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Exporting and Importing VM Settings with Azure Command-Line Tools
We've talked previously about the Windows Azure command-line tools, and have used them in a few posts such as Brian's Migrating Drupal to a Windows Azure VM. While the tools are generally useful for tons of stuff, one of the things that's been painful to do with the command-line is export the settings for a VM, and then recreate the VM from those settings. You might be wondering why you'd want to export a VM and then recreate it. For me, cost is the first thing that comes to mind. It costs more to keep a VM running than it does to just keep the disk in storage. So if I had something in a VM that I'm only using a few hours a day, I'd delete the VM when I'm not using it and recreate it when I need it again. Another potential reason is that you want to create a copy of the disk so that you can create a duplicate virtual machine. The export process used to be pretty arcane stuff; using the azure vm show command with a --json parameter and piping the output to file. Then hacking the .json file to fix it up so it could be used with the azure vm create-from command. It was bad. It was so bad, the developers added a new export command to create the .json file for you. Here's the basic process: Create a VM VM creation has been covered multiple ways already; you're either going to use the portal or command line tools, and you're either going to select an image from the library or upload a VHD. In my case, I used the following command: azure vm create larryubuntu CANONICAL__Canonical-Ubuntu-12-04-amd64-server-20120528.1.3-en-us-30GB.vhd larry NotaRe This command creates a new VM in the East US data center, enables SSH on port 22 and then stores a disk image for this VM in a blob. You can see the new disk image in blob storage by running: azure vm disk list The results should return something like: info: Executing command vm disk list + Fetching disk images data: Name OS data: ---------------------------------------- ------- data: larryubuntu-larryubuntu-0-20121019170709 Linux info: vm disk list command OK That's the actual disk image that is mounted by the VM. Export and Delete the VM Alright, I've done my work and it's the weekend. I need to export the VM settings so I can recreate it on Monday, then delete the VM so I won't get charged for the next 48 hours of not working. To export the settings for the VM, I use the following command: azure vm export larryubuntu c:\stuff\vminfo.json This tells Windows Azure to find the VM named larryubuntu and export its settings to c:\stuff\vminfo.json. The .json file will contain something like this: { "RoleName":"larryubuntu", "RoleType":"PersistentVMRole", "ConfigurationSets": [ { "ConfigurationSetType":"NetworkConfiguration", "InputEndpoints": [ { "LocalPort":"22", "Name":"ssh", "Port":"22", "Protocol":"tcp", "Vip":"168.62.177.227" } ], "SubnetNames":[] } ], "DataVirtualHardDisks":[], "OSVirtualHardDisk": { "HostCaching":"ReadWrite", "DiskName":"larryubuntu-larryubuntu-0-20121024155441", "OS":"Linux" }, "RoleSize":"Small" } If you're like me, you'll immediately start thinking "Hrmmm, I wonder if I can mess around with things like RoleSize." And yes, you can. If you wanted to bump this up to medium, you'd just change that parameter to medium. If you want to play around more with the various settings, it looks like the schema is maintained at https://github.com/WindowsAzure/azure-sdk-for-node/blob/master/lib/services/serviceManagement/models/roleschema.json. Once I've got the file, I can safely delete the VM by using the following command. azure vm delete larryubuntu It spins a bit and then no more VM. Recreate the VM Ugh, Monday. Time to go back to work, and I need my VM back up and running. So I run the following command: azure vm create-from larryubuntu c:\stuff\vminfo.json --location "East US" It takes only a minute or two to spin up the VM and it's ready for work. That's it - fast, simple, and far easier than the old process of generating the .json settings file. Note that I haven't played around much with the various settings described in the schema for the json file that I linked above. If you find anything useful or interesting that can be accomplished by hacking around with the .json, leave a comment about it.
October 29, 2012
by Larry Franks
· 6,566 Views
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Exploring the HTML5 Web Audio: Visualizing Sound
If you've read some of my other articles on this blog you probably know I'm a fan of HTML5. With HTML5 we get all this interesting functionality, directly in the browser, in a way that, eventually, is standard across browsers. One of the new HTML5 APIs that is slowly moving through the standardization process is the Web Audio API. With this API, currently only supported in Chrome, we get access to all kinds of interesting audio components you can use to create, modify and visualize sounds (such as the following spectrogram). So why do I start with visualizations? It looks nice, that's one reason, but not the important one. This API provides a number of more complex components, whose behavior is much easier to explain when you can see what happens. With a filter you can instantly see whether some frequencies are filtered, instead of trying to listen to the resulting audio for thse changes. There are many interesting examples that use this API. The problem is, though, that getting started with this API and with digital signal processing (DSP) usually isn't explained. In this article I'll walk you through a couple of steps that shows how to do the following: Create a signal volume meter Visualize the frequencies using a spectrum analyzer And show a time based spectrogram We start with the basic setup that we can use as the basis for the components we'll create. Setting up the basic If we want to experiment with sound, we need some sound source. We could use the microphone (as we'll do later in this series), but to keep it simple, for now we'll just use an mp3 as our input. To get this working using web audio we have to take the following steps: Load the data Read it in a buffer node and play the sound Load the data With the web audio we can use different types of audio sources. We've got a MediaElementAudioSourceNode that can be used to use the audio provided by a media element. There's also a MediaStreamAudioSourceNode. With this audio source node we can use the microphone as input (see my previous article on sound recognition). Finally there is the AudioBufferSourceNode. With this node we can load the data from an existing audio file (e.g mp3) and use that as input. For this example we'll use this last approach. // create the audio context (chrome only for now) var context = new webkitAudioContext(); var audioBuffer; var sourceNode; // load the sound setupAudioNodes(); loadSound("wagner-short.ogg"); function setupAudioNodes() { // create a buffer source node sourceNode = context.createBufferSource(); // and connect to destination sourceNode.connect(context.destination); } // load the specified sound function loadSound(url) { var request = new XMLHttpRequest(); request.open('GET', url, true); request.responseType = 'arraybuffer'; // When loaded decode the data request.onload = function() { // decode the data context.decodeAudioData(request.response, function(buffer) { // when the audio is decoded play the sound playSound(buffer); }, onError); } request.send(); } function playSound(buffer) { sourceNode.buffer = buffer; sourceNode.noteOn(0); } // log if an error occurs function onError(e) { console.log(e); } In this example you can see a couple of functions. The setupAudioNodes function creates a BufferSource audio node and connects it to the destination. The loadSound function shows how you can load an audio file. The buffer which is passed into the playSound function contains decoded audio that can be used by the web audio API. In this example I use an .ogg file, for a complete overview of the formats supported look at: https://sites.google.com/a/chromium.org/dev/audio-video Play the sound To play this audio file, all we have to do is turn the source node on, this is done in the playSound function: function playSound(buffer) { sourceNode.buffer = buffer; sourceNode.noteOn(0); } You can test this out at the following page: Example 1: Loading and playing a sound with Web Audio API. When you open that page, you'll hear some music. Nothing to spectacular for now, but nevertheless an easy way to load audio that'll use for the rest of this article. The first item on our list was the volume meter. Create a volume meter One of the basic scenario's, and often one of the first steps someone new to this API tries to create, is a simple signal volume meter (or an UV meter). I expected this to be a standard component in this API, where I could just read off the signal strength as a property. But, no such node exists. But not to worry, with the components that are available, it's pretty easy (not straightforward, but easy nevertheless) to get an indication of the signal strength of your audio file. Int this section we'll create the following simple volume meter: As you can see this is a simple volume meter where we measure the signal strength for the left and the right audio channel. This is drawn on the canvas, but you could have also used divs or svg to visualize this. Lets start with a single volume meter, instead of one for each channel. For this we need to do the following: Create an analyzer node: With this node we get realtime information about the data that is processed. This data we use to determine the signal strength Create a javascript node: We use this node as a timer to update the volume meters with new information Connect everything together Analyser node With the analyser node we can perform real-time frequency and time domain analysis. From the specification: a node which is able to provide real-time frequency and time-domain analysis information. The audio stream will be passed un-processed from input to output. I won't go into the mathematical details behind this node, since there are many articles out there that explain how this works (a good one is the chapter on fourier transformation from here). What you should now about this node is that it splits up the signal in frequency buckets and we get the amplitude (the signal strenght) for each set of frequencies (the bucket). The best way to understand this, is to skip a bit ahead in this article and look at the frequency distribution we'll create later on. This image plots the result from the analyser node. The frequencies increase from left to right, and the height of the bar shows the strength of that specific frequency bucket. More on this later on in the article. For now we don't want to see the strength of the separate frequency buckets, but the strength of the total signal. For this we'll just add all the strenghts from each bucket and divide it by the number of buckets. First we need to create an analyzer node // setup a analyzer analyser = context.createAnalyser(); analyser.smoothingTimeConstant = 0.3; analyser.fftSize = 1024; This creates an analyzer node whose result will be used to create the volume meter. We use a smoothingTimeConstant to make the meter less jittery. With this variable we use input from a longer time period to calculate the amplitudes, this results in a more smooth meter. The fftSize determine how many buckets we get containing frequency information. If we have a fftSize of 1024 we get 512 buckets (more info on this in the book on DPS and fourier transformations). When this node receives a stream of data, it analyzes this stream and provides us with