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  • PHP split content when a HTML element is found

    - by sea_1987
    Hello, I have a PHP variable that holds some HTML I wanting to be able to split the variable into two pieces, and I want the spilt to take place when a second bold <strong> or <b> is found, essentially if I have content that looks like this, My content This is my content. Some more bold content, that would spilt into another variable. is this at all possible?

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  • Using Table-Valued Parameters With SQL Server Reporting Services

    - by Jesse
    In my last post I talked about using table-valued parameters to pass a list of integer values to a stored procedure without resorting to using comma-delimited strings and parsing out each value into a TABLE variable. In this post I’ll extend the “Customer Transaction Summary” report example to see how we might leverage this same stored procedure from within an SQL Server Reporting Services (SSRS) report. I’ve worked with SSRS off and on for the past several years and have generally found it to be a very useful tool for building nice-looking reports for end users quickly and easily. That said, I’ve been frustrated by SSRS from time to time when seemingly simple things are difficult to accomplish or simply not supported at all. I thought that using table-valued parameters from within a SSRS report would be simple, but unfortunately I was wrong. Customer Transaction Summary Example Let’s take the “Customer Transaction Summary” report example from the last post and try to plug that same stored procedure into an SSRS report. Our report will have three parameters: Start Date – beginning of the date range for which the report will summarize customer transactions End Date – end of the date range for which the report will summarize customer transactions Customer Ids – One or more customer Ids representing the customers that will be included in the report The simplest way to get started with this report will be to create a new dataset and point it at our Customer Transaction Summary report stored procedure (note that I’m using SSRS 2012 in the screenshots below, but there should be little to no difference with SSRS 2008): When you initially create this dataset the SSRS designer will try to invoke the stored procedure to determine what the parameters and output fields are for you automatically. As part of this process the following dialog pops-up: Obviously I can’t use this dialog to specify a value for the ‘@customerIds’ parameter since it is of the IntegerListTableType user-defined type that we created in the last post. Unfortunately this really throws the SSRS designer for a loop, and regardless of what combination of Data Type, Pass Null Value, or Parameter Value I used here, I kept getting this error dialog with the message, "Operand type clash: nvarchar is incompatible with IntegerListTableType". This error message makes some sense considering that the nvarchar type is indeed incompatible with the IntegerListTableType, but there’s little clue given as to how to remedy the situation. I don’t know for sure, but I think that behind-the-scenes the SSRS designer is trying to give the @customerIds parameter an nvarchar-typed SqlParameter which is causing the issue. When I first saw this error I figured that this might just be a limitation of the dataset designer and that I’d be able to work around the issue by manually defining the parameters. I know that there are some special steps that need to be taken when invoking a stored procedure with a table-valued parameter from ADO .NET, so I figured that I might be able to use some custom code embedded in the report  to create a SqlParameter instance with the needed properties and value to make this work, but the “Operand type clash" error message persisted. The Text Query Approach Just because we’re using a stored procedure to create the dataset for this report doesn’t mean that we can’t use the ‘Text’ Query Type option and construct an EXEC statement that will invoke the stored procedure. In order for this to work properly the EXEC statement will also need to declare and populate an IntegerListTableType variable to pass into the stored procedure. Before I go any further I want to make one point clear: this is a really ugly hack and it makes me cringe to do it. Simply put, I strongly feel that it should not be this difficult to use a table-valued parameter with SSRS. With that said, let’s take a look at what we’ll have to do to make this work. Manually Define Parameters First, we’ll need to manually define the parameters for report by right-clicking on the ‘Parameters’ folder in the ‘Report Data’ window. We’ll need to define the ‘@startDate’ and ‘@endDate’ as simple date parameters. We’ll also create a parameter called ‘@customerIds’ that will be a mutli-valued Integer parameter: In the ‘Available Values’ tab we’ll point this parameter at a simple dataset that just returns the CustomerId and CustomerName of each row in the Customers table of the database or manually define a handful of Customer Id values to make available when the report runs. Once we have these parameters properly defined we can take another crack at creating the dataset that will invoke the ‘rpt_CustomerTransactionSummary’ stored procedure. This time we’ll choose the ‘Text’ query type option and put the following into the ‘Query’ text area: 1: exec('declare @customerIdList IntegerListTableType ' + @customerIdInserts + 2: ' EXEC rpt_CustomerTransactionSummary 3: @startDate=''' + @startDate + ''', 4: @endDate='''+ @endDate + ''', 5: @customerIds=@customerIdList')   By using the ‘Text’ query type we can enter any arbitrary SQL that we we want to and then use parameters and string concatenation to inject pieces of that query at run time. It can be a bit tricky to parse this out at first glance, but from the SSRS designer’s point of view this query defines three parameters: @customerIdInserts – This will be a Text parameter that we use to define INSERT statements that will populate the @customerIdList variable that is being declared in the SQL. This parameter won’t actually ever get passed into the stored procedure. I’ll go into how this will work in a bit. @startDate – This is a simple date parameter that will get passed through directly into the @startDate parameter of the stored procedure on line 3. @endDate – This is another simple data parameter that will get passed through into the @endDate parameter of the stored procedure on line 4. At this point the dataset designer will be able to correctly parse the query and should even be able to detect the fields that the stored procedure will return without needing to specify any values for query when prompted to. Once the dataset has been correctly defined we’ll have a @customerIdInserts parameter listed in the ‘Parameters’ tab of the dataset designer. We need to define an expression for this parameter that will take the values selected by the user for the ‘@customerIds’ parameter that we defined earlier and convert them into INSERT statements that will populate the @customerIdList variable that we defined in our Text query. In order to do this we’ll need to add some custom code to our report using the ‘Report Properties’ dialog: Any custom code defined in the Report Properties dialog gets embedded into the .rdl of the report itself and (unfortunately) must be written in VB .NET. Note that you can also add references to custom .NET assemblies (which could be written in any language), but that’s outside the scope of this post so we’ll stick with the “quick and dirty” VB .NET approach for now. Here’s the VB .NET code (note that any embedded code that you add here must be defined in a static/shared function, though you can define as many functions as you want): 1: Public Shared Function BuildIntegerListInserts(ByVal variableName As String, ByVal paramValues As Object()) As String 2: Dim insertStatements As New System.Text.StringBuilder() 3: For Each paramValue As Object In paramValues 4: insertStatements.AppendLine(String.Format("INSERT {0} VALUES ({1})", variableName, paramValue)) 5: Next 6: Return insertStatements.ToString() 7: End Function   This method takes a variable name and an array of objects. We use an array of objects here because that is how SSRS will pass us the values that were selected by the user at run-time. The method uses a StringBuilder to construct INSERT statements that will insert each value from the object array into the provided variable name. Once this method has been defined in the custom code for the report we can go back into the dataset designer’s Parameters tab and update the expression for the ‘@customerIdInserts’ parameter by clicking on the button with the “function” symbol that appears to the right of the parameter value. We’ll set the expression to: 1: =Code.BuildIntegerListInserts("@customerIdList ", Parameters!customerIds.Value)   In order to invoke our custom code method we simply need to invoke “Code.<method name>” and pass in any needed parameters. The first parameter needs to match the name of the IntegerListTableType variable that we used in the EXEC statement of our query. The second parameter will come from the Value property of the ‘@customerIds’ parameter (this evaluates to an object array at run time). Finally, we’ll need to edit the properties of the ‘@customerIdInserts’ parameter on the report to mark it as a nullable internal parameter so that users aren’t prompted to provide a value for it when running the report. Limitations And Final Thoughts When I first started looking into the text query approach described above I wondered if there might be an upper limit to the size of the string that can be used to run a report. Obviously, the size of the actual query could increase pretty dramatically if you have a parameter that has a lot of potential values or you need to support several different table-valued parameters in the same query. I tested the example Customer Transaction Summary report with 1000 selected customers without any issue, but your mileage may vary depending on how much data you might need to pass into your query. If you think that the text query hack is a lot of work just to use a table-valued parameter, I agree! I think that it should be a lot easier than this to use a table-valued parameter from within SSRS, but so far I haven’t found a better way. It might be possible to create some custom .NET code that could build the EXEC statement for a given set of parameters automatically, but exploring that will have to wait for another post. For now, unless there’s a really compelling reason or requirement to use table-valued parameters from SSRS reports I would probably stick with the tried and true “join-multi-valued-parameter-to-CSV-and-split-in-the-query” approach for using mutli-valued parameters in a stored procedure.

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  • Developer Dashboard in SharePoint 2010

    - by jcortez
    Introducing the Developer Dashboard As a SharePoint developer (or IT Professional), how many times have you had the pleasure of figuring out why a particular page on your site is taking too long to render? I'm sure one of the techniques you have employed in troubleshooting is the process of elimination - removing individual web parts from the page hoping to identify which web part is misbehaving. One of the new features of SharePoint 2010 is the Developer Dashboard. This dashboard provides tracing and performance information that can be useful when you are trying to troubleshoot pages that are loading too slow. The Developer Dashboard is turned off by default and I'll go over 3 different ways to display it. Here is a screenshot of what the Developer Dashboard looks like when displayed at the bottom of the page:   You can see on the left side the different events that fired during the page processing pipeline and how long these events took. This is where you will see individual web parts being processed and how long it took to complete (obviously the kind of processing depends on what the web part does). On the right side you would see the different database calls issued through the SharePoint Object Model to process the page. You will notice that each of these database queries are actually a hyperlink and clicking on it displays a pop-up window that shows the actual SQL Query Text, the Call Stack that triggered the database call, and the IO statistics of that query. Enabling the Developer Dashboard Option 1: Managed Code   The Developer Dashboard is a farm-wide setting and the code above won't work if it is used within a web part hosted on any non-Central Admin site. The SPDeveloperDashboardLevel enum has three possible values: On, Off, and OnDemand. Setting it to On will always display the Developer Dashboard at the bottom of the page. Setting it Off will hide the Developer Dashboard. Setting it to OnDemand will add an icon at the top right corner of the page (see screenshot below) where a Site Collection Admin can toggle the display of the Developer Dashboard for a particular site collection. In my opinion, OnDemand is the best setting when troubleshooting a page or during development since a Site Collection Admin can turn it on or off and for a particular site only. The first cool thing about this is that the Site Collection Admin that turned it on will be the only one to see the Developer Dashboard output. Everyday users won't see the Developer Dashboard output even if it was turned on by a Site Collection Admin. If you need more flexibility on who gets to see the Developer Dashboard output, you can set the SPDeveloperDashboardSettings.RequiredPermissions to control which group of users will have the permission to see the output. Option 2: Using stsadm Using stsadm, you can run the following command to configure the Developer Dashboard: STSADM –o setproperty –pn developer-dashboard –pv OnDemand To successfully execute this command, be sure you that are running as a Farm Admin. Option 3: Using PowerShell For all scripts in SharePoint 2010, I prefer writing them as PowerShell scripts. Though the stsadm command is less verbose, the PowerShell equivalent is pretty straightforward and uses the SharePoint Object Model: You can of course parameterized the value that gets assigned to the DisplayLevel property so you can turn it On, Off or OnDemand depending on the parameter. Events and the Developer Dashboard  Now, don't assume that all the code inside your web part or page will show up in the Developer Dashboard complete with all the great troubleshooting information. Only a finite set of events are monitored by default (for a web part it will events in the base web part class). Let's say you have a click event that could take some time, for example a web service call. And you want to include troubleshooting information for this event in the Developer Dashboard. Enter SPMonitoredScope which is also a new feature in SharePoint 2010. In SharePoint 2010, everything is executed within a "Monitored Scope". And each scope has a set of "Monitors" that measures and counts calls and timings which appears in the Developer Dashboard. Below is an example on how to get your custom code to get included in the Developer Dashboard by wrapping it inside a new monitored scope: The code above would include your new scope "My long web service call" into the Developer Dashboard and would log the time it took to complete processing. In my opinion, wrapping your custom code in a SPMonitoredScope is a SharePoint development best practice since it provides you visibility and a better understanding on the performance of your components.

