movies collection. Here, I am explicitly inserting the new movie into the client-side Movies collection. Meteor inserts the new movie into the server-side Movies collection behind the scenes. When Meteor inserts the movie into the server-side collection, the new movie is added to the MongoDB database associated with the Movies app automatically. If server-side insertion fails for whatever reasons – for example, your internet connection is lost – then Meteor will remove the movie from the client-side Movies collection automatically. In other words, Meteor takes care of keeping the client Movies collection and the server Movies collection in sync. If you open multiple browsers, and add movies, then you should notice that all of the movies appear on all of the open browser automatically. You don’t need to refresh individual browsers to update the client-side Movies collection. Meteor keeps everything synchronized between the browsers and server for you. Removing the Insecure Module To make it easier to develop and debug a new Meteor app, by default, you can modify the database directly from the client. For example, you can delete all of the data in the database by opening up your browser console window and executing multiple Movies.remove() commands. Obviously, enabling anyone to modify your database from the browser is not a good idea in a production application. Before you make a Meteor app public, you should first run the meteor remove insecure command from a command/terminal window: Running meteor remove insecure removes the insecure package from the Movie app. Unfortunately, it also breaks our Movie app. We’ll get an “Access denied” error in our browser console whenever we try to insert a new movie. No worries. I’ll fix this issue in the next section. Creating Meteor Methods By taking advantage of Meteor Methods, you can create methods which can be invoked on both the client and the server. By taking advantage of Meteor Methods you can: 1. Perform form validation on both the client and the server. For example, even if an evil hacker bypasses your client code, you can still prevent the hacker from submitting an invalid value for a form field by enforcing validation on the server. 2. Simulate database operations on the client but actually perform the operations on the server. Let me show you how we can modify our Movie app so it uses Meteor Methods to insert a new movie. First, we need to create a new file named common\methods.js which contains the definition of our Meteor Methods: Meteor.methods({ addMovie: function (newMovie) { // Perform form validation if (newMovie.title == "") { throw new Meteor.Error(413, "Missing title!"); } if (newMovie.director == "") { throw new Meteor.Error(413, "Missing director!"); } // Insert movie (simulate on client, do it on server) return Movies.insert(newMovie); } }); The addMovie() method is called from both the client and the server. This method does two things. First, it performs some basic validation. If you don’t enter a title or you don’t enter a director then an error is thrown. Second, the addMovie() method inserts the new movie into the Movies collection. When called on the client, inserting the new movie into the Movies collection just updates the collection. When called on the server, inserting the new movie into the Movies collection causes the database (MongoDB) to be updated with the new movie. You must add the common\methods.js file to the common folder so it will get executed on both the client and the server. Our folder structure now looks like this: We actually call the addMovie() method within our client code in the client\movies.js file. Here’s what the updated file looks like: // Declare client Movies collection Movies = new Meteor.Collection("movies"); // Bind moviesTemplate to Movies collection Template.moviesTemplate.movies = function () { return Movies.find(); }; // Handle movieForm events Template.movieForm.events = { 'submit': function (e, tmpl) { // Don't postback e.preventDefault(); // create the new movie var newMovie = { title: tmpl.find("#title").value, director: tmpl.find("#director").value }; // add the movie to the db Meteor.call( "addMovie", newMovie, function (err, result) { if (err) { alert("Could not add movie " + err.reason); } } ); } }; The addMovie() method is called – on both the client and the server – by calling the Meteor.call() method. This method accepts the following parameters: · The string name of the method to call. · The data to pass to the method (You can actually pass multiple params for the data if you like). · A callback function to invoke after the method completes. In the JavaScript code above, the addMovie() method is called with the new movie retrieved from the HTML form. The callback checks for an error. If there is an error then the error reason is displayed in an alert (please don’t use alerts for validation errors in a production app because they are ugly!). Summary The goal of this blog post was to provide you with a brief walk through of a simple Meteor app. I showed you how you can create a simple Movie Database app which enables you to display a list of movies and create new movies. I also explained why it is important to remove the Meteor insecure package from a production app. I showed you how to use Meteor Methods to insert data into the database instead of doing it directly from the client. I’m very impressed with the Meteor framework. The support for Live HTML and Latency Compensation are required features for many real world Single Page Apps but implementing these features by hand is not easy. Meteor makes it easy.

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Why should you choose Oracle WebLogic 12c instead of JBoss EAP 6?