information about the frequencies in that signal and their strengths. We now need a timer to update the meter at regular intervals. We could use the standard javascript setInterval function, but since we're looking at the Web Audio API lets use one of its nodes. The JavaScriptNode. The javascript node With the javascriptnode we can process the raw audio data directly from javascript. We can use this to write our own analyzers or complex components. We're not going to do that, though. When creating the javascript node, you can specify the interval at which it is called. We'll use that feature to update the meter at regulat intervals. Creating a javascript node is very easy. // setup a javascript node javascriptNode = context.createJavaScriptNode(2048, 1, 1); This will create a javascriptnode that is called whenever the 2048 frames have been sampled. Since our data is sampled at 44.1k, this function will be called approximately 21 times a second. Now what happens when this function is called: // when the javascript node is called // we use information from the analyzer node // to draw the volume javascriptNode.onaudioprocess = function() { // get the average, bincount is fftsize / 2 var array = new Uint8Array(analyser.frequencyBinCount); analyser.getByteFrequencyData(array); var average = getAverageVolume(array) // clear the current state ctx.clearRect(0, 0, 60, 130); // set the fill style ctx.fillStyle=gradient; // create the meters ctx.fillRect(0,130-average,25,130); } function getAverageVolume(array) { var values = 0; var average; var length = array.length; // get all the frequency amplitudes for (var i = 0; i < length; i++) { values += array[i]; } average = values / length; return average; } In these two functions we calculate the average and draw the meter directly on the canvas (using a gradient so we have nice colors). Now all we have to do is connect the output from the audiosource to the analyser, the analyser to the javasource node (and if we want audio to hear, we also need to connect something to the destionation). Connect everything together Connecting everything together is easy: function setupAudioNodes() { // setup a javascript node javascriptNode = context.createJavaScriptNode(2048, 1, 1); // connect to destination, else it isn't called javascriptNode.connect(context.destination); // setup a analyzer analyser = context.createAnalyser(); analyser.smoothingTimeConstant = 0.3; analyser.fftSize = 1024; // create a buffer source node sourceNode = context.createBufferSource(); // connect the source to the analyser sourceNode.connect(analyser); // we use the javascript node to draw at a specific interval. analyser.connect(javascriptNode); // and connect to destination, if you want audio sourceNode.connect(context.destination); } And that's it. This will draw a single volume meter, for the complete signal. Now what do we do when we want to have a volume meter for each channel. For this we use a ChannelSplitter. Let's dive right into the code to connect everything: function setupAudioNodes() { // setup a javascript node javascriptNode = context.createJavaScriptNode(2048, 1, 1); // connect to destination, else it isn't called javascriptNode.connect(context.destination); // setup a analyzer analyser = context.createAnalyser(); analyser.smoothingTimeConstant = 0.3; analyser.fftSize = 1024; analyser2 = context.createAnalyser(); analyser2.smoothingTimeConstant = 0.0; analyser2.fftSize = 1024; // create a buffer source node sourceNode = context.createBufferSource(); splitter = context.createChannelSplitter(); // connect the source to the analyser and the splitter sourceNode.connect(splitter); // connect one of the outputs from the splitter to // the analyser splitter.connect(analyser,0,0); splitter.connect(analyser2,1,0); // we use the javascript node to draw at a // specific interval. analyser.connect(javascriptNode); // and connect to destination sourceNode.connect(context.destination); } As you can see we don't really change much. We introduce a new node, the splitter node. This node splits the sound into a left and a right channel. These channels can be processed separately. With this layout the following happens: The audiosource creates a signal based on the buffered audio. This signal is sent to the splitter, who splits the signal into a left and right stream. Each of these two streams is processed by their own realtime analyser. From the javascript node, we now get the information from both analysers and plot both meters I've shown step 1 through 3, let's quickly move on the step 4. For this we simply add the following to the onaudioprocess node: javascriptNode.onaudioprocess = function() { // get the average for the first channel var array = new Uint8Array(analyser.frequencyBinCount); analyser.getByteFrequencyData(array); var average = getAverageVolume(array); // get the average for the second channel var array2 = new Uint8Array(analyser2.frequencyBinCount); analyser2.getByteFrequencyData(array2); var average2 = getAverageVolume(array2); // clear the current state ctx.clearRect(0, 0, 60, 130); // set the fill style ctx.fillStyle=gradient; // create the meters ctx.fillRect(0,130-average,25,130); ctx.fillRect(30,130-average2,25,130); } And now we've got two signal meters, one for each channel. Example 2: Visualize the signal strength with a volume meter. Or view the result on youtube: Now lets see how we can get the view of the frequencies I showed earlier. Create a frequency spectrum With all the work we already did in the previous section, creating a frequency spectrum overview is now very easy. We're going to aim for this: We set up the nodes just like we did in the first example: function setupAudioNodes() { // setup a javascript node javascriptNode = context.createJavaScriptNode(2048, 1, 1); // connect to destination, else it isn't called javascriptNode.connect(context.destination); // setup a analyzer analyser = context.createAnalyser(); analyser.smoothingTimeConstant = 0.3; analyser.fftSize = 512; // create a buffer source node sourceNode = context.createBufferSource(); sourceNode.connect(analyser); analyser.connect(javascriptNode); // sourceNode.connect(context.destination); } So this time we don't split the channels and we set the fftSize to 512. This means we get 256 bars that represent our frequency. We now just need to alter the onaudioprocess method and the gradient we use: var gradient = ctx.createLinearGradient(0,0,0,300); gradient.addColorStop(1,'#000000'); gradient.addColorStop(0.75,'#ff0000'); gradient.addColorStop(0.25,'#ffff00'); gradient.addColorStop(0,'#ffffff'); // when the javascript node is called // we use information from the analyzer node // to draw the volume javascriptNode.onaudioprocess = function() { // get the average for the first channel var array = new Uint8Array(analyser.frequencyBinCount); analyser.getByteFrequencyData(array); // clear the current state ctx.clearRect(0, 0, 1000, 325); // set the fill style ctx.fillStyle=gradient; drawSpectrum(array); } function drawSpectrum(array) { for ( var i = 0; i < (array.length); i++ ){ var value = array[i]; ctx.fillRect(i*5,325-value,3,325); } }; In the drawSpectrum function we iterate over the array, and draw a vertical bar based on the value. That's it. For a live example, click on the following link: Example 3: Visualize the frequency spectrum. Or view it on youtube: And then the final one. The spectrogram. Time based spectrogram When you run the previous demo you see the strength of the various frequency buckets in real time. While this is a nice visualization, it doesn't allow you to analyze information over a period of time. If you want to do that you can create a spectrogram. With a spectrogram we plot a single line for each measurement. The y-axis represents the frequency, the x-asis the time and the color of a pixel the strength of that frequency. It can be used to analyze the received audio, and also creates nice looking images. The good thing, is that to output this data we don't have to change much from what we've already got in place. The only function that'll change is the onaudioprocess node and we'll create a slightly different analyser. analyser = context.createAnalyser(); analyser.smoothingTimeConstant = 0; analyser.fftSize = 1024; The enalyser we create here has an fftSize of 1024, this means we get 512 frequency buckets with strengths. So we can draw a spectrogram that has a height of 512 pixels. Also note that the smoothingTimeConstant is set to 0. This means we don't use any of the previous results in the analysis. We want to show the real information, not provide a smooth volume meter or frequency spectrum analysis. The easiest way to draw a spectrogram is by just start drawing the line at the left, and for each new set of frequencies increase the x-coordinate by one. The problem is that this will quickly fill up our canvas, and we'll only be able to see the first half a minute of the audio. To fix this, we need some creative canvas copying. The complete code for drawing the spectrogram is shown here: // create a temp canvas we use for copying and scrolling var tempCanvas = document.createElement("canvas"), tempCtx = tempCanvas.getContext("2d"); tempCanvas.width=800; tempCanvas.height=512; // used for color distribution var hot = new chroma.ColorScale({ colors:['#000000', '#ff0000', '#ffff00', '#ffffff'], positions:[0, .25, .75, 1], mode:'rgb', limits:[0, 300] }); ... // when the javascript node is called // we use information from the analyzer node // to draw the volume javascriptNode.onaudioprocess = function () { // get the average for the first channel var array = new Uint8Array(analyser.frequencyBinCount); analyser.getByteFrequencyData(array); // draw the spectrogram if (sourceNode.playbackState == sourceNode.PLAYING_STATE) { drawSpectrogram(array); } } function drawSpectrogram(array) { // copy the current canvas onto the temp canvas var canvas = document.getElementById("canvas"); tempCtx.drawImage(canvas, 0, 0, 800, 512); // iterate over the elements from the array for (var i = 0; i < array.length; i++) { // draw each pixel with the specific color var value = array[i]; ctx.fillStyle = hot.getColor(value).hex(); // draw the line at the right side of the canvas ctx.fillRect(800 - 1, 512 - i, 1, 1); } // set translate on the canvas ctx.translate(-1, 0); // draw the copied image ctx.drawImage(tempCanvas, 0, 0, 800, 512, 0, 0, 800, 512); // reset the transformation matrix ctx.setTransform(1, 0, 0, 1, 0, 0); } To draw the spectrogram we do the following: We copy what is currently drawn to a hidden canvas Next we draw a line of the current values at the far right of the canvas We set the translate on the canvas to -1 We copy the copied information back to the original canvas (that is now drawn 1 pixel to the left) And reset the transformation matrix See a running example here: Example 4: Create a spectrogram Or view it here: Last thing I'd like to mention regarding the code is the chroma.js library I used for the colors. If you ever need to draw something color or gradient related (e.g maps, strengths, levels) you can easily create color scales with this library. Two final pointers, I know I'll get questions about: Volume could be represented as a magnitude, just didn't want to complicate matters for this. The spectogram doesn't use logarithmic scales. Once again, didn't want to complicate things
October 23, 2012
by Jos Dirksen
· 70,385 Views · 1 Like
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Understanding JVM Internals, from Basic Structure to Java SE 7 Features
Learn about the structure of JVM, how it works, executes Java bytecode, the order of execution, examples of common mistakes and their solutions, new Java SE 7 features.
