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• I owe you an explanation

  - by Blueberry Coder

  Welcome to my blog! I am Frédéric Desbiens, a new member of the ADF Product Management team.  I joined Oracle only a few weeks ago. My boss is Grant Ronald, and I have the privilege to work in the same team as Susan Duncan, Frank Nimphius, Lynn Munsinger and Chris Muir. I share with them a passion for all things Java and ADF. With this blog, I hope to help you be more successful with our products – whether you are a customer or a partner. You may have heard of me before. Maybe you have my book in your bookshelf; or maybe we met at a conference. I went to JavaOne, ODTUG Kaleidoscope and Oracle OpenWorld in the past, when I worked for a major consulting firm. I will spare you all the details of my career; you can have a look at my LinkedIn profile if you are curious about my past.  Usually, my posts will be of a technical nature, and will focus on Oracle ADF and Oracle JDeveloper. SOA and portals have always been two topics of interest for me, however, and I will write about them. Over time, you will probably get acquainted with my « strategic » side as well. I devour history books, and always had a tendency to look at the big picture. I will probably not resist to the temptation of mixing IT and history, but this will be occasional, I promise!  At this point, I owe you an explanation about the title of the blog. I am French-Canadian, and wanted to evoke my roots in an obvious yet unobtrusive way. I was born in Chicoutimi, which is one of the main cities found in the Saguenay-Lac-Saint-Jean region. Traditionally, a large part of the wild blueberry production of the province of Québec come from there. A common nickname for the inhabitants is thus Les Bleuets, « The Blueberries » in English. I hope to see you around. You can also follow me on Twitter under  @BlueberryCoder.

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• How to become a good team player?

  - by Nick

  I've been programming (obsessively) since I was 12. I am fairly knowledgeable across the spectrum of languages out there, from assembly, to C++, to Javascript, to Haskell, Lisp, and Qi. But all of my projects have been by myself. I got my degree in chemical engineering, not CS or computer engineering, but for the first time this fall I'll be working on a large programming project with other people, and I have no clue how to prepare. I've been using Windows all of my life, but this project is going to be very unix-y, so I purchased a Mac recently in the hopes of familiarizing myself with the environment. I was fortunate to participate in a hackathon with some friends this past year -- both CS majors -- and excitingly enough, we won. But I realized as I worked with them that their workflow was very different from mine. They used Git for version control. I had never used it at the time, but I've since learned all that I can about it. They also used a lot of frameworks and libraries. I had to learn what Rails was pretty much overnight for the hackathon (on the other hand, they didn't know what lexical scoping or closures were). All of our code worked well, but they didn't understand mine, and I didn't understand theirs. I hear references to things that real programmers do on a daily basis -- unit testing, code reviews, but I only have the vaguest sense of what these are. I normally don't have many bugs in my little projects, so I have never needed a bug tracking system or tests for them. And the last thing is that it takes me a long time to understand other people's code. Variable naming conventions (that vary with each new language) are difficult (__mzkwpSomRidicAbbrev), and I find the loose coupling difficult. That's not to say I don't loosely couple things -- I think I'm quite good at it for my own work, but when I download something like the Linux kernel or the Chromium source code to look at it, I spend hours trying to figure out how all of these oddly named directories and files connect. It's a programming sin to reinvent the wheel, but I often find it's just quicker to write up the functionality myself than to spend hours dissecting some library. Obviously, people who do this for a living don't have these problems, and I'll need to get to that point myself. Question: What are some steps that I can take to begin "integrating" with everyone else? Thanks!

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• What is this algorithm for converting strings into numbers called?

  - by CodexArcanum

  I've been doing some work in Parsec recently, and for my toy language I wanted multi-based fractional numbers to be expressible. After digging around in Parsec's source a bit, I found their implementation of a floating-point number parser, and copied it to make the needed modifications. So I understand what this code does, and vaguely why (I haven't worked out the math fully yet, but I think I get the gist). But where did it come from? This seems like a pretty clever way to turn strings into floats and ints, is there a name for this algorithm? Or is it just something basic that's a hole in my knowledge? Did the folks behind Parsec devise it? Here's the code, first for integers: number' :: Integer -> Parser Integer number' base = do { digits <- many1 ( oneOf ( sigilRange base )) ; let n = foldl (\x d -> base * x + toInteger (convertDigit base d)) 0 digits ; seq n (return n) } So the basic idea here is that digits contains the string representing the whole number part, ie "192". The foldl converts each digit individually into a number, then adds that to the running total multiplied by the base, which means that by the end each digit has been multiplied by the correct factor (in aggregate) to position it. The fractional part is even more interesting: fraction' :: Integer -> Parser Double fraction' base = do { digits <- many1 ( oneOf ( sigilRange base )) ; let base' = fromIntegral base ; let f = foldr (\d x -> (x + fromIntegral (convertDigit base d))/base') 0.0 digits ; seq f (return f) Same general idea, but now a foldr and using repeated division. I don't quite understand why you add first and then divide for the fraction, but multiply first then add for the whole. I know it works, just haven't sorted out why. Anyway, I feel dumb not working it out myself, it's very simple and clever looking at it. Is there a name for this algorithm? Maybe the imperative version using a loop would be more familiar?

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• Exalogic 2.0.1 Tea Break Snippets - Creating and using Distribution Groups

  - by The Old Toxophilist

  By default running your Exalogic in a Virtual provides you with, what to Cloud Users, is a single large resource and they can just create vServers and not care about how they are laid down on the the underlying infrastructure. All the Cloud Users will know is that they can create vServers. For example if we have a Quarter Rack (8 Nodes) and our Cloud User creates 8 vServers those 8 vServers may run on 8 distinct nodes or may all run on the same node. Although in many cases we, as Cloud Users, may not be to worried how the Virtualisation Algorithm decides where to place our vServers there are cases where it is extremely important that vServers run on distinct physical compute nodes. For example if we have a Weblogic Cluster we will want the Servers with in the cluster to run on distinct physical node to cover for the situation where one physical node is lost. To achieve this the Exalogic Virtualised implementation provides Distribution Groups that define and anti-aliasing policy that the underlying Virtualisation Algorithm will take into account when placing vServers. It should be noted that Distribution Groups must be created before you create vServers because a vServer can only be added to a Distribution Group at creation time. Creating A Distribution Group To create a Distribution Groups we will first need to select the Account in which we want the Distribution Group to be created. Once we have selected the account we will see the Interface update and Account specific Actions will be displayed within the Action Panes. From the Action pane (or Right-Click on the Account) select the "Create Distribution Group" action. This will initiate the create wizard as follows. Distribution Group Details Within the first Step of the Wizard we can specify the name of the distribution group and this should be unique. In addition we can provide a detailed description of the group. Distribution Group Configuration The second step of the configuration wizard allows you to specify the number of elements that are required within this group and will specify a maximum of the number of nodes within you Exalogic. At this point it is always better to specify a group with spare capacity allowing for future expansion. As vServers are added to group the available slots decrease. Summary Finally the last step of the wizard display a summary of the information entered.