- by Ricardo Ferreira

In this post, I will cover some technical differences between Oracle WebLogic 12c and JBoss EAP 6, which was released a couple days ago from Red Hat. This article claims to help you in the evaluation of key points that you should consider when choosing for an Java EE application server. In the following sections, I will present to you some important aspects that most customers ask us when they are seriously evaluating for an middleware infrastructure, specially if you are considering JBoss for some reason. I would suggest that you keep the following question in mind while you are reading the points: "Why should I choose JBoss instead of WebLogic?" 1) Multi Datacenter Deployment and Clustering - D/R ("Disaster & Recovery") architecture support is embedded on the WebLogic Server 12c product. JBoss EAP 6 on the other hand has no direct D/R support included, Red Hat relies on third-part tools with higher prices. When you consider a middleware solution to host your business critical application, you should worry with every architectural aspect that are related with the solution. Fail-over support is one little aspect of a truly reliable solution. If you do not worry about D/R, your solution will not be reliable. Having said that, with Red Hat and JBoss EAP 6, you have this extra cost that will increase considerably the total cost of ownership of the solution. As we commonly hear from analysts, open-source are not so cheaper when you start seeing the big picture. - WebLogic Server 12c supports advanced LAN clustering, detection of death servers and have a common alert framework. JBoss EAP 6 on the other hand has limited LAN clustering support with no server death detection. They do not generate any alerts when servers goes down (only if you buy JBoss ON which is a separated technology, but until now does not support JBoss EAP 6) and manual intervention are required when servers goes down. In most cases, admin people must rely on "kill -9", "tail -f someFile.log" and "ps ax | grep java" commands to manage failures and clustering anomalies. - WebLogic Server 12c supports the concept of Node Manager, which is a separated process that runs on the physical | virtual servers that allows extend the administration of the cluster to WebLogic managed servers that are often distributed across multiple machines and geographic locations. JBoss EAP 6 on the other hand has no equivalent technology. Whole server instances must be managed individually. - WebLogic Server 12c Node Manager supports Coherence to boost performance when managing servers. JBoss EAP 6 on the other hand has no similar technology. There is no way to coordinate JBoss and infiniband instances provided by JBoss using high throughput and low latency protocols like InfiniBand. The Node Manager feature also allows another very important feature that JBoss EAP lacks: secure the administration. When using WebLogic Node Manager, all the administration tasks are sent to the managed servers in a secure tunel protected by a certificate, which means that the transport layer that separates the WebLogic administration console from the managed servers are secured by SSL. - WebLogic Server 12c are now integrated with OTD ("Oracle Traffic Director") which is a web server technology derived from the former Sun iPlanet Web Server. This software complements the web server support offered by OHS ("Oracle HTTP Server"). Using OTD, WebLogic instances are load-balanced by a high powerful software that knows how to handle SDP ("Socket Direct Protocol") over InfiniBand, which boost performance when used with engineered systems technologies like Oracle Exalogic Elastic Cloud. JBoss EAP 6 on the other hand only offers support to Apache Web Server with custom modules created to deal with JBoss clusters, but only across standard TCP/IP networks. 2) Application and Runtime Diagnostics - WebLogic Server 12c have diagnostics capabilities embedded on the server called WLDF ("WebLogic Diagnostic Framework") so there is no need to rely on third-part tools. JBoss EAP 6 on the other hand has no diagnostics capabilities. Their only diagnostics tool is the log generated by the application server. Admin people are encouraged to analyse thousands of log lines to find out what is going on. - WebLogic Server 12c complement WLDF with JRockit MC ("Mission Control"), which provides to administrators and developers a complete insight about the JVM performance, behavior and possible bottlenecks. WebLogic Server 12c also have an classloader analysis tool embedded, and even a log analyzer tool that enables administrators and developers to view logs of multiple servers at the same time. JBoss EAP 6 on the other hand relies on third-part tools to do something similar. Again, only log searching are offered to find out whats going on. - WebLogic Server 12c offers end-to-end traceability and monitoring available through Oracle EM ("Enterprise Manager"), including monitoring of business transactions that flows through web servers, ESBs, application servers and database servers, all of this with high deep JVM analysis and diagnostics. JBoss EAP 6 on the other hand, even using JBoss ON ("Operations Network"), which is a separated technology, does not support those features. Red Hat relies on third-part tools to provide direct Oracle database traceability across JVMs. One of those tools are Oracle EM for non-Oracle middleware that manage JBoss, Tomcat, Websphere and IIS transparently. - WebLogic Server 12c with their JRockit support offers a tool called JRockit Flight Recorder, which can give developers a complete visibility of a certain period of application production monitoring with zero extra overhead. This automatic recording allows you to deep analyse threads latency, memory leaks, thread contention, resource utilization, stack overflow damages and GC ("Garbage Collection") cycles, to observe in real time stop-the-world phenomenons, generational, reference count and parallel collects and mutator threads analysis. JBoss EAP 6 don't even dream to support something similar, even because they don't have their own JVM. 3) Application Server Administration - WebLogic Server 12c offers a complete administration console complemented with scripting and macro-like recording capabilities. A single WebLogic console can managed up to hundreds of WebLogic servers belonging to the same domain. JBoss EAP 6 on the other hand has a limited console and provides a XML centric administration. JBoss, after ten years, started the development of a rudimentary centralized administration that still leave a lot of administration tasks aside, so admin people and developers must touch scripts and XML configuration files for most advanced and even simple administration tasks. This lead applications to error prone and risky deployments. Even using JBoss ON, JBoss EAP are not able to offer decent administration features for admin people which must be high skilled in JBoss internal architecture and its managing capabilities. - Oracle EM is available to manage multiple domains, databases, application servers, operating systems and virtualization, with a complete end-to-end visibility. JBoss ON does not provide management capabilities across the complete architecture, only basic monitoring. Even deployment must be done aside JBoss ON which does no integrate well with others softwares than JBoss. Until now, JBoss ON does not supports JBoss EAP 6, so even their minimal support for JBoss are not available for JBoss EAP 6 leaving customers uncovered and subject to high skilled JBoss admin people. - WebLogic Server 12c has the same administration model whatever is the topology selected by the customer. JBoss EAP 6 on the other hand differentiates between two operational models: standalone-mode and domain-mode, that are not consistent with each other. Depending on the mode used, the administration skill is different. - WebLogic Server 12c has no point-of-failures processes, and it does not need to define any specialized server. Domain model in WebLogic is available for years (at least ten years or more) and is production proven. JBoss EAP 6 on the other hand needs special processes to garantee JBoss integrity, the PC ("Process-Controller") and the HC ("Host-Controller"). Different from WebLogic, the domain model in JBoss is quite new (one year at tops) of maturity, and need to mature considerably until start doing things like WebLogic domain model does. - WebLogic Server 12c supports parallel deployment model which enables some artifacts being deployed at the same time. JBoss EAP 6 on the other hand does not have any similar feature. Every deployment are done atomically in the containers. This means that if you have a huge EAR (an EAR of 120 MB of size for instance) and deploy onto JBoss EAP 6, this EAR will take some minutes in order to starting accept thread requests. The same EAR deployed onto WebLogic Server 12c will reduce the deployment time at least in 2X compared to JBoss. 4) Support and Upgrades - WebLogic Server 12c has patch management available. JBoss EAP 6 on the other hand has no patch management available, each JBoss EAP instance should be patched manually. To achieve such feature, you need to buy a separated technology called JBoss ON ("Operations Network") that manage this type of stuff. But until now, JBoss ON does not support JBoss EAP 6 so, in practice, JBoss EAP 6 does not have this feature. - WebLogic Server 12c supports previuous WebLogic domains without any reconfiguration since its kernel is robust and mature since its creation in 1995. JBoss EAP 6 on the other hand has a proven lack of supportability between JBoss AS 4, 5, 6 and 7. Different kernels and messaging engines were implemented in JBoss stack in the last five years reveling their incapacity to create a well architected and proven middleware technology. - WebLogic Server 12c has patch prescription based on customer configuration. JBoss EAP 6 on the other hand has no such capability. People need to create ticket supports and have their installations revised by Red Hat support guys to gain some patch prescription from them. - Oracle WebLogic Server independent of the version has 8 years of support of new patches and has lifetime release of existing patches beyond that. JBoss EAP 6 on the other hand provides patches for a specific application server version up to 5 years after the release date. JBoss EAP 4 and previous versions had only 4 years. A good question that Red Hat will argue to answer is: "what happens when you find issues after year 5"? 5) RAC ("Real Application Clusters") Support - WebLogic Server 12c ships with a specific JDBC driver to leverage Oracle RAC clustering capabilities (Fast-Application-Notification, Transaction Affinity, Fast-Connection-Failover, etc). Oracle JDBC thin driver are also available. JBoss EAP 6 on the other hand ships only the standard Oracle JDBC thin driver. Load balancing with Oracle RAC are not supported. Manual intervention in case of planned or unplanned RAC downtime are necessary. In JBoss EAP 6, situation does not reestablish automatically after downtime. - WebLogic Server 12c has a feature called Active GridLink for Oracle RAC which provides up to 3X performance on OLTP applications. This seamless integration between WebLogic and Oracle database enable more value added to critical business applications leveraging their investments in Oracle database technology and Oracle middleware. JBoss EAP 6 on the other hand has no performance gains at all, even when admin people implement some kind of connection-pooling tuning. - WebLogic Server 12c also supports transaction and web session affinity to the Oracle RAC, which provides aditional gains of performance. This is particularly interesting if you are creating a reliable solution that are distributed not only in an LAN cluster, but into a different data center. JBoss EAP 6 on the other hand has no such support. 6) Standards and Technology Support - WebLogic Server 12c is fully Java EE 6 compatible and production ready since december of 2011. JBoss EAP 6 on the other hand became fully compatible with Java EE 6 only in the community version after three months, and production ready only in a few days considering that this article was written in June of 2012. Red Hat says that they are the masters of innovation and technology proliferation, but compared with Oracle and even other proprietary vendors like IBM, they historically speaking are lazy to deliver the most newest technologies and standards adherence. - Oracle is the steward of Java, driving innovation into the platform from commercial and open-source vendors. Red Hat on the other hand does not have its own JVM and relies on third-part JVMs to complete their application server offer. 95% of Red Hat customers are using Oracle HotSpot as JVM, which means that without Oracle involvement, their support are limited exclusively to the application server layer and we all know that most problems are happens in the JVM layer. - WebLogic Server 12c supports natively JDK 7, which empower developers to explore the maximum of the Java platform productivity when writing code. This feature differentiate WebLogic from others application servers (except GlassFish that are also managed by Oracle) because the usage of JDK 7 introduce such remarkable productivity features like the "try-with-resources" enhancement, catching multiple exceptions with one try block, Strings in the switch statements, JVM improvements in terms of JDBC, I/O, networking, security, concurrency and of course, the most important feature of Java 7: native support for multiple non-Java languages. More features regarding JDK 7 can be found here. JBoss EAP 6 on the other hand does not support JDK 7 officially, they comment in their community version that "Java SE 7 can be used with JBoss 7" which does not gives you any guarantees of enterprise support for JDK 7. - Oracle WebLogic Server 12c supports integration with Spring framework allowing Spring applications to use WebLogic special transaction manager, exposing bean interfaces to WebLogic MBeans to take advantage of all WebLogic monitoring and administration advantages. JBoss EAP 6 on the other hand has no special integration with Spring. In fact, Red Hat offers a suspicious package called "JBoss Web Platform" that in theory supports Spring, but in practice this package does not offers any special integration. It is just a facility for Red Hat customers to have support from both JBoss and Spring technology using the same customer support. 7) Lightweight Development - Oracle WebLogic Server 12c and Oracle GlassFish are completely integrated and can share applications without any modifications. Starting with the 12c version, WebLogic now understands natively GlassFish deployment descriptors and specific configurations in order to offer you a truly and reliable migration path from a community Java EE application server to a enterprise middleware product like WebLogic. JBoss EAP 6 on the other hand has no support to natively reuse an existing (or still in development) application from JBoss AS community server. Users of JBoss suffer of critical issues during deployment time that includes: changing the libraries and dependencies of the application, patching the DTD or XSD deployment descriptors, refactoring of the application layers due classloading issues and anomalies, rebuilding of persistence, business and web layers due issues with "usage of the certified version of an certain dependency" or "frameworks that Red Hat potentially does not recommend" etc. If you have the culture or enterprise IT directive of developing Java EE applications using community middleware to in a certain future, transition to enterprise (supported by a vendor) middleware, Oracle WebLogic plus Oracle GlassFish offers you a more sustainable solution. - WebLogic Server 12c has a very light ZIP distribution (less than 165 MB). JBoss EAP 6 ZIP size is around 130 MB, together with JBoss ON you have more 100 MB resulting in a higher download footprint. This is particularly interesting if you plan to use automated setup of application server instances (for example, to rapidly setup a development or staging environment) using Maven or Hudson. - WebLogic Server 12c has a complete integration with Maven allowing developers to setup WebLogic domains with few commands. Tasks like downloading WebLogic, installation, domain creation, data sources deployment are completely integrated. JBoss EAP 6 on the other hand has a limited offer integration with those tools. - WebLogic Server 12c has a startup mode called WLX that turns-off EJB, JMS and JCA containers leaving enabled only the web container with Java EE 6 web profile. JBoss EAP 6 on the other hand has no such feature, you need to disable manually the containers that you do not want to use. - WebLogic Server 12c supports fastswap, which enables you to change classes without redeployment. This is particularly interesting if you are developing patches for the application that is already deployed and you do not want to redeploy the entire application. This is the same behavior that most application servers offers to JSP pages, but with WebLogic Server 12c, you have the same feature for Java classes in general. JBoss EAP 6 on the other hand has no such support. Even JBoss EAP 5 does not support this until now. 8) JMS and Messaging - WebLogic Server 12c has a proven and high scalable JMS implementation since its initial release in 1995. JBoss EAP 6 on the other hand has a still immature technology called HornetQ, which was introduced in JBoss EAP 5 replacing everything that was implemented in the previous versions. Red Hat loves to introduce new technologies across JBoss versions, playing around with customers and their investments. And when they are asked about why they have changed the implementation and caused such a mess, their answer is always: "the previous implementation was inadequate and not aligned with the community strategy so we are creating a new a improved one". This Red Hat practice leads to uncomfortable investments that in a near future (sometimes less than a year) will be affected in someway. - WebLogic Server 12c has troubleshooting and monitoring features included on the WebLogic console and WLDF. JBoss EAP 6 on the other hand has no direct monitoring on the console, activity is reflected only on the logs, no debug logs available in case of JMS issues. - WebLogic Server 12c has extremely good performance and scalability. JBoss EAP 6 on the other hand has a JMS storage mechanism relying on Oracle database or MySQL. This means that if an issue in production happens and Red Hat affirms that an performance issue is happening due to database problems, they will not support you on the performance issue. They will orient you to call Oracle instead. - WebLogic Server 12c supports messaging enterprise features like SAF ("Store and Forward"), Distributed Queues/Topics and Foreign JMS providers support that leverage JMS implementations without compromise developer code making things completely transparent. JBoss EAP 6 on the other hand do not even dream to support such features. 9) Caching and Grid - Coherence, which is the leading and most mature data grid technology from Oracle, is available since early 2000 and was integrated with WebLogic in 2009. Coherence and WebLogic clusters can be both managed from WebLogic administrative console. Even Node Manager supports Coherence. JBoss on the other hand discontinued JBoss Cache, which was their caching implementation just like they did with the messaging implementation (JBossMQ) which was a issue for long term customers. JBoss EAP 6 ships InfiniSpan version 1.0 which is immature and lack a proven record of successful cases and reliability. - WebLogic Server 12c has a feature called ActiveCache which uses Coherence to, without any code changes, replicate HTTP sessions from both WebLogic and other application servers like JBoss, Tomcat, Websphere, GlassFish and even Microsoft IIS. JBoss EAP 6 on the other hand does have such support and even when they do in the future, they probably will support only their own application server. - Coherence can be used to manage both L1 and L2 cache levels, providing support to Oracle TopLink and others JPA compliant implementations, even Hibernate. JBoss EAP 6 and Infinispan on the other hand supports only Hibernate. And most important of all: Infinispan does not have any successful case of L1 or L2 caching level support using Hibernate, which lead us to reflect about its viability. 10) Performance - WebLogic Server 12c is certified with Oracle Exalogic Elastic Cloud and can run unchanged applications at this engineered system. This approach can benefit customers from Exalogic optimization's of both kernel and JVM layers to boost performance in terms of 10X for web, OLTP, JMS and grid applications. JBoss EAP 6 on the other hand has no investment on engineered systems: customers do not have the choice to deploy on a Java ultra fast system if their project becomes relevant and performance issues are detected. - WebLogic Server 12c maintains a performance gain across each new release: starting on WebLogic 5.1, the overall performance gain has been close to 4X, which close to a 20% gain release by release. JBoss on the other hand does not provide SPECJAppServer or SPECJEnterprise performance benchmarks. Their so called "performance gains" remains hidden in their customer environments, which lead us to think if it is true or not since we will never get access to those environments. - WebLogic Server 12c has industry performance benchmarks with submissions across platforms and configurations leading SPECJ. Oracle WebLogic leads SPECJAppServer performance in multiple categories, fitting all customer topologies like: dual-node, single-node, multi-node and multi-node with RAC. JBoss... again, does not provide any SPECJAppServer performance benchmarks. - WebLogic Server 12c has a feature called work manager which allows your application to embrace new performance levels based on critical resource utilization of the CPUs usage. Work managers prioritizes work and allocates threads based on an execution model that takes into account administrator-defined parameters and actual run-time performance and throughput. JBoss EAP 6 on the other hand has no compared feature and probably they never will. Not supporting such feature like work managers, JBoss EAP 6 forces admin people and specially developers to uncover performance gains in a intrusive way, rewriting the code and doing performance refactorings. 11) Professional Services Support - WebLogic Server 12c and any other technology sold by Oracle give customers the possibility of hire OCS ("Oracle Consulting Services") to manage critical scenarios, deployment assistance of new applications, high skilled consultancy of architecture, best practices and people allocation together with customer teams. All OCS services are available without any restrictions, having the customer bought software from Oracle or just starting their implementation before any acquisition. JBoss EAP 6 or Red Hat to be more specifically, only offers professional services if you buy subscriptions from them. If you are developing a new critical application for your business and need the help of Red Hat for a serious issue or architecture decision, they will probably say: "OK... I can help you but after you buy subscriptions from me". Red Hat also does not allows their professional services consultants to manage environments that uses community based software. They will probably force you to first buy a subscription, download their "enterprise" version and them, optionally hire their consultants. - Oracle provides you our university to educate your team into our technologies, including of course specialized trainings of WebLogic application server. At any time and location, you can hire Oracle to train your team so you get trustful knowledge according to your specific needs. Certifications for the products are also available if your technical people desire to differentiate themselves as professionals. Red Hat on the other hand have a limited pool of resources to train your team in their technologies. Basically they are selling training and certification for RHEL ("Red Hat Enterprise Linux") but if you demand more specialized training in JBoss middleware, they will probably connect you to some "certified" partner localized training since they are apparently discontinuing their education center, at least here in Brazil. They were not able to reproduce their success with RHEL education to their middleware division since they need first sell the subscriptions to after gives you specialized training. And again, they only offer you specialized training based on their enterprise version (EAP in the case of JBoss) which means that the courses will be a quite outdated. There are reports of developers that took official training's from Red Hat at this year (2012) and in a certain JBoss advanced course, Red Hat supposedly covered JBossMQ as the messaging subsystem, and even the printed material provided was based on JBossMQ since the training was created for JBoss EAP 4.3. 12) Encouraging Transparency without Ulterior Motives - WebLogic Server 12c like any other software from Oracle can be downloaded any time from anywhere, you should only possess an OTN ("Oracle Technology Network") credential and you can download any enterprise software how many times you want. And is not some kind of "trial" version. It is the official binaries that will be running for ever in your data center. Oracle does not encourages the usage of "specific versions" of our software. The binaries you buy from Oracle are the same binaries anyone in the world could download and use for testing and personal education. JBoss EAP 6 on the other hand are not available for download unless you buy a subscription and get access to the Red Hat enterprise repositories. If you need to test, learn or just start creating your application using Red Hat's middleware software, you should download it from the community website. You are not allowed to download the enterprise version that, according to Red Hat are more secure, reliable and robust. But no one of us want to start the development of a software with an unsecured, unreliable and not scalable middleware right? So what you do? You are "invited" by Red Hat to buy subscriptions from them to get access to the "cool" version of the software. - WebLogic Server 12c prices are publicly available in the Oracle website. If you want to know right now how much WebLogic will cost to your organization, just click here and get access to our price list. In the case of WebLogic, check out the "US Oracle Technology Commercial Price List". Oracle also encourages you to get in touch with a sales representative to discuss discounts that would make possible the investment into our technology. But you are not required to do this, only if you are interested in buying our technology or maybe you want to discuss some discount scenarios. JBoss EAP 6 on the other hand does not have its cost publicly available in Red Hat's website or in any other media, at least is not so easy to get such information. The only link you will possibly find in their website is a "Contact a Sales Representative" link. This is not a very good relationship between an customer and an vendor. This is not an example of transparency, mainly when the software are sold as open. In this situations, customers expects to see the software prices publicly available, so they can have the chance to decide, based on the existing features of the software, if the cost is fair or not. Conclusion Oracle WebLogic is the most mature, secure, reliable and scalable Java EE application server of the market, and have a proven record of success around the globe to prove it's majority. Don't lose the chance to discover today how WebLogic could fit your needs and sustain your global IT middleware strategy, no matter if your strategy are completely based on the Cloud or not.