October 19, 2012
by Esen Sagynov
· 180,476 Views · 20 Likes
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PartitionKey and RowKey in Windows Azure Table Storage
For the past few months, I’ve been coaching a “Microsoft Student Partner” (who has a great blog on Kinect for Windows by the way!) on Windows Azure. One of the questions he recently had was around PartitionKey and RowKey in Windows Azure Table Storage. What are these for? Do I have to specify them manually? Let’s explain… Windows Azure storage partitions All Windows Azure storage abstractions (Blob, Table, Queue) are built upon the same stack (whitepaper here). While there’s much more to tell about it, the reason why it scales is because of its partitioning logic. Whenever you store something on Windows Azure storage, it is located on some partition in the system. Partitions are used for scale out in the system. Imagine that there’s only 3 physical machines that are used for storing data in Windows Azure storage: Based on the size and load of a partition, partitions are fanned out across these machines. Whenever a partition gets a high load or grows in size, the Windows Azure storage management can kick in and move a partition to another machine: By doing this, Windows Azure can ensure a high throughput as well as its storage guarantees. If a partition gets busy, it’s moved to a server which can support the higher load. If it gets large, it’s moved to a location where there’s enough disk space available. Partitions are different for every storage mechanism: In blob storage, each blob is in a separate partition. This means that every blob can get the maximal throughput guaranteed by the system. In queues, every queue is a separate partition. In tables, it’s different: you decide how data is co-located in the system. PartitionKey in Table Storage In Table Storage, you have to decide on the PartitionKey yourself. In essence, you are responsible for the throughput you’ll get on your system. If you put every entity in the same partition (by using the same partition key), you’ll be limited to the size of the storage machines for the amount of storage you can use. Plus, you’ll be constraining the maximal throughput as there’s lots of entities in the same partition. Should you set the PartitionKey to the same value for every entity stored? No. You’ll end up with scaling issues at some point. Should you set the PartitionKey to a unique value for every entity stored? No. You can do this and every entity stored will end up in its own partition, but you’ll find that querying your data becomes more difficult. And that’s where our next concept kicks in… RowKey in Table Storage A RowKey in Table Storage is a very simple thing: it’s your “primary key” within a partition. PartitionKey + RowKey form the composite unique identifier for an entity. Within one PartitionKey, you can only have unique RowKeys. If you use multiple partitions, the same RowKey can be reused in every partition. So in essence, a RowKey is just the identifier of an entity within a partition. PartitionKey and RowKey and performance Before building your code, it’s a good idea to think about both properties. Don’t just assign them a guid or a random string as it does matter for performance. The fastest way of querying? Specifying both PartitionKey and RowKey. By doing this, table storage will immediately know which partition to query and can simply do an ID lookup on RowKey within that partition. Less fast but still fast enough will be querying by specifying PartitionKey: table storage will know which partition to query. Less fast: querying on only RowKey. Doing this will give table storage no pointer on which partition to search in, resulting in a query that possibly spans multiple partitions, possibly multiple storage nodes as well. Wihtin a partition, searching on RowKey is still pretty fast as it’s a unique index. Slow: searching on other properties (again, spans multiple partitions and properties). Note that Windows Azure storage may decide to group partitions in so-called "Range partitions" - see http://msdn.microsoft.com/en-us/library/windowsazure/hh508997.aspx. In order to improve query performance, think about your PartitionKey and RowKey upfront, as they are the fast way into your datasets. Deciding on PartitionKey and RowKey Here’s an exercise: say you want to store customers, orders and orderlines. What will you choose as the PartitionKey (PK) / RowKey (RK)? Let’s use three tables: Customer, Order and Orderline. An ideal setup may be this one, depending on how you want to query everything: Customer (PK: sales region, RK: customer id) – it enables fast searches on region and on customer id Order (PK: customer id, RK; order id) – it allows me to quickly fetch all orders for a specific customer (as they are colocated in one partition), it still allows fast querying on a specific order id as well) Orderline (PK: order id, RK: order line id) – allows fast querying on both order id as well as order line id. Of course, depending on the system you are building, the following may be a better setup: Customer (PK: customer id, RK: display name) – it enables fast searches on customer id and display name Order (PK: customer id, RK; order id) – it allows me to quickly fetch all orders for a specific customer (as they are colocated in one partition), it still allows fast querying on a specific order id as well) Orderline (PK: order id, RK: item id) – allows fast querying on both order id as well as the item bought, of course given that one order can only contain one order line for a specific item (PK + RK should be unique) You see? Choose them wisely, depending on your queries. And maybe an important sidenote: don’t be afraid of denormalizing your data and storing data twice in a different format, supporting more query variations. There’s one additional “index” That’s right! People have been asking Microsoft for a secondary index. And it’s already there… The table name itself! Take our customer – order – orderline sample again… Having a Customer table containing all customers may be interesting to search within that data. But having an Orders table containing every order for every customer may not be the ideal solution. Maybe you want to create an order table per customer? Doing that, you can easily query the order id (it’s the table name) and within the order table, you can have more detail in PK and RK. And there's one more: your account name. Split data over multiple storage accounts and you have yet another "partition". Conclusion In conclusion? Choose PartitionKey and RowKey wisely. The more meaningful to your application or business domain, the faster querying will be and the more efficient table storage will work in the long run.
October 19, 2012
by Maarten Balliauw
· 57,926 Views · 10 Likes
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Debugging Hibernate Envers - Historical Data
recently in our project we reported a strange bug. in one report where we display historical data provided by hibernate envers , users encountered duplicated records in the dropdown used for filtering. we tried to find the source of this bug, but after spending a few hours looking at the code responsible for this functionality we had to give up and ask for a dump from production database to check what actually is stored in one table. and when we got it and started investigating, it turned out that there is a bug in hibernate envers 3.6 that is a cause of our problems. but luckily after some investigation and invaluable help from adam warski (author of envers) we were able to fix this issue. bug itself let’s consider following scenario: a transaction is started. we insert some audited entities during it and then it is rolled back. the same entitymanager is reused to start another transaction second transaction is committed but when we check audit tables for entities that were created and then rolled back in step one, we will notice that they are still there and were not rolled back as we expected. we were able to reproduce it in a failing test in our project, so the next step was to prepare failing test in envers so we could verify if our fix is working. failing test the simplest test cases already present in envers are located in simple.java class and they look quite straightforward: public class simple extends abstractentitytest { private integer id1; public void configure(ejb3configuration cfg) { cfg.addannotatedclass(inttestentity.class); } @test public void initdata() { entitymanager em = getentitymanager(); em.gettransaction().begin(); inttestentity ite = new inttestentity(10); em.persist(ite); id1 = ite.getid(); em.gettransaction().commit(); em.gettransaction().begin(); ite = em.find(inttestentity.class, id1); ite.setnumber(20); em.gettransaction().commit(); } @test(dependsonmethods = "initdata") public void testrevisionscounts() { assert arrays.aslist(1, 2).equals(getauditreader().getrevisions(inttestentity.class, id1)); } @test(dependsonmethods = "initdata") public void testhistoryofid1() { inttestentity ver1 = new inttestentity(10, id1); inttestentity ver2 = new inttestentity(20, id1); assert getauditreader().find(inttestentity.class, id1, 1).equals(ver1); assert getauditreader().find(inttestentity.class, id1, 2).equals(ver2); } } so preparing my failing test executing scenario described above wasn’t a rocket science: /** * @author tomasz dziurko (tdziurko at gmail dot com) */ public class transactionrollbackbehaviour extends abstractentitytest { public void configure(ejb3configuration cfg) { cfg.addannotatedclass(inttestentity.class); } @test public void testauditrecordsrollback() { // given entitymanager em = getentitymanager(); em.gettransaction().begin(); inttestentity itetorollback = new inttestentity(30); em.persist(itetorollback); integer rollbackediteid = itetorollback.getid(); em.gettransaction().rollback(); // when em.gettransaction().begin(); inttestentity ite2 = new inttestentity(50); em.persist(ite2); integer ite2id = ite2.getid(); em.gettransaction().commit(); // then list revisionsforsavedclass = getauditreader().getrevisions(inttestentity.class, ite2id); assertequals(revisionsforsavedclass.size(), 1, "there should be one revision for inserted entity"); list revisionsforrolledbackclass = getauditreader().getrevisions(inttestentity.class, rollbackediteid); assertequals(revisionsforrolledbackclass.size(), 0, "there should be no revisions for insert that was rolled back"); } } now i could verify that tests are failing on the forked 3.6 branch and check if the fix that we had is making this test green. the fix after writing a failing test in our project, i placed several breakpoints in envers code to understand better what is wrong there. but imagine being thrown in a project developed for a few years by many programmers smarter than you. i felt overwhelmed and had no idea where the fix should be applied and what exactly is not working as expected. luckily in my company we have adam warski on board. he is the initial author of envers and actually he pointed us the solution. the fix itself contains only one check that registers audit processes that will be executed on transaction completion only when such processes iare still in the map for the given transaction. it sounds complicated, but if you look at the class auditprocessmanager in this commit it should be more clear what is happening there. official path besides locating a problem and fixing it, there are some more official steps that must be performed to have fix included in envers. step 1. create jira issue with bug - https://hibernate.onjira.com/browse/hhh-7682 step 2: create local branch envers-bugfix-hhh-7682 of forked hibernate 3.6 step 3: commit and push failing test and fix to your local and remote repository on github step 4: create pull request - https://github.com/hibernate/hibernate-orm/pull/393 step 5: wait for merge and that’s all. now fix is merged into main repository and we have one bug less in the world of open source
October 17, 2012
by Tomasz Dziurko
· 7,955 Views
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Create a Java App Server on a Virtual Machine