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  • Metro: Namespaces and Modules

    - by Stephen.Walther
    The goal of this blog entry is to describe how you can use the Windows JavaScript (WinJS) library to create namespaces. In particular, you learn how to use the WinJS.Namespace.define() and WinJS.Namespace.defineWithParent() methods. You also learn how to hide private methods by using the module pattern. Why Do We Need Namespaces? Before we do anything else, we should start by answering the question: Why do we need namespaces? What function do they serve? Do they just add needless complexity to our Metro applications? After all, plenty of JavaScript libraries do just fine without introducing support for namespaces. For example, jQuery has no support for namespaces and jQuery is the most popular JavaScript library in the universe. If jQuery can do without namespaces, why do we need to worry about namespaces at all? Namespaces perform two functions in a programming language. First, namespaces prevent naming collisions. In other words, namespaces enable you to create more than one object with the same name without conflict. For example, imagine that two companies – company A and company B – both want to make a JavaScript shopping cart control and both companies want to name the control ShoppingCart. By creating a CompanyA namespace and CompanyB namespace, both companies can create a ShoppingCart control: a CompanyA.ShoppingCart and a CompanyB.ShoppingCart control. The second function of a namespace is organization. Namespaces are used to group related functionality even when the functionality is defined in different physical files. For example, I know that all of the methods in the WinJS library related to working with classes can be found in the WinJS.Class namespace. Namespaces make it easier to understand the functionality available in a library. If you are building a simple JavaScript application then you won’t have much reason to care about namespaces. If you need to use multiple libraries written by different people then namespaces become very important. Using WinJS.Namespace.define() In the WinJS library, the most basic method of creating a namespace is to use the WinJS.Namespace.define() method. This method enables you to declare a namespace (of arbitrary depth). The WinJS.Namespace.define() method has the following parameters: · name – A string representing the name of the new namespace. You can add nested namespace by using dot notation · members – An optional collection of objects to add to the new namespace For example, the following code sample declares two new namespaces named CompanyA and CompanyB.Controls. Both namespaces contain a ShoppingCart object which has a checkout() method: // Create CompanyA namespace with ShoppingCart WinJS.Namespace.define("CompanyA"); CompanyA.ShoppingCart = { checkout: function (){ return "Checking out from A"; } }; // Create CompanyB.Controls namespace with ShoppingCart WinJS.Namespace.define( "CompanyB.Controls", { ShoppingCart: { checkout: function(){ return "Checking out from B"; } } } ); // Call CompanyA ShoppingCart checkout method console.log(CompanyA.ShoppingCart.checkout()); // Writes "Checking out from A" // Call CompanyB.Controls checkout method console.log(CompanyB.Controls.ShoppingCart.checkout()); // Writes "Checking out from B" In the code above, the CompanyA namespace is created by calling WinJS.Namespace.define(“CompanyA”). Next, the ShoppingCart is added to this namespace. The namespace is defined and an object is added to the namespace in separate lines of code. A different approach is taken in the case of the CompanyB.Controls namespace. The namespace is created and the ShoppingCart object is added to the namespace with the following single line of code: WinJS.Namespace.define( "CompanyB.Controls", { ShoppingCart: { checkout: function(){ return "Checking out from B"; } } } ); Notice that CompanyB.Controls is a nested namespace. The top level namespace CompanyB contains the namespace Controls. You can declare a nested namespace using dot notation and the WinJS library handles the details of creating one namespace within the other. After the namespaces have been defined, you can use either of the two shopping cart controls. You call CompanyA.ShoppingCart.checkout() or you can call CompanyB.Controls.ShoppingCart.checkout(). Using WinJS.Namespace.defineWithParent() The WinJS.Namespace.defineWithParent() method is similar to the WinJS.Namespace.define() method. Both methods enable you to define a new namespace. The difference is that the defineWithParent() method enables you to add a new namespace to an existing namespace. The WinJS.Namespace.defineWithParent() method has the following parameters: · parentNamespace – An object which represents a parent namespace · name – A string representing the new namespace to add to the parent namespace · members – An optional collection of objects to add to the new namespace The following code sample demonstrates how you can create a root namespace named CompanyA and add a Controls child namespace to the CompanyA parent namespace: WinJS.Namespace.define("CompanyA"); WinJS.Namespace.defineWithParent(CompanyA, "Controls", { ShoppingCart: { checkout: function () { return "Checking out"; } } } ); console.log(CompanyA.Controls.ShoppingCart.checkout()); // Writes "Checking out" One significant advantage of using the defineWithParent() method over the define() method is the defineWithParent() method is strongly-typed. In other words, you use an object to represent the base namespace instead of a string. If you misspell the name of the object (CompnyA) then you get a runtime error. Using the Module Pattern When you are building a JavaScript library, you want to be able to create both public and private methods. Some methods, the public methods, are intended to be used by consumers of your JavaScript library. The public methods act as your library’s public API. Other methods, the private methods, are not intended for public consumption. Instead, these methods are internal methods required to get the library to function. You don’t want people calling these internal methods because you might need to change them in the future. JavaScript does not support access modifiers. You can’t mark an object or method as public or private. Anyone gets to call any method and anyone gets to interact with any object. The only mechanism for encapsulating (hiding) methods and objects in JavaScript is to take advantage of functions. In JavaScript, a function determines variable scope. A JavaScript variable either has global scope – it is available everywhere – or it has function scope – it is available only within a function. If you want to hide an object or method then you need to place it within a function. For example, the following code contains a function named doSomething() which contains a nested function named doSomethingElse(): function doSomething() { console.log("doSomething"); function doSomethingElse() { console.log("doSomethingElse"); } } doSomething(); // Writes "doSomething" doSomethingElse(); // Throws ReferenceError You can call doSomethingElse() only within the doSomething() function. The doSomethingElse() function is encapsulated in the doSomething() function. The WinJS library takes advantage of function encapsulation to hide all of its internal methods. All of the WinJS methods are defined within self-executing anonymous functions. Everything is hidden by default. Public methods are exposed by explicitly adding the public methods to namespaces defined in the global scope. Imagine, for example, that I want a small library of utility methods. I want to create a method for calculating sales tax and a method for calculating the expected ship date of a product. The following library encapsulates the implementation of my library in a self-executing anonymous function: (function (global) { // Public method which calculates tax function calculateTax(price) { return calculateFederalTax(price) + calculateStateTax(price); } // Private method for calculating state tax function calculateStateTax(price) { return price * 0.08; } // Private method for calculating federal tax function calculateFederalTax(price) { return price * 0.02; } // Public method which returns the expected ship date function calculateShipDate(currentDate) { currentDate.setDate(currentDate.getDate() + 4); return currentDate; } // Export public methods WinJS.Namespace.define("CompanyA.Utilities", { calculateTax: calculateTax, calculateShipDate: calculateShipDate } ); })(this); // Show expected ship date var shipDate = CompanyA.Utilities.calculateShipDate(new Date()); console.log(shipDate); // Show price + tax var price = 12.33; var tax = CompanyA.Utilities.calculateTax(price); console.log(price + tax); In the code above, the self-executing anonymous function contains four functions: calculateTax(), calculateStateTax(), calculateFederalTax(), and calculateShipDate(). The following statement is used to expose only the calcuateTax() and the calculateShipDate() functions: // Export public methods WinJS.Namespace.define("CompanyA.Utilities", { calculateTax: calculateTax, calculateShipDate: calculateShipDate } ); Because the calculateTax() and calcuateShipDate() functions are added to the CompanyA.Utilities namespace, you can call these two methods outside of the self-executing function. These are the public methods of your library which form the public API. The calculateStateTax() and calculateFederalTax() methods, on the other hand, are forever hidden within the black hole of the self-executing function. These methods are encapsulated and can never be called outside of scope of the self-executing function. These are the internal methods of your library. Summary The goal of this blog entry was to describe why and how you use namespaces with the WinJS library. You learned how to define namespaces using both the WinJS.Namespace.define() and WinJS.Namespace.defineWithParent() methods. We also discussed how to hide private members and expose public members using the module pattern.

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  • The Incremental Architect&acute;s Napkin &ndash; #3 &ndash; Make Evolvability inevitable

    - by Ralf Westphal
    Originally posted on: http://geekswithblogs.net/theArchitectsNapkin/archive/2014/06/04/the-incremental-architectacutes-napkin-ndash-3-ndash-make-evolvability-inevitable.aspxThe easier something to measure the more likely it will be produced. Deviations between what is and what should be can be readily detected. That´s what automated acceptance tests are for. That´s what sprint reviews in Scrum are for. It´s no small wonder our software looks like it looks. It has all the traits whose conformance with requirements can easily be measured. And it´s lacking traits which cannot easily be measured. Evolvability (or Changeability) is such a trait. If an operation is correct, if an operation if fast enough, that can be checked very easily. But whether Evolvability is high or low, that cannot be checked by taking a measure or two. Evolvability might correlate with certain traits, e.g. number of lines of code (LOC) per function or Cyclomatic Complexity or test coverage. But there is no threshold value signalling “evolvability too low”; also Evolvability is hardly tangible for the customer. Nevertheless Evolvability is of great importance - at least in the long run. You can get away without much of it for a short time. Eventually, though, it´s needed like any other requirement. Or even more. Because without Evolvability no other requirement can be implemented. Evolvability is the foundation on which all else is build. Such fundamental importance is in stark contrast with its immeasurability. To compensate this, Evolvability must be put at the very center of software development. It must become the hub around everything else revolves. Since we cannot measure Evolvability, though, we cannot start watching it more. Instead we need to establish practices to keep it high (enough) at all times. Chefs have known that for long. That´s why everybody in a restaurant kitchen is constantly seeing after cleanliness. Hygiene is important as is to have clean tools at standardized locations. Only then the health of the patrons can be guaranteed and production efficiency is constantly high. Still a kitchen´s level of cleanliness is easier to measure than software Evolvability. That´s why important practices like reviews, pair programming, or TDD are not enough, I guess. What we need to keep Evolvability in focus and high is… to continually evolve. Change must not be something to avoid but too embrace. To me that means the whole change cycle from requirement analysis to delivery needs to be gone through more often. Scrum´s sprints of 4, 2 even 1 week are too long. Kanban´s flow of user stories across is too unreliable; it takes as long as it takes. Instead we should fix the cycle time at 2 days max. I call that Spinning. No increment must take longer than from this morning until tomorrow evening to finish. Then it should be acceptance checked by the customer (or his/her representative, e.g. a Product Owner). For me there are several resasons for such a fixed and short cycle time for each increment: Clear expectations Absolute estimates (“This will take X days to complete.”) are near impossible in software development as explained previously. Too much unplanned research and engineering work lurk in every feature. And then pervasive interruptions of work by peers and management. However, the smaller the scope the better our absolute estimates become. That´s because we understand better what really are the requirements and what the solution should look like. But maybe more importantly the shorter the timespan the more we can control how we use our time. So much can happen over the course of a week and longer timespans. But if push comes to shove I can block out all distractions and interruptions for a day or possibly two. That´s why I believe we can give rough absolute estimates on 3 levels: Noon Tonight Tomorrow Think of a meeting with a Product Owner at 8:30 in the morning. If she asks you, how long it will take you to implement a user story or bug fix, you can say, “It´ll be fixed by noon.”, or you can say, “I can manage to implement it until tonight before I leave.”, or you can say, “You´ll get it by tomorrow night at latest.” Yes, I believe all else would be naive. If you´re not confident to get something done by tomorrow night (some 34h from now) you just cannot reliably commit to any timeframe. That means you should not promise anything, you should not even start working on the issue. So when estimating use these four categories: Noon, Tonight, Tomorrow, NoClue - with NoClue meaning the requirement needs to be broken down further so each aspect can be assigned to one of the first three categories. If you like absolute estimates, here you go. But don´t do deep estimates. Don´t estimate dozens of issues; don´t think ahead (“Issue A is a Tonight, then B will be a Tomorrow, after that it´s C as a Noon, finally D is a Tonight - that´s what I´ll do this week.”). Just estimate so Work-in-Progress (WIP) is 1 for everybody - plus a small number of buffer issues. To be blunt: Yes, this makes promises impossible as to what a team will deliver in terms of scope at a certain date in the future. But it will give a Product Owner a clear picture of what to pull for acceptance feedback tonight and tomorrow. Trust through reliability Our trade is lacking trust. Customers don´t trust software companies/departments much. Managers don´t trust developers much. I find that perfectly understandable in the light of what we´re trying to accomplish: delivering software in the face of uncertainty by means of material good production. Customers as well as managers still expect software development to be close to production of houses or cars. But that´s a fundamental misunderstanding. Software development ist development. It´s basically research. As software developers we´re constantly executing experiments to find out what really provides value to users. We don´t know what they need, we just have mediated hypothesises. That´s why we cannot reliably deliver on preposterous demands. So trust is out of the window in no time. If we switch to delivering in short cycles, though, we can regain trust. Because estimates - explicit or implicit - up to 32 hours at most can be satisfied. I´d say: reliability over scope. It´s more important to reliably deliver what was promised then to cover a lot of requirement area. So when in doubt promise less - but deliver without delay. Deliver on scope (Functionality and Quality); but also deliver on Evolvability, i.e. on inner quality according to accepted principles. Always. Trust will be the reward. Less complexity of communication will follow. More goodwill buffer will follow. So don´t wait for some Kanban board to show you, that flow can be improved by scheduling smaller stories. You don´t need to learn that the hard way. Just start with small batch sizes of three different sizes. Fast feedback What has been finished can be checked for acceptance. Why wait for a sprint of several weeks to end? Why let the mental model of the issue and its solution dissipate? If you get final feedback after one or two weeks, you hardly remember what you did and why you did it. Resoning becomes hard. But more importantly youo probably are not in the mood anymore to go back to something you deemed done a long time ago. It´s boring, it´s frustrating to open up that mental box again. Learning is harder the longer it takes from event to feedback. Effort can be wasted between event (finishing an issue) and feedback, because other work might go in the wrong direction based on false premises. Checking finished issues for acceptance is the most important task of a Product Owner. It´s even more important than planning new issues. Because as long as work started is not released (accepted) it´s potential waste. So before starting new work better make sure work already done has value. By putting the emphasis on acceptance rather than planning true pull is established. As long as planning and starting work is more important, it´s a push process. Accept a Noon issue on the same day before leaving. Accept a Tonight issue before leaving today or first thing tomorrow morning. Accept a Tomorrow issue tomorrow night before leaving or early the day after tomorrow. After acceptance the developer(s) can start working on the next issue. Flexibility As if reliability/trust and fast feedback for less waste weren´t enough economic incentive, there is flexibility. After each issue the Product Owner can change course. If on Monday morning feature slices A, B, C, D, E were important and A, B, C were scheduled for acceptance by Monday evening and Tuesday evening, the Product Owner can change her mind at any time. Maybe after A got accepted she asks for continuation with D. But maybe, just maybe, she has gotten a completely different idea by then. Maybe she wants work to continue on F. And after B it´s neither D nor E, but G. And after G it´s D. With Spinning every 32 hours at latest priorities can be changed. And nothing is lost. Because what got accepted is of value. It provides an incremental value to the customer/user. Or it provides internal value to the Product Owner as increased knowledge/decreased uncertainty. I find such reactivity over commitment economically very benefical. Why commit a team to some workload for several weeks? It´s unnecessary at beast, and inflexible and wasteful at worst. If we cannot promise delivery of a certain scope on a certain date - which is what customers/management usually want -, we can at least provide them with unpredecented flexibility in the face of high uncertainty. Where the path is not clear, cannot be clear, make small steps so you´re able to change your course at any time. Premature completion Customers/management are used to premeditating budgets. They want to know exactly how much to pay for a certain amount of requirements. That´s understandable. But it does not match with the nature of software development. We should know that by now. Maybe there´s somewhere in the world some team who can consistently deliver on scope, quality, and time, and budget. Great! Congratulations! I, however, haven´t seen such a team yet. Which does not mean it´s impossible, but I think it´s nothing I can recommend to strive for. Rather I´d say: Don´t try this at home. It might hurt you one way or the other. However, what we can do, is allow customers/management stop work on features at any moment. With spinning every 32 hours a feature can be declared as finished - even though it might not be completed according to initial definition. I think, progress over completion is an important offer software development can make. Why think in terms of completion beyond a promise for the next 32 hours? Isn´t it more important to constantly move forward? Step by step. We´re not running sprints, we´re not running marathons, not even ultra-marathons. We´re in the sport of running forever. That makes it futile to stare at the finishing line. The very concept of a burn-down chart is misleading (in most cases). Whoever can only think in terms of completed requirements shuts out the chance for saving money. The requirements for a features mostly are uncertain. So how does a Product Owner know in the first place, how much is needed. Maybe more than specified is needed - which gets uncovered step by step with each finished increment. Maybe less than specified is needed. After each 4–32 hour increment the Product Owner can do an experient (or invite users to an experiment) if a particular trait of the software system is already good enough. And if so, she can switch the attention to a different aspect. In the end, requirements A, B, C then could be finished just 70%, 80%, and 50%. What the heck? It´s good enough - for now. 33% money saved. Wouldn´t that be splendid? Isn´t that a stunning argument for any budget-sensitive customer? You can save money and still get what you need? Pull on practices So far, in addition to more trust, more flexibility, less money spent, Spinning led to “doing less” which also means less code which of course means higher Evolvability per se. Last but not least, though, I think Spinning´s short acceptance cycles have one more effect. They excert pull-power on all sorts of practices known for increasing Evolvability. If, for example, you believe high automated test coverage helps Evolvability by lowering the fear of inadverted damage to a code base, why isn´t 90% of the developer community practicing automated tests consistently? I think, the answer is simple: Because they can do without. Somehow they manage to do enough manual checks before their rare releases/acceptance checks to ensure good enough correctness - at least in the short term. The same goes for other practices like component orientation, continuous build/integration, code reviews etc. None of that is compelling, urgent, imperative. Something else always seems more important. So Evolvability principles and practices fall through the cracks most of the time - until a project hits a wall. Then everybody becomes desperate; but by then (re)gaining Evolvability has become as very, very difficult and tedious undertaking. Sometimes up to the point where the existence of a project/company is in danger. With Spinning that´s different. If you´re practicing Spinning you cannot avoid all those practices. With Spinning you very quickly realize you cannot deliver reliably even on your 32 hour promises. Spinning thus is pulling on developers to adopt principles and practices for Evolvability. They will start actively looking for ways to keep their delivery rate high. And if not, management will soon tell them to do that. Because first the Product Owner then management will notice an increasing difficulty to deliver value within 32 hours. There, finally there emerges a way to measure Evolvability: The more frequent developers tell the Product Owner there is no way to deliver anything worth of feedback until tomorrow night, the poorer Evolvability is. Don´t count the “WTF!”, count the “No way!” utterances. In closing For sustainable software development we need to put Evolvability first. Functionality and Quality must not rule software development but be implemented within a framework ensuring (enough) Evolvability. Since Evolvability cannot be measured easily, I think we need to put software development “under pressure”. Software needs to be changed more often, in smaller increments. Each increment being relevant to the customer/user in some way. That does not mean each increment is worthy of shipment. It´s sufficient to gain further insight from it. Increments primarily serve the reduction of uncertainty, not sales. Sales even needs to be decoupled from this incremental progress. No more promises to sales. No more delivery au point. Rather sales should look at a stream of accepted increments (or incremental releases) and scoup from that whatever they find valuable. Sales and marketing need to realize they should work on what´s there, not what might be possible in the future. But I digress… In my view a Spinning cycle - which is not easy to reach, which requires practice - is the core practice to compensate the immeasurability of Evolvability. From start to finish of each issue in 32 hours max - that´s the challenge we need to accept if we´re serious increasing Evolvability. Fortunately higher Evolvability is not the only outcome of Spinning. Customer/management will like the increased flexibility and “getting more bang for the buck”.