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Cold Start

- by antony.reynolds

Well we had snow drifts 3ft deep on Saturday so it must be spring time. In preparation for Spring we decided to move the lawn tractor. Of course after sitting in the garage all winter it refused to start. I then come into the office and need to start my 11g SOA Suite installation. I thought about this and decided my tractor might be cranky but at least I can script the startup of my SOA Suite 11g installation. So with this in mind I created 6 scripts. I created them for Linux but they should translate to Windows without too many problems. This is left as an exercise to the reader, note you will have to hardcode more than I did in the Linux scripts and create separate script files for the sqlplus and WLST sections. Order to start things I believe there should be order in all things, especially starting the SOA Suite. So here is my preferred order. Start Database This is need by EM and the rest of SOA Suite so best to start it before the Admin Server and managed servers. Start Node Manager on all machines This is needed if you want the scripts to work across machines. Start Admin Server Once this is done in theory you can manually stat the managed servers using WebLogic console. But then you have to wait for console to be available. Scripting it all is quicker and easier way of starting. Start Managed Servers & Clusters Best to start them one per physical machine at a time to avoid undue load on the machines. Non-clustered install will have just soa_server1 and bam_serv1 by default. Clusters will have at least SOA and BAM clusters that can be started as a group or individually. I have provided scripts for standalone servers, but easy to change them to work with clusters. Starting Database I have provided a very primitive script (available here) to start the database, the listener and the DB console. The section highlighted in red needs to match your database name. #!/bin/sh echo "##############################" echo "# Setting Oracle Environment #" echo "##############################" . oraenv <<-EOF orcl EOF echo "#####################" echo "# Starting Database #" echo "#####################" sqlplus / as sysdba <<-EOF startup exit EOF echo "#####################" echo "# Starting Listener #" echo "#####################" lsnrctl start echo "######################" echo "# Starting dbConsole #" echo "######################" emctl start dbconsole read -p "Hit <enter> to continue" Starting SOA Suite My script for starting the SOA Suite (available here) breaks the task down into five sections. Setting the Environment First set up the environment variables. The variables highlighted in red probably need changing for your environment. #!/bin/sh echo "###########################" echo "# Setting SOA Environment #" echo "###########################" export MW_HOME=~oracle/Middleware11gPS1 export WL_HOME=$MW_HOME/wlserver_10.3 export ORACLE_HOME=$MW_HOME/Oracle_SOA export DOMAIN_NAME=soa_std_domain export DOMAIN_HOME=$MW_HOME/user_projects/domains/$DOMAIN_NAME Starting the Node Manager I start node manager with a nohup to stop it exiting when the script terminates and I redirect the standard output and standard error to a file in a logs directory. cd $DOMAIN_HOME echo "#########################" echo "# Starting Node Manager #" echo "#########################" nohup $WL_HOME/server/bin/startNodeManager.sh >logs/NodeManager.out 2>&1 & Starting the Admin Server I had problems starting the Admin Server from Node Manager so I decided to start it using the command line script. I again use nohup and redirect output. echo "#########################" echo "# Starting Admin Server #" echo "#########################" nohup ./startWebLogic.sh >logs/AdminServer.out 2>&1 & Starting the Managed Servers I then used WLST (WebLogic Scripting Tool) to start the managed servers. First I waited for the Admin Server to come up by putting a connect command in a loop. I could have put the WLST commands into a separate script file but I wanted to reduce the number of files I was using and so used redirected input (here syntax). $ORACLE_HOME/common/bin/wlst.sh <<-EOF import time sleep=time.sleep print "#####################################" print "# Waiting for Admin Server to Start #" print "#####################################" while True: try: connect(adminServerName="AdminServer") break except: sleep(10) I then start the SOA server and tell WLST to wait until it is started before returning. If starting a cluster then the start command would be modified accordingly to start the SOA cluster. print "#######################" print "# Starting SOA Server #" print "#######################" start(name="soa_server1", block="true") I then start the BAM server in the same way as the SOA server. print "#######################" print "# Starting BAM Server #" print "#######################" start(name="bam_server1", block="true") EOF Finally I let people know the servers are up and wait for input in case I am running in a separate window, in which case the result would be lost without the read command. echo "#####################" echo "# SOA Suite Started #" echo "#####################" read -p "Hit <enter> to continue" Stopping the SOA Suite My script for shutting down the SOA Suite (available here) is basically the reverse of my startup script. After setting the environment I connect to the Admin Server using WLST and shut down the managed servers and the admin server. Again the script would need modifying for a cluster. Stopping the Servers If I cannot connect to the Admin Server I try to connect to the node manager, in case the Admin Server is down but the managed servers are up. #!/bin/sh echo "###########################" echo "# Setting SOA Environment #" echo "###########################" export MW_HOME=~oracle/Middleware11gPS1 export WL_HOME=$MW_HOME/wlserver_10.3 export ORACLE_HOME=$MW_HOME/Oracle_SOA export DOMAIN_NAME=soa_std_domain export DOMAIN_HOME=$MW_HOME/user_projects/domains/$DOMAIN_NAME cd $DOMAIN_HOME $MW_HOME/Oracle_SOA/common/bin/wlst.sh <<-EOF try: print("#############################") print("# Connecting to AdminServer #") print("#############################") connect(username='weblogic',password='welcome1',url='t3://localhost:7001') except: print "#########################################" print "# Unable to connect to Admin Server #" print "# Attempting to connect to Node Manager #" print "#########################################" nmConnect(domainName=os.getenv("DOMAIN_NAME")) print "#######################" print "# Stopping BAM Server #" print "#######################" shutdown('bam_server1') print "#######################" print "# Stopping SOA Server #" print "#######################" shutdown('soa_server1') print "#########################" print "# Stopping Admin Server #" print "#########################" shutdown('AdminServer') disconnect() nmDisconnect() EOF Stopping the Node Manager I stopped the node manager by searching for the java node manager process using the ps command and then killing that process. echo "#########################" echo "# Stopping Node Manager #" echo "#########################" kill -9 `ps -ef | grep java | grep NodeManager | awk '{print $2;}'` echo "#####################" echo "# SOA Suite Stopped #" echo "#####################" read -p "Hit <enter> to continue" Stopping the Database Again my script for shutting down the database is the reverse of my start script. It is available here. The only change needed might be to the database name. #!/bin/sh echo "##############################" echo "# Setting Oracle Environment #" echo "##############################" . oraenv <<-EOF orcl EOF echo "######################" echo "# Stopping dbConsole #" echo "######################" emctl stop dbconsole echo "#####################" echo "# Stopping Listener #" echo "#####################" lsnrctl stop echo "#####################" echo "# Stopping Database #" echo "#####################" sqlplus / as sysdba <<-EOF shutdown immediate exit EOF read -p "Hit <enter> to continue" Cleaning Up Cleaning SOA Suite I often run tests and want to clean up all the log files. The following script (available here) does this for the WebLogic servers in a given domain on a machine. After setting the domain I just remove all files under the servers logs directories. It also cleans up the log files I created with my startup scripts. These scripts could be enhanced to copy off the log files if you needed them but in my test environments I don’t need them and would prefer to reclaim the disk space. #!/bin/sh echo "###########################" echo "# Setting SOA Environment #" echo "###########################" export MW_HOME=~oracle/Middleware11gPS1 export WL_HOME=$MW_HOME/wlserver_10.3 export ORACLE_HOME=$MW_HOME/Oracle_SOA export DOMAIN_NAME=soa_std_domain export DOMAIN_HOME=$MW_HOME/user_projects/domains/$DOMAIN_NAME echo "##########################" echo "# Cleaning SOA Log Files #" echo "##########################" cd $DOMAIN_HOME rm -Rf logs/* servers/*/logs/* read -p "Hit <enter> to continue" Cleaning Database I also created a script to clean up the dump files of an Oracle database instance and also the EM log files (available here). This relies on the machine name being correct as the EM log files are stored in a directory that is based on the hostname and the Oracle SID. #!/bin/sh echo "##############################" echo "# Setting Oracle Environment #" echo "##############################" . oraenv <<-EOF orcl EOF echo "#############################" echo "# Cleaning Oracle Log Files #" echo "#############################" rm -Rf $ORACLE_BASE/admin/$ORACLE_SID/*dump/* rm -Rf $ORACLE_HOME/`hostname`_$ORACLE_SID/sysman/log/* read -p "Hit <enter> to continue" Summary Hope you find the above scripts useful. They certainly stop me hanging around waiting for things to happen on my test machine and make it easy to run a test, change parameters, bounce the SOA Suite and clean the logs between runs so I can see exactly what is happening. Now I need to get that mower started…

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Insert Record by Drag & Drop from ADF Tree to ADF Tree Table

- by arul.wilson(at)oracle.com

If you want to create record based on the values Dragged from ADF Tree and Dropped on a ADF Tree Table, then here you go.UseCase DescriptionUser Drags a tree node from ADF Tree and Drops it on a ADF Tree Table node. A new row gets added in the Tree Table based on the source tree node, subsequently a record gets added to the database table on which Tree table in based on.Following description helps to achieve this using ADF BC.Run the DragDropSchema.sql to create required tables.Create Business Components from tables (PRODUCTS, COMPONENTS, SUB_COMPONENTS, USERS, USER_COMPONENTS) created above.Add custom method to App Module Impl, this method will be used to insert record from view layer. public String createUserComponents(String p_bugdbId, String p_productId, String p_componentId, String p_subComponentId){ Row newUserComponentsRow = this.getUserComponentsView1().createRow(); try { newUserComponentsRow.setAttribute("Bugdbid", p_bugdbId); newUserComponentsRow.setAttribute("ProductId", new oracle.jbo.domain.Number(p_productId)); newUserComponentsRow.setAttribute("Component1", p_componentId); newUserComponentsRow.setAttribute("SubComponent", p_subComponentId); } catch (Exception e) { e.printStackTrace(); return "Failure"; } return "Success"; }Expose this method to client interface.To display the root node we need a custom VO which can be achieved using below query. SELECT Users.ACTIVE, Users.BUGDB_ID, Users.EMAIL, Users.FIRSTNAME, Users.GLOBAL_ID, Users.LASTNAME, Users.MANAGER_ID, Users.MANAGER_PRIVILEGEFROM USERS UsersWHERE Users.MANAGER_ID is NULLCreate VL between UsersView and UsersRootNodeView VOs.Drop ProductsView from DC as ADF Tree to jspx page.Add Tree Level Rule based on ComponentsView and SubComponentsView.Drop UsersRootNodeView as ADF Tree TableAdd Tree Level Rules based on UserComponentsView and UsersView.Add DragSource to ADF Tree and CollectionDropTarget to ADF Tree Table respectively.Bind CollectionDropTarget's DropTarget to backing bean and implement method of signature DnDAction (DropEvent), this method gets invoked when Tree Table encounters a drop action, here details required for creating new record are captured from the drag source and passed to 'createUserComponents' method. public DnDAction onTreeDrop(DropEvent dropEvent) { String newBugdbId = ""; String msgtxt=""; try { // Getting the target node bugdb id Object serverRowKey = dropEvent.getDropSite(); if (serverRowKey != null) { //Code for Tree Table as target String dropcomponent = dropEvent.getDropComponent().toString(); dropcomponent = (String)dropcomponent.subSequence(0, dropcomponent.indexOf("[")); if (dropcomponent.equals("RichTreeTable")){ RichTreeTable richTreeTable = (RichTreeTable)dropEvent.getDropComponent(); richTreeTable.setRowKey(serverRowKey); int rowIndexTreeTable = richTreeTable.getRowIndex(); //Drop Target Logic if (((JUCtrlHierNodeBinding)richTreeTable.getRowData(rowIndexTreeTable)).getAttributeValue()==null) { //Get Parent newBugdbId = (String)((JUCtrlHierNodeBinding)richTreeTable.getRowData(rowIndexTreeTable)).getParent().getAttributeValue(); } else { if (isNum(((JUCtrlHierNodeBinding)richTreeTable.getRowData(rowIndexTreeTable)).getAttributeValue().toString())) { //Get Parent's parent newBugdbId = (String)((JUCtrlHierNodeBinding)richTreeTable.getRowData(rowIndexTreeTable)).getParent().getParent().getAttributeValue(); } else{ //Dropped on USER newBugdbId = (String)((JUCtrlHierNodeBinding)richTreeTable.getRowData(rowIndexTreeTable)).getAttributeValue(); } } } } DataFlavor<RowKeySet> df = DataFlavor.getDataFlavor(RowKeySet.class); RowKeySet droppedValue = dropEvent.getTransferable().getData(df); Object[] keys = droppedValue.toArray(); Key componentKey = null; Key subComponentKey = null; // binding for createUserComponents method defined in AppModuleImpl class to insert record in database. operationBinding = bindings.getOperationBinding("createUserComponents"); // get the Product, Component, Subcomponent details and insert to UserComponents table. // loop through the keys if more than one comp/subcomponent is select. for (int i = 0; i < keys.length; i++) { System.out.println("in for :"+i); List list = (List)keys[i]; System.out.println("list "+i+" : "+list); System.out.println("list size "+list.size()); if (list.size() == 1) { // we cannot drag and drop the highest node ! msgtxt="You cannot drop Products, please drop Component or SubComponent from the Tree."; System.out.println(msgtxt); this.showInfoMessage(msgtxt); } else { if (list.size() == 2) { // were doing the first branch, in this case all components. componentKey = (Key)list.get(1); Object[] droppedProdCompValues = componentKey.getAttributeValues(); operationBinding.getParamsMap().put("p_bugdbId",newBugdbId); operationBinding.getParamsMap().put("p_productId",droppedProdCompValues[0]); operationBinding.getParamsMap().put("p_componentId",droppedProdCompValues[1]); operationBinding.getParamsMap().put("p_subComponentId","ALL"); Object result = operationBinding.execute(); } else { subComponentKey = (Key)list.get(2); Object[] droppedProdCompSubCompValues = subComponentKey.getAttributeValues(); operationBinding.getParamsMap().put("p_bugdbId",newBugdbId); operationBinding.getParamsMap().put("p_productId",droppedProdCompSubCompValues[0]); operationBinding.getParamsMap().put("p_componentId",droppedProdCompSubCompValues[1]); operationBinding.getParamsMap().put("p_subComponentId",droppedProdCompSubCompValues[2]); Object result = operationBinding.execute(); } } } /* this.getCil1().setDisabled(false); this.getCil1().setPartialSubmit(true); */ return DnDAction.MOVE; } catch (Exception ex) { System.out.println("drop failed with : " + ex.getMessage()); ex.printStackTrace(); /* this.getCil1().setDisabled(true); */ return DnDAction.NONE; } } Run jspx page and drop a Component or Subcomponent from Products Tree to UserComponents Tree Table.