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• Facebook Game database design

  - by facebook-100000781341887

  Hi, I'm currently develop a facebook mafia like PHP game(of course, a light weight version), here is a simplify database(MySQL) of the game id-a <int3> <for index> uid <chr15> <facebook uid> HP <int3> <health point> exp <int3> <experience> money <int3> <money> list_inventory <chr5> <the inventory user hold...some special here, talk next> ... and 20 other fields just like reputation, num of combat... *the number next to the type is the size(byte) of the type For the list_inventory, there have 40 inventorys in my game, (actually, I have 5 these kind of list in my database), and each user can only contain 1 qty of each inventory, therefore, I assign 5 char for this field and each bit of char as 1 item(5 char * 8 bit = 40 slot), and I will do some manipulation by PHP to extract the data from this 5 byte. OK, I was thinking on this, if this game contains 100,000 user, and only 10% are active, therefore, if use my method, for the space use, 5 byte * 100,000 = 500 KB if I use another method, create a table user_hold_inventory, if the user have the inventory, then insert a record into this table, so, for 10,000 active user, I assume they got all item, but for other, I assume they got no item, here is the fields of the new table id-b <int3> <for index> id-a <int3> <id of the user table> inv_no <int1> <inventory that user hold> for the space use, ([id] (3+3) byte + [inv_no] 1 byte ) * [active user] 10,000 * [all inventory] * 40 = 2.8 MB seems method 2 have use more space, but it consume less CPU power. Please comment these 2 method or please correct me if there have another better method rather than what I think. Another question is, my database contain 26 fields, but I counted 5 of them are not change frquently, should I need to separate it on the other table or not? So many words, thanks for reading :)

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• How to handle this unfortunately non hypothetical situation with end-users?

  - by User Smith

  I work in a medium sized company but with a very small IT force. Last year (2011), I wrote an application that is very popular with a large group of end-users. We hit a deadline at the end of last year and some functionality (I will call funcA from now on) was not added into the application that was wanted at the very end. So, this application has been running in live/production since the end of 2011, I might add without issue. Yesterday, a whole group of end-users started complaining that funcA that was never in the application is no longer working. Our priority at this company is that if an application is broken it must be fixed first prior to prioritized projects. I have compared code and queries and there is no difference since 2011, which is proofA. I then was able to get one of the end-users to admit that it never worked proofB, but since then that end-user has went back and said that it was working previously......I believe the horde of end-users has assimilated her. I have also reviewed my notes for this project which has requirements and daily updates regarding the project which specifically states, "funcA not achieved due to time constraints", proofC. I have spoken with many of them and I can see where they could be confused as they are very far from a programming background, but I also know they are intelligent enough to act in a group in order to bypass project prioritization orders in order to get functionality that they want to make their job easier. The worst part is is that now group think is setting in and my boss and the head of IT is actually starting to believe them, even though there is no code or query changes. As far as reviewing the state of the logic it is very cut and dry to the point of if 1 = 1, funcA will not work. So, this is the end of the description of my scenario, but I am trying not to get severally dinged on my performance metrics due to this which would essentially have me moved to fixing a production problem that doesn't exist that will probably take over 1 month. I am looking for direct answers to this question. This question is not for rants, polling, or discussions as this is not the format for StackExchange. Please don't downvote me too terribly it is pretty common on this specific site of stack, I am looking for honest answers to this situation and I couldn't find a forum more appropriate.

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• How to remove the boundary effects arising due to zero padding in scipy/numpy fft?

  - by Omkar

  I have made a python code to smoothen a given signal using the Weierstrass transform, which is basically the convolution of a normalised gaussian with a signal. The code is as follows: #Importing relevant libraries from __future__ import division from scipy.signal import fftconvolve import numpy as np def smooth_func(sig, x, t= 0.002): N = len(x) x1 = x[-1] x0 = x[0] # defining a new array y which is symmetric around zero, to make the gaussian symmetric. y = np.linspace(-(x1-x0)/2, (x1-x0)/2, N) #gaussian centered around zero. gaus = np.exp(-y**(2)/t) #using fftconvolve to speed up the convolution; gaus.sum() is the normalization constant. return fftconvolve(sig, gaus/gaus.sum(), mode='same') If I run this code for say a step function, it smoothens the corner, but at the boundary it interprets another corner and smoothens that too, as a result giving unnecessary behaviour at the boundary. I explain this with a figure shown in the link below. Boundary effects This problem does not arise if we directly integrate to find convolution. Hence the problem is not in Weierstrass transform, and hence the problem is in the fftconvolve function of scipy. To understand why this problem arises we first need to understand the working of fftconvolve in scipy. The fftconvolve function basically uses the convolution theorem to speed up the computation. In short it says: convolution(int1,int2)=ifft(fft(int1)*fft(int2)) If we directly apply this theorem we dont get the desired result. To get the desired result we need to take the fft on a array double the size of max(int1,int2). But this leads to the undesired boundary effects. This is because in the fft code, if size(int) is greater than the size(over which to take fft) it zero pads the input and then takes the fft. This zero padding is exactly what is responsible for the undesired boundary effects. Can you suggest a way to remove this boundary effects? I have tried to remove it by a simple trick. After smoothening the function I am compairing the value of the smoothened signal with the original signal near the boundaries and if they dont match I replace the value of the smoothened func with the input signal at that point. It is as follows: i = 0 eps=1e-3 while abs(smooth[i]-sig[i])> eps: #compairing the signals on the left boundary smooth[i] = sig[i] i = i + 1 j = -1 while abs(smooth[j]-sig[j])> eps: # compairing on the right boundary. smooth[j] = sig[j] j = j - 1 There is a problem with this method, because of using an epsilon there are small jumps in the smoothened function, as shown below: jumps in the smooth func Can there be any changes made in the above method to solve this boundary problem?