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  • JMS Step 5 - How to Create an 11g BPEL Process Which Reads a Message Based on an XML Schema from a JMS Queue

    - by John-Brown.Evans
    JMS Step 5 - How to Create an 11g BPEL Process Which Reads a Message Based on an XML Schema from a JMS Queue .jblist{list-style-type:disc;margin:0;padding:0;padding-left:0pt;margin-left:36pt} ol{margin:0;padding:0} .c12_5{vertical-align:top;width:468pt;border-style:solid;background-color:#f3f3f3;border-color:#000000;border-width:1pt;padding:5pt 5pt 5pt 5pt} .c8_5{vertical-align:top;border-style:solid;border-color:#000000;border-width:1pt;padding:5pt 5pt 0pt 5pt} .c10_5{vertical-align:top;width:207pt;border-style:solid;border-color:#000000;border-width:1pt;padding:5pt 5pt 5pt 5pt} .c14_5{vertical-align:top;border-style:solid;border-color:#000000;border-width:1pt;padding:0pt 5pt 0pt 5pt} .c21_5{background-color:#ffffff} .c18_5{color:#1155cc;text-decoration:underline} .c16_5{color:#666666;font-size:12pt} .c5_5{background-color:#f3f3f3;font-weight:bold} .c19_5{color:inherit;text-decoration:inherit} .c3_5{height:11pt;text-align:center} .c11_5{font-weight:bold} .c20_5{background-color:#00ff00} .c6_5{font-style:italic} .c4_5{height:11pt} .c17_5{background-color:#ffff00} .c0_5{direction:ltr} .c7_5{font-family:"Courier New"} .c2_5{border-collapse:collapse} .c1_5{line-height:1.0} .c13_5{background-color:#f3f3f3} .c15_5{height:0pt} .c9_5{text-align:center} .title{padding-top:24pt;line-height:1.15;text-align:left;color:#000000;font-size:36pt;font-family:"Arial";font-weight:bold;padding-bottom:6pt} .subtitle{padding-top:18pt;line-height:1.15;text-align:left;color:#666666;font-style:italic;font-size:24pt;font-family:"Georgia";padding-bottom:4pt} li{color:#000000;font-size:10pt;font-family:"Arial"} p{color:#000000;font-size:10pt;margin:0;font-family:"Arial"} h1{padding-top:0pt;line-height:1.15;text-align:left;color:#888;font-size:24pt;font-family:"Arial";font-weight:normal} h2{padding-top:0pt;line-height:1.15;text-align:left;color:#888;font-size:18pt;font-family:"Arial";font-weight:normal} h3{padding-top:0pt;line-height:1.15;text-align:left;color:#888;font-size:14pt;font-family:"Arial";font-weight:normal} h4{padding-top:0pt;line-height:1.15;text-align:left;color:#888;font-size:12pt;font-family:"Arial";font-weight:normal} h5{padding-top:0pt;line-height:1.15;text-align:left;color:#888;font-size:11pt;font-family:"Arial";font-weight:normal} h6{padding-top:0pt;line-height:1.15;text-align:left;color:#888;font-size:10pt;font-family:"Arial";font-weight:normal} Welcome to another post in the series of blogs which demonstrates how to use JMS queues in a SOA context. The previous posts were: JMS Step 1 - How to Create a Simple JMS Queue in Weblogic Server 11g JMS Step 2 - Using the QueueSend.java Sample Program to Send a Message to a JMS Queue JMS Step 3 - Using the QueueReceive.java Sample Program to Read a Message from a JMS Queue JMS Step 4 - How to Create an 11g BPEL Process Which Writes a Message Based on an XML Schema to a JMS Queue Today we will create a BPEL process which will read (dequeue) the message from the JMS queue, which we enqueued in the last example. The JMS adapter will dequeue the full XML payload from the queue. 1. Recap and Prerequisites In the previous examples, we created a JMS Queue, a Connection Factory and a Connection Pool in the WebLogic Server Console. Then we designed and deployed a BPEL composite, which took a simple XML payload and enqueued it to the JMS queue. In this example, we will read that same message from the queue, using a JMS adapter and a BPEL process. As many of the configuration steps required to read from that queue were done in the previous samples, this one will concentrate on the new steps. A summary of the required objects is listed below. To find out how to create them please see the previous samples. They also include instructions on how to verify the objects are set up correctly. WebLogic Server Objects Object Name Type JNDI Name TestConnectionFactory Connection Factory jms/TestConnectionFactory TestJMSQueue JMS Queue jms/TestJMSQueue eis/wls/TestQueue Connection Pool eis/wls/TestQueue Schema XSD File The following XSD file is used for the message format. It was created in the previous example and will be copied to the new process. stringPayload.xsd <?xml version="1.0" encoding="windows-1252" ?> <xsd:schema xmlns:xsd="http://www.w3.org/2001/XMLSchema"                 xmlns="http://www.example.org"                 targetNamespace="http://www.example.org"                 elementFormDefault="qualified">   <xsd:element name="exampleElement" type="xsd:string">   </xsd:element> </xsd:schema> JMS Message After executing the previous samples, the following XML message should be in the JMS queue located at jms/TestJMSQueue: <?xml version="1.0" encoding="UTF-8" ?><exampleElement xmlns="http://www.example.org">Test Message</exampleElement> JDeveloper Connection You will need a valid Application Server Connection in JDeveloper pointing to the SOA server which the process will be deployed to. 2. Create a BPEL Composite with a JMS Adapter Partner Link In the previous example, we created a composite in JDeveloper called JmsAdapterWriteSchema. In this one, we will create a new composite called JmsAdapterReadSchema. There are probably many ways of incorporating a JMS adapter into a SOA composite for incoming messages. One way is design the process in such a way that the adapter polls for new messages and when it dequeues one, initiates a SOA or BPEL instance. This is possibly the most common use case. Other use cases include mid-flow adapters, which are activated from within the BPEL process. In this example we will use a polling adapter, because it is the most simple to set up and demonstrate. But it has one disadvantage as a demonstrative model. When a polling adapter is active, it will dequeue all messages as soon as they reach the queue. This makes it difficult to monitor messages we are writing to the queue, because they will disappear from the queue as soon as they have been enqueued. To work around this, we will shut down the composite after deploying it and restart it as required. (Another solution for this would be to pause the consumption for the queue and resume consumption again if needed. This can be done in the WLS console JMS-Modules -> queue -> Control -> Consumption -> Pause/Resume.) We will model the composite as a one-way incoming process. Usually, a BPEL process will do something useful with the message after receiving it, such as passing it to a database or file adapter, a human workflow or external web service. But we only want to demonstrate how to dequeue a JMS message using BPEL and a JMS adapter, so we won’t complicate the design with further activities. However, we do want to be able to verify that we have read the message correctly, so the BPEL process will include a small piece of embedded java code, which will print the message to standard output, so we can view it in the SOA server’s log file. Alternatively, you can view the instance in the Enterprise Manager and verify the message. The following steps are all executed in JDeveloper. Create the project in the same JDeveloper application used for the previous examples or create a new one. Create a SOA Project Create a new project and choose SOA Tier > SOA Project as its type. Name it JmsAdapterReadSchema. When prompted for the composite type, choose Empty Composite. Create a JMS Adapter Partner Link In the composite editor, drag a JMS adapter over from the Component Palette to the left-hand swim lane, under Exposed Services. This will start the JMS Adapter Configuration Wizard. Use the following entries: Service Name: JmsAdapterRead Oracle Enterprise Messaging Service (OEMS): Oracle WebLogic JMS AppServer Connection: Use an application server connection pointing to the WebLogic server on which the JMS queue and connection factory mentioned under Prerequisites above are located. Adapter Interface > Interface: Define from operation and schema (specified later) Operation Type: Consume Message Operation Name: Consume_message Consume Operation Parameters Destination Name: Press the Browse button, select Destination Type: Queues, then press Search. Wait for the list to populate, then select the entry for TestJMSQueue , which is the queue created in a previous example. JNDI Name: The JNDI name to use for the JMS connection. As in the previous example, this is probably the most common source of error. This is the JNDI name of the JMS adapter’s connection pool created in the WebLogic Server and which points to the connection factory. JDeveloper does not verify the value entered here. If you enter a wrong value, the JMS adapter won’t find the queue and you will get an error message at runtime, which is very difficult to trace. In our example, this is the value eis/wls/TestQueue . (See the earlier step on how to create a JMS Adapter Connection Pool in WebLogic Server for details.) Messages/Message SchemaURL: We will use the XSD file created during the previous example, in the JmsAdapterWriteSchema project to define the format for the incoming message payload and, at the same time, demonstrate how to import an existing XSD file into a JDeveloper project. Press the magnifying glass icon to search for schema files. In the Type Chooser, press the Import Schema File button. Select the magnifying glass next to URL to search for schema files. Navigate to the location of the JmsAdapterWriteSchema project > xsd and select the stringPayload.xsd file. Check the “Copy to Project” checkbox, press OK and confirm the following Localize Files popup. Now that the XSD file has been copied to the local project, it can be selected from the project’s schema files. Expand Project Schema Files > stringPayload.xsd and select exampleElement: string . Press Next and Finish, which will complete the JMS Adapter configuration.Save the project. Create a BPEL Component Drag a BPEL Process from the Component Palette (Service Components) to the Components section of the composite designer. Name it JmsAdapterReadSchema and select Template: Define Service Later and press OK. Wire the JMS Adapter to the BPEL Component Now wire the JMS adapter to the BPEL process, by dragging the arrow from the adapter to the BPEL process. A Transaction Properties popup will be displayed. Set the delivery mode to async.persist. This completes the steps at the composite level. 3 . Complete the BPEL Process Design Invoke the BPEL Flow via the JMS Adapter Open the BPEL component by double-clicking it in the design view of the composite.xml, or open it from the project navigator by selecting the JmsAdapterReadSchema.bpel file. This will display the BPEL process in the design view. You should see the JmsAdapterRead partner link in the left-hand swim lane. Drag a Receive activity onto the BPEL flow diagram, then drag a wire (left-hand yellow arrow) from it to the JMS adapter. This will open the Receive activity editor. Auto-generate the variable by pressing the green “+” button and check the “Create Instance” checkbox. This will result in a BPEL instance being created when a new JMS message is received. At this point it would actually be OK to compile and deploy the composite and it would pick up any messages from the JMS queue. In fact, you can do that to test it, if you like. But it is very rudimentary and would not be doing anything useful with the message. Also, you could only verify the actual message payload by looking at the instance’s flow in the Enterprise Manager. There are various other possibilities; we could pass the message to another web service, write it to a file using a file adapter or to a database via a database adapter etc. But these will all introduce unnecessary complications to our sample. So, to keep it simple, we will add a small piece of Java code to the BPEL process which will write the payload to standard output. This will be written to the server’s log file, which will be easy to monitor. Add a Java Embedding Activity First get the full name of the process’s input variable, as this will be needed for the Java code. Go to the Structure pane and expand Variables > Process > Variables. Then expand the input variable, for example, "Receive1_Consume_Message_InputVariable > body > ns2:exampleElement”, and note variable’s name and path, if they are different from this one. Drag a Java Embedding activity from the Component Palette (Oracle Extensions) to the BPEL flow, after the Receive activity, then open it to edit. Delete the example code and replace it with the following, replacing the variable parts with those in your sample, if necessary.: System.out.println("JmsAdapterReadSchema process picked up a message"); oracle.xml.parser.v2.XMLElement inputPayload =    (oracle.xml.parser.v2.XMLElement)getVariableData(                           "Receive1_Consume_Message_InputVariable",                           "body",                           "/ns2:exampleElement");   String inputString = inputPayload.getFirstChild().getNodeValue(); System.out.println("Input String is " + inputPayload.getFirstChild().getNodeValue()); Tip. If you are not sure of the exact syntax of the input variable, create an Assign activity in the BPEL process and copy the variable to another, temporary one. Then check the syntax created by the BPEL designer. This completes the BPEL process design in JDeveloper. Save, compile and deploy the process to the SOA server. 3. Test the Composite Shut Down the JmsAdapterReadSchema Composite After deploying the JmsAdapterReadSchema composite to the SOA server it is automatically activated. If there are already any messages in the queue, the adapter will begin polling them. To ease the testing process, we will deactivate the process first Log in to the Enterprise Manager (Fusion Middleware Control) and navigate to SOA > soa-infra (soa_server1) > default (or wherever you deployed your composite to) and click on JmsAdapterReadSchema [1.0] . Press the Shut Down button to disable the composite and confirm the following popup. Monitor Messages in the JMS Queue In a separate browser window, log in to the WebLogic Server Console and navigate to Services > Messaging > JMS Modules > TestJMSModule > TestJMSQueue > Monitoring. This is the location of the JMS queue we created in an earlier sample (see the prerequisites section of this sample). Check whether there are any messages already in the queue. If so, you can dequeue them using the QueueReceive Java program created in an earlier sample. This will ensure that the queue is empty and doesn’t contain any messages in the wrong format, which would cause the JmsAdapterReadSchema to fail. Send a Test Message In the Enterprise Manager, navigate to the JmsAdapterWriteSchema created earlier, press Test and send a test message, for example “Message from JmsAdapterWriteSchema”. Confirm that the message was written correctly to the queue by verifying it via the queue monitor in the WLS Console. Monitor the SOA Server’s Output A program deployed on the SOA server will write its standard output to the terminal window in which the server was started, unless this has been redirected to somewhere else, for example to a file. If it has not been redirected, go to the terminal session in which the server was started, otherwise open and monitor the file to which it was redirected. Re-Enable the JmsAdapterReadSchema Composite In the Enterprise Manager, navigate to the JmsAdapterReadSchema composite again and press Start Up to re-enable it. This should cause the JMS adapter to dequeue the test message and the following output should be written to the server’s standard output: JmsAdapterReadSchema process picked up a message. Input String is Message from JmsAdapterWriteSchema Note that you can also monitor the payload received by the process, by navigating to the the JmsAdapterReadSchema’s Instances tab in the Enterprise Manager. Then select the latest instance and view the flow of the BPEL component. The Receive activity will contain and display the dequeued message too. 4 . Troubleshooting This sample demonstrates how to dequeue an XML JMS message using a BPEL process and no additional functionality. For example, it doesn’t contain any error handling. Therefore, any errors in the payload will result in exceptions being written to the log file or standard output. If you get any errors related to the payload, such as Message handle error ... ORABPEL-09500 ... XPath expression failed to execute. An error occurs while processing the XPath expression; the expression is /ns2:exampleElement. ... etc. check that the variable used in the Java embedding part of the process was entered correctly. Possibly follow the tip mentioned in previous section. If this doesn’t help, you can delete the Java embedding part and simply verify the message via the flow diagram in the Enterprise Manager. Or use a different method, such as writing it to a file via a file adapter. This concludes this example. In the next post, we will begin with an AQ JMS example, which uses JMS to write to an Advanced Queue stored in the database. Best regards John-Brown Evans Oracle Technology Proactive Support Delivery