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How to find and fix performance problems in ORM powered applications

- by FransBouma

Once in a while we get requests about how to fix performance problems with our framework. As it comes down to following the same steps and looking into the same things every single time, I decided to write a blogpost about it instead, so more people can learn from this and solve performance problems in their O/R mapper powered applications. In some parts it's focused on LLBLGen Pro but it's also usable for other O/R mapping frameworks, as the vast majority of performance problems in O/R mapper powered applications are not specific for a certain O/R mapper framework. Too often, the developer looks at the wrong part of the application, trying to fix what isn't a problem in that part, and getting frustrated that 'things are so slow with <insert your favorite framework X here>'. I'm in the O/R mapper business for a long time now (almost 10 years, full time) and as it's a small world, we O/R mapper developers know almost all tricks to pull off by now: we all know what to do to make task ABC faster and what compromises (because there are almost always compromises) to deal with if we decide to make ABC faster that way. Some O/R mapper frameworks are faster in X, others in Y, but you can be sure the difference is mainly a result of a compromise some developers are willing to deal with and others aren't. That's why the O/R mapper frameworks on the market today are different in many ways, even though they all fetch and save entities from and to a database. I'm not suggesting there's no room for improvement in today's O/R mapper frameworks, there always is, but it's not a matter of 'the slowness of the application is caused by the O/R mapper' anymore. Perhaps query generation can be optimized a bit here, row materialization can be optimized a bit there, but it's mainly coming down to milliseconds. Still worth it if you're a framework developer, but it's not much compared to the time spend inside databases and in user code: if a complete fetch takes 40ms or 50ms (from call to entity object collection), it won't make a difference for your application as that 10ms difference won't be noticed. That's why it's very important to find the real locations of the problems so developers can fix them properly and don't get frustrated because their quest to get a fast, performing application failed. Performance tuning basics and rules Finding and fixing performance problems in any application is a strict procedure with four prescribed steps: isolate, analyze, interpret and fix, in that order. It's key that you don't skip a step nor make assumptions: these steps help you find the reason of a problem which seems to be there, and how to fix it or leave it as-is. Skipping a step, or when you assume things will be bad/slow without doing analysis will lead to the path of premature optimization and won't actually solve your problems, only create new ones. The most important rule of finding and fixing performance problems in software is that you have to understand what 'performance problem' actually means. Most developers will say "when a piece of software / code is slow, you have a performance problem". But is that actually the case? If I write a Linq query which will aggregate, group and sort 5 million rows from several tables to produce a resultset of 10 rows, it might take more than a couple of milliseconds before that resultset is ready to be consumed by other logic. If I solely look at the Linq query, the code consuming the resultset of the 10 rows and then look at the time it takes to complete the whole procedure, it will appear to me to be slow: all that time taken to produce and consume 10 rows? But if you look closer, if you analyze and interpret the situation, you'll see it does a tremendous amount of work, and in that light it might even be extremely fast. With every performance problem you encounter, always do realize that what you're trying to solve is perhaps not a technical problem at all, but a perception problem. The second most important rule you have to understand is based on the old saying "Penny wise, Pound Foolish": the part which takes e.g. 5% of the total time T for a given task isn't worth optimizing if you have another part which takes a much larger part of the total time T for that same given task. Optimizing parts which are relatively insignificant for the total time taken is not going to bring you better results overall, even if you totally optimize that part away. This is the core reason why analysis of the complete set of application parts which participate in a given task is key to being successful in solving performance problems: No analysis -> no problem -> no solution. One warning up front: hunting for performance will always include making compromises. Fast software can be made maintainable, but if you want to squeeze as much performance out of your software, you will inevitably be faced with the dilemma of compromising one or more from the group {readability, maintainability, features} for the extra performance you think you'll gain. It's then up to you to decide whether it's worth it. In almost all cases it's not. The reason for this is simple: the vast majority of performance problems can be solved by implementing the proper algorithms, the ones with proven Big O-characteristics so you know the performance you'll get plus you know the algorithm will work. The time taken by the algorithm implementing code is inevitable: you already implemented the best algorithm. You might find some optimizations on the technical level but in general these are minor. Let's look at the four steps to see how they guide us through the quest to find and fix performance problems. Isolate The first thing you need to do is to isolate the areas in your application which are assumed to be slow. For example, if your application is a web application and a given page is taking several seconds or even minutes to load, it's a good candidate to check out. It's important to start with the isolate step because it allows you to focus on a single code path per area with a clear begin and end and ignore the rest. The rest of the steps are taken per identified problematic area. Keep in mind that isolation focuses on tasks in an application, not code snippets. A task is something that's started in your application by either another task or the user, or another program, and has a beginning and an end. You can see a task as a piece of functionality offered by your application. Analyze Once you've determined the problem areas, you have to perform analysis on the code paths of each area, to see where the performance problems occur and which areas are not the problem. This is a multi-layered effort: an application which uses an O/R mapper typically consists of multiple parts: there's likely some kind of interface (web, webservice, windows etc.), a part which controls the interface and business logic, the O/R mapper part and the RDBMS, all connected with either a network or inter-process connections provided by the OS or other means. Each of these parts, including the connectivity plumbing, eat up a part of the total time it takes to complete a task, e.g. load a webpage with all orders of a given customer X. To understand which parts participate in the task / area we're investigating and how much they contribute to the total time taken to complete the task, analysis of each participating task is essential. Start with the code you wrote which starts the task, analyze the code and track the path it follows through your application. What does the code do along the way, verify whether it's correct or not. Analyze whether you have implemented the right algorithms in your code for this particular area. Remember we're looking at one area at a time, which means we're ignoring all other code paths, just the code path of the current problematic area, from begin to end and back. Don't dig in and start optimizing at the code level just yet. We're just analyzing. If your analysis reveals big architectural stupidity, it's perhaps a good idea to rethink the architecture at this point. For the rest, we're analyzing which means we collect data about what could be wrong, for each participating part of the complete application. Reviewing the code you wrote is a good tool to get deeper understanding of what is going on for a given task but ultimately it lacks precision and overview what really happens: humans aren't good code interpreters, computers are. We therefore need to utilize tools to get deeper understanding about which parts contribute how much time to the total task, triggered by which other parts and for example how many times are they called. There are two different kind of tools which are necessary: .NET profilers and O/R mapper / RDBMS profilers. .NET profiling .NET profilers (e.g. dotTrace by JetBrains or Ants by Red Gate software) show exactly which pieces of code are called, how many times they're called, and the time it took to run that piece of code, at the method level and sometimes even at the line level. The .NET profilers are essential tools for understanding whether the time taken to complete a given task / area in your application is consumed by .NET code, where exactly in your code, the path to that code, how many times that code was called by other code and thus reveals where hotspots are located: the areas where a solution can be found. Importantly, they also reveal which areas can be left alone: remember our penny wise pound foolish saying: if a profiler reveals that a group of methods are fast, or don't contribute much to the total time taken for a given task, ignore them. Even if the code in them is perhaps complex and looks like a candidate for optimization: you can work all day on that, it won't matter. As we're focusing on a single area of the application, it's best to start profiling right before you actually activate the task/area. Most .NET profilers support this by starting the application without starting the profiling procedure just yet. You navigate to the particular part which is slow, start profiling in the profiler, in your application you perform the actions which are considered slow, and afterwards you get a snapshot in the profiler. The snapshot contains the data collected by the profiler during the slow action, so most data is produced by code in the area to investigate. This is important, because it allows you to stay focused on a single area. O/R mapper and RDBMS profiling .NET profilers give you a good insight in the .NET side of things, but not in the RDBMS side of the application. As this article is about O/R mapper powered applications, we're also looking at databases, and the software making it possible to consume the database in your application: the O/R mapper. To understand which parts of the O/R mapper and database participate how much to the total time taken for task T, we need different tools. There are two kind of tools focusing on O/R mappers and database performance profiling: O/R mapper profilers and RDBMS profilers. For O/R mapper profilers, you can look at LLBLGen Prof by hibernating rhinos or the Linq to Sql/LLBLGen Pro profiler by Huagati. Hibernating rhinos also have profilers for other O/R mappers like NHibernate (NHProf) and Entity Framework (EFProf) and work the same as LLBLGen Prof. For RDBMS profilers, you have to look whether the RDBMS vendor has a profiler. For example for SQL Server, the profiler is shipped with SQL Server, for Oracle it's build into the RDBMS, however there are also 3rd party tools. Which tool you're using isn't really important, what's important is that you get insight in which queries are executed during the task / area we're currently focused on and how long they took. Here, the O/R mapper profilers have an advantage as they collect the time it took to execute the query from the application's perspective so they also collect the time it took to transport data across the network. This is important because a query which returns a massive resultset or a resultset with large blob/clob/ntext/image fields takes more time to get transported across the network than a small resultset and a database profiler doesn't take this into account most of the time. Another tool to use in this case, which is more low level and not all O/R mappers support it (though LLBLGen Pro and NHibernate as well do) is tracing: most O/R mappers offer some form of tracing or logging system which you can use to collect the SQL generated and executed and often also other activity behind the scenes. While tracing can produce a tremendous amount of data in some cases, it also gives insight in what's going on. Interpret After we've completed the analysis step it's time to look at the data we've collected. We've done code reviews to see whether we've done anything stupid and which parts actually take place and if the proper algorithms have been implemented. We've done .NET profiling to see which parts are choke points and how much time they contribute to the total time taken to complete the task we're investigating. We've performed O/R mapper profiling and RDBMS profiling to see which queries were executed during the task, how many queries were generated and executed and how long they took to complete, including network transportation. All this data reveals two things: which parts are big contributors to the total time taken and which parts are irrelevant. Both aspects are very important. The parts which are irrelevant (i.e. don't contribute significantly to the total time taken) can be ignored from now on, we won't look at them. The parts which contribute a lot to the total time taken are important to look at. We now have to first look at the .NET profiler results, to see whether the time taken is consumed in our own code, in .NET framework code, in the O/R mapper itself or somewhere else. For example if most of the time is consumed by DbCommand.ExecuteReader, the time it took to complete the task is depending on the time the data is fetched from the database. If there was just 1 query executed, according to tracing or O/R mapper profilers / RDBMS profilers, check whether that query is optimal, uses indexes or has to deal with a lot of data. Interpret means that you follow the path from begin to end through the data collected and determine where, along the path, the most time is contributed. It also means that you have to check whether this was expected or is totally unexpected. My previous example of the 10 row resultset of a query which groups millions of rows will likely reveal that a long time is spend inside the database and almost no time is spend in the .NET code, meaning the RDBMS part contributes the most to the total time taken, the rest is compared to that time, irrelevant. Considering the vastness of the source data set, it's expected this will take some time. However, does it need tweaking? Perhaps all possible tweaks are already in place. In the interpret step you then have to decide that further action in this area is necessary or not, based on what the analysis results show: if the analysis results were unexpected and in the area where the most time is contributed to the total time taken is room for improvement, action should be taken. If not, you can only accept the situation and move on. In all cases, document your decision together with the analysis you've done. If you decide that the perceived performance problem is actually expected due to the nature of the task performed, it's essential that in the future when someone else looks at the application and starts asking questions you can answer them properly and new analysis is only necessary if situations changed. Fix After interpreting the analysis results you've concluded that some areas need adjustment. This is the fix step: you're actively correcting the performance problem with proper action targeted at the real cause. In many cases related to O/R mapper powered applications it means you'll use different features of the O/R mapper to achieve the same goal, or apply optimizations at the RDBMS level. It could also mean you apply caching inside your application (compromise memory consumption over performance) to avoid unnecessary re-querying data and re-consuming the results. After applying a change, it's key you re-do the analysis and interpretation steps: compare the results and expectations with what you had before, to see whether your actions had any effect or whether it moved the problem to a different part of the application. Don't fall into the trap to do partly analysis: do the full analysis again: .NET profiling and O/R mapper / RDBMS profiling. It might very well be that the changes you've made make one part faster but another part significantly slower, in such a way that the overall problem hasn't changed at all. Performance tuning is dealing with compromises and making choices: to use one feature over the other, to accept a higher memory footprint, to go away from the strict-OO path and execute queries directly onto the RDBMS, these are choices and compromises which will cross your path if you want to fix performance problems with respect to O/R mappers or data-access and databases in general. In most cases it's not a big issue: alternatives are often good choices too and the compromises aren't that hard to deal with. What is important is that you document why you made a choice, a compromise: which analysis data, which interpretation led you to the choice made. This is key for good maintainability in the years to come. Most common performance problems with O/R mappers Below is an incomplete list of common performance problems related to data-access / O/R mappers / RDBMS code. It will help you with fixing the hotspots you found in the interpretation step. SELECT N+1: (Lazy-loading specific). Lazy loading triggered performance bottlenecks. Consider a list of Orders bound to a grid. You have a Field mapped onto a related field in Order, Customer.CompanyName. Showing this column in the grid will make the grid fetch (indirectly) for each row the Customer row. This means you'll get for the single list not 1 query (for the orders) but 1+(the number of orders shown) queries. To solve this: use eager loading using a prefetch path to fetch the customers with the orders. SELECT N+1 is easy to spot with an O/R mapper profiler or RDBMS profiler: if you see a lot of identical queries executed at once, you have this problem. Prefetch paths using many path nodes or sorting, or limiting. Eager loading problem. Prefetch paths can help with performance, but as 1 query is fetched per node, it can be the number of data fetched in a child node is bigger than you think. Also consider that data in every node is merged on the client within the parent. This is fast, but it also can take some time if you fetch massive amounts of entities. If you keep fetches small, you can use tuning parameters like the ParameterizedPrefetchPathThreshold setting to get more optimal queries. Deep inheritance hierarchies of type Target Per Entity/Type. If you use inheritance of type Target per Entity / Type (each type in the inheritance hierarchy is mapped onto its own table/view), fetches will join subtype- and supertype tables in many cases, which can lead to a lot of performance problems if the hierarchy has many types. With this problem, keep inheritance to a minimum if possible, or switch to a hierarchy of type Target Per Hierarchy, which means all entities in the inheritance hierarchy are mapped onto the same table/view. Of course this has its own set of drawbacks, but it's a compromise you might want to take. Fetching massive amounts of data by fetching large lists of entities. LLBLGen Pro supports paging (and limiting the # of rows returned), which is often key to process through large sets of data. Use paging on the RDBMS if possible (so a query is executed which returns only the rows in the page requested). When using paging in a web application, be sure that you switch server-side paging on on the datasourcecontrol used. In this case, paging on the grid alone is not enough: this can lead to fetching a lot of data which is then loaded into the grid and paged there. Keep note that analyzing queries for paging could lead to the false assumption that paging doesn't occur, e.g. when the query contains a field of type ntext/image/clob/blob and DISTINCT can't be applied while it should have (e.g. due to a join): the datareader will do DISTINCT filtering on the client. this is a little slower but it does perform paging functionality on the data-reader so it won't fetch all rows even if the query suggests it does. Fetch massive amounts of data because blob/clob/ntext/image fields aren't excluded. LLBLGen Pro supports field exclusion for queries. You can exclude fields (also in prefetch paths) per query to avoid fetching all fields of an entity, e.g. when you don't need them for the logic consuming the resultset. Excluding fields can greatly reduce the amount of time spend on data-transport across the network. Use this optimization if you see that there's a big difference between query execution time on the RDBMS and the time reported by the .NET profiler for the ExecuteReader method call. Doing client-side aggregates/scalar calculations by consuming a lot of data. If possible, try to formulate a scalar query or group by query using the projection system or GetScalar functionality of LLBLGen Pro to do data consumption on the RDBMS server. It's far more efficient to process data on the RDBMS server than to first load it all in memory, then traverse the data in-memory to calculate a value. Using .ToList() constructs inside linq queries. It might be you use .ToList() somewhere in a Linq query which makes the query be run partially in-memory. Example: var q = from c in metaData.Customers.ToList() where c.Country=="Norway" select c; This will actually fetch all customers in-memory and do an in-memory filtering, as the linq query is defined on an IEnumerable<T>, and not on the IQueryable<T>. Linq is nice, but it can often be a bit unclear where some parts of a Linq query might run. Fetching all entities to delete into memory first. To delete a set of entities it's rather inefficient to first fetch them all into memory and then delete them one by one. It's more efficient to execute a DELETE FROM ... WHERE query on the database directly to delete the entities in one go. LLBLGen Pro supports this feature, and so do some other O/R mappers. It's not always possible to do this operation in the context of an O/R mapper however: if an O/R mapper relies on a cache, these kind of operations are likely not supported because they make it impossible to track whether an entity is actually removed from the DB and thus can be removed from the cache. Fetching all entities to update with an expression into memory first. Similar to the previous point: it is more efficient to update a set of entities directly with a single UPDATE query using an expression instead of fetching the entities into memory first and then updating the entities in a loop, and afterwards saving them. It might however be a compromise you don't want to take as it is working around the idea of having an object graph in memory which is manipulated and instead makes the code fully aware there's a RDBMS somewhere. Conclusion Performance tuning is almost always about compromises and making choices. It's also about knowing where to look and how the systems in play behave and should behave. The four steps I provided should help you stay focused on the real problem and lead you towards the solution. Knowing how to optimally use the systems participating in your own code (.NET framework, O/R mapper, RDBMS, network/services) is key for success as well as knowing what's going on inside the application you built. I hope you'll find this guide useful in tracking down performance problems and dealing with them in a useful way.