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• 4.8M wasn't enough so we went for 5.055M tpmc with Unbreakable Enterprise Kernel r2 :-)

  - by wcoekaer

  We released a new set of benchmarks today. One is an updated tpc-c from a few months ago where we had just over 4.8M tpmc at $0.98 and we just updated it to go to 5.05M and $0.89. The other one is related to Java Middleware performance. You can find the press release here. Now, I don't want to talk about the actual relevance of the benchmark numbers, as I am not in the benchmark team. I want to talk about why these numbers and these efforts, unrelated to what they mean to your workload, matter to customers. The actual benchmark effort is a very big, long, expensive undertaking where many groups work together as a big virtual team. Having the virtual team be within a single company of course helps tremendously... We already start with a very big server setup with tons of storage, many disks, lots of ram, lots of cpu's, cores, threads, large database setups. Getting the whole setup going to start tuning, by itself, is no easy task, but then the real fun starts with tuning the system for optimal performance -and- stability. A benchmark is not just revving an engine at high rpm, it's actually hitting the circuit. The tests require long runs, require surviving availability tests, such as surviving crashes -and- recovery under load. In the TPC-C example, the x4800 system had 4TB ram, 160 threads (8 sockets, hyperthreaded, 10 cores/socket), tons of storage attached, tons of luns visible to the OS. flash storage, non flash storage... many things at high scale that all have to be perfectly synchronized. During this process, we find bugs, we fix bugs, we find performance issues, we fix performance issues, we find interesting potential features to investigate for the future, we start new development projects for future releases and all this goes back into the products. As more and more customers, for Oracle Linux, are running larger and larger, faster and faster, more mission critical, higher available databases..., these things are just absolutely critical. Unrelated to what anyone's specific opinion is about tpc-c or tpc-h or specjenterprise etc, there is a ton of effort that the customer benefits from. All this work makes Oracle Linux and/or Oracle Solaris better platforms. Whether it's faster, more stable, more scalable, more resilient. It helps. Another point that I always like to re-iterate around UEK and UEK2 : we have our kernel source git repository online. Complete changelog of the mainline kernel, and our changes, easy to pull, easy to dissect, easy to know what went in when, why and where. No need to go log into a website and manually click through pages to hopefully discover changes or patches. No need to untar 2 tar balls and run a diff.

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• Why would more CPU cores on virtual machine slow compile times?

  - by Sid

  [edit#2] If anyone from VMWare can hit me up with a copy of VMWare Fusion, I'd be more than happy to do the same as a VirtualBox vs VMWare comparison. Somehow I suspect the VMWare hypervisor will be better tuned for hyperthreading (see my answer too) I'm seeing something curious. As I increase the number of cores on my Windows 7 x64 virtual machine, the overall compile time increases instead of decreasing. Compiling is usually very well suited for parallel processing as in the middle part (post dependency mapping) you can simply call a compiler instance on each of your .c/.cpp/.cs/whatever file to build partial objects for the linker to take over. So I would have imagined that compiling would actually scale very well with # of cores. But what I'm seeing is: 8 cores: 1.89 sec 4 cores: 1.33 sec 2 cores: 1.24 sec 1 core: 1.15 sec Is this simply a design artifact due to a particular vendor's hypervisor implementation (type2:virtualbox in my case) or something more pervasive across more VMs to make hypervisor implementations more simpler? With so many factors, I seem to be able to make arguments both for and against this behavior - so if someone knows more about this than me, I'd be curious to read your answer. Thanks Sid [edit:addressing comments] @MartinBeckett: Cold compiles were discarded. @MonsterTruck: Couldn't find an opensource project to compile directly. Would be great but can't screwup my dev env right now. @Mr Lister, @philosodad: Have 8 hw threads, using VirtualBox, so should be 1:1 mapping without emulation @Thorbjorn: I have 6.5GB for the VM and a smallish VS2012 project - it's quite unlikely that I'm swapping in/out trashing the page file. @All: If someone can point to an open source VS2010/VS2012 project, that might be a better community reference than my (proprietary) VS2012 project. Orchard and DNN seem to need environment tweaking to compile in VS2012. I really would like to see if someone with VMWare Fusion also sees this (for VMWare vs VirtualBox compartmentalization) Test details: Hardware: Macbook Pro Retina CPU : Core i7 @ 2.3Ghz (quad core, hyper threaded = 8 cores in windows task manager) Memory : 16 GB Disk : 256GB SSD Host OS: Mac OS X 10.8 VM type: VirtualBox 4.1.18 (type 2 hypervisor) Guest OS: Windows 7 x64 SP1 Compiler: VS2012 compiling a solution with 3 C# Azure projects Compile times measure by VS2012 plugin called 'VSCommands' All tests run 5 times, first 2 runs discarded, last 3 averaged

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• With AMD style modules in JavaScript is there any benefit to namespaces?

  - by gman

  Coming from C++ originally and seeing lots of Java programmers doing the same we brought namespaces to JavaScript. See Google's closure library as an example where they have a main namespace, goog and under that many more namespaces like goog.async, goog.graphics But now, having learned the AMD style of requiring modules it seems like namespaces are kind of pointless in JavaScript. Not only pointless but even arguably an anti-pattern. What is AMD? It's a way of defining and including modules that removes all direct dependencies. Effectively you do this // some/module.js define([ 'name/of/needed/module', 'name/of/someother/needed/module', ], function( RefToNeededModule, RefToSomeOtherNeededModule) { ...code... return object or function }); This format lets the AMD support code know that this module needs name/of/needed/module.js and name/of/someother/needed/module.js loaded. The AMD code can load all the modules and then, assuming no circular dependencies, call the define function on each module in the correct order, record the object/function returned by the module as it calls them, and then call any other modules' define function with references to those modules. This seems to remove any need for namespaces. In your own code you can call the reference to any other module anything you want. For example if you had 2 string libraries, even if they define similar functions, as long as they follow the AMD pattern you can easily use both in the same module. No need for namespaces to solve that. It also means there's no hard coded dependencies. For example in Google's closure any module could directly reference another module with something like var value = goog.math.someMathFunc(otherValue) and if you're unlucky it will magically work where as with AMD style you'd have to explicitly include the math library otherwise the module wouldn't have a reference to it since there are no globals with AMD. On top of that dependency injection for testing becomes easy. None of the code in the AMD module references things by namespace so there is no hardcoded namespace paths, you can easily mock classes at testing time. Is there any other point to namespaces or is that something that C++ / Java programmers are bringing to JavaScript that arguably doesn't really belong?