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  • SQL SERVER – Quiz and Video – Introduction to SQL Error Actions

    - by pinaldave
    This blog post is inspired from SQL Programming Joes 2 Pros: Programming and Development for Microsoft SQL Server 2008 – SQL Exam Prep Series 70-433 – Volume 4. [Amazon] | [Flipkart] | [Kindle] | [IndiaPlaza] This is follow up blog post of my earlier blog post on the same subject - SQL SERVER – Introduction to SQL Error Actions – A Primer. In the article we discussed various basics terminology of the error handling. The article further covers following important concepts of error handling. Introduction to SQL Error Actions Statement Termination Scope Abortion Batch Termination Above three are the most important concepts related to error handling and SQL Server.  There are many more things one has to learn but without beginners fundamentals one can’t learn the advanced concepts. Let us have small quiz and check how many of you get the fundamentals right. Quiz 1.) Which SQL Server error action happens for errors with a severity of 11-16 when you set the XACT_ABORT setting to ON? You will get Statement Termination. You will get Scope Abortion. You will get Batch Abortion. You will get Connection Termination. SQL Server will pick the error action. 2.) Which SQL Server error action happens for errors with a severity of 11-16 when you set the XACT_ABORT setting to OFF? You will get Statement Termination You will get Scope Abortion You will get Batch Abortion You will get Connection Termination SQL Server will pick the error action Now make sure that you write down all the answers on the piece of paper. Watch following video and read earlier article over here. If you want to change the answer you still have chance. Solution 1) 3 2) 5 Now compare let us check the answers and compare your answers to following answers. I am very confident you will get them correct. Available at USA: Amazon India: Flipkart | IndiaPlaza Volume: 1, 2, 3, 4, 5 Please leave your feedback in the comment area for the quiz and video. Did you know all the answers of the quiz? Reference: Pinal Dave (http://blog.sqlauthority.com) Filed under: Joes 2 Pros, PostADay, SQL, SQL Authority, SQL Query, SQL Server, SQL Tips and Tricks, T SQL, Technology

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  • Two Wifi Icons in Panel [Solved]

    - by Alex
    I have the exact problem in 13.10 as this user Two Wifi indicators in panel. Here are some screenshots: Here are some screenshots from another user: http://ubuntuforums.org/showthread.php?t=2183020&p=12825563 ifconfig and iwconfig outputs $ ifconfig lo Link encap:Local Loopback inet addr:XXXXXX Mask:XXXXXXX inet6 addr: ::1/128 Scope:Host UP LOOPBACK RUNNING MTU:65536 Metric:1 RX packets:2243 errors:0 dropped:0 overruns:0 frame:0 TX packets:2243 errors:0 dropped:0 overruns:0 carrier:0 collisions:0 txqueuelen:0 RX bytes:209889 (209.8 KB) TX bytes:209889 (209.8 KB) wlan0 Link encap:Ethernet HWaddr XXXXXXXXX inet addr:XXXXXX Bcast:XXXXXXXX Mask:XXXXXXX inet6 addr: XXXXXXX Scope:Link UP BROADCAST RUNNING MULTICAST MTU:1500 Metric:1 RX packets:5925 errors:0 dropped:0 overruns:0 frame:0 TX packets:3361 errors:0 dropped:0 overruns:0 carrier:0 collisions:0 txqueuelen:1000 RX bytes:2951818 (2.9 MB) TX bytes:630579 (630.5 KB) $ iwconfig lo no wireless extensions. wlan0 IEEE 802.11abgn ESSID:"XXXXX" Mode:Managed Frequency:2.437 GHz Access Point: XXXXXXXX Bit Rate=72.2 Mb/s Tx-Power=15 dBm Retry long limit:7 RTS thr:off Fragment thr:off Power Management:on Link Quality=49/70 Signal level=-61 dBm Rx invalid nwid:0 Rx invalid crypt:0 Rx invalid frag:0 Tx excessive retries:153 Invalid misc:472 Missed beacon:0

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  • Configure IPv6 on your Linux system (Ubuntu)

    After the presentation on IPv6 at the first event of the Emtel Knowledge Series and some recent discussion on social media networks with other geeks and Linux interested IT people here in Mauritius, I thought that I should give it a try (finally) and tweak my local network infrastructure. Honestly, I have been to busy with contractual project work and it never really occurred to me to set up IPv6 in my LAN. Well, the following paragraphs are going to shed some light on those aspects of modern computer and network technology. This is the first article in a series on IPv6 configuration: Configure IPv6 on your Linux system DHCPv6: Provide IPv6 information in your local network Enabling DNS for IPv6 infrastructure Accessing your web server via IPv6 Piece of advice: This is based on my findings on the internet while reading other people's helpful articles and going through a couple of man-pages on my local system. Let's embrace IPv6 The basic configuration on Linux is actually very simple as the kernel, operating system, and user-space programs support that protocol natively. If your system is ready to go for IP (aka: IPv4), then you are good to go for anything else. At least, I didn't have to install any additional packages on my system(s). We are going to assign a static IPv6 address to the system. Hence, we have to modify the definition of interfaces and check whether we have an inet6 entry specified. Open your favourite text editor and check the following entries (it should be at least similar to this): $ sudo nano /etc/network/interfaces auto eth0# IPv4 configurationiface eth0 inet static  address 192.168.1.2  network 192.168.1.0  netmask 255.255.255.0  broadcast 192.168.1.255# IPv6 configurationiface eth0 inet6 static  pre-up modprobe ipv6  address 2001:db8:bad:a55::2  netmask 64 Of course, you might have to adjust your interface device (eth0) or you might be interested to have multiple directives for additional devices (eth1, eth2, etc.). The auto instruction takes care that your device is enabled and configured during the booting phase. The use of the pre-up directive depends on your kernel configuration but in most scenarios this might be an optional line. Anyways, it doesn't hurt to have it enabled after all - just to be on the safe side. Next, either restart your network subsystem like so: $ sudo service networking restart Or you might prefer to do it manually with identical parameters, like so: $ sudo ifconfig eth0 inet6 add 2001:db8:bad:a55::2/64 In case that you're logged in remotely into your PC (ie. via ssh), it is highly advised to opt for the second choice and add the device manually. You can check your configuration afterwards with one of the following commands (depends on whether it is installed): $ sudo ifconfig eth0eth0      Link encap:Ethernet  HWaddr 00:21:5a:50:d7:94            inet addr:192.168.160.2  Bcast:192.168.160.255  Mask:255.255.255.0          inet6 addr: fe80::221:5aff:fe50:d794/64 Scope:Link          inet6 addr: 2001:db8:bad:a55::2/64 Scope:Global          UP BROADCAST RUNNING MULTICAST  MTU:1500  Metric:1 $ sudo ip -6 address show eth03: eth0: <BROADCAST,MULTICAST,UP,LOWER_UP> mtu 1500 qlen 1000    inet6 2001:db8:bad:a55::2/64 scope global        valid_lft forever preferred_lft forever    inet6 fe80::221:5aff:fe50:d794/64 scope link        valid_lft forever preferred_lft forever In both cases, it confirms that our network device has been assigned a valid IPv6 address. That's it in general for your setup on one system. But of course, you might be interested to enable more services for IPv6, especially if you're already running a couple of them in your IP network. More details are available on the official Ubuntu Wiki. Continue to configure your network to provide IPv6 address information automatically in your local infrastructure.

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  • Segmentation fault 11 in MacOS X- C++ [migrated]

    - by Marcos Cesar Vargas Magana
    all. I have a "segmentation fault 11" error when I run the following code. The code actually compiles but I get the error at run time. //** Terror.h ** #include <iostream> #include <string> #include <map> using std::map; using std::pair; using std::string; template<typename Tsize> class Terror { public: //Inserts a message in the map. static Tsize insertMessage(const string& message) { mErrorMessages.insert( pair<Tsize, string>(mErrorMessages.size()+1, message) ); return mErrorMessages.size(); } private: static map<Tsize, string> mErrorMessages; } template<typename Tsize> map<Tsize,string> Terror<Tsize>::mErrorMessages; //** error.h ** #include <iostream> #include "Terror.h" typedef unsigned short errorType; typedef Terror<errorType> error; errorType memoryAllocationError=error::insertMessage("ERROR: out of memory."); //** main.cpp ** #include <iostream> #include "error.h" using namespace std; int main() { try { throw error(memoryAllocationError); } catch(error& err) { } } I have kind of debugging the code and the error happens when the message is being inserted in the static map member. An observation is that if I put the line: errorType memoryAllocationError=error::insertMessage("ERROR: out of memory."); inside the "main()" function instead of at global scope, then everything works fine. But I would like to extend the error messages at global scope, not at local scope. The map is defined static so that all instances of "error" share the same error codes and messages. Do you know how can I get this or something similar. Thank you very much.

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  • Including slick2d or slick-util in maven build?