Curator's note: This tutorial originally appeared at the Windows Azure Java Developer Center. With Windows Azure, you can use a virtual machine to provide server capabilities. As an example, a virtual machine running on Windows Azure can be configured to host a Java application server, such as Apache Tomcat. On completing this guide, you will have an understanding of how to create a virtual machine running on Windows Azure and configure it to run a Java application server. You will learn: How to create a virtual machine. How to remotely log in to your virtual machine. How to install a JDK on your virtual machine. How to install a Java application server on your virtual machine. How to create an endpoint for your virtual machine. How to open a port in the firewall for your application server. For purposes of this tutorial, an Apache Tomcat application server will be installed on a virtual machine. The completed installation will result in a Tomcat installation such as the following. Note To complete this tutorial, you need a Windows Azure account that has the Windows Azure Virtual Machines feature enabled. You can create a free trial account and enable preview features in just a couple of minutes. For details, see Create a Windows Azure account and enable preview features. To create a virtual machine Log in to the Windows Azure Preview Management Portal. Click New. Click Virtual machine. Click Quick create. In the Create virtual machine screen, enter a value for DNS name. From the Image dropdown list, select an image, such as Windows Server 2008 R2 SP1. Enter a password in the New password field, and re-enter it in the Confirm field. This is the Administrator account password. Remember this password, you will use it when you remotely log in to the virtual machine. From the Location drop down list, select the data center location for your virtual machine; for example, West US. Your screen will look similar to the following. Click Create virtual machine. Your virtual machine will be created. You can monitor the status in the Virtual machines section of the management portal. To remotely log in to your virtual machine Log in to the Preview Management Portal. Click Virtual Machines, and then select the MyTestVM1 virtual machine that you previously created. On the command bar, click Connect. Click Open to use the remote desktop protocol file that was automatically created for the virtual machine Click Connect to proceed with the connection process. Type the password that you specified as the password of the Administrator account when you created the virtual machine, and then click OK. Click Yes to verify the identity of the virtual machine. To install a JDK on your virtual machine You can copy a Java Developer Kit (JDK) to your virtual machine, or install a JDK through an installer. For purposes of this tutorial, a JDK will be installed from Oracle's site. Log in to your virtual machine. Within your browser, open http://www.oracle.com/technetwork/java/javase/downloads/index.html. Click the Download button for the JDK that you want to download. For purposes of this tutorial, the Download button for the Java SE 6 Update 32 JDK was used. Accept the license agreement. Click the download executable for Windows x64 (64-bit). Follow the prompts and respond as needed to install the JDK to your virtual machine. To install a Java application server on your virtual machine You can copy a Java application server to your virtual machine, or install a Java application server through an installer. For purposes of this tutorial, a Java application server will be installed by copying a zip file from Apache's site. Log in to your virtual machine. Within your browser, open http://tomcat.apache.org/download-70.cgi. Double-click 64-bit Windows zip. (This tutorial used the zip for Tomcat Apache 7.0.27.) When prompted, choose to save the zip. When the zip is saved, open the folder that contains the zip and double-click the zip. Extract the zip. For purposes of this tutorial, the path used was C:\program files\apache-tomcat-7.0.27-windows-x64. To run the Java application server privately on your virtual machine The following steps show you how to run the Java application server and test it within the virtual machine's browser. It won't be usable by external computers until you create an endpoint and open a port (those steps are described later). Log in to your virtual machine. Add the JDK bin folder to the Pathenvironment variable: Click Windows Start. Right-click Computer. Click Properties. Click Advanced system settings. Click Advanced. Click Environment variables. In the System variables section, click the Path variable and then click Edit. Add a trailing ; to the Path variable value (if there is not one already) and then add c:\program files\java\jdk\bin to the end of the Path variable value (adjust the path as needed if you did not use c:\program files\java\jdk as the path for your JDK installation). Press OK on the opened dialogs to save your Path change. Set the JAVA_HOMEenvironment variable: Click Windows Start. Right-click Computer. Click Properties. Click Advanced system settings. Click Advanced. Click Environment variables. In the System variables section, click New. Create a variable named JRE_HOME and set its value to c:\program files\java\jdk\jre (adjust the path as needed if you did not use c:\program files\java\jdk as the path for your JDK installation). Press OK on the open dialogs to save your JRE_HOME environment variable. Start Tomcat: Open a command prompt. Change the current directory to the Apache Tomcat binfolder. For example: cd c:\program files\apache-tomcat-7.0.27-windows-x64\apache-tomcat-7.0.27\bin (Adjust the path as needed if you used a differrent installation path for Tomcat.) Run catalina.bat start. You should now see Tomcat running if you run the virtual machine's browser and open http://localhost:8080. To see Tomcat running from external machines, you'll need to create an endpoint and open a port. To create an endpoint for your virtual machine Log in to the Preview Management Portal. Click Virtual machines. Click the name of the virtual machine that is running your Java application server. Click Endpoints. Click Add endpoint. In the Add endpoint dialog, ensure Add endpoint is checked and click the Next button. In the New endpoint detailsdialog Specify a name for the endpoint; for example, HttpIn. Specify TCP for the protocol. Specify 80 for the public port. Specify 8080for the private port. Your screen should look similar to the following: Click the Check button to close the dialog. Your endpoint will now be created. To open a port in the firewall for your virtual machine Log in to your virtual machine. Click Windows Start. Click Control Panel. Click System and Security, click Windows Firewall, and then click Advanced Settings. Click Inbound Rules and then click New Rule. For the new rule, select Port for the Rule type and click Next. Select TCP for the protocol and specify 8080 for the port, and click Next. Choose Allow the connection and click Next. Ensure Domain, Private, and Public are checked for the profile and click Next. Specify a name for the rule, such as HttpIn (the rule name is not required to match the endpoint name, however), and then click Finish. At this point, your Tomcat web site should now be viewable from an external browser, using a URL of the form http://your_DNS_name.cloudapp.net, where your_DNS_name is the DNS name you specified when you created the virtual machine. Application lifecycle considerations You could create your own application web archive (WAR) and add it to the webapps folder. For example, create a basic Java Service Page (JSP) dynamic web project and export it as a WAR file, copy the WAR to the Apache Tomcat webapps folder on the virtual machine, then run it in a browser. This tutorial runs Tomcat through a command prompt where catalina.bat start was called. You may instead want to run Tomcat as a service, a key benefit being to have it automatically start if the virtual machine is rebooted. To run Tomcat as a service, you can install it as a service via the service.bat file in the Apache Tomcat bin folder, and then you could set it up to run automatically via the Services snap-in. You can start the Services snap-in by clicking Windows Start, Administrative Tools, and then Services. If you run service.bat install MyTomcat in the Apache Tomcat bin folder, then within the Services snap-in, your service name will appear as Apache Tomcat MyTomcat. By default when the service is installed, it will be set to start manually. To set it to start automatically, double-click the service in the Services snap-in and set Startup Type to Automatic, as shown in the following. You'll need to start the service the first time, which you can do through the Services snap-in (alternatively, you can reboot the virtual machine). Close the running occurrence of catalina.bat start if it is still running before starting the service.
October 15, 2012
by Eric Gregory
· 31,627 Views
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Implementing Repository Pattern with Entity Framework
When working with Entity Framework - Code First model approach, a developer creates POCO entities for database tables. The benefit of using Code First model is to have POCO entity for each table that can be used as either WCF Data Contracts or you can apply your own custom attributes to handle Security, Logging, etc. and there is no mapping needed as we used to do in Entity Framework (Model First) approach if the application architecture is n-tier based. Considering the Data Access layer, we will implement a repository pattern that encapsulates the persistence logic in a separate class. This class will be responsible to perform database operations. Let’s suppose the application is based on n-tier architecture and having 3 tiers namely Presentation, Business and Data Access. Common library contains all our POCO entities that will be used by all the layers. Presentation Layer: Contains Views, Forms Business Layer: Managers that handle logic functionality Data Access Layer: Contains Repository class that handles CRUD operations Common Library: Contain POCO entities. We will implement an interface named “IRepository” that defines the signature of all the appropriate generic methods needed to perform CRUD operation and then implement the Repository class that defines the actual implementation of each method. We can also instantiate Repository object using Dependency Injection or apply Factory pattern. Code Snippet: IRepository public interface IRepository : IDisposable { /// /// Gets all objects from database /// /// IQueryable All() where T : class; /// /// Gets objects from database by filter. /// /// Specified a filter /// IQueryable Filter(Expression> predicate) where T : class; /// /// Gets objects from database with filting and paging. /// /// /// Specified a filter /// Returns the total records count of the filter. /// Specified the page index. /// Specified the page size /// IQueryable Filter(Expression> filter, out int total, int index = 0, int size = 50) where T : class; /// /// Gets the object(s) is exists in database by specified filter. /// /// Specified the filter expression /// bool Contains(Expression> predicate) where T : class; /// /// Find object by keys. /// /// Specified the search keys. /// T Find(params object[] keys) where T : class; /// /// Find object by specified expression. /// /// /// T Find(Expression> predicate) where T : class; /// /// Create a new object to database. /// /// Specified a new object to create. /// T Create(T t) where T : class; /// /// Delete the object from database. /// /// Specified a existing object to delete. int Delete(T t) where T : class; /// /// Delete objects from database by specified filter expression. /// /// /// int Delete(Expression> predicate) where T : class; /// /// Update object changes and save to database. /// /// Specified the object to save. /// int Update(T t) where T : class; /// /// Select Single Item by specified expression. /// /// /// /// T Single(Expression> expression) where T : class; void SaveChanges(); void ExecuteProcedure(String procedureCommand, params SqlParameter[] sqlParams); } Code Snippet: Repository public class Repository : IRepository { DbContext Context; public Repository() { Context = new DBContext(); } public Repository(DBContext context) { Context = context; } public void CommitChanges() { Context.SaveChanges(); } public T Single(Expression> expression) where T : class { return All().FirstOrDefault(expression); } public IQueryable All() where T : class { return Context.Set().AsQueryable(); } public virtual IQueryable Filter(Expression> predicate) where T : class { return Context.Set().Where(predicate).AsQueryable(); } public virtual IQueryable Filter(Expression> filter, out int total, int index = 0, int size = 50) where T : class { int skipCount = index * size; var _resetSet = filter != null ? Context.Set().Where(filter).AsQueryable() : Context.Set().AsQueryable(); _resetSet = skipCount == 0 ? _resetSet.Take(size) : _resetSet.Skip(skipCount).Take(size); total = _resetSet.Count(); return _resetSet.AsQueryable(); } public virtual T Create(T TObject) where T : class { var newEntry = Context.Set().Add(TObject); Context.SaveChanges(); return newEntry; } public virtual int Delete(T TObject) where T : class { Context.Set().Remove(TObject); return Context.SaveChanges(); } public virtual int Update(T TObject) where T : class { try { var entry = Context.Entry(TObject); Context.Set().Attach(TObject); entry.State = EntityState.Modified; return Context.SaveChanges(); } catch (OptimisticConcurrencyException ex) { throw ex; } } public virtual int Delete(Expression> predicate) where T : class { var objects = Filter(predicate); foreach (var obj in objects) Context.Set().Remove(obj); return Context.SaveChanges(); } public bool Contains(Expression> predicate) where T : class { return Context.Set().Count(predicate) > 0; } public virtual T Find(params object[] keys) where T : class { return (T)Context.Set().Find(keys); } public virtual T Find(Expression> predicate) where T : class { return Context.Set().FirstOrDefault(predicate); } public virtual void ExecuteProcedure(String procedureCommand, params SqlParameter[] sqlParams){ Context.Database.ExecuteSqlCommand(procedureCommand, sqlParams); } public virtual void SaveChanges() { Context.SaveChanges(); } public void Dispose() { if (Context != null) Context.Dispose(); } } The benefit of using Repository pattern is that all the database operations will be managed centrally and in future if you want to change the underlying database connector you can add another Repository class and defines its own implementation or change the existing one.