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• LWJGL - Mixing 2D and 3D

  - by nathan

  I'm trying to mix 2D and 3D using LWJGL. I have wrote 2D little method that allow me to easily switch between 2D and 3D. protected static void make2D() { glEnable(GL_BLEND); GL11.glMatrixMode(GL11.GL_PROJECTION); GL11.glLoadIdentity(); glOrtho(0.0f, SCREEN_WIDTH, SCREEN_HEIGHT, 0.0f, 0.0f, 1.0f); GL11.glMatrixMode(GL11.GL_MODELVIEW); GL11.glLoadIdentity(); } protected static void make3D() { glDisable(GL_BLEND); GL11.glMatrixMode(GL11.GL_PROJECTION); GL11.glLoadIdentity(); // Reset The Projection Matrix GLU.gluPerspective(45.0f, ((float) SCREEN_WIDTH / (float) SCREEN_HEIGHT), 0.1f, 100.0f); // Calculate The Aspect Ratio Of The Window GL11.glMatrixMode(GL11.GL_MODELVIEW); glLoadIdentity(); } The in my rendering code i would do something like: make2D(); //draw 2D stuffs here make3D(); //draw 3D stuffs here What i'm trying to do is to draw a 3D shape (in my case a quad) and i 2D image. I found this example and i took the code from TextureLoader, Texture and Sprite to load and render a 2D image. Here is how i load the image. TextureLoader loader = new TextureLoader(); Sprite s = new Sprite(loader, "player.png") And how i render it: make2D(); s.draw(0, 0); It works great. Here is how i render my quad: glTranslatef(0.0f, 0.0f, 30.0f); glScalef(12.0f, 9.0f, 1.0f); DrawUtils.drawQuad(); Once again, no problem, the quad is properly rendered. DrawUtils is a simple class i wrote containing utility method to draw primitives shapes. Now my problem is when i want to mix both of the above, loading/rendering the 2D image, rendering the quad. When i try to load my 2D image with the following: s = new Sprite(loader, "player.png); My quad is not rendered anymore (i'm not even trying to render the 2D image at this point). Only the fact of creating the texture create the issue. After looking a bit at the code of Sprite and TextureLoader i found that the problem appears after the call of the glTexImage2d. In the TextureLoader class: glTexImage2D(target, 0, dstPixelFormat, get2Fold(bufferedImage.getWidth()), get2Fold(bufferedImage.getHeight()), 0, srcPixelFormat, GL_UNSIGNED_BYTE, textureBuffer); Commenting this like make the problem disappear. My question is then why? Is there anything special to do after calling this function to do 3D? Does this function alter the render part, the projection matrix?

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• Should a server "be lenient" in what it accepts and "discard faulty input silently"?

  - by romkyns

  I was under the impression that by now everyone agrees this maxim was a mistake. But I recently saw this answer which has a "be lenient" comment upvoted 137 times (as of today). In my opinion, the leniency in what browsers accept was the direct cause of the utter mess that HTML and some other web standards were a few years ago, and have only recently begun to properly crystallize out of that mess. The way I see it, being lenient in what you accept will lead to this. The second part of the maxim is "discard faulty input silently, without returning an error message unless this is required by the specification", and this feels borderline offensive. Any programmer who has banged their head on the wall when something fails silently will know what I mean. So, am I completely wrong about this? Should my program be lenient in what it accepts and swallow errors silently? Or am I mis-interpreting what this is supposed to mean? The original question said "program", and I take everyone's point about that. It can make sense for programs to be lenient. What I really meant, however, is APIs: interfaces exposed to other programs, rather than people. HTTP is an example. The protocol is an interface that only other programs use. People never directly provide the dates that go into headers like "If-Modified-Since". So, the question is: should the server implementing a standard be lenient and allow dates in several other formats, in addition to the one that's actually required by the standard? I believe the "be lenient" is supposed to apply to this situation, rather than human interfaces. If the server is lenient, it might seem like an overall improvement, but I think in practice it only leads to client implementations that end up relying on the leniency and thus failing to work with another server that's lenient in slightly different ways. So, should a server exposing some API be lenient or is that a very bad idea? Now onto lenient handling of user input. Consider YouTrack (a bug tracking software). It uses a language for text entry that is reminiscent of Markdown. Except that it's "lenient". For example, writing - foo - bar - baz is not a documented way of creating a bulleted list, and yet it worked. Consequently, it ended up being used a lot throughout our internal bugtracker. Next version comes out, and this lenient feature starts working slightly differently, breaking a bunch of lists that (mis)used this (non)feature. The documented way to create bulleted lists still works, of course. So, should my software be lenient in what user inputs it accepts?

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• How do you handle objects that need custom behavior, and need to exist as an entity in the database?

  - by Scott Whitlock

  For a simple example, assume your application sends out notifications to users when various events happen. So in the database I might have the following tables: TABLE Event EventId uniqueidentifier EventName varchar TABLE User UserId uniqueidentifier Name varchar TABLE EventSubscription EventUserId EventId UserId The events themselves are generated by the program. So there are hard-coded points in the application where an event instance is generated, and it needs to notify all the subscribed users. So, the application itself doesn't edit the Event table, except during initial installation, and during an update where a new Event might be created. At some point, when an event is generated, the application needs to lookup the Event and get a list of Users. What's the best way to link the event in the source code to the event in the database? Option 1: Store the EventName in the program as a fixed constant, and look it up by name. Option 2: Store the EventId in the program as a static Guid, and look it up by ID. Extra Credit In other similar circumstances I may want to include custom behavior with the event type. That is, I'll want subclasses of my Event entity class with different behaviors, and when I lookup an event, I want it to return an instance of my subclass. For instance: class Event { public Guid Id { get; } public Guid EventName { get; } public ReadOnlyCollection<EventSubscription> EventSubscriptions { get; } public void NotifySubscribers() { foreach(var eventSubscription in EventSubscriptions) { eventSubscription.Notify(); } this.OnSubscribersNotified(); } public virtual void OnSubscribersNotified() {} } class WakingEvent : Event { private readonly IWaker waker; public WakingEvent(IWaker waker) { if(waker == null) throw new ArgumentNullException("waker"); this.waker = waker; } public override void OnSubscribersNotified() { this.waker.Wake(); base.OnSubscribersNotified(); } } So, that means I need to map WakingEvent to whatever key I'm using to look it up in the database. Let's say that's the EventId. Where do I store this relationship? Does it go in the event repository class? Should the WakingEvent know declare its own ID in a static member or method? ...and then, is this all backwards? If all events have a subclass, then instead of retrieving events by ID, should I be asking my repository for the WakingEvent like this: public T GetEvent<T>() where T : Event { ... // what goes here? ... } I can't be the first one to tackle this. What's the best practice?

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• Is there an API for determining congressional districts?