    - by BotskoNet
    I'm converting a project to lwjgl and trying to use slick-util as well. There's no slick-util maven repo anywhere (nor slick2d itself anymore). I've included local dependancies before using <dependency> <groupId>org.newdawn</groupId> <artifactId>slick</artifactId> <version>237</version> <scope>system</scope> <systemPath>${project.basedir}/lib/slick-util.jar</systemPath> </dependency> The maven package process runs without issue, but when I try to run the jar, it errors out with a ClassNotFoundException. There's no mention of slick-util in the manifest and I can't find out how to make my game load that jar properly. Side question: how do I ensure when I distribute my applications, the game properly installs these libraries?

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  • NUMA-aware placement of communication variables

    - by Dave
    For classic NUMA-aware programming I'm typically most concerned about simple cold, capacity and compulsory misses and whether we can satisfy the miss by locally connected memory or whether we have to pull the line from its home node over the coherent interconnect -- we'd like to minimize channel contention and conserve interconnect bandwidth. That is, for this style of programming we're quite aware of where memory is homed relative to the threads that will be accessing it. Ideally, a page is collocated on the node with the thread that's expected to most frequently access the page, as simple misses on the page can be satisfied without resorting to transferring the line over the interconnect. The default "first touch" NUMA page placement policy tends to work reasonable well in this regard. When a virtual page is first accessed, the operating system will attempt to provision and map that virtual page to a physical page allocated from the node where the accessing thread is running. It's worth noting that the node-level memory interleaving granularity is usually a multiple of the page size, so we can say that a given page P resides on some node N. That is, the memory underlying a page resides on just one node. But when thinking about accesses to heavily-written communication variables we normally consider what caches the lines underlying such variables might be resident in, and in what states. We want to minimize coherence misses and cache probe activity and interconnect traffic in general. I don't usually give much thought to the location of the home NUMA node underlying such highly shared variables. On a SPARC T5440, for instance, which consists of 4 T2+ processors connected by a central coherence hub, the home node and placement of heavily accessed communication variables has very little impact on performance. The variables are frequently accessed so likely in M-state in some cache, and the location of the home node is of little consequence because a requester can use cache-to-cache transfers to get the line. Or at least that's what I thought. Recently, though, I was exploring a simple shared memory point-to-point communication model where a client writes a request into a request mailbox and then busy-waits on a response variable. It's a simple example of delegation based on message passing. The server polls the request mailbox, and having fetched a new request value, performs some operation and then writes a reply value into the response variable. As noted above, on a T5440 performance is insensitive to the placement of the communication variables -- the request and response mailbox words. But on a Sun/Oracle X4800 I noticed that was not the case and that NUMA placement of the communication variables was actually quite important. For background an X4800 system consists of 8 Intel X7560 Xeons . Each package (socket) has 8 cores with 2 contexts per core, so the system is 8x8x2. Each package is also a NUMA node and has locally attached memory. Every package has 3 point-to-point QPI links for cache coherence, and the system is configured with a twisted ladder "mobius" topology. The cache coherence fabric is glueless -- there's not central arbiter or coherence hub. The maximum distance between any two nodes is just 2 hops over the QPI links. For any given node, 3 other nodes are 1 hop distant and the remaining 4 nodes are 2 hops distant. Using a single request (client) thread and a single response (server) thread, a benchmark harness explored all permutations of NUMA placement for the two threads and the two communication variables, measuring the average round-trip-time and throughput rate between the client and server. In this benchmark the server simply acts as a simple transponder, writing the request value plus 1 back into the reply field, so there's no particular computation phase and we're only measuring communication overheads. In addition to varying the placement of communication variables over pairs of nodes, we also explored variations where both variables were placed on one page (and thus on one node) -- either on the same cache line or different cache lines -- while varying the node where the variables reside along with the placement of the threads. The key observation was that if the client and server threads were on different nodes, then the best placement of variables was to have the request variable (written by the client and read by the server) reside on the same node as the client thread, and to place the response variable (written by the server and read by the client) on the same node as the server. That is, if you have a variable that's to be written by one thread and read by another, it should be homed with the writer thread. For our simple client-server model that means using split request and response communication variables with unidirectional message flow on a given page. This can yield up to twice the throughput of less favorable placement strategies. Our X4800 uses the QPI 1.0 protocol with source-based snooping. Briefly, when node A needs to probe a cache line it fires off snoop requests to all the nodes in the system. Those recipients then forward their response not to the original requester, but to the home node H of the cache line. H waits for and collects the responses, adjudicates and resolves conflicts and ensures memory-model ordering, and then sends a definitive reply back to the original requester A. If some node B needed to transfer the line to A, it will do so by cache-to-cache transfer and let H know about the disposition of the cache line. A needs to wait for the authoritative response from H. So if a thread on node A wants to write a value to be read by a thread on node B, the latency is dependent on the distances between A, B, and H. We observe the best performance when the written-to variable is co-homed with the writer A. That is, we want H and A to be the same node, as the writer doesn't need the home to respond over the QPI link, as the writer and the home reside on the very same node. With architecturally informed placement of communication variables we eliminate at least one QPI hop from the critical path. Newer Intel processors use the QPI 1.1 coherence protocol with home-based snooping. As noted above, under source-snooping a requester broadcasts snoop requests to all nodes. Those nodes send their response to the home node of the location, which provides memory ordering, reconciles conflicts, etc., and then posts a definitive reply to the requester. In home-based snooping the snoop probe goes directly to the home node and are not broadcast. The home node can consult snoop filters -- if present -- and send out requests to retrieve the line if necessary. The 3rd party owner of the line, if any, can respond either to the home or the original requester (or even to both) according to the protocol policies. There are myriad variations that have been implemented, and unfortunately vendor terminology doesn't always agree between vendors or with the academic taxonomy papers. The key is that home-snooping enables the use of a snoop filter to reduce interconnect traffic. And while home-snooping might have a longer critical path (latency) than source-based snooping, it also may require fewer messages and less overall bandwidth. It'll be interesting to reprise these experiments on a platform with home-based snooping. While collecting data I also noticed that there are placement concerns even in the seemingly trivial case when both threads and both variables reside on a single node. Internally, the cores on each X7560 package are connected by an internal ring. (Actually there are multiple contra-rotating rings). And the last-level on-chip cache (LLC) is partitioned in banks or slices, which with each slice being associated with a core on the ring topology. A hardware hash function associates each physical address with a specific home bank. Thus we face distance and topology concerns even for intra-package communications, although the latencies are not nearly the magnitude we see inter-package. I've not seen such communication distance artifacts on the T2+, where the cache banks are connected to the cores via a high-speed crossbar instead of a ring -- communication latencies seem more regular.

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  • Key ATG architecture principles

    - by Glen Borkowski
    Overview The purpose of this article is to describe some of the important foundational concepts of ATG.  This is not intended to cover all areas of the ATG platform, just the most important subset - the ones that allow ATG to be extremely flexible, configurable, high performance, etc.  For more information on these topics, please see the online product manuals. Modules The first concept is called the 'ATG Module'.  Simply put, you can think of modules as the building blocks for ATG applications.  The ATG development team builds the out of the box product using modules (these are the 'out of the box' modules).  Then, when a customer is implementing their site, they build their own modules that sit 'on top' of the out of the box ATG modules.  Modules can be very simple - containing minimal definition, and perhaps a small amount of configuration.  Alternatively, a module can be rather complex - containing custom logic, database schema definitions, configuration, one or more web applications, etc.  Modules generally will have dependencies on other modules (the modules beneath it).  For example, the Commerce Reference Store module (CRS) requires the DCS (out of the box commerce) module. Modules have a ton of value because they provide a way to decouple a customers implementation from the out of the box ATG modules.  This allows for a much easier job when it comes time to upgrade the ATG platform.  Modules are also a very useful way to group functionality into a single package which can be leveraged across multiple ATG applications. One very important thing to understand about modules, or more accurately, ATG as a whole, is that when you start ATG, you tell it what module(s) you want to start.  One of the first things ATG does is to look through all the modules you specified, and for each one, determine a list of modules that are also required to start (based on each modules dependencies).  Once this final, ordered list is determined, ATG continues to boot up.  One of the outputs from the ordered list of modules is that each module can contain it's own classes and configuration.  During boot, the ordered list of modules drives the unified classpath and configpath.  This is what determines which classes override others, and which configuration overrides other configuration.  Think of it as a layered approach. The structure of a module is well defined.  It simply looks like a folder in a filesystem that has certain other folders and files within it.  Here is a list of items that can appear in a module: MyModule: META-INF - this is required, along with a file called MANIFEST.MF which describes certain properties of the module.  One important property is what other modules this module depends on. config - this is typically present in most modules.  It defines a tree structure (folders containing properties files, XML, etc) that maps to ATG components (these are described below). lib - this contains the classes (typically in jarred format) for any code defined in this module j2ee - this is where any web-apps would be stored. src - in case you want to include the source code for this module, it's standard practice to put it here sql - if your module requires any additions to the database schema, you should place that schema here Here's a screenshots of a module: Modules can also contain sub-modules.  A dot-notation is used when referring to these sub-modules (i.e. MyModule.Versioned, where Versioned is a sub-module of MyModule). Finally, it is important to completely understand how modules work if you are going to be able to leverage them effectively.  There are many different ways to design modules you want to create, some approaches are better than others, especially if you plan to share functionality between multiple different ATG applications. Components A component in ATG can be thought of as a single item that performs a certain set of related tasks.  An example could be a ProductViews component - used to store information about what products the current customer has viewed.  Components have properties (also called attributes).  The ProductViews component could have properties like lastProductViewed (stores the ID of the last product viewed) or productViewList (stores the ID's of products viewed in order of their being viewed).  The previous examples of component properties would typically also offer get and set methods used to retrieve and store the property values.  Components typically will also offer other types of useful methods aside from get and set.  In the ProductViewed component, we might want to offer a hasViewed method which will tell you if the customer has viewed a certain product or not. Components are organized in a tree like hierarchy called 'nucleus'.  Nucleus is used to locate and instantiate ATG Components.  So, when you create a new ATG component, it will be able to be found 'within' nucleus.  Nucleus allows ATG components to reference one another - this is how components are strung together to perform meaningful work.  It's also a mechanism to prevent redundant configuration - define it once and refer to it from everywhere. Here is a screenshot of a component in nucleus:  Components can be extremely simple (i.e. a single property with a get method), or can be rather complex offering many properties and methods.  To be an ATG component, a few things are required: a class - you can reference an existing out of the box class or you could write your own a properties file - this is used to define your component the above items must be located 'within' nucleus by placing them in the correct spot in your module's config folder Within the properties file, you will need to point to the class you want to use: $class=com.mycompany.myclass You may also want to define the scope of the class (request, session, or global): $scope=session In summary, ATG Components live in nucleus, generally have links to other components, and provide some meaningful type of work.  You can configure components as well as extend their functionality by writing code. Repositories Repositories (a.k.a. Data Anywhere Architecture) is the mechanism that ATG uses to access data primarily stored in relational databases, but also LDAP or other backend systems.  ATG applications are required to be very high performance, and data access is critical in that if not handled properly, it could create a bottleneck.  ATG's repository functionality has been around for a long time - it's proven to be extremely scalable.  Developers new to ATG need to understand how repositories work as this is a critical aspect of the ATG architecture.   Repositories essentially map relational tables to objects in ATG, as well as handle caching.  ATG defines many repositories out of the box (i.e. user profile, catalog, orders, etc), and this is comprised of both the underlying database schema along with the associated repository definition files (XML).  It is fully expected that implementations will extend / change the out of the box repository definitions, so there is a prescribed approach to doing this.  The first thing to be sure of is to encapsulate your repository definition additions / changes within your own module (as described above).  The other important best practice is to never modify the out of the box schema - in other words, don't add columns to existing ATG tables, just create your own new tables.  These will help ensure you can easily upgrade your application at a later date. xml-combination As mentioned earlier, when you start ATG, the order of the modules will determine the final configpath.  Files within this configpath are 'layered' such that modules on top can override configuration of modules below it.  This is the same concept for repository definition files.  If you want to add a few properties to the out of the box user profile, you simply need to create an XML file containing only your additions, and place it in the correct location in your module.  At boot time, your definition will be combined (hence the term xml-combination) with the lower, out of the box modules, with the result being a user profile that contains everything (out of the box, plus your additions).  Aside from just adding properties, there are also ways to remove and change properties. types of properties Aside from the normal 'database backed' properties, there are a few other interesting types: transient properties - these are properties that are in memory, but not backed by any database column.  These are useful for temporary storage. java-backed properties - by nature, these are transient, but in addition, when you access this property (by called the get method) instead of looking up a piece of data, it performs some logic and returns the results.  'Age' is a good example - if you're storing a birth date on the profile, but your business rules are defined in terms of someones age, you could create a simple java-backed property to look at the birth date and compare it to the current date, and return the persons age. derived properties - this is what allows for inheritance within the repository structure.  You could define a property at the category level, and have the product inherit it's value as well as override it.  This is useful for setting defaults, with the ability to override. caching There are a number of different caching modes which are useful at different times depending on the nature of the data being cached.  For example, the simple cache mode is useful for things like user profiles.  This is because the user profile will typically only be used on a single instance of ATG at one time.  Simple cache mode is also useful for read-only types of data such as the product catalog.  Locked cache mode is useful when you need to ensure that only one ATG instance writes to a particular item at a time - an example would be a customers order.  There are many options in terms of configuring caching which are outside the scope of this article - please refer to the product manuals for more details. Other important concepts - out of scope for this article There are a whole host of concepts that are very important pieces to the ATG platform, but are out of scope for this article.  Here's a brief description of some of them: formhandlers - these are ATG components that handle form submissions by users. pipelines - these are configurable chains of logic that are used for things like handling a request (request pipeline) or checking out an order. special kinds of repositories (versioned, files, secure, ...) - there are a couple different types of repositories that are used in various situations.  See the manuals for more information. web development - JSP/ DSP tag library - ATG provides a traditional approach to developing web applications by providing a tag library called the DSP library.  This library is used throughout your JSP pages to interact with all the ATG components. messaging - a message sub-system used as another way for components to interact. personalization - ability for business users to define a personalized user experience for customers.  See the other blog posts related to personalization.

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  • New code base, what experiences/recommendations do you have?