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When is a SQL function not a function?

- by Rob Farley

Should SQL Server even have functions? (Oh yeah – this is a T-SQL Tuesday post, hosted this month by Brad Schulz) Functions serve an important part of programming, in almost any language. A function is a piece of code that is designed to return something, as opposed to a piece of code which isn’t designed to return anything (which is known as a procedure). SQL Server is no different. You can call stored procedures, even from within other stored procedures, and you can call functions and use these in other queries. Stored procedures might query something, and therefore ‘return data’, but a function in SQL is considered to have the type of the thing returned, and can be used accordingly in queries. Consider the internal GETDATE() function. SELECT GETDATE(), SomeDatetimeColumn FROM dbo.SomeTable; There’s no logical difference between the field that is being returned by the function and the field that’s being returned by the table column. Both are the datetime field – if you didn’t have inside knowledge, you wouldn’t necessarily be able to tell which was which. And so as developers, we find ourselves wanting to create functions that return all kinds of things – functions which look up values based on codes, functions which do string manipulation, and so on. But it’s rubbish. Ok, it’s not all rubbish, but it mostly is. And this isn’t even considering the SARGability impact. It’s far more significant than that. (When I say the SARGability aspect, I mean “because you’re unlikely to have an index on the result of some function that’s applied to a column, so try to invert the function and query the column in an unchanged manner”) I’m going to consider the three main types of user-defined functions in SQL Server: Scalar Inline Table-Valued Multi-statement Table-Valued I could also look at user-defined CLR functions, including aggregate functions, but not today. I figure that most people don’t tend to get around to doing CLR functions, and I’m going to focus on the T-SQL-based user-defined functions. Most people split these types of function up into two types. So do I. Except that most people pick them based on ‘scalar or table-valued’. I’d rather go with ‘inline or not’. If it’s not inline, it’s rubbish. It really is. Let’s start by considering the two kinds of table-valued function, and compare them. These functions are going to return the sales for a particular salesperson in a particular year, from the AdventureWorks database. CREATE FUNCTION dbo.FetchSales_inline(@salespersonid int, @orderyear int) RETURNS TABLE AS RETURN ( SELECT e.LoginID as EmployeeLogin, o.OrderDate, o.SalesOrderID FROM Sales.SalesOrderHeader AS o LEFT JOIN HumanResources.Employee AS e ON e.EmployeeID = o.SalesPersonID WHERE o.SalesPersonID = @salespersonid AND o.OrderDate >= DATEADD(year,@orderyear-2000,'20000101') AND o.OrderDate < DATEADD(year,@orderyear-2000+1,'20000101') ) ; GO CREATE FUNCTION dbo.FetchSales_multi(@salespersonid int, @orderyear int) RETURNS @results TABLE ( EmployeeLogin nvarchar(512), OrderDate datetime, SalesOrderID int ) AS BEGIN INSERT @results (EmployeeLogin, OrderDate, SalesOrderID) SELECT e.LoginID, o.OrderDate, o.SalesOrderID FROM Sales.SalesOrderHeader AS o LEFT JOIN HumanResources.Employee AS e ON e.EmployeeID = o.SalesPersonID WHERE o.SalesPersonID = @salespersonid AND o.OrderDate >= DATEADD(year,@orderyear-2000,'20000101') AND o.OrderDate < DATEADD(year,@orderyear-2000+1,'20000101') ; RETURN END ; GO You’ll notice that I’m being nice and responsible with the use of the DATEADD function, so that I have SARGability on the OrderDate filter. Regular readers will be hoping I’ll show what’s going on in the execution plans here. Here I’ve run two SELECT * queries with the “Show Actual Execution Plan” option turned on. Notice that the ‘Query cost’ of the multi-statement version is just 2% of the ‘Batch cost’. But also notice there’s trickery going on. And it’s nothing to do with that extra index that I have on the OrderDate column. Trickery. Look at it – clearly, the first plan is showing us what’s going on inside the function, but the second one isn’t. The second one is blindly running the function, and then scanning the results. There’s a Sequence operator which is calling the TVF operator, and then calling a Table Scan to get the results of that function for the SELECT operator. But surely it still has to do all the work that the first one is doing... To see what’s actually going on, let’s look at the Estimated plan. Now, we see the same plans (almost) that we saw in the Actuals, but we have an extra one – the one that was used for the TVF. Here’s where we see the inner workings of it. You’ll probably recognise the right-hand side of the TVF’s plan as looking very similar to the first plan – but it’s now being called by a stack of other operators, including an INSERT statement to be able to populate the table variable that the multi-statement TVF requires. And the cost of the TVF is 57% of the batch! But it gets worse. Let’s consider what happens if we don’t need all the columns. We’ll leave out the EmployeeLogin column. Here, we see that the inline function call has been simplified down. It doesn’t need the Employee table. The join is redundant and has been eliminated from the plan, making it even cheaper. But the multi-statement plan runs the whole thing as before, only removing the extra column when the Table Scan is performed. A multi-statement function is a lot more powerful than an inline one. An inline function can only be the result of a single sub-query. It’s essentially the same as a parameterised view, because views demonstrate this same behaviour of extracting the definition of the view and using it in the outer query. A multi-statement function is clearly more powerful because it can contain far more complex logic. But a multi-statement function isn’t really a function at all. It’s a stored procedure. It’s wrapped up like a function, but behaves like a stored procedure. It would be completely unreasonable to expect that a stored procedure could be simplified down to recognise that not all the columns might be needed, but yet this is part of the pain associated with this procedural function situation. The biggest clue that a multi-statement function is more like a stored procedure than a function is the “BEGIN” and “END” statements that surround the code. If you try to create a multi-statement function without these statements, you’ll get an error – they are very much required. When I used to present on this kind of thing, I even used to call it “The Dangers of BEGIN and END”, and yes, I’ve written about this type of thing before in a similarly-named post over at my old blog. Now how about scalar functions... Suppose we wanted a scalar function to return the count of these. CREATE FUNCTION dbo.FetchSales_scalar(@salespersonid int, @orderyear int) RETURNS int AS BEGIN RETURN ( SELECT COUNT(*) FROM Sales.SalesOrderHeader AS o LEFT JOIN HumanResources.Employee AS e ON e.EmployeeID = o.SalesPersonID WHERE o.SalesPersonID = @salespersonid AND o.OrderDate >= DATEADD(year,@orderyear-2000,'20000101') AND o.OrderDate < DATEADD(year,@orderyear-2000+1,'20000101') ); END ; GO Notice the evil words? They’re required. Try to remove them, you just get an error. That’s right – any scalar function is procedural, despite the fact that you wrap up a sub-query inside that RETURN statement. It’s as ugly as anything. Hopefully this will change in future versions. Let’s have a look at how this is reflected in an execution plan. Here’s a query, its Actual plan, and its Estimated plan: SELECT e.LoginID, y.year, dbo.FetchSales_scalar(p.SalesPersonID, y.year) AS NumSales FROM (VALUES (2001),(2002),(2003),(2004)) AS y (year) CROSS JOIN Sales.SalesPerson AS p LEFT JOIN HumanResources.Employee AS e ON e.EmployeeID = p.SalesPersonID; We see here that the cost of the scalar function is about twice that of the outer query. Nicely, the query optimizer has worked out that it doesn’t need the Employee table, but that’s a bit of a red herring here. There’s actually something way more significant going on. If I look at the properties of that UDF operator, it tells me that the Estimated Subtree Cost is 0.337999. If I just run the query SELECT dbo.FetchSales_scalar(281,2003); we see that the UDF cost is still unchanged. You see, this 0.0337999 is the cost of running the scalar function ONCE. But when we ran that query with the CROSS JOIN in it, we returned quite a few rows. 68 in fact. Could’ve been a lot more, if we’d had more salespeople or more years. And so we come to the biggest problem. This procedure (I don’t want to call it a function) is getting called 68 times – each one between twice as expensive as the outer query. And because it’s calling it in a separate context, there is even more overhead that I haven’t considered here. The cheek of it, to say that the Compute Scalar operator here costs 0%! I know a number of IT projects that could’ve used that kind of costing method, but that’s another story that I’m not going to go into here. Let’s look at a better way. Suppose our scalar function had been implemented as an inline one. Then it could have been expanded out like a sub-query. It could’ve run something like this: SELECT e.LoginID, y.year, (SELECT COUNT(*) FROM Sales.SalesOrderHeader AS o LEFT JOIN HumanResources.Employee AS e ON e.EmployeeID = o.SalesPersonID WHERE o.SalesPersonID = p.SalesPersonID AND o.OrderDate >= DATEADD(year,y.year-2000,'20000101') AND o.OrderDate < DATEADD(year,y.year-2000+1,'20000101') ) AS NumSales FROM (VALUES (2001),(2002),(2003),(2004)) AS y (year) CROSS JOIN Sales.SalesPerson AS p LEFT JOIN HumanResources.Employee AS e ON e.EmployeeID = p.SalesPersonID; Don’t worry too much about the Scan of the SalesOrderHeader underneath a Nested Loop. If you remember from plenty of other posts on the matter, execution plans don’t push the data through. That Scan only runs once. The Index Spool sucks the data out of it and populates a structure that is used to feed the Stream Aggregate. The Index Spool operator gets called 68 times, but the Scan only once (the Number of Executions property demonstrates this). Here, the Query Optimizer has a full picture of what’s being asked, and can make the appropriate decision about how it accesses the data. It can simplify it down properly. To get this kind of behaviour from a function, we need it to be inline. But without inline scalar functions, we need to make our function be table-valued. Luckily, that’s ok. CREATE FUNCTION dbo.FetchSales_inline2(@salespersonid int, @orderyear int) RETURNS table AS RETURN (SELECT COUNT(*) as NumSales FROM Sales.SalesOrderHeader AS o LEFT JOIN HumanResources.Employee AS e ON e.EmployeeID = o.SalesPersonID WHERE o.SalesPersonID = @salespersonid AND o.OrderDate >= DATEADD(year,@orderyear-2000,'20000101') AND o.OrderDate < DATEADD(year,@orderyear-2000+1,'20000101') ); GO But we can’t use this as a scalar. Instead, we need to use it with the APPLY operator. SELECT e.LoginID, y.year, n.NumSales FROM (VALUES (2001),(2002),(2003),(2004)) AS y (year) CROSS JOIN Sales.SalesPerson AS p LEFT JOIN HumanResources.Employee AS e ON e.EmployeeID = p.SalesPersonID OUTER APPLY dbo.FetchSales_inline2(p.SalesPersonID, y.year) AS n; And now, we get the plan that we want for this query. All we’ve done is tell the function that it’s returning a table instead of a single value, and removed the BEGIN and END statements. We’ve had to name the column being returned, but what we’ve gained is an actual inline simplifiable function. And if we wanted it to return multiple columns, it could do that too. I really consider this function to be superior to the scalar function in every way. It does need to be handled differently in the outer query, but in many ways it’s a more elegant method there too. The function calls can be put amongst the FROM clause, where they can then be used in the WHERE or GROUP BY clauses without fear of calling the function multiple times (another horrible side effect of functions). So please. If you see BEGIN and END in a function, remember it’s not really a function, it’s a procedure. And then fix it. @rob_farley