  - by ardavis

  I'm looking to determine the congressional district based on an address my user is providing. This will avoid having the user to look it up themselves. Does an API of this sort exist? Note Through my attempts to find one, I've only come across these: http://www.govtrack.us/developers/api (not sure how to submit an an address or zip code however) The following resources are available in the API ...Bills and resolutions in the U.S. Congress since 1973 (the 93rd Congress). ...A (bill, person) pair indicating cosponsorship, with join and withdrawn dates. ...Members of Congress and U.S. Presidents since the founding of the nation. ...Terms held in office by Members of Congress and U.S. Presidents. Each term corresponds with an election, meaning each term in the House covers two years (one 'Congress'), as President four years, and in the Senate six years (three 'Congresses'). ...Roll call votes in the U.S. Congress since 1789. How people voted is accessed through the Vote_voter API. ...How people voted on roll call votes in the U.S. Congress since 1789. See the Vote API. Filter on the vote field to get the results of a particular vote... http://www.opencongress.org/api (seems to be a way to find congress information, but not districts) This API provides programmers with structured access to all the data on OpenCongress, everything from official bill info to news and blog coverage to user-generated votes on bills and much more... This API defaults to returning XML. All queries can also return JSON... https://groups.google.com/forum/?fromgroups=#!topic/opendems-discuss/CeKyi_aANaE (similar question, no resolution) I've been looking over Open Dems, and seeing what's exposed at this point and what isn't. I work with Democrats Abroad, and am interested in using stuff from the lab for their sites. I quickly looked over the Precinct API, which does both more and less than what I'd need. An ideal resource would be any way of translating addresses into CD at the very least (getting state district data would be good as well), since that would make it easier for DA's membership to make a difference in races like last month's NY26 race... Update I'm looking at the source for the govtrack.us website and the 'doGeoCode' function may be useful. view-source:http://www.govtrack.us/congress/members If no one has any suggestions, I will try to go off of what they are doing.

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• Using the @ in SQL Azure Connections

  - by BuckWoody

  The other day I was working with a client on an application they were changing to a hybrid architecture – some data on-premise and other data in SQL Azure and Windows Azure Blob storage. I had them make a couple of corrections - the first was that all communications to SQL Azure need to be encrypted. It’s a simple addition to the connection string, depending on the library you use. Which brought up another interesting point. They had been using something that looked like this, using the .NET provider: Server=tcp:[serverName].database.windows.net;Database=myDataBase; User ID=LoginName;Password=myPassword; Trusted_Connection=False;Encrypt=True; This includes most of the formatting needed for SQL Azure. It specifies TCP as the transport mechanism, the database name is included, Trusted_Connection is off, and encryption is on. But it needed one more change: Server=tcp:[serverName].database.windows.net;Database=myDataBase; User ID=[LoginName]@[serverName];Password=myPassword; Trusted_Connection=False;Encrypt=True; Notice the difference? It’s the User ID parameter. It includes the @ symbol and the name of the server – not the whole DNS name, just the server name itself. The developers were a bit surprised, since it had been working with the first format that just used the user name. Why did both work, and why is one better than the other? It has to do with the connection library you use. For most libraries, the user name is enough. But for some libraries (subject to change so I don’t list them here) the server name parameter isn’t sent in the way the load balancer understands, so you need to include the server name right in the login, so the system can parse it correctly. Keep in mind, the string limit for that is 128 characters – so take the @ symbol and the server name into consideration for user names. The user connection info is detailed here: http://msdn.microsoft.com/en-us/library/ee336268.aspx Upshot? Include the @servername on your connection string just to be safe. And plan for that extra space…  

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• Problem upgrading 11.04

  - by Krazy_Kaos

  I've been trying to upgrade my ubuntu 11.04 desktop computer, but when I click on the ugrade button: I get this error: I've tryied to change my repositories, but it changes nothing in the error((on the "setting new software channel"). Can someone point me in the right direction? This is my sources.list: # deb http://ppa.launchpad.net/ailurus/ppa/ubuntu karmic main # disabled on upgrade to karmic # deb-src http://ppa.launchpad.net/ailurus/ppa/ubuntu karmic main # disabled on upgrade to karmic # deb cdrom:[Ubuntu 9.04 _Jaunty Jackalope_ - Release i386 (20090421.3)]/ jaunty main restricted # See http://help.ubuntu.com/community/UpgradeNotes for how to upgrade to # newer versions of the distribution. deb http://us.archive.ubuntu.com/ubuntu/ natty main restricted multiverse universe ## Major bug fix updates produced after the final release of the ## distribution. deb http://us.archive.ubuntu.com/ubuntu/ natty-updates main restricted multiverse universe ## N.B. software from this repository is ENTIRELY UNSUPPORTED by the Ubuntu ## team. Also, please note that software in universe WILL NOT receive any ## review or updates from the Ubuntu security team. ## N.B. software from this repository is ENTIRELY UNSUPPORTED by the Ubuntu ## team, and may not be under a free licence. Please satisfy yourself as to ## your rights to use the software. Also, please note that software in ## multiverse WILL NOT receive any review or updates from the Ubuntu ## security team. ## Uncomment the following two lines to add software from the 'backports' ## repository. ## N.B. software from this repository may not have been tested as ## extensively as that contained in the main release, although it includes ## newer versions of some applications which may provide useful features. ## Also, please note that software in backports WILL NOT receive any review ## or updates from the Ubuntu security team. deb-src http://pt.archive.ubuntu.com/ubuntu/ jaunty-backports main restricted universe multiverse ## Uncomment the following two lines to add software from Canonical's ## 'partner' repository. ## This software is not part of Ubuntu, but is offered by Canonical and the ## respective vendors as a service to Ubuntu users. deb http://archive.canonical.com/ubuntu natty partner deb-src http://archive.canonical.com/ubuntu natty partner deb http://us.archive.ubuntu.com/ubuntu/ natty-security main restricted multiverse universe deb http://us.archive.ubuntu.com/ubuntu/ natty-proposed restricted main multiverse universe # deb http://deb.torproject.org/torproject.org karmic main # disabled on upgrade to maverick # deb-src http://deb.torproject.org/torproject.org karmic main # disabled on upgrade to maverick deb http://extras.ubuntu.com/ubuntu natty main #Third party developers repository

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• Quaternion based rotation and pivot position