    - by hlovdal
    I will later this year start on a project (embedded hardware, C, small company) where I believe that most (if not all) code will be new. So what experiences do you have to share as advice to starting a new code base? What have you been missing in projects that you have been working on? What has worked really well? What has not worked? Let's limit this question to be about things that relate directly to the code (e.g "banning the use of gets()": in scope, version control: border line, build system: out of scope).

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  • WinPE, Startnet.CMD and passing variables to second batch file not working

    - by user140892
    I don't know scripting or PowerShell (yes I need to learn something). I'm not an expert batch file maker either. I have a WinPE flash drive which I used to deploy OS images. I have the WIM, drivers and anything needed else outside the WinPE environment to ensure that Updates, changes are easier for me to make. I use the "STARTNET.CMD" batch file which is part of the WinPE. The reason to go through the letter drives is that the WinPE always gets the X letter drive assigned. The flash drive itself can receive a random letter which always changes. My deployment menu is located on the flash drive it self and not inside the WinPE. This is so that if I need to make a change I don't have to re-do the WinPE. I am able to locate the "menu.bat" batch file and launch it. I use a variable to capture the letter drive. I call the second batch file named "menu.bat" and pass the variable to it. When the second batch file loads, I believe that I am calling the variable correctly. If I break out of the batch file I can echo the variable and see the expected reply. The issue is that I can't use the variable to work with anything on the second batch file. In my test, I can get this to work over and over. When it runs from the real USB flash drive it does not work. I removed comments from the second batch file to make it smaller. My issue is that files below all get a message stating that the system cannot find the path specified. Diskpart Imagex.exe bcdboot.exe Why can't I get the varible to properly function when I try to using example "ImageX.exe"? Contents of the Startnet.cmd @echo off for %%p in (a b c d e f g h i j k l m n o p q r s t u v w x y z) do if exist %%p:\Tools\ set w=%%p Set execpatch=%w%\Tools\ call %w%:\Menu.bat \Tools\ Contents of the Menu.BAT @echo off set SecondPath=%1 cls :Start cls Echo. Echo.============================================================== Echo. Windows 7 64 Bit Ent Basic Desktops Echo.============================================================== Echo. Echo A. 790 Windows 7 - Basic Echo. Echo. Echo I. Exit Echo. Echo. set /p choice=Choose your option = if not '%choice%'=='' set choice=%choice:~0,1% if '%choice%'=='a' goto 790_Windows_7_Basic echo "%choice%" is not a valid (answer/command) echo. goto start :790_Windows_7_Basic REM DISKPART /s %SecondPath%BatchFiles\Make-Partition.txt %SecondPath%imagex.exe /apply %SecondPath%Images\Win7-64b-Ent-Basic-SysPreped.wim 1 o:\ /verify %SecondPath%bcdboot.exe o:\Windows /s S: Copy %SecondPath%Unattended\unattend.XML o:\Windows\System32\sysprep\unattend.XML /y xcopy %SecondPath%Drivers\790\*.* o:\Windows\INF\790\ /E /Q /Y MD o:\Windows\Setup\Scripts\ Copy %SecondPath%BatchFiles\SetupComplete.cmd o:\Windows\Setup\Scripts\ /y Goto Done :Done Exit

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  • Cool Tools You Can Use: Validation Templates for PeopleSoft Contracts Processes

    - by Mark Rosenberg
    This is the first in a series of postings we’ll be making under the heading of Cool Tools You Can Use. Our PeopleSoft product management team identified the need for this series after reflecting on the many conversations we have each year with our PeopleSoft community members. During these conversations, we were discovering that customers and implementation partners were often not aware that solutions exist to the problems they were trying to address and that the solutions were readily available at no additional charge. Thus, the Cool Tools You Can Use series will describe the business challenge we’ve heard, the PeopleSoft solution to the challenge, and how you can learn more about the solution so that everyone can be sure to make full use of what PeopleSoft applications have to offer. The first cool tool we’ll look at is the Validation Template for PeopleSoft Contracts Process Requests, which was first released in December 2013 as part of PeopleSoft Contracts 9.2 Update Image 4. The business issue our customers highlighted to us is the need to tightly control but easily configure and manage the scope of data that any user can process when initiating a process. Control of each user’s span of impact is essential to reducing billing reconciliation issues, passing span of authority audits, and reducing (or even eliminating) the frequency of unexpected process results.  Setting Up the Validation Template for a PeopleSoft Contracts Process With the validation template, organizations can easily and quickly ensure the software restricts the scope of transactions a user can affect and gives organizations the confidence to know that business processes are being governed effectively. Additionally, this control of PeopleSoft Contracts process requests can be applied and easily maintained and adjusted from a web browser thereby enabling analysts to administer the rules without having to engage software developers to customize the software. During the field validation template setup, an analyst specifies the combinations of fields that must contain values when a user tries to setup a run control and initiate a PeopleSoft Contracts process from a process request page. For example, for the Process Limits component, an organization could require that users enter a valid combination of values for the business unit, contract, and contract type fields or a value in the contract administrator field. Until the user enters a valid combination of entries on the process request page, he cannot launch the process. With the validation template activated for process request pages, organizations can be confident that PeopleSoft Contracts users will not accidentally begin generating invoices or triggering other revenue management processes for transactions beyond their scope of authority. To learn more about the Validation Template, please review the Defining Validation Templates section of the PeopleSoft Contracts PeopleBooks. 

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  • How to get wireless (Alfa) operate in full power and speed up wireless Internet?

    - by MahboobeAlam
    I am using Wireless to connect to the internet (router). My laptop has Atheros wireless (AR 242x/542x) adapter but the router is a little bit far-away from my room so I have to use an external wireless adapter i.e. Alfa (Realtek 8187) for connectivity. However, since I have started using Ubuntu I noticed that Alfa is not working in full power, internet speed in Ubuntu is much slower than in Windows on my laptop. When I am using Windows 7 the LED (bulb) on Alfa blinks as it should, but when in Ubuntu, it doesn't blink rather it is on but very dim. Connection using Atheros adapter is also the same (slow)... I have tried 4 methods (I found on the Internet) to troubleshoot this matter but no success. It seems to me that Ubuntu/Linux is not letting these wireless adapters to operate in full strength. iwconfig shows that power management is off for both. So what's the problem? Details: ifconfig eth0 Link encap:Ethernet HWaddr 00:1f:16:1e:36:ec UP BROADCAST MULTICAST MTU:1500 Metric:1 RX packets:0 errors:0 dropped:0 overruns:0 frame:0 TX packets:0 errors:0 dropped:0 overruns:0 carrier:0 collisions:0 txqueuelen:1000 RX bytes:0 (0.0 B) TX bytes:0 (0.0 B) Interrupt:45 Base address:0x6000 lo Link encap:Local Loopback inet addr:127.0.0.1 Mask:255.0.0.0 inet6 addr: ::1/128 Scope:Host UP LOOPBACK RUNNING MTU:16436 Metric:1 RX packets:1657 errors:0 dropped:0 overruns:0 frame:0 TX packets:1657 errors:0 dropped:0 overruns:0 carrier:0 collisions:0 txqueuelen:0 RX bytes:241697 (241.6 KB) TX bytes:241697 (241.6 KB) wlan0 Link encap:Ethernet HWaddr 00:22:5f:9b:24:b5 inet6 addr: fe80::222:5fff:fe9b:24b5/64 Scope:Link UP BROADCAST MULTICAST MTU:1500 Metric:1 RX packets:715460 errors:0 dropped:0 overruns:0 frame:0 TX packets:694246 errors:0 dropped:0 overruns:0 carrier:0 collisions:0 txqueuelen:1000 RX bytes:493539292 (493.5 MB) TX bytes:235159393 (235.1 MB) wlan1 Link encap:Ethernet HWaddr 00:c0:ca:42:14:62 inet addr:192.168.100.102 Bcast:192.168.100.255 Mask:255.255.255.0 inet6 addr: fe80::2c0:caff:fe42:1462/64 Scope:Link UP BROADCAST RUNNING MULTICAST MTU:1500 Metric:1 RX packets:171053 errors:0 dropped:0 overruns:0 frame:0 TX packets:181363 errors:0 dropped:0 overruns:0 carrier:0 collisions:0 txqueuelen:1000 RX bytes:74094659 (74.0 MB) TX bytes:59474204 (59.4 MB) iwconfig lo no wireless extensions. eth0 no wireless extensions. wlan0 IEEE 802.11bg ESSID:off/any Mode:Managed Access Point: Not-Associated Tx-Power=20 dBm Retry long limit:7 RTS thr:off Fragment thr:off Power Management:off wlan1 IEEE 802.11bg ESSID:"Zia" Mode:Managed Frequency:2.412 GHz Access Point: 00:0D:F0:9C:A6:18 Bit Rate=54 Mb/s Tx-Power=20 dBm Retry long limit:7 RTS thr:off Fragment thr:off Power Management:off Link Quality=70/70 Signal level=-31 dBm Rx invalid nwid:0 Rx invalid crypt:0 Rx invalid frag:0 Tx excessive retries:204 Invalid misc:6610 Missed beacon:0

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  • OpenVPN Server Ethernet Bridging Question

    - by Hooplad
    Hello All, I am having a difficult time properly configuring an ethernet bridge using OpenVPN 2.0.9 install on CentOS 5 ( VPN server ). The goal that I am trying to complete is to connect a VM ( instance running on the same CentOS machine ) acting as a Microsoft Business Contact Manager server. I would then like this "BCM server" to serve Windows XP clients on 192.168.1.0/24 network as well as clients connecting from VPN ( 10.8.0.0/24 ). The setup as it is now was based off a known working configuration. The problem with the working configuration was that it would allow to the client to connect and access everything running on the VPN server ( SVN, Samba, VM Server ) but not any computers on the 192.168.1.0/24 network. I must disclose that the VPN server is behind a router/firewall. Ports are being forwarded correctly ( again, clients were able to connect to the VPN server with no problem. netcat confirms the udp port is open as well ). current ifconfig output br0 Link encap:Ethernet HWaddr 00:21:5E:4D:3A:C2 inet addr:192.168.1.169 Bcast:192.168.1.255 Mask:255.255.255.0 inet6 addr: fe80::221:5eff:fe4d:3ac2/64 Scope:Link UP BROADCAST RUNNING MULTICAST MTU:1500 Metric:1 RX packets:846890 errors:0 dropped:0 overruns:0 frame:0 TX packets:3072351 errors:0 dropped:0 overruns:0 carrier:0 collisions:0 txqueuelen:0 RX bytes:42686842 (40.7 MiB) TX bytes:4540654180 (4.2 GiB) eth0 Link encap:Ethernet HWaddr 00:21:5E:4D:3A:C2 UP BROADCAST RUNNING SLAVE MULTICAST MTU:1500 Metric:1 RX packets:882641 errors:0 dropped:0 overruns:0 frame:0 TX packets:1781383 errors:0 dropped:0 overruns:0 carrier:0 collisions:0 txqueuelen:1000 RX bytes:82342803 (78.5 MiB) TX bytes:2614727660 (2.4 GiB) Interrupt:169 eth1 Link encap:Ethernet HWaddr 00:21:5E:4D:3A:C3 UP BROADCAST RUNNING SLAVE MULTICAST MTU:1500 Metric:1 RX packets:650 errors:0 dropped:0 overruns:0 frame:0 TX packets:1347223 errors:0 dropped:0 overruns:0 carrier:0 collisions:0 txqueuelen:1000 RX bytes:67403 (65.8 KiB) TX bytes:1959529142 (1.8 GiB) Interrupt:233 lo Link encap:Local Loopback inet addr:127.0.0.1 Mask:255.0.0.0 inet6 addr: ::1/128 Scope:Host UP LOOPBACK RUNNING MTU:16436 Metric:1 RX packets:17452058 errors:0 dropped:0 overruns:0 frame:0 TX packets:17452058 errors:0 dropped:0 overruns:0 carrier:0 collisions:0 txqueuelen:0 RX bytes:94020256229 (87.5 GiB) TX bytes:94020256229 (87.5 GiB) tap0 Link encap:Ethernet HWaddr DE:18:C6:D7:01:63 inet6 addr: fe80::dc18:c6ff:fed7:163/64 Scope:Link UP BROADCAST RUNNING PROMISC MULTICAST MTU:1500 Metric:1 RX packets:0 errors:0 dropped:0 overruns:0 frame:0 TX packets:3086 errors:0 dropped:166 overruns:0 carrier:0 collisions:0 txqueuelen:100 RX bytes:0 (0.0 b) TX bytes:315099 (307.7 KiB) vmnet1 Link encap:Ethernet HWaddr 00:50:56:C0:00:01 inet addr:192.168.177.1 Bcast:192.168.177.255 Mask:255.255.255.0 inet6 addr: fe80::250:56ff:fec0:1/64 Scope:Link UP BROADCAST RUNNING MULTICAST MTU:1500 Metric:1 RX packets:0 errors:0 dropped:0 overruns:0 frame:0 TX packets:4224 errors:0 dropped:0 overruns:0 carrier:0 collisions:0 txqueuelen:1000 RX bytes:0 (0.0 b) TX bytes:0 (0.0 b) vmnet8 Link encap:Ethernet HWaddr 00:50:56:C0:00:08 inet addr:192.168.55.1 Bcast:192.168.55.255 Mask:255.255.255.0 inet6 addr: fe80::250:56ff:fec0:8/64 Scope:Link UP BROADCAST RUNNING MULTICAST MTU:1500 Metric:1 RX packets:0 errors:0 dropped:0 overruns:0 frame:0 TX packets:4226 errors:0 dropped:0 overruns:0 carrier:0 collisions:0 txqueuelen:1000 RX bytes:0 (0.0 b) TX bytes:0 (0.0 b) current route table Kernel IP routing table Destination Gateway Genmask Flags Metric Ref Use Iface 192.168.55.0 * 255.255.255.0 U 0 0 0 vmnet8 192.168.177.0 * 255.255.255.0 U 0 0 0 vmnet1 192.168.1.0 * 255.255.255.0 U 0 0 0 br0 current iptables output Chain INPUT (policy ACCEPT) target prot opt source destination ACCEPT all -- anywhere anywhere ACCEPT all -- anywhere anywhere Chain FORWARD (policy ACCEPT) target prot opt source destination ACCEPT all -- anywhere anywhere Chain OUTPUT (policy ACCEPT) target prot opt source destination server_known_working.conf local banshee port 1194 proto udp dev tap0 ca ca.crt cert banshee_server.crt key banshee_server.key dh dh1024.pem server 10.8.0.0 255.255.255.0 ifconfig-pool-persist ipp.txt push "route 192.168.1.0 255.255.255.0" client-to-client keepalive 10 120 tls-auth ta.key 0 user nobody group nobody persist-key persist-tun status openvpn-status.log verb 4 The following is the current CentOS server config file. server_ethernet_bridged.conf ( current ) local 192.168.1.169 port 1194 proto udp dev tap0 ca ca.crt cert server.crt key server.key dh dh1024.pem ifconfig-pool-persist ipp.txt server-bridge 192.168.1.169 255.255.255.0 192.168.1.200 192.168.1.210 push "route 192.168.1.0 255.255.255.0 192.168.1.1" client-to-client keepalive 10 120 tls-auth ta.key 0 user nobody group nobody persist-key persist-tun status openvpn-status.log verb 6 The following is one of the client's config file that was used with the known working configuration. client.opvn client dev tap proto udp remote XXX.XXX.XXX 1194 resolv-retry infinite nobind persist-key persist-tun ca client.crt cert client.crt key client.key tls-auth client.key 1 verb 3 I have tried the HOWTO provided by OpenVPN as well as others http://www.thebakershome.net/openvpn%5Ftutorial?page=1 with no success. Any help or suggestions would be appreciated.