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Beware Sneaky Reads with Unique Indexes

- by Paul White NZ

A few days ago, Sandra Mueller (twitter | blog) asked a question using twitter’s #sqlhelp hash tag: “Might SQL Server retrieve (out-of-row) LOB data from a table, even if the column isn’t referenced in the query?” Leaving aside trivial cases (like selecting a computed column that does reference the LOB data), one might be tempted to say that no, SQL Server does not read data you haven’t asked for. In general, that’s quite correct; however there are cases where SQL Server might sneakily retrieve a LOB column… Example Table Here’s a T-SQL script to create that table and populate it with 1,000 rows: CREATE TABLE dbo.LOBtest ( pk INTEGER IDENTITY NOT NULL, some_value INTEGER NULL, lob_data VARCHAR(MAX) NULL, another_column CHAR(5) NULL, CONSTRAINT [PK dbo.LOBtest pk] PRIMARY KEY CLUSTERED (pk ASC) ); GO DECLARE @Data VARCHAR(MAX); SET @Data = REPLICATE(CONVERT(VARCHAR(MAX), 'x'), 65540); WITH Numbers (n) AS ( SELECT ROW_NUMBER() OVER (ORDER BY (SELECT 0)) FROM master.sys.columns C1, master.sys.columns C2 ) INSERT LOBtest WITH (TABLOCKX) ( some_value, lob_data ) SELECT TOP (1000) N.n, @Data FROM Numbers N WHERE N.n <= 1000; Test 1: A Simple Update Let’s run a query to subtract one from every value in the some_value column: UPDATE dbo.LOBtest WITH (TABLOCKX) SET some_value = some_value - 1; As you might expect, modifying this integer column in 1,000 rows doesn’t take very long, or use many resources. The STATITICS IO and TIME output shows a total of 9 logical reads, and 25ms elapsed time. The query plan is also very simple: Looking at the Clustered Index Scan, we can see that SQL Server only retrieves the pk and some_value columns during the scan: The pk column is needed by the Clustered Index Update operator to uniquely identify the row that is being changed. The some_value column is used by the Compute Scalar to calculate the new value. (In case you are wondering what the Top operator is for, it is used to enforce SET ROWCOUNT). Test 2: Simple Update with an Index Now let’s create a nonclustered index keyed on the some_value column, with lob_data as an included column: CREATE NONCLUSTERED INDEX [IX dbo.LOBtest some_value (lob_data)] ON dbo.LOBtest (some_value) INCLUDE ( lob_data ) WITH ( FILLFACTOR = 100, MAXDOP = 1, SORT_IN_TEMPDB = ON ); This is not a useful index for our simple update query; imagine that someone else created it for a different purpose. Let’s run our update query again: UPDATE dbo.LOBtest WITH (TABLOCKX) SET some_value = some_value - 1; We find that it now requires 4,014 logical reads and the elapsed query time has increased to around 100ms. The extra logical reads (4 per row) are an expected consequence of maintaining the nonclustered index. The query plan is very similar to before (click to enlarge): The Clustered Index Update operator picks up the extra work of maintaining the nonclustered index. The new Compute Scalar operators detect whether the value in the some_value column has actually been changed by the update. SQL Server may be able to skip maintaining the nonclustered index if the value hasn’t changed (see my previous post on non-updating updates for details). Our simple query does change the value of some_data in every row, so this optimization doesn’t add any value in this specific case. The output list of columns from the Clustered Index Scan hasn’t changed from the one shown previously: SQL Server still just reads the pk and some_data columns. Cool. Overall then, adding the nonclustered index hasn’t had any startling effects, and the LOB column data still isn’t being read from the table. Let’s see what happens if we make the nonclustered index unique. Test 3: Simple Update with a Unique Index Here’s the script to create a new unique index, and drop the old one: CREATE UNIQUE NONCLUSTERED INDEX [UQ dbo.LOBtest some_value (lob_data)] ON dbo.LOBtest (some_value) INCLUDE ( lob_data ) WITH ( FILLFACTOR = 100, MAXDOP = 1, SORT_IN_TEMPDB = ON ); GO DROP INDEX [IX dbo.LOBtest some_value (lob_data)] ON dbo.LOBtest; Remember that SQL Server only enforces uniqueness on index keys (the some_data column). The lob_data column is simply stored at the leaf-level of the non-clustered index. With that in mind, we might expect this change to make very little difference. Let’s see: UPDATE dbo.LOBtest WITH (TABLOCKX) SET some_value = some_value - 1; Whoa! Now look at the elapsed time and logical reads: Scan count 1, logical reads 2016, physical reads 0, read-ahead reads 0, lob logical reads 36015, lob physical reads 0, lob read-ahead reads 15992. CPU time = 172 ms, elapsed time = 16172 ms. Even with all the data and index pages in memory, the query took over 16 seconds to update just 1,000 rows, performing over 52,000 LOB logical reads (nearly 16,000 of those using read-ahead). Why on earth is SQL Server reading LOB data in a query that only updates a single integer column? The Query Plan The query plan for test 3 looks a bit more complex than before: In fact, the bottom level is exactly the same as we saw with the non-unique index. The top level has heaps of new stuff though, which I’ll come to in a moment. You might be expecting to find that the Clustered Index Scan is now reading the lob_data column (for some reason). After all, we need to explain where all the LOB logical reads are coming from. Sadly, when we look at the properties of the Clustered Index Scan, we see exactly the same as before: SQL Server is still only reading the pk and some_value columns – so what’s doing the LOB reads? Updates that Sneakily Read Data We have to go as far as the Clustered Index Update operator before we see LOB data in the output list: [Expr1020] is a bit flag added by an earlier Compute Scalar. It is set true if the some_value column has not been changed (part of the non-updating updates optimization I mentioned earlier). The Clustered Index Update operator adds two new columns: the lob_data column, and some_value_OLD. The some_value_OLD column, as the name suggests, is the pre-update value of the some_value column. At this point, the clustered index has already been updated with the new value, but we haven’t touched the nonclustered index yet. An interesting observation here is that the Clustered Index Update operator can read a column into the data flow as part of its update operation. SQL Server could have read the LOB data as part of the initial Clustered Index Scan, but that would mean carrying the data through all the operations that occur prior to the Clustered Index Update. The server knows it will have to go back to the clustered index row to update it, so it delays reading the LOB data until then. Sneaky! Why the LOB Data Is Needed This is all very interesting (I hope), but why is SQL Server reading the LOB data? For that matter, why does it need to pass the pre-update value of the some_value column out of the Clustered Index Update? The answer relates to the top row of the query plan for test 3. I’ll reproduce it here for convenience: Notice that this is a wide (per-index) update plan. SQL Server used a narrow (per-row) update plan in test 2, where the Clustered Index Update took care of maintaining the nonclustered index too. I’ll talk more about this difference shortly. The Split/Sort/Collapse combination is an optimization, which aims to make per-index update plans more efficient. It does this by breaking each update into a delete/insert pair, reordering the operations, removing any redundant operations, and finally applying the net effect of all the changes to the nonclustered index. Imagine we had a unique index which currently holds three rows with the values 1, 2, and 3. If we run a query that adds 1 to each row value, we would end up with values 2, 3, and 4. The net effect of all the changes is the same as if we simply deleted the value 1, and added a new value 4. By applying net changes, SQL Server can also avoid false unique-key violations. If we tried to immediately update the value 1 to a 2, it would conflict with the existing value 2 (which would soon be updated to 3 of course) and the query would fail. You might argue that SQL Server could avoid the uniqueness violation by starting with the highest value (3) and working down. That’s fine, but it’s not possible to generalize this logic to work with every possible update query. SQL Server has to use a wide update plan if it sees any risk of false uniqueness violations. It’s worth noting that the logic SQL Server uses to detect whether these violations are possible has definite limits. As a result, you will often receive a wide update plan, even when you can see that no violations are possible. Another benefit of this optimization is that it includes a sort on the index key as part of its work. Processing the index changes in index key order promotes sequential I/O against the nonclustered index. A side-effect of all this is that the net changes might include one or more inserts. In order to insert a new row in the index, SQL Server obviously needs all the columns – the key column and the included LOB column. This is the reason SQL Server reads the LOB data as part of the Clustered Index Update. In addition, the some_value_OLD column is required by the Split operator (it turns updates into delete/insert pairs). In order to generate the correct index key delete operation, it needs the old key value. The irony is that in this case the Split/Sort/Collapse optimization is anything but. Reading all that LOB data is extremely expensive, so it is sad that the current version of SQL Server has no way to avoid it. Finally, for completeness, I should mention that the Filter operator is there to filter out the non-updating updates. Beating the Set-Based Update with a Cursor One situation where SQL Server can see that false unique-key violations aren’t possible is where it can guarantee that only one row is being updated. Armed with this knowledge, we can write a cursor (or the WHILE-loop equivalent) that updates one row at a time, and so avoids reading the LOB data: SET NOCOUNT ON; SET STATISTICS XML, IO, TIME OFF; DECLARE @PK INTEGER, @StartTime DATETIME; SET @StartTime = GETUTCDATE(); DECLARE curUpdate CURSOR LOCAL FORWARD_ONLY KEYSET SCROLL_LOCKS FOR SELECT L.pk FROM LOBtest L ORDER BY L.pk ASC; OPEN curUpdate; WHILE (1 = 1) BEGIN FETCH NEXT FROM curUpdate INTO @PK; IF @@FETCH_STATUS = -1 BREAK; IF @@FETCH_STATUS = -2 CONTINUE; UPDATE dbo.LOBtest SET some_value = some_value - 1 WHERE CURRENT OF curUpdate; END; CLOSE curUpdate; DEALLOCATE curUpdate; SELECT DATEDIFF(MILLISECOND, @StartTime, GETUTCDATE()); That completes the update in 1280 milliseconds (remember test 3 took over 16 seconds!) I used the WHERE CURRENT OF syntax there and a KEYSET cursor, just for the fun of it. One could just as well use a WHERE clause that specified the primary key value instead. Clustered Indexes A clustered index is the ultimate index with included columns: all non-key columns are included columns in a clustered index. Let’s re-create the test table and data with an updatable primary key, and without any non-clustered indexes: IF OBJECT_ID(N'dbo.LOBtest', N'U') IS NOT NULL DROP TABLE dbo.LOBtest; GO CREATE TABLE dbo.LOBtest ( pk INTEGER NOT NULL, some_value INTEGER NULL, lob_data VARCHAR(MAX) NULL, another_column CHAR(5) NULL, CONSTRAINT [PK dbo.LOBtest pk] PRIMARY KEY CLUSTERED (pk ASC) ); GO DECLARE @Data VARCHAR(MAX); SET @Data = REPLICATE(CONVERT(VARCHAR(MAX), 'x'), 65540); WITH Numbers (n) AS ( SELECT ROW_NUMBER() OVER (ORDER BY (SELECT 0)) FROM master.sys.columns C1, master.sys.columns C2 ) INSERT LOBtest WITH (TABLOCKX) ( pk, some_value, lob_data ) SELECT TOP (1000) N.n, N.n, @Data FROM Numbers N WHERE N.n <= 1000; Now here’s a query to modify the cluster keys: UPDATE dbo.LOBtest SET pk = pk + 1; The query plan is: As you can see, the Split/Sort/Collapse optimization is present, and we also gain an Eager Table Spool, for Halloween protection. In addition, SQL Server now has no choice but to read the LOB data in the Clustered Index Scan: The performance is not great, as you might expect (even though there is no non-clustered index to maintain): Table 'LOBtest'. Scan count 1, logical reads 2011, physical reads 0, read-ahead reads 0, lob logical reads 36015, lob physical reads 0, lob read-ahead reads 15992. Table 'Worktable'. Scan count 1, logical reads 2040, physical reads 0, read-ahead reads 0, lob logical reads 34000, lob physical reads 0, lob read-ahead reads 8000. SQL Server Execution Times: CPU time = 483 ms, elapsed time = 17884 ms. Notice how the LOB data is read twice: once from the Clustered Index Scan, and again from the work table in tempdb used by the Eager Spool. If you try the same test with a non-unique clustered index (rather than a primary key), you’ll get a much more efficient plan that just passes the cluster key (including uniqueifier) around (no LOB data or other non-key columns): A unique non-clustered index (on a heap) works well too: Both those queries complete in a few tens of milliseconds, with no LOB reads, and just a few thousand logical reads. (In fact the heap is rather more efficient). There are lots more fun combinations to try that I don’t have space for here. Final Thoughts The behaviour shown in this post is not limited to LOB data by any means. If the conditions are met, any unique index that has included columns can produce similar behaviour – something to bear in mind when adding large INCLUDE columns to achieve covering queries, perhaps. Paul White Email: [email protected] Twitter: @PaulWhiteNZ