  - by Michael IV

  I can't figure out how to perform matrix rotation using Quaternion while taking into account pivot position in OpenGL.What I am currently getting is rotation of the object around some point in the space and not a local pivot which is what I want. Here is the code [Using Java] Quaternion rotation method: public void rotateTo3(float xr, float yr, float zr) { _rotation.x = xr; _rotation.y = yr; _rotation.z = zr; Quaternion xrotQ = Glm.angleAxis((xr), Vec3.X_AXIS); Quaternion yrotQ = Glm.angleAxis((yr), Vec3.Y_AXIS); Quaternion zrotQ = Glm.angleAxis((zr), Vec3.Z_AXIS); xrotQ = Glm.normalize(xrotQ); yrotQ = Glm.normalize(yrotQ); zrotQ = Glm.normalize(zrotQ); Quaternion acumQuat; acumQuat = Quaternion.mul(xrotQ, yrotQ); acumQuat = Quaternion.mul(acumQuat, zrotQ); Mat4 rotMat = Glm.matCast(acumQuat); _model = new Mat4(1); scaleTo(_scaleX, _scaleY, _scaleZ); _model = Glm.translate(_model, new Vec3(_pivot.x, _pivot.y, 0)); _model =rotMat.mul(_model);//_model.mul(rotMat); //rotMat.mul(_model); _model = Glm.translate(_model, new Vec3(-_pivot.x, -_pivot.y, 0)); translateTo(_x, _y, _z); notifyTranformChange(); } Model matrix scale method: public void scaleTo(float x, float y, float z) { _model.set(0, x); _model.set(5, y); _model.set(10, z); _scaleX = x; _scaleY = y; _scaleZ = z; notifyTranformChange(); } Translate method: public void translateTo(float x, float y, float z) { _x = x - _pivot.x; _y = y - _pivot.y; _z = z; _position.x = _x; _position.y = _y; _position.z = _z; _model.set(12, _x); _model.set(13, _y); _model.set(14, _z); notifyTranformChange(); } But this method in which I don't use Quaternion works fine: public void rotate(Vec3 axis, float angleDegr) { _rotation.add(axis.scale(angleDegr)); // change to GLM: Mat4 backTr = new Mat4(1.0f); backTr = Glm.translate(backTr, new Vec3(_pivot.x, _pivot.y, 0)); backTr = Glm.rotate(backTr, angleDegr, axis); backTr = Glm.translate(backTr, new Vec3(-_pivot.x, -_pivot.y, 0)); _model =_model.mul(backTr);///backTr.mul(_model); notifyTranformChange(); }

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• How granular should a command be in a CQ[R]S model?

  - by Aaronaught

  I'm considering a project to migrate part of our WCF-based SOA over to a service bus model (probably nServiceBus) and using some basic pub-sub to achieve Command-Query Separation. I'm not new to SOA, or even to service bus models, but I confess that until recently my concept of "separation" was limited to run-of-the-mill database mirroring and replication. Still, I'm attracted to the idea because it seems to provide all the benefits of an eventually-consistent system while sidestepping many of the obvious drawbacks (most notably the lack of proper transactional support). I've read a lot on the subject from Udi Dahan who is basically the guru on ESB architectures (at least in the Microsoft world), but one thing he says really puzzles me: As we get larger entities with more fields on them, we also get more actors working with those same entities, and the higher the likelihood that something will touch some attribute of them at any given time, increasing the number of concurrency conflicts. [...] A core element of CQRS is rethinking the design of the user interface to enable us to capture our users’ intent such that making a customer preferred is a different unit of work for the user than indicating that the customer has moved or that they’ve gotten married. Using an Excel-like UI for data changes doesn’t capture intent, as we saw above. -- Udi Dahan, Clarified CQRS From the perspective described in the quotation, it's hard to argue with that logic. But it seems to go against the grain with respect to SOAs. An SOA (and really services in general) are supposed to deal with coarse-grained messages so as to minimize network chatter - among many other benefits. I realize that network chatter is less of an issue when you've got highly-distributed systems with good message queuing and none of the baggage of RPC, but it doesn't seem wise to dismiss the issue entirely. Udi almost seems to be saying that every attribute change (i.e. field update) ought to be its own command, which is hard to imagine in the context of one user potentially updating hundreds or thousands of combined entities and attributes as it often is with a traditional web service. One batch update in SQL Server may take a fraction of a second given a good highly-parameterized query, table-valued parameter or bulk insert to a staging table; processing all of these updates one at a time is slow, slow, slow, and OLTP database hardware is the most expensive of all to scale up/out. Is there some way to reconcile these competing concerns? Am I thinking about it the wrong way? Does this problem have a well-known solution in the CQS/ESB world? If not, then how does one decide what the "right level" of granularity in a Command should be? Is there some "standard" one can use as a starting point - sort of like 3NF in databases - and only deviate when careful profiling suggests a potentially significant performance benefit? Or is this possibly one of those things that, despite several strong opinions being expressed by various experts, is really just a matter of opinion?

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• Headaches using distributed version control for traditional teams?

  - by J Cooper

  Though I use and like DVCS for my personal projects, and can totally see how it makes managing contributions to your project from others easier (e.g. your typical Github scenario), it seems like for a "traditional" team there could be some problems over the centralized approach employed by solutions like TFS, Perforce, etc. (By "traditional" I mean a team of developers in an office working on one project that no one person "owns", with potentially everyone touching the same code.) A couple of these problems I've foreseen on my own, but please chime in with other considerations. In a traditional system, when you try to check your change in to the server, if someone else has previously checked in a conflicting change then you are forced to merge before you can check yours in. In the DVCS model, each developer checks in their changes locally and at some point pushes to some other repo. That repo then has a branch of that file that 2 people changed. It seems that now someone must be put in charge of dealing with that situation. A designated person on the team might not have sufficient knowledge of the entire codebase to be able to handle merging all conflicts. So now an extra step has been added where someone has to approach one of those developers, tell him to pull and do the merge and then push again (or you have to build an infrastructure that automates that task). Furthermore, since DVCS tends to make working locally so convenient, it is probable that developers could accumulate a few changes in their local repos before pushing, making such conflicts more common and more complicated. Obviously if everyone on the team only works on different areas of the code, this isn't an issue. But I'm curious about the case where everyone is working on the same code. It seems like the centralized model forces conflicts to be dealt with quickly and frequently, minimizing the need to do large, painful merges or have anyone "police" the main repo. So for those of you who do use a DVCS with your team in your office, how do you handle such cases? Do you find your daily (or more likely, weekly) workflow affected negatively? Are there any other considerations I should be aware of before recommending a DVCS at my workplace?

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• DTLoggedExec 1.1.2008.4 Released!