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  • How can I install the Unity Photo Lens

    - by Chuqui
    I can't install the Photo Lens. After running these commands: sudo add-apt-repository ppa:scopes-packagers/ppa sudo apt-get update && sudo apt-get install unity-lens-photo unity-scope-shotwell unity-scope-flickr I get this: Package unity-lens-photo is not available, but is referred to by another package. This may mean that the package is missing, has been obsoleted, or is only available from another source E: Package 'unity-lens-photo' has no installation candidate Is there a fix for that? Thank you very much.

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  • Bridging LXC containers to host eth0 so they can have a public IP

    - by Vianney Stroebel
    UPDATE: I found the solution there: http://www.linuxfoundation.org/collaborate/workgroups/networking/bridge#No_traffic_gets_trough_.28except_ARP_and_STP.29 # cd /proc/sys/net/bridge # ls bridge-nf-call-arptables bridge-nf-call-iptables bridge-nf-call-ip6tables bridge-nf-filter-vlan-tagged # for f in bridge-nf-*; do echo 0 $f; done But I'd like to have expert opinions on this: is it safe to disable all bridge-nf-*? What are they here for? END OF UPDATE I need to bridge LXC containers to the physical interface (eth0) of my host, reading numerous tutorials, documents and blog posts on the subject. I need the containers to have their own public IP (which I've previously done KVM/libvirt). After two days of searching and trying, I still can't make it work with LXC containers. The host runs a freshly installed Ubuntu Server Quantal (12.10) with only libvirt (which I'm not using here) and lxc installed. I created the containers with : lxc-create -t ubuntu -n mycontainer So they also run Ubuntu 12.10. Content of /var/lib/lxc/mycontainer/config is: lxc.utsname = mycontainer lxc.mount = /var/lib/lxc/test/fstab lxc.rootfs = /var/lib/lxc/test/rootfs lxc.network.type = veth lxc.network.flags = up lxc.network.link = br0 lxc.network.name = eth0 lxc.network.veth.pair = vethmycontainer lxc.network.ipv4 = 179.43.46.233 lxc.network.hwaddr= 02:00:00:86:5b:11 lxc.devttydir = lxc lxc.tty = 4 lxc.pts = 1024 lxc.arch = amd64 lxc.cap.drop = sys_module mac_admin mac_override lxc.pivotdir = lxc_putold # uncomment the next line to run the container unconfined: #lxc.aa_profile = unconfined lxc.cgroup.devices.deny = a # Allow any mknod (but not using the node) lxc.cgroup.devices.allow = c *:* m lxc.cgroup.devices.allow = b *:* m # /dev/null and zero lxc.cgroup.devices.allow = c 1:3 rwm lxc.cgroup.devices.allow = c 1:5 rwm # consoles lxc.cgroup.devices.allow = c 5:1 rwm lxc.cgroup.devices.allow = c 5:0 rwm #lxc.cgroup.devices.allow = c 4:0 rwm #lxc.cgroup.devices.allow = c 4:1 rwm # /dev/{,u}random lxc.cgroup.devices.allow = c 1:9 rwm lxc.cgroup.devices.allow = c 1:8 rwm lxc.cgroup.devices.allow = c 136:* rwm lxc.cgroup.devices.allow = c 5:2 rwm # rtc lxc.cgroup.devices.allow = c 254:0 rwm #fuse lxc.cgroup.devices.allow = c 10:229 rwm #tun lxc.cgroup.devices.allow = c 10:200 rwm #full lxc.cgroup.devices.allow = c 1:7 rwm #hpet lxc.cgroup.devices.allow = c 10:228 rwm #kvm lxc.cgroup.devices.allow = c 10:232 rwm Then I changed my host /etc/network/interfaces to: auto lo iface lo inet loopback auto br0 iface br0 inet static bridge_ports eth0 bridge_fd 0 address 92.281.86.226 netmask 255.255.255.0 network 92.281.86.0 broadcast 92.281.86.255 gateway 92.281.86.254 dns-nameservers 213.186.33.99 dns-search ovh.net When I try command line configuration ("brctl addif", "ifconfig eth0", etc.) my remote host becomes inaccessible and I have to hard reboot it. I changed the content of /var/lib/lxc/mycontainer/rootfs/etc/network/interfaces to: auto lo iface lo inet loopback auto eth0 iface eth0 inet static address 179.43.46.233 netmask 255.255.255.255 broadcast 178.33.40.233 gateway 92.281.86.254 It takes several minutes for mycontainer to start (lxc-start -n mycontainer). I tried replacing gateway 92.281.86.254 by : post-up route add 92.281.86.254 dev eth0 post-up route add default gw 92.281.86.254 post-down route del 92.281.86.254 dev eth0 post-down route del default gw 92.281.86.254 My container then starts instantly. But whatever configuration I set in /var/lib/lxc/mycontainer/rootfs/etc/network/interfaces, I cannot ping from mycontainer to any IP (including the host's) : ubuntu@mycontainer:~$ ping 92.281.86.226 PING 92.281.86.226 (92.281.86.226) 56(84) bytes of data. ^C --- 92.281.86.226 ping statistics --- 6 packets transmitted, 0 received, 100% packet loss, time 5031ms And my host cannot ping the container: root@host:~# ping 179.43.46.233 PING 179.43.46.233 (179.43.46.233) 56(84) bytes of data. ^C --- 179.43.46.233 ping statistics --- 5 packets transmitted, 0 received, 100% packet loss, time 4000ms My container's ifconfig: ubuntu@mycontainer:~$ ifconfig eth0 Link encap:Ethernet HWaddr 02:00:00:86:5b:11 inet addr:179.43.46.233 Bcast:255.255.255.255 Mask:0.0.0.0 inet6 addr: fe80::ff:fe79:5a31/64 Scope:Link UP BROADCAST RUNNING MULTICAST MTU:1500 Metric:1 RX packets:64 errors:0 dropped:6 overruns:0 frame:0 TX packets:54 errors:0 dropped:0 overruns:0 carrier:0 collisions:0 txqueuelen:1000 RX bytes:4070 (4.0 KB) TX bytes:4168 (4.1 KB) lo Link encap:Local Loopback inet addr:127.0.0.1 Mask:255.0.0.0 inet6 addr: ::1/128 Scope:Host UP LOOPBACK RUNNING MTU:16436 Metric:1 RX packets:32 errors:0 dropped:0 overruns:0 frame:0 TX packets:32 errors:0 dropped:0 overruns:0 carrier:0 collisions:0 txqueuelen:0 RX bytes:2496 (2.4 KB) TX bytes:2496 (2.4 KB) My host's ifconfig: root@host:~# ifconfig br0 Link encap:Ethernet HWaddr 4c:72:b9:43:65:2b inet addr:92.281.86.226 Bcast:91.121.67.255 Mask:255.255.255.0 inet6 addr: fe80::4e72:b9ff:fe43:652b/64 Scope:Link UP BROADCAST RUNNING MULTICAST MTU:1500 Metric:1 RX packets:1453 errors:0 dropped:18 overruns:0 frame:0 TX packets:1630 errors:0 dropped:0 overruns:0 carrier:0 collisions:0 txqueuelen:0 RX bytes:145125 (145.1 KB) TX bytes:299943 (299.9 KB) eth0 Link encap:Ethernet HWaddr 4c:72:b9:43:65:2b UP BROADCAST RUNNING MULTICAST MTU:1500 Metric:1 RX packets:3178 errors:0 dropped:0 overruns:0 frame:0 TX packets:1637 errors:0 dropped:0 overruns:0 carrier:0 collisions:0 txqueuelen:1000 RX bytes:298263 (298.2 KB) TX bytes:309167 (309.1 KB) Interrupt:20 Memory:fe500000-fe520000 lo Link encap:Local Loopback inet addr:127.0.0.1 Mask:255.0.0.0 inet6 addr: ::1/128 Scope:Host UP LOOPBACK RUNNING MTU:16436 Metric:1 RX packets:6 errors:0 dropped:0 overruns:0 frame:0 TX packets:6 errors:0 dropped:0 overruns:0 carrier:0 collisions:0 txqueuelen:0 RX bytes:300 (300.0 B) TX bytes:300 (300.0 B) vethtest Link encap:Ethernet HWaddr fe:0d:7f:3e:70:88 inet6 addr: fe80::fc0d:7fff:fe3e:7088/64 Scope:Link UP BROADCAST RUNNING MULTICAST MTU:1500 Metric:1 RX packets:54 errors:0 dropped:0 overruns:0 frame:0 TX packets:67 errors:0 dropped:0 overruns:0 carrier:0 collisions:0 txqueuelen:1000 RX bytes:4168 (4.1 KB) TX bytes:4250 (4.2 KB) virbr0 Link encap:Ethernet HWaddr de:49:c5:66:cf:84 inet addr:192.168.122.1 Bcast:192.168.122.255 Mask:255.255.255.0 UP BROADCAST MULTICAST MTU:1500 Metric:1 RX packets:0 errors:0 dropped:0 overruns:0 frame:0 TX packets:0 errors:0 dropped:0 overruns:0 carrier:0 collisions:0 txqueuelen:0 RX bytes:0 (0.0 B) TX bytes:0 (0.0 B) I have disabled lxcbr0 (USE_LXC_BRIDGE="false" in /etc/default/lxc). root@host:~# brctl show bridge name bridge id STP enabled interfaces br0 8000.4c72b943652b no eth0 vethtest I have configured the IP 179.43.46.233 to point to 02:00:00:86:5b:11 in my hosting provider (OVH) config panel. (The IPs in this post are not the real ones.) Thanks for reading this long question! :-) Vianney

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  • Intel Centrino Wireless-N 1000 Again ! Ubuntu 13.04 x64