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Know more about Enqueue Deadlock Detection

- by Liu Maclean(???)

??? ORACLE ALLSTAR???????????????????,??????? ???????enqueue lock?????????3 ??????,????????????????????????????ora-00060 dead lock??process???3s: SQL> select * from v$version; BANNER ---------------------------------------------------------------- Oracle Database 10g Enterprise Edition Release 10.2.0.5.0 - 64bi PL/SQL Release 10.2.0.5.0 - Production CORE 10.2.0.5.0 Production TNS for Linux: Version 10.2.0.5.0 - Production NLSRTL Version 10.2.0.5.0 - Production SQL> select * from global_name; GLOBAL_NAME -------------------------------------------------------------------------------- www.oracledatabase12g.com PROCESS A: set timing on; update maclean1 set t1=t1+1; PROCESS B: update maclean2 set t1=t1+1; PROCESS A: update maclean2 set t1=t1+1; PROCESS B: update maclean1 set t1=t1+1; ??3s? PROCESS A ?? ERROR at line 1: ORA-00060: deadlock detected while waiting for resource Elapsed: 00:00:03.02 ????Process A????????????? 3s,?????????????,??????? ?????????? ???????: SQL> col name for a30 SQL> col value for a5 SQL> col DESCRIB for a50 SQL> set linesize 140 pagesize 1400 SQL> SELECT x.ksppinm NAME, y.ksppstvl VALUE, x.ksppdesc describ 2 FROM SYS.x$ksppi x, SYS.x$ksppcv y 3 WHERE x.inst_id = USERENV ('Instance') 4 AND y.inst_id = USERENV ('Instance') 5 AND x.indx = y.indx 6 AND x.ksppinm='_enqueue_deadlock_scan_secs'; NAME VALUE DESCRIB ------------------------------ ----- -------------------------------------------------- _enqueue_deadlock_scan_secs 0 deadlock scan interval SQL> alter system set "_enqueue_deadlock_scan_secs"=18 scope=spfile; System altered. Elapsed: 00:00:00.01 SQL> startup force; ORACLE instance started. Total System Global Area 851443712 bytes Fixed Size 2100040 bytes Variable Size 738198712 bytes Database Buffers 104857600 bytes Redo Buffers 6287360 bytes Database mounted. Database opened. PROCESS A: SQL> set timing on; SQL> update maclean1 set t1=t1+1; 1 row updated. Elapsed: 00:00:00.06 Process B SQL> update maclean2 set t1=t1+1; 1 row updated. SQL> update maclean1 set t1=t1+1; Process A: SQL> SQL> alter session set events '10704 trace name context forever,level 10:10046 trace name context forever,level 8'; Session altered. SQL> update maclean2 set t1=t1+1; update maclean2 set t1=t1+1 * ERROR at line 1: ORA-00060: deadlock detected while waiting for resource Elapsed: 00:00:18.05 ksqcmi: TX,90011,4a9 mode=6 timeout=21474836 WAIT #12: nam='enq: TX - row lock contention' ela= 2930070 name|mode=1415053318 usn<<16 | slot=589841 sequence=1193 obj#=56810 tim=1308114759849120 WAIT #12: nam='enq: TX - row lock contention' ela= 2930636 name|mode=1415053318 usn<<16 | slot=589841 sequence=1193 obj#=56810 tim=1308114762779801 WAIT #12: nam='enq: TX - row lock contention' ela= 2930439 name|mode=1415053318 usn<<16 | slot=589841 sequence=1193 obj#=56810 tim=1308114765710430 *** 2012-06-12 09:58:43.089 WAIT #12: nam='enq: TX - row lock contention' ela= 2931698 name|mode=1415053318 usn<<16 | slot=589841 sequence=1193 obj#=56810 tim=1308114768642192 WAIT #12: nam='enq: TX - row lock contention' ela= 2930428 name|mode=1415053318 usn<<16 | slot=589841 sequence=1193 obj#=56810 tim=1308114771572755 WAIT #12: nam='enq: TX - row lock contention' ela= 2931408 name|mode=1415053318 usn<<16 | slot=589841 sequence=1193 obj#=56810 tim=1308114774504207 DEADLOCK DETECTED ( ORA-00060 ) [Transaction Deadlock] The following deadlock is not an ORACLE error. It is a deadlock due to user error in the design of an application or from issuing incorrect ad-hoc SQL. The following information may aid in determining the deadlock: ??????Process A?’enq: TX – row lock contention’ ?????ORA-00060 deadlock detected????3s ??? 18s , ???hidden parameter “_enqueue_deadlock_scan_secs”?????,????????0? ??????????: SQL> alter system set "_enqueue_deadlock_scan_secs"=4 scope=spfile; System altered. Elapsed: 00:00:00.01 SQL> alter system set "_enqueue_deadlock_time_sec"=9 scope=spfile; System altered. Elapsed: 00:00:00.00 SQL> startup force; ORACLE instance started. Total System Global Area 851443712 bytes Fixed Size 2100040 bytes Variable Size 738198712 bytes Database Buffers 104857600 bytes Redo Buffers 6287360 bytes Database mounted. Database opened. SQL> set linesize 140 pagesize 1400 SQL> show parameter dead NAME TYPE VALUE ------------------------------------ -------------------------------- ------------------------------ _enqueue_deadlock_scan_secs integer 4 _enqueue_deadlock_time_sec integer 9 SQL> set timing on SQL> select * from maclean1 for update wait 8; T1 ---------- 11 Elapsed: 00:00:00.01 PROCESS B SQL> select * from maclean2 for update wait 8; T1 ---------- 3 SQL> select * from maclean1 for update wait 8; select * from maclean1 for update wait 8 PROCESS A SQL> select * from maclean2 for update wait 8; select * from maclean2 for update wait 8 * ERROR at line 1: ORA-30006: resource busy; acquire with WAIT timeout expired Elapsed: 00:00:08.00 ???????? ??? select for update wait?enqueue request timeout ?????8s? ,???????”_enqueue_deadlock_scan_secs”=4(deadlock scan interval),?4s???deadlock detected,????Process A????deadlock ???, ??????? ??Process A?????8s?raised??”ORA-30006: resource busy; acquire with WAIT timeout expired”??,??ORA-00060,?????process A???????? ????????”_enqueue_deadlock_time_sec”(requests with timeout <= this will not have deadlock detection)???,?enqueue request time < “_enqueue_deadlock_time_sec”?Server process?????dead lock detection,?????????enqueue request ??????timeout??????(_enqueue_deadlock_time_sec????5,?timeout<5s),???????????????;??????timeout>”_enqueue_deadlock_time_sec”???,Oracle????????????????????? ??????????: SQL> show parameter dead NAME TYPE VALUE ------------------------------------ -------------------------------- ------------------------------ _enqueue_deadlock_scan_secs integer 4 _enqueue_deadlock_time_sec integer 9 Process A: SQL> set timing on; SQL> select * from maclean1 for update wait 10; T1 ---------- 11 Process B: SQL> select * from maclean2 for update wait 10; T1 ---------- 3 SQL> select * from maclean1 for update wait 10; PROCESS A: SQL> select * from maclean2 for update wait 10; select * from maclean2 for update wait 10 * ERROR at line 1: ORA-00060: deadlock detected while waiting for resource Elapsed: 00:00:06.02 ??????? select for update wait 10?10s??, ?? 10s?????_enqueue_deadlock_time_sec???(9s),??Process A???????? ???????????????6s ???????_enqueue_deadlock_scan_secs?4s ? ???????????,???????????_enqueue_deadlock_scan_secs?????????3???? ??: enqueue lock?????????????? 1. ?????????deadlock detection??3s????, ????????_enqueue_deadlock_scan_secs(deadlock scan interval)???,??????0,????????_enqueue_deadlock_scan_secs?????????3???, ?_enqueue_deadlock_scan_secs=0 ??3s??, ?_enqueue_deadlock_scan_secs=4??6s??,????? 2. ???????_enqueue_deadlock_time_sec(requests with timeout <= this will not have deadlock detection)???,?enqueue request timeout< _enqueue_deadlock_time_sec(????5),?Server process?????????enqueue request timeout>_enqueue_deadlock_time_sec ????_enqueue_deadlock_scan_secs???????, ??request timeout??????select for update wait [TIMEOUT]??? ??: ???10.2.0.1?????????2?hidden parameter , ???patchset 10.2.0.3????? _enqueue_deadlock_time_sec, ?patchset 10.2.0.5??????_enqueue_deadlock_scan_secs? ?????RAC???????????10s, ???????_lm_dd_interval(dd time interval in seconds) ,????????8.0.6???? ???????????????,??????, ?10g???????60s,?11g???????10s? ???????11g??_lm_dd_interval?????????????,?????11g??LMD????????????,??????????RAC?LMD?Deadlock Detection???????CPU,???11g?Oracle????Team???LMD????????CPU????: ????????11g?LMD???????,???????11g??? UTS TRACE ????? DD???: SQL> select * from v$version; BANNER -------------------------------------------------------------------------------- Oracle Database 11g Enterprise Edition Release 11.2.0.3.0 - 64bit Production PL/SQL Release 11.2.0.3.0 - Production CORE 11.2.0.3.0 Production TNS for Linux: Version 11.2.0.3.0 - Production NLSRTL Version 11.2.0.3.0 - Production SQL> SQL> select * from global_name 2 ; GLOBAL_NAME -------------------------------------------------------------------------------- www.oracledatabase12g.com SQL> alter system set "_lm_dd_interval"=20 scope=spfile; System altered. SQL> startup force; ORACLE instance started. Total System Global Area 1570009088 bytes Fixed Size 2228704 bytes Variable Size 1325403680 bytes Database Buffers 234881024 bytes Redo Buffers 7495680 bytes Database mounted. Database opened. SQL> set linesize 140 pagesize 1400 SQL> show parameter lm_dd NAME TYPE VALUE ------------------------------------ -------------------------------- ------------------------------ _lm_dd_interval integer 20 SQL> select count(*) from gv$instance; COUNT(*) ---------- 2 instance 1: SQL> oradebug setorapid 12 Oracle pid: 12, Unix process pid: 8608, image: [email protected] (LMD0) ? LMD0??? UTS TRACE??RAC???????????? SQL> oradebug event 10046 trace name context forever,level 8:10708 trace name context forever,level 103: trace[rac.