  - by Davide Mauri

  Today I've relased the latest version of my DTExec replacement tool, DTLoggedExec. The main changes are the following: Used a new strategy for version numbers. Now it will follow the following pattern Major.Minor.TargetSQLServerVersion.Revision Added support for Auto Configurations Fixed a bug that reported incorrect number of errors and warnings to Log Providers Fixed a buf that prevented correct casting of values when using /Set and /Param options Errors and Warnings are now counted more precisely. Updated database and log import scripts to categorize logs by projects and sections. E.g.: Project: MyBIProject; Sections: Staging, Datawarehouse Removed unused report stored procedures from database Updated Samples: 12 samples are now available to show ALL DTLoggedExec features From this version only SSIS 2008 will be supported http://dtloggedexec.codeplex.com/releases/view/62218  It useful to say something more on a couple of specific points: From this version only SSIS 2008 will be supportedYes, Integration Services 2005 are not supported anymore. The latest version capable of running SSIS 2005 Packages is the 1.0.0.2. Updated database and log import scripts to categorize logs by projects and sectionsWhen you import a log file, you can now assign it to a Project and to a Section of that project. In this way it's easier to gather statistical information for an entire project or a subsection of it. This also allows to store logged data of package belonging to different projects in the same database. For example:  Updated SamplesA complete set of samples that shows how to use all DTLoggedExec features are now shipped with the product. Enjoy! Added support for Auto ConfigurationsThis point will have a post on its own, since it's quite important and is by far the biggest new feature introduced in this release. To explain it in a few words, I can just say that you don't need to waste time with complex DTS configuration files or options, since a package will configure itself automatically. You just need to write a single statement as a parameter for DTLoggedExec. This feature can simplify deployment *a lot* :)   I the next days I'll write the mentioned post on Auto-Configurations and i'll update the documentation available on theDTLoggedExec website:   http://dtloggedexec.davidemauri.it/MainPage.ashx

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• Attaching two objects and changing their world matrices accordingly

  - by A-Type

  I'm having a hard time wrapping my head around the transformations required to bind two objects together in either a two-way or one-way relationship. I will need to implement both types. For the first case, I want to be able to 'couple' two ships together in space. The ships have different mass, of course. Forces applied to either ship will use combined mass and moment of inertia to calculate and move both ships. The trick is, being sure that the point at which they are coupled remains the same, and they don't move at all relative to each other. The second case is similar: I want a ship to be able to enter the atmosphere of a planet and move relative to the planet. The planet will be orbiting the sun, which is fixed at 0,0,0. Essentially, when the ship is sitting still outside of the atmosphere, the planet will move past it on its course-- but when the ship is sitting still inside the atmosphere, it moves and rotates with the planet, so that it is always relative to the horizon. Essentially, the vertices which make up the ship are now transformed just like the ones that make up the planet, except that the ship can move itself around relative to the planet. I get the feeling I can implement both of these with the same code. Essentially, I am thinking of giving each object (which I call Fixtures) a list of "slave" Fixtures onto which that Fixture's world matrix is imposed. So, this would be the planet imposing its world on any contained ships. In the case of coupling, I would simply make each ship a slave of the other, somehow. Obviously I can't just multiply the ship's world matrix by the planet's, or each ship by the others. What I'd like some help with is what calculations to make in order to get a nice, seamless relative world to the other object. I was thinking maybe I could just multiply the world of the slave by the inverse of the master, but then when you couple two ships you would lose all that world data. So, perhaps I need an intermediate "world" which is the absolute world, but use a secondary "final world" to actually transform the objects?

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• No wireless connection using a conceptronic c54i (RT2561/RT61 rev B)

  - by jrosell

  Detected but not working. New install on ubuntu 11.10 using coneptronic C54Ri. As documentation says it uses Ralink drivers.... Any ideas why my wireless does not work? $ lspci -nn | grep -i 'ralink' 01:05.0 Network controller: Ralink corp. RT2561/RT61 rev B 802.11g ifconfig eth0 Link encap:Ethernet HWaddr 00:1e:90:e5:af:13 inet addr:192.168.0.197 Bcast:192.168.0.255 Mask:255.255.255.0 inet6 addr: fe80::21e:90ff:fee5:af13/64 Scope:Link UP BROADCAST RUNNING MULTICAST MTU:1500 Metric:1 RX packets:28361 errors:0 dropped:0 overruns:0 frame:0 TX packets:16858 errors:0 dropped:0 overruns:0 carrier:0 collisions:0 txqueuelen:1000 RX bytes:39812172 (39.8 MB) TX bytes:1633405 (1.6 MB) Interrupt:43 Base address:0xc000 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:80 errors:0 dropped:0 overruns:0 frame:0 TX packets:80 errors:0 dropped:0 overruns:0 carrier:0 collisions:0 txqueuelen:0 RX bytes:6608 (6.6 KB) TX bytes:6608 (6.6 KB) iwconfig wlan0 wlan0 IEEE 802.11abg ESSIDff/any Mode:Managed Access Point: Not-Associated Tx-Power=0 dBm Retry long limit:7 RTS thrff Fragment thrff Power Managementff lsmod | grep rt rt61pci 27493 0 crc_itu_t 12627 1 rt61pci rt2x00pci 14202 1 rt61pci rt2x00lib 48114 2 rt61pci,rt2x00pci mac80211 272785 2 rt2x00pci,rt2x00lib cfg80211 172392 2 rt2x00lib,mac80211 eeprom_93cx6 12653 1 rt61pci parport_pc 32114 1 parport 40930 3 ppdev,parport_pc,lp lsmod | grep rt [ 2497.816989] phy0 -> rt2x00pci_regbusy_read: Error - Indirect register access failed: offset=0x0000308c, value=0xffffffff [ 2497.827112] phy0 -> rt2x00pci_regbusy_read: Error - Indirect register access failed: offset=0x0000308c, value=0xffffffff [ 2497.837430] phy0 -> rt2x00pci_regbusy_read: Error - Indirect register access failed: offset=0x0000308c, value=0xffffffff [ 2497.847528] phy0 -> rt2x00pci_regbusy_read: Error - Indirect register access failed: offset=0x0000308c, value=0xffffffff [ 2497.847632] phy0 -> rt61pci_wait_bbp_ready: Error - BBP register access faile d, aborting. [ 2497.847637] phy0 -> rt61pci_set_device_state: Error - Device failed to enter state 4 (-5). sudo lshw -C network *-network DISABLED description: Wireless interface product: RT2561/RT61 rev B 802.11g vendor: Ralink corp. physical id: 5 bus info: pci@0000:01:05.0 logical name: wlan0 version: 00 serial: fa:b8:14:58:62:35 width: 32 bits clock: 33MHz capabilities: pm cap_list ethernet physical wireless configuration: broadcast=yes driver=rt61pci driverversion=3.0.0-12-generic firmware=0.8 latency=0 link=no multicast=yes wireless=IEEE 802.11abg resources: irq:16 memory:fdef8000-fdefffff iwlist scan lo Interface doesn't support scanning. eth0 Interface doesn't support scanning. wlan0 Failed to read scan data : Network is down uname -mr 3.0.0-12-generic i686 Edit 1 $ rfkill list all 0: phy0: Wireless LAN Soft blocked: no Hard blocked: no On reboot, sudo lshw -C network returns network is ok. Hovever, WPA keeps on asking the wireless key