October 13, 2012
by Ovais Mehboob Ahmed Khan
· 31,955 Views
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Bug Fixing: To Estimate, or Not to Estimate: That is The Question
According to Steve McConnell in Code Complete (data from 1975-1992) most bugs don’t take long to fix. About 85% of errors can be fixed in less than a few hours. Some more can be fixed in a few hours to a few days. But the rest take longer, sometimes much longer – as I talked about in an earlier post. Given all of these factors and uncertainty, how to you estimate a bug fix? Or should you bother? Block out some time for bug fixing Some teams don’t estimate bug fixes upfront. Instead they allocate a block of time, some kind of buffer for bug fixing as a regular part of the team’s work, especially if they are working in time boxes. Developers come back with an estimate only if it looks like the fix will require a substantial change – after they’ve dug into the code and found out that the fix isn’t going to be easy, that it may require a redesign or require changes to complex or critical code that needs careful review and testing. Use a rule of thumb placeholder for each bug fix Another approach is to use a rough rule of thumb, a standard place holder for every bug fix. Estimate ½ day of development work for each bug, for example. According to this post on Stack Overflow the ½ day suggestion comes from Jeff Sutherland, one of the inventors of Scrum. This place holder should work for most bugs. If it takes a developer more than ½ day to come up with a fix, then they probably need help and people need to know anyways. Pick a place holder and use it for a while. If it seems too small or too big, change it. Iterate. You will always have bugs to fix. You might get better at fixing them over time, or they might get harder to find and fix once you’ve got past the obvious ones. Or you could use the data earlier from Capers Jones on how long it takes to fix a bug by the type of bug. A day or half day works well on average, especially since most bugs are coding bugs (on average 3 hours) or data bugs (6.5 hours). Even design bugs on average only take little more than a day to resolve. Collect some data – and use it Steve McConnell, In Software Estimation: Demystifying the Black Art says that it’s always better to use data than to guess. He suggests collecting time data for as little as a few weeks or maybe a couple of months on how long on average it takes to fix a bug, and use this as a guide for estimating bug fixes going forward. If you have enough defect data, you can be smarter about how to use it. If you are tracking bugs in a bug database like Jira, and if programmers are tracking how much time they spend on fixing each bug for billing or time accounting purposes (which you can also do in Jira), then you can mine the bug database for similar bugs and see how long they took to fix – and maybe get some ideas on how to fix the bug that you are working on by reviewing what other people did on other bugs before you. You can group different bugs into buckets (by size – small, medium, large, x-large – or type) and then come up with an average estimate, and maybe even a best case, worst case and most likely for each type. Use Benchmarks For a maintenance team (a sustaining engineering or break/fix team responsible for software repairs only), you could use industry productivity benchmarks to project how many bugs your team can handle. Capers Jones in Estimating Software Costs says that the average programmer (in the US, in 2009), can fix 8-10 bugs per month (of course, if you’re an above-average programmer working in Canada in 2012, you’ll have to set these numbers much higher). Inexperienced programmers can be expected to fix 6 a month, while experienced developers using good tools can fix up to 20 per month. If you’re focusing on fixing security vulnerabilities reported by a pen tester or a scan, check out the remediation statistical data that Denim Group has started to collect, to get an idea on how long it might take to fix a SQL injection bug or an XSS vulnerability. So, do you estimate bug fixes, or not? Because you can’t estimate how long it will take to fix a bug until you’ve figured out what’s wrong, and most of the work in fixing a bug involves figuring out what’s wrong, it doesn’t make sense to try to do an in-depth estimate of how long it will take to fix each bug as they come up. Using simple historical data, a benchmark, or even a rough guess place holder as a rule-of-thumb all seem to work just as well. Whatever you do, do it in the simplest and most efficient way possible, don’t waste time trying to get it perfect – and realize that you won’t always be able to depend on it. Remember the 10x rule – some outlier bugs can take up to 10x as long to find and fix than an average bug. And some bugs can’t be found or fixed at all – or at least not with the information that you have today. When you’re wrong (and sometimes you’re going to be wrong), you can be really wrong, and even careful estimating isn’t going to help. So stick with a simple, efficient approach, and be prepared when you hit a hard problem, because it's gonna happen.
October 12, 2012
by Jim Bird
· 23,293 Views
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MongoDB Aggregation Framework Examples in C#
MongoDB version 2.2 was released in late August and the biggest change it brought was the addition of the Aggregation Framework. Previously the aggregations required the usage of map/reduce, which in MongoDB doesn’t perform that well, mainly because of the single-threaded Javascript-based execution. The aggregation framework steps away from the Javascript and is implemented in C++, with an aim to accelerate performance of analytics and reporting up to 80 percent compared to using MapReduce. The aim of this post is to show examples of running the MongoDB Aggregation Framework with the official MongoDB C# drivers. Aggregation Framework and Linq Even though the current version of the MongoDB C# drivers (1.6) supports Linq, the support doesn’t extend to the aggregation framework. It’s highly probable that the Linq-support will be added later on and there’s already some hints about this in the driver’s source code. But at this point the execution of the aggregations requires the usage of the BsonDocument-objects. Aggregation Framework and GUIDs If you use GUIDs in your documents, the aggregation framework doesn’t work. This is because by default the GUIDs are stored in binary format and the aggregations won’t work against documents which contain binary data.. The solution is to store the GUIDs as strings. You can force the C# drivers to make this conversion automatically by configuring the mapping. Given that your C# class has Id-property defined as a GUID, the following code tells the driver to serialize the GUID as a string: BsonClassMap.RegisterClassMap(cm => { cm.AutoMap(); cm.GetMemberMap(c => c.Id) .SetRepresentation( BsonType.String); }); The example data These examples use the following documents: > db.examples.find() { "_id" : "1", "User" : "Tom", "Country" : "Finland", "Count" : 1 } { "_id" : "2", "User" : "Tom", "Country" : "Finland", "Count" : 3 } { "_id" : "3", "User" : "Tom", "Country" : "Finland", "Count" : 2 } { "_id" : "4", "User" : "Mary", "Country" : "Sweden", "Count" : 1 } { "_id" : "5", "User" : "Mary", "Country" : "Sweden", "Count" : 7 } Example 1: Aggregation Framework Basic usage This example shows how the aggregation framework can be executed through C#. We’re not going run any calculations to the data, we’re just going to filter it by the User. To run the aggregations, you can use either the MongoDatabase.RunCommand –method or the helper MongoCollection.Aggregate. We’re going to use the latter: var coll = localDb.GetCollection("examples"); ... coll.Aggregate(pipeline); The hardest part when working with Aggregation Framework through C# is building the pipeline. The pipeline is similar concept to the piping in PowerShell. Each operation in the pipeline will make modifications to the data: the operations can for example filter, group and project the data. In C#, the pipeline is a collection of BsonDocument object. Each document represents one operation. In our first example we need to do only one operation: $match. This operator will filter out the given documents. The following BsonDocument is a pipeline operation which filters out all the documents which don’t have User-field set to “Tom”. var match = new BsonDocument { { "$match", new BsonDocument { {"User", "Tom"} } } }; To execute this operation we add it to an array and pass the array to the MongoCollection.Aggregate-method: var pipeline = new[] { match }; var result = coll.Aggregate(pipeline); The MongoCollection.Aggregate-method returns an AggregateResult-object. It’s ResultDocuments-property (IEnumarable) contains the documents which are the output of the aggregation. To check how many results there were, we can get the Count: var result = coll.Aggregate(pipeline); Console.WriteLine(result.ResultDocuments.Count()); The result documents are BsonDocument-objects. If you have a C#-class which represent the documents, you can cast the results: var matchingExamples = result.ResultDocuments .Select(BsonSerializer.Deserialize) .ToList(); foreach (var example in matchingExamples) { var message = string.Format("{0} - {1}", example.User, example.Count); Console.WriteLine(message); } Another alternative is to use C#’s dynamic type. The following extension method uses JSON.net to convert a BsonDocument into a dynamic: public static class MongoExtensions { public static dynamic ToDynamic(this BsonDocument doc) { var json = doc.ToJson(); dynamic obj = JToken.Parse(json); return obj; } } Here’s a way to convert all the result documents into dynamic objects: var matchingExamples = result.ResultDocuments .Select(x => x.ToDynamic()) .ToList(); Example 2: Multiple filters & comparison operators This example filters the data with the following criteria: User: Tom Count: >= 2 var match = new BsonDocument { { "$match", new BsonDocument { {"User", "Tom"}, {"Count", new BsonDocument { { "$gte", 2 } } } } }; The execution of this operation is identical to the first example: var pipeline = new[] { match }; var result = coll.Aggregate(pipeline); var matchingExamples = result.ResultDocuments .Select(x => x.ToDynamic()) .ToList(); Also the result are as expected: foreach (var example in matchingExamples) { var message = string.Format("{0} - {1}", example.User, example.Count); Console.WriteLine(message); } Example 3: Multiple operations In our first two examples, the pipeline was as simple as possible: It contained only one operation. This example will filter the data with the same exact criteria as the second example, but this time using two $match operations: User: Tom Count: >= 2 var match = new BsonDocument { { "$match", new BsonDocument { {"User", "Tom"} } } }; var match2 = new BsonDocument { { "$match", new