    - by vafa
    First I have to say that I tried everything written about this concept. The problem is that it stops working randomly in 3 main forms : 1 - sometimes it disconnect from wireless network and reconnect automatically 2 - sometimes it disconnect and wont connect no matter what (needs reboot) 3 - some times it's still connected but cannot ping or surf or whatever. I already tried disabling N mod using these commands : sudo modprobe -r iwlwifi modprobe iwlwifi 11n_disable=1 (or 0, whatever) it didn't help . these are the results of lspci, sudo lshw -C network, ifconfig, iwconfig, rfkill list when it disconnected and didn't connect till reboot : ifconfig : eth0 Link encap:Ethernet HWaddr c8:0a:a9:34:65:77 UP BROADCAST MULTICAST MTU:1500 Metric:1 RX packets:1563213476557380 errors:9379306629148050 dropped:3126435543049350 overruns:1563217771524675 frame:7816088857623375 TX packets:1563217771524675 errors:6252871086098700 dropped:0 overruns:1563217771524675 carrier:3126435543049350 collisions:7816088857623375 txqueuelen:1000 RX bytes:1563217771524675 (1.5 PB) TX bytes:1563217771524675 (1.5 PB) ham0 Link encap:Ethernet HWaddr 7a:79:19:a5:e4:93 inet addr:25.165.228.147 Bcast:25.255.255.255 Mask:255.0.0.0 inet6 addr: fe80::7879:19ff:fea5:e493/64 Scope:Link inet6 addr: 2620:9b::19a5:e493/96 Scope:Global UP BROADCAST RUNNING MULTICAST MTU:1404 Metric:1 RX packets:7743 errors:0 dropped:0 overruns:0 frame:0 TX packets:1250 errors:0 dropped:0 overruns:0 carrier:0 collisions:0 txqueuelen:500 RX bytes:665642 (665.6 KB) TX bytes:204056 (204.0 KB) lo Link encap:Local Loopback inet addr:127.0.0.1 Mask:255.0.0.0 inet6 addr: ::1/128 Scope:Host UP LOOPBACK RUNNING MTU:65536 Metric:1 RX packets:41138 errors:0 dropped:0 overruns:0 frame:0 TX packets:41138 errors:0 dropped:0 overruns:0 carrier:0 collisions:0 txqueuelen:0 RX bytes:6420962 (6.4 MB) TX bytes:6420962 (6.4 MB) wlan0 Link encap:Ethernet HWaddr 00:1e:64:45:fb:70 inet6 addr: fe80::21e:64ff:fe45:fb70/64 Scope:Link UP BROADCAST MULTICAST MTU:1500 Metric:1 RX packets:286999 errors:0 dropped:0 overruns:0 frame:0 TX packets:226966 errors:0 dropped:0 overruns:0 carrier:0 collisions:0 txqueuelen:1000 RX bytes:324386887 (324.3 MB) TX bytes:30674804 (30.6 MB) iwconfig : ham0 no wireless extensions. eth0 no wireless extensions. lo no wireless extensions. wlan0 IEEE 802.11bg ESSID:off/any Mode:Managed Access Point: Not-Associated Tx-Power=14 dBm Retry long limit:7 RTS thr:off Fragment thr:off Power Management:off sudo lshw -C network: *-network description: Wireless interface product: Centrino Wireless-N 1000 [Condor Peak] vendor: Intel Corporation physical id: 0 bus info: pci@0000:07:00.0 logical name: wlan0 version: 00 serial: 00:1e:64:45:fb:70 width: 64 bits clock: 33MHz capabilities: pm msi pciexpress bus_master cap_list ethernet physical wireless configuration: broadcast=yes driver=iwlwifi driverversion=3.8.0-30-generic firmware=39.31.5.1 build 35138 latency=0 link=no multicast=yes wireless=IEEE 802.11bg resources: irq:46 memory:c0400000-c0401fff *-network description: Ethernet interface product: AR8131 Gigabit Ethernet vendor: Qualcomm Atheros physical id: 0 bus info: pci@0000:09:00.0 logical name: eth0 version: c0 serial: c8:0a:a9:34:65:77 capacity: 1Gbit/s width: 64 bits clock: 33MHz capabilities: pm msi pciexpress vpd cap_list ethernet physical tp 10bt 10bt-fd 100bt 100bt-fd 1000bt-fd autonegotiation configuration: autonegotiation=on broadcast=yes driver=atl1c driverversion=1.0.1.1-NAPI latency=0 link=no multicast=yes port=twisted pair resources: irq:47 memory:c0900000-c093ffff ioport:5000(size=128) *-network description: Ethernet interface physical id: 2 logical name: ham0 serial: 7a:79:19:a5:e4:93 size: 10Mbit/s capabilities: ethernet physical configuration: autonegotiation=off broadcast=yes driver=tun driverversion=1.6 duplex=full ip=25.165.228.147 link=yes multicast=yes port=twisted pair speed=10Mbit/s lspci: 00:00.0 Host bridge: Intel Corporation Mobile 4 Series Chipset Memory Controller Hub (rev 07) 00:01.0 PCI bridge: Intel Corporation Mobile 4 Series Chipset PCI Express Graphics Port (rev 07) 00:1a.0 USB controller: Intel Corporation 82801I (ICH9 Family) USB UHCI Controller #4 (rev 03) 00:1a.1 USB controller: Intel Corporation 82801I (ICH9 Family) USB UHCI Controller #5 (rev 03) 00:1a.7 USB controller: Intel Corporation 82801I (ICH9 Family) USB2 EHCI Controller #2 (rev 03) 00:1b.0 Audio device: Intel Corporation 82801I (ICH9 Family) HD Audio Controller (rev 03) 00:1c.0 PCI bridge: Intel Corporation 82801I (ICH9 Family) PCI Express Port 1 (rev 03) 00:1c.3 PCI bridge: Intel Corporation 82801I (ICH9 Family) PCI Express Port 4 (rev 03) 00:1c.5 PCI bridge: Intel Corporation 82801I (ICH9 Family) PCI Express Port 6 (rev 03) 00:1d.0 USB controller: Intel Corporation 82801I (ICH9 Family) USB UHCI Controller #1 (rev 03) 00:1d.1 USB controller: Intel Corporation 82801I (ICH9 Family) USB UHCI Controller #2 (rev 03) 00:1d.2 USB controller: Intel Corporation 82801I (ICH9 Family) USB UHCI Controller #3 (rev 03) 00:1d.3 USB controller: Intel Corporation 82801I (ICH9 Family) USB UHCI Controller #6 (rev 03) 00:1d.7 USB controller: Intel Corporation 82801I (ICH9 Family) USB2 EHCI Controller #1 (rev 03) 00:1e.0 PCI bridge: Intel Corporation 82801 Mobile PCI Bridge (rev 93) 00:1f.0 ISA bridge: Intel Corporation ICH9M LPC Interface Controller (rev 03) 00:1f.2 SATA controller: Intel Corporation 82801IBM/IEM (ICH9M/ICH9M-E) 4 port SATA Controller [AHCI mode] (rev 03) 00:1f.3 SMBus: Intel Corporation 82801I (ICH9 Family) SMBus Controller (rev 03) 01:00.0 VGA compatible controller: NVIDIA Corporation G98M [GeForce G 105M] (rev a1) 07:00.0 Network controller: Intel Corporation Centrino Wireless-N 1000 [Condor Peak] 09:00.0 Ethernet controller: Qualcomm Atheros AR8131 Gigabit Ethernet (rev c0) rfkill list : 1: acer-wireless: Wireless LAN Soft blocked: no Hard blocked: no 2: acer-bluetooth: Bluetooth Soft blocked: yes Hard blocked: no 9: phy0: Wireless LAN Soft blocked: no Hard blocked: no any help will be REALLLYYYY appreciated

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  • Two Wifi Icons in Panel

    - by Alex
    I have the exact problem in 13.10 as this user Two Wifi indicators in panel. Here are some screenshots: Here are some screenshots from another user: http://ubuntuforums.org/showthread.php?t=2183020&p=12825563 ifconfig and iwconfig outputs $ ifconfig lo Link encap:Local Loopback inet addr:XXXXXX Mask:XXXXXXX inet6 addr: ::1/128 Scope:Host UP LOOPBACK RUNNING MTU:65536 Metric:1 RX packets:2243 errors:0 dropped:0 overruns:0 frame:0 TX packets:2243 errors:0 dropped:0 overruns:0 carrier:0 collisions:0 txqueuelen:0 RX bytes:209889 (209.8 KB) TX bytes:209889 (209.8 KB) wlan0 Link encap:Ethernet HWaddr XXXXXXXXX inet addr:XXXXXX Bcast:XXXXXXXX Mask:XXXXXXX inet6 addr: XXXXXXX Scope:Link UP BROADCAST RUNNING MULTICAST MTU:1500 Metric:1 RX packets:5925 errors:0 dropped:0 overruns:0 frame:0 TX packets:3361 errors:0 dropped:0 overruns:0 carrier:0 collisions:0 txqueuelen:1000 RX bytes:2951818 (2.9 MB) TX bytes:630579 (630.5 KB) $ iwconfig lo no wireless extensions. wlan0 IEEE 802.11abgn ESSID:"XXXXX" Mode:Managed Frequency:2.437 GHz Access Point: XXXXXXXX Bit Rate=72.2 Mb/s Tx-Power=15 dBm Retry long limit:7 RTS thr:off Fragment thr:off Power Management:on Link Quality=49/70 Signal level=-61 dBm Rx invalid nwid:0 Rx invalid crypt:0 Rx invalid frag:0 Tx excessive retries:153 Invalid misc:472 Missed beacon:0

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  • Restoring databases to a set drive and directory

    - by okeofs
     Restoring databases to a set drive and directory Introduction Often people say that necessity is the mother of invention. In this case I was faced with the dilemma of having to restore several databases, with multiple ‘ndf’ files, and having to restore them with different physical file names, drives and directories on servers other than the servers from which they originated. As most of us would do, I went to Google to see if I could find some code to achieve this task and found some interesting snippets on Pinal Dave’s website. Naturally, I had to take it further than the code snippet, HOWEVER it was a great place to start. Creating a temp table to hold database file details First off, I created a temp table which would hold the details of the individual data files within the database. Although there are a plethora of fields (within the temp table below), I utilize LogicalName only within this example. The temporary table structure may be seen below:   create table #tmp ( LogicalName nvarchar(128)  ,PhysicalName nvarchar(260)  ,Type char(1)  ,FileGroupName nvarchar(128)  ,Size numeric(20,0)  ,MaxSize numeric(20,0), Fileid tinyint, CreateLSN numeric(25,0), DropLSN numeric(25, 0), UniqueID uniqueidentifier, ReadOnlyLSN numeric(25,0), ReadWriteLSN numeric(25,0), BackupSizeInBytes bigint, SourceBlocSize int, FileGroupId int, LogGroupGUID uniqueidentifier, DifferentialBaseLSN numeric(25,0), DifferentialBaseGUID uniqueidentifier, IsReadOnly bit, IsPresent bit,  TDEThumbPrint varchar(50) )    We now declare and populate a variable(@path), setting the variable to the path to our SOURCE database backup. declare @path varchar(50) set @path = 'P:\DATA\MYDATABASE.bak'   From this point, we insert the file details of our database into the temp table. Note that we do so by utilizing a restore statement HOWEVER doing so in ‘filelistonly’ mode.   insert #tmp EXEC ('restore filelistonly from disk = ''' + @path + '''')   At this point, I depart from what I gleaned from Pinal Dave.   I now instantiate a few more local variables. The use of each variable will be evident within the cursor (which follows):   Declare @RestoreString as Varchar(max) Declare @NRestoreString as NVarchar(max) Declare @LogicalName  as varchar(75) Declare @counter as int Declare @rows as int set @counter = 1 select @rows = COUNT(*) from #tmp  -- Count the number of records in the temp                                    -- table   Declaring and populating the cursor At this point I do realize that many people are cringing about the use of a cursor. Being an Oracle professional as well, I have learnt that there is a time and place for cursors. I would remind the reader that the data that will be read into the cursor is from a local temp table and as such, any locking of the records (within the temp table) is not really an issue.   DECLARE MY_CURSOR Cursor  FOR  Select LogicalName  From #tmp   Parsing the logical names from within the cursor. A small caveat that works in our favour,  is that the first logical name (of our database) is the logical name of the primary data file (.mdf). Other files, except for the very last logical name, belong to secondary data files. The last logical name is that of our database log file.   I now open my cursor and populate the variable @RestoreString Open My_Cursor  set @RestoreString =  'RESTORE DATABASE [MYDATABASE] FROM DISK = N''P:\DATA\ MYDATABASE.bak''' + ' with  '   We now fetch the first record from the temp table.   Fetch NEXT FROM MY_Cursor INTO @LogicalName   While there are STILL records left within the cursor, we dynamically build our restore string. Note that we are using concatenation to create ‘one big restore executable string’.   Note also that the target physical file name is hardwired, as is the target directory.   While (@@FETCH_STATUS <> -1) BEGIN IF (@@FETCH_STATUS <> -2) -- As long as there are no rows missing select @RestoreString = case  when @counter = 1 then -- This is the mdf file    @RestoreString + 'move  N''' + @LogicalName + '''' + ' TO N’’X:\DATA1\'+ @LogicalName + '.mdf' + '''' + ', '   -- OK, if it passes through here we are dealing with an .ndf file -- Note that Counter must be greater than 1 and less than the number of rows.   when @counter > 1 and @counter < @rows then -- These are the ndf file(s)    @RestoreString + 'move  N''' + @LogicalName + '''' + ' TO N’’X:\DATA1\'+ @LogicalName + '.ndf' + '''' + ', '   -- OK, if it passes through here we are dealing with the log file When @LogicalName like '%log%' then    @RestoreString + 'move  N''' + @LogicalName + '''' + ' TO N’’X:\DATA1\'+ @LogicalName + '.ldf' +'''' end --Increment the counter   set @counter = @counter + 1 FETCH NEXT FROM MY_CURSOR INTO @LogicalName END   At this point we have populated the varchar(max) variable @RestoreString with a concatenation of all the necessary file names. What we now need to do is to run the sp_executesql stored procedure, to effect the restore.   First, we must place our ‘concatenated string’ into an nvarchar based variable. Obviously this will only work as long as the length of @RestoreString is less than varchar(max) / 2.   set @NRestoreString = @RestoreString EXEC sp_executesql @NRestoreString   Upon completion of this step, the database should be restored to the server. I now close and deallocate the cursor, and to be clean, I would also drop my temp table.   CLOSE MY_CURSOR DEALLOCATE MY_CURSOR GO   Conclusion Restoration of databases on different servers with different physical names and on different drives are a fact of life. Through the use of a few variables and a simple cursor, we may achieve an efficient and effective way to achieve this task.

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  • Software engineering project idea feedback [on hold]

    - by Chris Sewell
    I'm a third year student currently undergoing my project/dissertation section of my degree. I have drafted a proposal for my final year project and would appreciate any feedback. The feedback can be anything constructive either specific to this proposal, the area that I will be working and researching in or my ideas. I will accept all input. Aims My aim is to attempt a proof of concept and prototype a runtime-as-a-service (RaaS). This cloud based runtime will allow clients to dynamically offload tasks or create cloud applications. Currently software-as-a-service (SaaS) cloud applications are purpose built and have a predefined scope in which they can assist or serve the client; this scope cannot be changed without physical alteration to the client and server software. With RaaS the client potentially could define any task it wanted at any time depending on its environment variables, the client and server would then communicate parameters and returns for that task. For the client to utilize a RaaS it must be able to conceive and then define a task using an appropriate XML vocabulary. As the scope of the cloud solution is defined by the client at its runtime, the cloud solution only has to exist for as long as the client requires it to as opposed to a client using a dedicated service. Deliverables The crux of the project will require an XML vocabulary in which the client and server will communicate. I’ll prototype the server application that will dynamically create and manage cloud solutions. The solution will be coded using an interpreted language, such as python or javascript, which can evaluate expressions in runtime or a language that can dynamically compile such as C# or Java. As a further proof of concept I will also produce a mock client that offloads tasks to the server. The report will attempt to explain the different flavours of cloud computing solutions including infrastructure-as-a-service (IaaS), platform-as-a-service (PaaS) and SaaS including real world examples and where the use of a RaaS could have improved the overall example solution. Disclaimer: I'm not requesting stakeholders in my project nor am I delegating work. Any materials other than feedback, advice or directions will not be utilized.

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