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• Behaviour Trees with irregular updates

  - by Robominister

  I'm interested in behaviour trees that aren't iterated every game tick, but every so often. (Edit: the tree could specify how many frames within the main game loop to wait before running its tick function again). Every theoretical implementation I have seen of behaviour trees talks of the tree search being carried out every game update - which seems necessary, because a leaf node (eg a behaviour, like 'return to base') needs to be constantly checked to see if is still running, failed or completed. Can anyone suggest how I might start implementing a tree that isnt run every tick, or point me in the direction of good material specific to this case (I am struggling to find anything)? My thoughts so far: action leaf nodes (when they start) must only push some kind of action object onto a list for an entity, rather than directly calling any code that makes the entity do something. The list of actions for the entity would be run every frame (update any that need to run, pop any that have completed from the list). the return state from a given action must be fed back into the tree, so that when we run the tree iteration again (and reach the same action leaf node - so the tree has so far determined that we ought to still be trying this action) - that the action has completed, or is still running etc. If my actual action code is running from an action list on an entity, then I possibly need to cancel previously running actions in the list - i am thinking that I can just delete the entire stack of queued up actions. I've seen the idea of ActionLists which block lower priority actions when a higher priority one is added, but this seems like very close logic to behaviour trees, and I dont want to be duplicating behaviour. This leaves me with some questions 1) How would I feed the action return state back into the tree? Its obvious I need to store some information relating to 'currently executing actions' on the entity, and check that in the tree tick, but I can't imagine how. 2) Does having a seperate behaviour tree (for deciding behaviour) and action list (for carrying out actual queued up actions) sound like a reasonable approach? 3) Is the approach of updating a behaviour tree irregularly actually used by anyone? It seems like a nice idea for budgeting ai search time when you have a lot of ai entities to process. (Edit) - I am also thinking about storing a single instance of a given behaviour tree in memory, and providing it by reference to any entity that uses it. So any information about what action was last selected for execution on an entity must be stored in a data context relative to the entity (which the tree can check). (I am probably answering my own questions as i go!) I hope I have expressed my questions adequately! Thanks in advance for any help :)

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• PeopleSoft New Design Solves Navigation Problem

  - by Applications User Experience

  Anna Budovsky, User Experience Principal Designer, Applications User Experience In PeopleSoft we strive to improve User Experience on all levels. Simplifying navigation and streamlining access to the most important pages is always an important goal. No one likes to waste time waiting for pages to load and watching a spinning glass going on and on. Those performance-affecting server trips, page-load waits and just-too-many clicks were complained about for a long time. Something had to be done. A few new designs came in PeopleSoft 9.2 helping users to access their everyday work areas easier and faster. For example, Dashboard and Work Center aggregate most accessed information sections on a single page; Related Information allows users to complete transaction-related-research without interrupting a transaction and Secure Search gets users to a specific page directly. Today we’ll talk about the Actions menu. Most PeopleSoft pages are shared between individual products and product lines. It means changing the content on a single page involves Oracle development and quality assurance time for making and testing the changes. In order to streamline the navigation and cut down on accessing PeopleSoft pages one-page-at-a-time, we introduced a new menu design. The new menu allows accessing shared pages without the Oracle development team making any local changes, and it works as an additional one-click-path to specific high-traffic actionable pages. Let’s look at how many steps it took to Change Salary for an employee in HCM 9.1 before: Figure 1. BEFORE: The 6 steps a user would take to Change Salary in PeopleSoft HCM 9.1 In PeopleSoft 9.1 it took 5 steps + page loading time + additional verification time for making sure a correct employee is selected from the table. In PeopleSoft 9.2 it only takes 2 steps. To complete Ad Hoc Change Salary action, the user can start from the HCM Manager's Dashboard, click the Action menu within a table, choose a menu option, and access a correct employee’s details page to take an action. Figure 2. AFTER: The 2 steps a user would take to Change Salary in PeopleSoft HCM 9.2 The new menu is placed on a row level which ensures the user accesses the correct employee’s details page. The Actions menu separates menu options into hierarchical sections which help to scan and access the correct option quickly. The new menu’s small size and its structure enabled users to access high-traffic pages from any page and from any part of the page. No more spinning hourglass, no more multiple pages upload. The flexible design fits anywhere on a page and provides a fast and reliable path to the correct destination within the product. Now users can: Access any target page no matter how far it is buried from the starting point; Reduce navigation and page-load time; Improve productivity and reduce errors. The new menu design is available and widely used in all PeopleSoft 9.2 product lines.

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• What are the differences between abstract classes, interfaces, and when to use them

  - by user66662

  Recently I have started to wrap my head around OOP, and I am now to the point where the more I read about the differences between Abstract classes and Interfaces the more confused I become. So far, neither can be instantiated. Interfaces are more or less structural blueprints that determine the skeleton and abstracts are different by being able to partially develop code. I would like to learn more about these through my specific situation. Here is a link to my first question if you would like a little more background information: What is a good design model for my new class? Here are two classes I created: class Ad { $title; $description $price; function get_data($website){ } function validate_price(){ } } class calendar_event { $title; $description $start_date; function get_data($website){ //guts } function validate_dates(){ //guts } } So, as you can see these classes are almost identical. Not shown here, but there are other functions, like get_zip(), save_to_database() that are common across my classes. I have also added other classes Cars and Pets which have all the common methods and of course properties specific to those objects (mileage, weight, for example). Now I have violated the DRY principle and I am managing and changing the same code across multiple files. I intend on having more classes like boats, horses, or whatever. So is this where I would use an interface or abstract class? From what I understand about abstract classes I would use a super class as a template with all of the common elements built into the abstract class, and then add only the items specifically needed in future classes. For example: abstract class content { $title; $description function get_data($website){ } function common_function2() { } function common_function3() { } } class calendar_event extends content { $start_date; function validate_dates(){ } } Or would I use an interface and, because these are so similar, create a structure that each of the subclasses are forced to use for integrity reasons, and leave it up to the end developer who fleshes out that class to be responsible for each of the details of even the common functions. my thinking there is that some 'common' functions may need to be tweaked in the future for the needs of their specific class. Despite all that above, if you believe I am misunderstanding the what and why of abstracts and interfaces altogether, by all means let a valid answer to be stop thinking in this direction and suggest the proper way to move forward! Thanks!

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