BsonDocument { {"Count", new BsonDocument { { "$gte", 2 } } } } }; var pipeline = new[] { match, match2 }; The output stays the same: The first operation “match” takes all the documents from the examples collection and removes every document which doesn’t match the criteria User = Tom. The output of this operation (3 documents) then moves to the second operation “match2” of the pipeline. This operation only sees those 3 documents, not the original collection. The operation filters out these documents based on its criteria and moves the result (2 documents) forward. This is where our pipeline ends and this is also our result. Example 4: Group and sum Thus far we’ve used the aggregation framework to just filter out the data. The true strength of the framework is its ability to run calculations on the documents. This example shows how we can calculate how many documents there are in the collection, grouped by the user. This is done using the $group-operator: var group = new BsonDocument { { "$group", new BsonDocument { { "_id", new BsonDocument { { "MyUser","$User" } } }, { "Count", new BsonDocument { { "$sum", 1 } } } } } }; The grouping key (in our case the User-field) is defined with the _id. The above example states that the grouping key has one field (“MyUser”) and the value for that field comes from the document’s User-field ($User). In the $group operation the other fields are aggregate functions. This example defines the field “Count” and adds 1 to it for every document that matches the group key (_id). var pipeline = new[] { group }; var result = coll.Aggregate(pipeline); var matchingExamples = result.ResultDocuments .Select(x => x.ToDynamic()) .ToList(); foreach (var example in matchingExamples) { var message = string.Format("{0} - {1}", example._id.MyUser, example.Count); Console.WriteLine(message); } Note the format in which the results are outputted: The user’s name is accessed through _id.MyUser-property. Example 5: Group and sum by field This example is similar to example 4. But instead of calculating the amount of documents, we calculate the sum of the Count-fields by the user: var group = new BsonDocument { { "$group", new BsonDocument { { "_id", new BsonDocument { { "MyUser","$User" } } }, { "Count", new BsonDocument { { "$sum", "$Count" } } } } } }; The only change is that instead of adding 1, we add the value from the Count-field (“$Count”). Example 6: Projections This example shows how the $project operator can be used to change the format of the output. The grouping in example 5 works well, but to access the user’s name we currently have to point to the _id.MyUser-property. Let’s change this so that user’s name is available directly through UserName-property: var group = new BsonDocument { { "$group", new BsonDocument { { "_id", new BsonDocument { { "MyUser","$User" } } }, { "Count", new BsonDocument { { "$sum", "$Count" } } } } } }; var project = new BsonDocument { { "$project", new BsonDocument { {"_id", 0}, {"UserName","$_id.MyUser"}, {"Count", 1}, } } }; var pipeline = new[] { group, project }; The code removes the _id –property from the output. It adds the UserName-property, which value is accessed from field _id.MyUser. The projection operations also states that the Count-value should stay as it is. var matchingExamples = result.ResultDocuments .Select(x => x.ToDynamic()) .ToList(); foreach (var example in matchingExamples) { var message = string.Format("{0} - {1}", example.UserName, example.Count); Console.WriteLine(message); } Example 7: Group with multiple fields in the keys For this example we add a new row into our document collection, leaving us with the following: { "_id" : "1", "User" : "Tom", "Country" : "Finland", "Count" : 1 } { "_id" : "2", "User" : "Tom", "Country" : "Finland", "Count" : 3 } { "_id" : "3", "User" : "Tom", "Country" : "Finland", "Count" : 2 } { "_id" : "4", "User" : "Mary", "Country" : "Sweden", "Count" : 1 } { "_id" : "5", "User" : "Mary", "Country" : "Sweden", "Count" : 7 } { "_id" : "6", "User" : "Tom", "Country" : "England", "Count" : 3 } This example shows how you can group the data by using multiple fields in the grouping key: var group = new BsonDocument { { "$group", new BsonDocument { { "_id", new BsonDocument { { "MyUser","$User" }, { "Country","$Country" }, } }, { "Count", new BsonDocument { { "$sum", "$Count" } } } } } }; var project = new BsonDocument { { "$project", new BsonDocument { {"_id", 0}, {"UserName","$_id.MyUser"}, {"Country", "$_id.Country"}, {"Count", 1}, } } }; var pipeline = new[] { group, project }; var result = coll.Aggregate(pipeline); var matchingExamples = result.ResultDocuments .Select(x => x.ToDynamic()) .ToList(); foreach (var example in matchingExamples) { var message = string.Format("{0} - {1} - {2}", example.UserName, example.Country, example.Count); Console.WriteLine(message); } Example 8: Match, group and project This example shows how you can combine many different pipeline operations. The data is first filtered ($match) by User=Tom, then grouped by the Country (“$group”) and finally the output is formatted into a readable format ($project). Match: var match = new BsonDocument { { "$match", new BsonDocument { {"User", "Tom"} } } }; Group: var group = new BsonDocument { { "$group", new BsonDocument { { "_id", new BsonDocument { { "Country","$Country" }, } }, { "Count", new BsonDocument { { "$sum", "$Count" } } } } } }; Project: var project = new BsonDocument { { "$project", new BsonDocument { {"_id", 0}, {"Country", "$_id.Country"}, {"Count", 1}, } } }; Result: var pipeline = new[] { match, group, project }; var result = coll.Aggregate(pipeline); var matchingExamples = result.ResultDocuments .Select(x => x.ToDynamic()) .ToList(); foreach (var example in matchingExamples) { var message = string.Format("{0} - {1}", example.Country, example.Count); Console.WriteLine(message); } More There are many other interesting operators in the MongoDB Aggregation Framework, like $unwind and $sort. The usage of these operators is identical to ones we used above so it should be possible to copy-paste one of the examples and use it as a basis for these other operations. Links MongoDB C# Language Center MongoDB Aggregation Framework Easy to follow blog post about the aggregation framework
October 11, 2012
by Mikael Koskinen
· 47,886 Views · 2 Likes
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How to Create and Deploy a Website with Windows Azure
Curator's note: This article originally appeared at WindowsAzure.com. To use this feature and other new Windows Azure capabilities, sign up for the free preview. Just as you can quickly create and deploy a web application created from the gallery, you can also deploy a website created on a workstation with traditional developer tools from Microsoft or other companies. Table of Contents Deployment Options How to: Create a Website Using the Management Portal How to: Create a Website from the Gallery How to: Delete a Website Next Steps Deployment Options Windows Azure supports deploying websites from remote computers using WebDeploy, FTP, GIT or TFS. Many development tools provide integrated support for publication using one or more of these methods and may only require that you provide the necessary credentials, site URL and hostname or URL for your chosen deployment method. Credentials and deployment URLs for all enabled deployment methods are stored in the website's publish profile, a file which can be downloaded in the Windows Azure (Preview) Management Portal from the Quick Start page or the quick glance section of the Dashboard page. If you prefer to deploy your website with a separate client application, high quality open source GIT and FTP clients are available for download on the Internet for this purpose. How to: Create a Website Using the Management Portal Follow these steps to create a website in Windows Azure. Login to the Windows Azure (Preview) Management Portal. Click the Create New icon on the bottom left of the Management Portal. Click the Web Site icon, click the Quick Create icon, enter a value for URL and then click the check mark next to create web site on the bottom right corner of the page. When the website has been created you will see the text Creation of Web Site '[SITENAME]' Completed. Click the name of the website displayed in the list of websites to open the website's Quick Start management page. On the Quick Start page you are provided with options to set up TFS or GIT publishing if you would like to deploy your finished website to Windows Azure using these methods. FTP publishing is set up by default for websites and the FTP Host name is displayed under FTP Hostname on the Quick Start and Dashboard pages. Before publishing with FTP or GIT choose the option to Reset deployment credentials on the Dashboard page. Then specify the new credentials (username and password) to authenticate against the FTP Host or the Git Repository when deploying content to the website. The Configure management page exposes several configurable application settings in the following sections: Framework: Set the version of .NET framework or PHP required by your web application. Diagnostics: Set logging options for gathering diagnostic information for your website in this section. App Settings: Specify name/value pairs that will be loaded by your web application on start up. For .NET sites, these settings will be injected into your .NET configuration AppSettings at runtime, overriding existing settings. For PHP and Node sites these settings will be available as environment variables at runtime. Connection Strings: View connection strings for linked resources. For .NET sites, these connection strings will be injected into your .NET configuration connectionStrings settings at runtime, overriding existing entries where the key equals the linked database name. For PHP and Node sites these settings will be available as environment variables at runtime. Default Documents: Add your web application's default document to this list if it is not already in the list. If your web application contains more than one of the files in the list then make sure your website's default document appears at the top of the list. How to: Create a Website from the Gallery The gallery makes available a wide range of popular web applications developed by Microsoft, third party companies, and open source software initiatives. Web applications created from the gallery do not require installation of any software other than the browser used to connect to the Windows Azure Management Portal. In this tutorial, you'll learn: How to create a new site through the gallery. How to deploy the site through the Windows Azure Portal. You'll build a Word press blog that uses a default template. The following illustration shows the completed application: Note To complete this tutorial, you need a Windows Azure account that has the Windows Azure Web Sites feature enabled. You can create a free trial account and