*] disk high; Statement processed. Elapsed: 00:00:00.00 SQL> update maclean1 set t1=t1+1; 1 row updated. instance 2: SQL> update maclean2 set t1=t1+1; 1 row updated. SQL> update maclean1 set t1=t1+1; Instance 1: SQL> update maclean2 set t1=t1+1; update maclean2 set t1=t1+1 * ERROR at line 1: ORA-00060: deadlock detected while waiting for resource Elapsed: 00:00:20.51 LMD0???UTS TRACE 2012-06-12 22:27:00.929284 : [kjmpbmsg:process][type 22][msg 0x7fa620ac85a8][from 1][seq 8148.0][len 192] 2012-06-12 22:27:00.929346 : [kjmxmpm][type 22][seq 0.0][msg 0x7fa620ac85a8][from 1] *** 2012-06-12 22:27:00.929 * kjddind: received DDIND msg with subtype x6 * reqp->dd_master_inst_kjxmddi == 1 * kjddind: dump sgh: 2012-06-12 22:27:00.929346*: kjddind: req->timestamp [0.15], kjddt [0.13] 2012-06-12 22:27:00.929346*: >> DDmsg:KJX_DD_REMOTE,TS[0.15],Inst 1->2,ddxid[id1,id2,inst:2097153,31,1],ddlock[0x95023930,829],ddMasterInst 1 2012-06-12 22:27:00.929346*: lock [0x95023930,829], op = [mast] 2012-06-12 22:27:00.929346*: reqp->timestamp [0.15], kjddt [0.13] 2012-06-12 22:27:00.929346*: kjddind: updated local timestamp [0.15] * kjddind: case KJX_DD_REMOTE 2012-06-12 22:27:00.929346*: ADD IO NODE WFG: 0 frame pointer 2012-06-12 22:27:00.929346*: PUSH: type=res, enqueue(0xffffffff.0xffffffff)=0xbbb9af40, block=KJUSEREX, snode=1 2012-06-12 22:27:00.929346*: PROCESS: type=res, enqueue(0xffffffff.0xffffffff)=0xbbb9af40, block=KJUSEREX, snode=1 2012-06-12 22:27:00.929346*: POP: type=res, enqueue(0xffffffff.0xffffffff)=0xbbb9af40, block=KJUSEREX, snode=1 2012-06-12 22:27:00.929346*: kjddopr[TX 0xe000c.0x32][ext 0x5,0x0]: blocking lock 0xbbb9a800, owner 2097154 of inst 2 2012-06-12 22:27:00.929346*: PUSH: type=txn, enqueue(0xffffffff.0xffffffff)=0xbbb9a800, block=KJUSEREX, snode=1 2012-06-12 22:27:00.929346*: PROCESS: type=txn, enqueue(0xffffffff.0xffffffff)=0xbbb9a800, block=KJUSEREX, snode=1 2012-06-12 22:27:00.929346*: ADD NODE TO WFG: type=txn, enqueue(0xffffffff.0xffffffff)=0xbbb9a800, block=KJUSEREX, snode=1 2012-06-12 22:27:00.929346*: POP: type=txn, enqueue(0xffffffff.0xffffffff)=0xbbb9a800, block=KJUSEREX, snode=1 2012-06-12 22:27:00.929346*: kjddopt: converting lock 0xbbce92f8 on 'TX' 0x80016.0x5d4,txid [2097154,34]of inst 2 2012-06-12 22:27:00.929346*: PUSH: type=res, enqueue(0xffffffff.0xffffffff)=0xbbce92f8, block=KJUSEREX, snode=1 2012-06-12 22:27:00.929346*: PROCESS: type=res, enqueue(0xffffffff.0xffffffff)=0xbbce92f8, block=KJUSEREX, snode=1 2012-06-12 22:27:00.929346*: ADD NODE TO WFG: type=res, enqueue(0xffffffff.0xffffffff)=0xbbce92f8, block=KJUSEREX, snode=1 2012-06-12 22:27:00.929855 : GSIPC:AMBUF: rcv buff 0x7fa620aa8cd8, pool rcvbuf, rqlen 1102 2012-06-12 22:27:00.929878 : GSIPC:GPBMSG: new bmsg 0x7fa620aa8d48 mb 0x7fa620aa8cd8 msg 0x7fa620aa8d68 mlen 192 dest x100 flushsz -1 2012-06-12 22:27:00.929878*: << DDmsg:KJX_DD_REMOTE,TS[0.15],Inst 2->1,ddxid[id1,id2,inst:2097153,31,1],ddlock[0x95023930,829],ddMasterInst 1 2012-06-12 22:27:00.929878*: lock [0xbbce92f8,287], op = [mast] 2012-06-12 22:27:00.929878*: ADD IO NODE WFG: 0 frame pointer 2012-06-12 22:27:00.929923 : [kjmpbmsg:compl][msg 0x7fa620ac8588][typ p][nmsgs 1][qtime 0][ptime 0] 2012-06-12 22:27:00.929947 : GSIPC:PBAT: flush start. flag 0x79 end 0 inc 4.4 2012-06-12 22:27:00.929963 : GSIPC:PBAT: send bmsg 0x7fa620aa8d48 blen 224 dest 1.0 2012-06-12 22:27:00.929979 : GSIPC:SNDQ: enq msg 0x7fa620aa8d48, type 65521 seq 8325, inst 1, receiver 0, queued 1 012-06-12 22:27:00.929979 : GSIPC:SNDQ: enq msg 0x7fa620aa8d48, type 65521 seq 8325, inst 1, receiver 0, queued 1 2012-06-12 22:27:00.929996 : GSIPC:BSEND: flushing sndq 0xb491dd28, id 0, dcx 0xbc517770, inst 1, rcvr 0 qlen 0 1 2012-06-12 22:27:00.930014 : GSIPC:BSEND: no batch1 msg 0x7fa620aa8d48 type 65521 len 224 dest (1:0) 2012-06-12 22:27:00.930088 : kjbsentscn[0x0.3f72dc][to 1] 2012-06-12 22:27:00.930144 : GSIPC:SENDM: send msg 0x7fa620aa8d48 dest x10000 seq 8325 type 65521 tkts x1 mlen xe00110 2012-06-12 22:27:00.930531 : GSIPC:KSXPCB: msg 0x7fa620aa8d48 status 30, type 65521, dest 1, rcvr 0 WAIT #0: nam='ges remote message' ela= 1372 waittime=80 loop=0 p3=74 obj#=-1 tim=1339554420931640 2012-06-12 22:27:00.931728 : GSIPC:RCVD: ksxp msg 0x7fa620af6490 sndr 1 seq 0.8149 type 65521 tkts 1 2012-06-12 22:27:00.931746 : GSIPC:RCVD: watq msg 0x7fa620af6490 sndr 1, seq 8149, type 65521, tkts 1 2012-06-12 22:27:00.931763 : GSIPC:RCVD: seq update (0.8148)->(0.8149) tp -15 fg 0x4 from 1 pbattr 0x0 2012-06-12 22:27:00.931779 : GSIPC:TKT: collect msg 0x7fa620af6490 from 1 for rcvr 0, tickets 1 2012-06-12 22:27:00.931794 : kjbrcvdscn[0x0.3f72dc][from 1][idx 2012-06-12 22:27:00.931810 : kjbrcvdscn[no bscn dd_master_inst_kjxmddi == 1 * kjddind: dump sgh: NXTIN (nil) 0 wq 0 cvtops x0 0x0.0x0(ext 0x0,0x0)[0000-0000-00000000] inst 1 BLOCKER 0xbbb9a800 5 wq 1 cvtops x28 TX 0xe000c.0x32(ext 0x5,0x0)[20000-0002-00000022] inst 2 BLOCKED 0xbbce92f8 5 wq 2 cvtops x1 TX 0x80016.0x5d4(ext 0x2,0x0)[20000-0002-00000022] inst 2 NXTOUT (nil) 0 wq 0 cvtops x0 0x0.0x0(ext 0x0,0x0)[0000-0000-00000000] inst 1 2012-06-12 22:27:00.932058*: kjddind: req->timestamp [0.15], kjddt [0.15] 2012-06-12 22:27:00.932058*: >> DDmsg:KJX_DD_VALIDATE,TS[0.15],Inst 1->2,ddxid[id1,id2,inst:2097153,31,1],ddlock[0x95023930,829],ddMasterInst 1 2012-06-12 22:27:00.932058*: lock [(nil),0], op = [vald_dd] 2012-06-12 22:27:00.932058*: kjddind: updated local timestamp [0.15] * kjddind: case KJX_DD_VALIDATE *** 2012-06-12 22:27:00.932 * kjddvald called: kjxmddi stuff: * cont_lockp (nil) * dd_lockp 0x95023930 * dd_inst 1 * dd_master_inst 1 * sgh graph: NXTIN (nil) 0 wq 0 cvtops x0 0x0.0x0(ext 0x0,0x0)[0000-0000-00000000] inst 1 BLOCKER 0xbbb9a800 5 wq 1 cvtops x28 TX 0xe000c.0x32(ext 0x5,0x0)[20000-0002-00000022] inst 2 BLOCKED 0xbbce92f8 5 wq 2 cvtops x1 TX 0x80016.0x5d4(ext 0x2,0x0)[20000-0002-00000022] inst 2 NXTOUT (nil) 0 wq 0 cvtops x0 0x0.0x0(ext 0x0,0x0)[0000-0000-00000000] inst 1 POP WFG NODE: lock=(nil) * kjddvald: dump the PRQ: BLOCKER 0xbbb9a800 5 wq 1 cvtops x28 TX 0xe000c.0x32(ext 0x5,0x0)[20000-0002-00000022] inst 2 BLOCKED 0xbbce92f8 5 wq 2 cvtops x1 TX 0x80016.0x5d4(ext 0x2,0x0)[20000-0002-00000022] inst 2 * kjddvald: KJDD_NXTONOD ->node_kjddsg.dinst_kjddnd =1 * kjddvald: ... which is not my node, my subgraph is validated but the cycle is not complete Global blockers dump start:--------------------------------- DUMP LOCAL BLOCKER/HOLDER: block level 5 res [0x80016][0x5d4],[TX][ext 0x2,0x0] ??dead lock!!! ???????11.2.0.3???? RAC LMD???????????”_lm_dd_interval”????????????20s? ???????10g?_lm_dd_interval???60s,??????Processes?????????????????,????????????Server Process????????60s??????11g?????(??????LMD???????)???????,???????????10s??? Enqueue Deadlock Detection? ?11g??? RAC?LMD???????hidden parameter ????”_lm_dd_interval”???,RAC????????????????,???????????: SQL> col name for a50 SQL> col describ for a60 SQL> col value for a20 SQL> set linesize 140 pagesize 1400 SQL> SELECT x.ksppinm NAME, y.ksppstvl VALUE, x.ksppdesc describ 2 FROM SYS.x$ksppi x, SYS.x$ksppcv y 3 WHERE x.inst_id = USERENV ('Instance') 4 AND y.inst_id = USERENV ('Instance') 5 AND x.indx = y.indx 6 AND x.ksppinm like '_lm_dd%'; NAME VALUE DESCRIB -------------------------------------------------- -------------------- ------------------------------------------------------------ _lm_dd_interval 20 dd time interval in seconds _lm_dd_scan_interval 5 dd scan interval in seconds _lm_dd_search_cnt 3 number of dd search per token get _lm_dd_max_search_time 180 max dd search time per token _lm_dd_maxdump 50 max number of locks to be dumped during dd validation _lm_dd_ignore_nodd FALSE if TRUE nodeadlockwait/nodeadlockblock options are ignored 6 rows selected.

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