enable preview features in just a couple of minutes. For details, see Create a Windows Azure account and enable preview features. Create a web site in the portal Login to the Windows Azure Management Portal. Click the New icon on the bottom left of the dashboard. Click the Web Site icon, and click From Gallery. Locate and click the WordPress icon in list, and then click Next. On the Configure Your App page, enter or select values for all fields: Enter a URL name of your choice Leave Create a new MySQL database selected in the Database field Select the region closest to you Then click Next. On the Create New Database page, you can specify a name for your new MySQL database or use the default name. Select the region closest to you as the hosting location. Select the box at the bottom of the screen to agree to ClearDB's usage terms for your hosted MySQL database. Then click the check to complete the site creation. After you click Complete Windows Azure will initiate build and deploy operations. While the web site is being built and deployed the status of these operations is displayed at the bottom of the Web Sites page. After all operations are performed, A final status message when the site has been successfully deployed. Launch and manage your WordPress site Click on your new site from the Web Sites page to open the dashboard for the site. On the Dashboard management page, scroll down and click the link on the left under Site Url to open the site’s welcome page. Enter appropriate configuration information required by WordPress and click Install WordPress to finalize configuration and open the web site’s login page. Login to the new WordPress web site by entering the username and password that you specified on the Welcome page. You'll have a new WordPress site that looks similar to the site below. How to: Delete a Website Websites are deleted using the Delete icon in the Windows Azure Management Portal. The Delete icon is available in the Windows Azure Portal when you click Web Sites to list all of your websites and at the bottom of each of the website management pages. Next Steps For more information about Websites, see the following: Walkthrough: Troubleshooting a Website on Windows Azure
October 9, 2012
by Eric Gregory
· 85,476 Views
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Spring 3.1: Caching and EhCache
If you look around the web for examples of using Spring 3.1’s built in caching then you’ll usually bump into Spring’s SimpleCacheManager, which the Guys at Spring say is “Useful for testing or simple caching declarations”. I actually prefer to think of SimpleCacheManager as lightweight rather than simple; useful in those situations where you want a small in memory cache on a per JVM basis. If the Guys at Spring were running a supermarket then SimpleCacheManagerwould be in their own brand ‘basics’ product range. If, on the other hand, you need a heavy duty cache, one that’s scalable, persistent and distributed, then Spring also comes with a built in ehCache wrapper. The good news is that swapping between Spring's caching implementations is easy. In theory it’s all a matter of configuration and, to prove the theory correct, I took the sample code from my Caching and @Cacheable blog and ran it using an EhCache implementation. The configuration steps are similar to those described in my last blog Caching and Config in that you still need to specify: ...in your Spring config file to switch caching on. You also need to define a bean with an id of cacheManager, only this time you reference Spring’s EhCacheCacheManager class instead of SimpleCacheManager. The example above demonstrates an EhCacheCacheManager configuration. Notice that it references a second bean with an id of 'ehcache'. This is configured as follows: "ehcache" has two properties: configLocation and shared. 'configLocation' is an optional attribute that’s used to specify the location of an ehcache configuration file. In my test code I used the following example file: ...which creates two caches: a default cache and one named “employee”. If this file is missing then the EhCacheManagerFactoryBean simply picks up a default ehcache config file: ehcache-failsafe.xml, which is located in ehcache’s ehcache-core jar file. The other EhCacheManagerFactoryBean attribute is 'shared'. This is supposed to be optional as the documentation states that it defines "whether the EHCache CacheManager should be shared (as a singleton at the VM level) or independent (typically local within the application). Default is 'false', creating an independent instance.” However, if this is set to false then you’ll get the following exception: org.springframework.beans.factory.BeanCreationException: Error creating bean with name 'org.springframework.cache.interceptor.CacheInterceptor#0': Cannot resolve reference to bean 'cacheManager' while setting bean property 'cacheManager'; nested exception is org.springframework.beans.factory.BeanCreationException: Error creating bean with name 'cacheManager' defined in class path resource [ehcache-example.xml]: Cannot resolve reference to bean 'ehcache' while setting bean property 'cacheManager'; nested exception is org.springframework.beans.factory.BeanCreationException: Error creating bean with name 'ehcache' defined in class path resource [ehcache-example.xml]: Invocation of init method failed; nested exception is net.sf.ehcache.CacheException: Another unnamed CacheManager already exists in the same VM. Please provide unique names for each CacheManager in the config or do one of following: 1. Use one of the CacheManager.create() static factory methods to reuse same CacheManager with same name or create one if necessary 2. Shutdown the earlier cacheManager before creating new one with same name. The source of the existing CacheManager is: InputStreamConfigurationSource [stream=java.io.BufferedInputStream@424c414] at org.springframework.beans.factory.support.BeanDefinitionValueResolver.resolveReference(BeanDefinitionValueResolver.java:328) at org.springframework.beans.factory.support.BeanDefinitionValueResolver.resolveValueIfNecessary(BeanDefinitionValueResolver.java:106) at org.springframework.beans.factory.support.AbstractAutowireCapableBeanFactory.applyPropertyValues(AbstractAutowireCapableBeanFactory.java:1360) at org.springframework.beans.factory.support.AbstractAutowireCapableBeanFactory.populateBean(AbstractAutowireCapableBeanFactory.java:1118) at org.springframework.beans.factory.support.AbstractAutowireCapableBeanFactory.doCreateBean(AbstractAutowireCapableBeanFactory.java:517) at org.springframework.beans.factory.support.AbstractAutowireCapableBeanFactory.createBean(AbstractAutowireCapableBeanFactory.java:456) ... stack trace shortened for clarity at org.eclipse.jdt.internal.junit.runner.RemoteTestRunner.runTests(RemoteTestRunner.java:683) at org.eclipse.jdt.internal.junit.runner.RemoteTestRunner.run(RemoteTestRunner.java:390) at org.eclipse.jdt.internal.junit.runner.RemoteTestRunner.main(RemoteTestRunner.java:197) Caused by: org.springframework.beans.factory.BeanCreationException: Error creating bean with name 'cacheManager' defined in class path resource [ehcache-example.xml]: Cannot resolve reference to bean 'ehcache' while setting bean property 'cacheManager'; nested exception is org.springframework.beans.factory.BeanCreationException: Error creating bean with name 'ehcache' defined in class path resource [ehcache-example.xml]: Invocation of init method failed; nested exception is net.sf.ehcache.CacheException: Another unnamed CacheManager already exists in the same VM. Please provide unique names for each CacheManager in the config or do one of following: 1. Use one of the CacheManager.create() static factory methods to reuse same CacheManager with same name or create one if necessary 2. Shutdown the earlier cacheManager before creating new one with same name. The source of the existing CacheManager is: InputStreamConfigurationSource [stream=java.io.BufferedInputStream@424c414] at org.springframework.beans.factory.support.BeanDefinitionValueResolver.resolveReference(BeanDefinitionValueResolver.java:328) at org.springframework.beans.factory.support.BeanDefinitionValueResolver.resolveValueIfNecessary(BeanDefinitionValueResolver.java:106) at org.springframework.beans.factory.support.AbstractAutowireCapableBeanFactory.applyPropertyValues(AbstractAutowireCapableBeanFactory.java:1360) ... stack trace shortened for clarity at org.springframework.beans.factory.support.AbstractBeanFactory.getBean(AbstractBeanFactory.java:193) at org.springframework.beans.factory.support.BeanDefinitionValueResolver.resolveReference(BeanDefinitionValueResolver.java:322) ... 38 more Caused by: org.springframework.beans.factory.BeanCreationException: Error creating bean with name 'ehcache' defined in class path resource [ehcache-example.xml]: Invocation of init method failed; nested exception is net.sf.ehcache.CacheException: Another unnamed CacheManager already exists in the same VM. Please provide unique names for each CacheManager in the config or do one of following: 1. Use one of the CacheManager.create() static factory methods to reuse same CacheManager with same name or create one if necessary 2. Shutdown the earlier cacheManager before creating new one with same name. The source of the existing CacheManager is: InputStreamConfigurationSource [stream=java.io.BufferedInputStream@424c414] at org.springframework.beans.factory.support.AbstractAutowireCapableBeanFactory.initializeBean(AbstractAutowireCapableBeanFactory.java:1455) at org.springframework.beans.factory.support.AbstractAutowireCapableBeanFactory.doCreateBean(AbstractAutowireCapableBeanFactory.java:519) at org.springframework.beans.factory.support.AbstractAutowireCapableBeanFactory.createBean(AbstractAutowireCapableBeanFactory.java:456) at org.springframework.beans.factory.support.AbstractBeanFactory$1.getObject(AbstractBeanFactory.java:294) at org.springframework.beans.factory.support.DefaultSingletonBeanRegistry.getSingleton(DefaultSingletonBeanRegistry.java:225) at org.springframework.beans.factory.support.AbstractBeanFactory.doGetBean(AbstractBeanFactory.java:291) at org.springframework.beans.factory.support.AbstractBeanFactory.getBean(AbstractBeanFactory.java:193) at org.springframework.beans.factory.support.BeanDefinitionValueResolver.resolveReference(BeanDefinitionValueResolver.java:322) ... 48 more Caused by: net.sf.ehcache.CacheException: Another unnamed CacheManager already exists in the same VM. Please provide unique names for each CacheManager in the config or do one of following: 1. Use one of the CacheManager.create() static factory methods to reuse same CacheManager with same name or create one if necessary 2. Shutdown the earlier cacheManager before creating new one with same name. The source of the existing CacheManager is: InputStreamConfigurationSource [stream=java.io.BufferedInputStream@424c414] at net.sf.ehcache.CacheManager.assertNoCacheManagerExistsWithSameName(CacheManager.java:521) at net.sf.ehcache.CacheManager.init(CacheManager.java:371) at net.sf.ehcache.CacheManager.(CacheManager.java:339) at org.springframework.cache.ehcache.EhCacheManagerFactoryBean.afterPropertiesSet(EhCacheManagerFactoryBean.java:104) at org.springframework.beans.factory.support.AbstractAutowireCapableBeanFactory.invokeInitMethods(AbstractAutowireCapableBeanFactory.java:1514) at org.springframework.beans.factory.support.AbstractAutowireCapableBeanFactory.initializeBean(AbstractAutowireCapableBeanFactory.java:1452) ... 55 more ...when you try to run a bunch of unit tests. I think that this comes down to a simple bug Spring’s the ehcache manager factory as it’s trying to create multiple cache instances using new() rather than using, as the exception states, “one of the CacheManager.create() static factory methods" which allows it to reuse same CacheManager with same name. Hence, my first JUnit test works okay, but all others fail. The offending line of code is: this.cacheManager = (this.shared ? CacheManager.create() : new CacheManager()); My full XML config file is listed below for completeness: In using ehcache, the only other configuration details to consider are the Maven dependencies. These are pretty straight forward as the Guys at Ehcache have combined all the various ehcache jars into one Maven POM module. This POM module can be added to your project's POM file using the XML below: net.sf.ehcache ehcache 2.6.0 pom test Finally, the ehcache Jar files are available from both the Maven Central and Sourceforge repositories: sourceforge http://oss.sonatype.org/content/groups/sourceforge/ true true
October 5, 2012
by Roger Hughes
· 106,309 Views
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Record Audio Using webrtc in Chrome and Speech Recognition With Websockets
there are many different web api standards that are turning the web browser into a complete application platform. with websockets we get nice asynchronous communication, various standards allow us access to sensors in laptops and mobile devices and we can even determine how full the battery is. one of the standards i'm really interested in is webrtc. with webrtc we can get real-time audio and video communication between browsers without needing plugins or additional tools. a couple of months ago i wrote about how you can use webrtc to access the webcam and use it for face recognition . at that time, none of the browser allowed you to access the microphone. a couple of months later though, and both the developer version of firefox and developer version of chrome, allow you to access the microphone! so let's see what we can do with this. most of the examples i've seen so far focus on processing the input directly, within the browser, using the web audio api . you get synthesizers, audio visualizations, spectrometers etc. what was missing, however, was a means of recording the audio data and storing it for further processing at the server side. in this article i'll show you just that. i'm going to show you how you can create the following (you might need to enlarge it to read the response from the server): in this screencast you can see the following: a simple html page that access your microphone the speech is recorded and using websockets is sent to a backend the backend combines the audio data and sends it to google's speech to text api the result from this api call is returned to the browser and all this is done without any plugins in the browser! so what's involved to accomplish all this. allowing access to your microphone the first thing you need to do is make sure you've got an up to date version of chrome. i use the dev build, and am currently on this version: since this is still an experimental feature we need to enable this using the chrome flags. make sure the "web audio input" flag is enabled. with this configuration out of the way we can start to access our microphone. access the audio stream from the microphone this is actually very easy: function callback(stream) { var context = new webkitaudiocontext(); var mediastreamsource = context.createmediastreamsource(stream); ... } $(document).ready(function() { navigator.webkitgetusermedia({audio:true}, callback); ... } as you can see i use the webkit prefix functions directly, you could, of course, also use a shim so it is browser independent. what happens in the code above is rather straightforward. we ask, using getusermedia, for access to the microphone. if this is successful our callback gets called with the audio stream as its parameter. in this callback we use the web audio specification to create a mediastreamsource from our microphone. with this mediastreamsource we can do all the nice web audio tricks you can see here . but we don't want that, we want to record the stream and send it to a backend server for further processing. in future versions this will probably be possible directly from the webrtc api, at this time, however, this isn't possible yet. luckily, though, we can use a feature from the web audio api to get access to the raw data. with the javascriptaudionode we can create a custom node, which we can use to access the raw data (which is pcm encoded). before i started my own work on this i searched around a bit and came across the recoder.js project from here: https://github.com/mattdiamond/recorderjs . matt created a recorder that can record the output from web audio nodes, and that's exactly what i needed. all i needed to do now was connect the stream we just created to the recorder library: function callback(stream) { var context = new webkitaudiocontext(); var mediastreamsource = context.createmediastreamsource(stream); rec = new recorder(mediastreamsource); } with this code, we create a recorder from our stream. this recorder provides the following functions: record: start recording from the input stop: stop recording clear: clear the current recording exportwav: export the data as a wav file connect the recorder to the buttons i've created a simple webpage with an output for the text and two buttons to control the recording: the 'record' button starts the recording, and once you hit the 'export' button the recording stops, and is sent to the backend for processing. record button: $('#record').click(function() { rec.record(); ws.send("start"); $("#message").text("click export to stop recording and analyze the input"); // export a wav every second, so we can send it using websockets intervalkey = setinterval(function() { rec.exportwav(function(blob) { rec.clear(); ws.send(blob); }); }, 1000); }); this function (using jquery to connect it to the button) when clicked starts the recording. it also uses a websocket (ws), see further down on how to setup the websocket, to indicate to the backend server to expect a new recording (more on this later). finally when the button is clicked an interval is created that passes the data to the backend, encoded as wav file, every second. we do this to avoid sending too large chunks of data to the backend and improve performance. export button: $('#export').click(function() { // first send the stop command rec.stop(); ws.send("stop"); clearinterval(intervalkey); ws.send("analyze"); $("#message").text(""); }); the export button, bad naming i think when i'm writing this, stops the recording, the interval and informs the backend server that it can send the received data to the google api for further processing. connecting the frontend to the backend to connect the webapplication to the backend server we use websockets. in the previous code fragments you've already seen how they are used. we create them with the following: var ws = new websocket("ws://127.0.0.1:9999"); ws.onopen = function () { console.log("openened connection to websocket"); }; ws.onmessage = function(e) { var jsonresponse = jquery.parsejson(e.data ); console.log(jsonresponse); if (jsonresponse.hypotheses.length > 0) { var bestmatch = jsonresponse.hypotheses[0].utterance; $("#outputtext").text(bestmatch); } } we create a connection, and when we receive a message from the backend we just assume it contains the response to our speech analysis. and that's it for the complete front end of the application. we use getusermedia to access the microphone, use the web audio api to get access to the raw data and communicate with websockets with the backend server. the backend server our backend server needs to do a couple of things. it first needs to combine the incoming chunks to a single audio file, next it needs to convert this to a format google apis expect, which is flac. finally we make a call to the google api and return the response. i've used jetty as the websocket server for this example. if you want to know the details about setting this up, look at the facedetection example. in this article i'll only show the code to process the incoming messages. first step, combine the incoming data the data we receive is encoded as wav (thanks to the recorder.js library we don't have to do this ourselves). in our backend we thus receive sound fragments with a length of one second. we can't just concatenate these together, since wav files have a header that tells how long the fragment is (amongst other things), so we have to combine them, and rewrite the header. lets first look at the code (ugly code, but works good enough for now :) public void onmessage(byte[] data, int offset, int length) { if (currentcommand.equals("start")) { try { // the temporary file that contains our captured audio stream file f = new file("out.wav"); // if the file already exists we append it. if (f.exists()) { log.info("adding received block to existing file."); // two clips are used to concat the data audioinputstream clip1 = audiosystem.getaudioinputstream(f); audioinputstream clip2 = audiosystem.getaudioinputstream(new bytearrayinputstream(data)); // use a sequenceinput to cat them together audioinputstream appendedfiles = new audioinputstream( new sequenceinputstream(clip1, clip2), clip1.getformat(), clip1.getframelength() + clip2.getframelength()); // write out the output to a temporary file audiosystem.write(appendedfiles, audiofileformat.type.wave, new file("out2.wav")); // rename the files and delete the old one file f1 = new file("out.wav"); file f2 = new file("out2.wav"); f1.delete(); f2.renameto(new file("out.wav")); } else { log.info("starting new recording."); fileoutputstream fout = new fileoutputstream("out.wav",true); fout.write(data); fout.close(); } } catch (exception e) { ...} } } this method gets called for each chunk of audio we receive from the browser. what we do here is the following: first, we check whether we have a temp audio file, if not we create it if the file exists we use java's audiosystem to create an audio sequence this sequence is then written to another file the original is deleted and the new one is renamed. we repeat this for each chunk so at this point we have a wav file that keeps on growing for each added chunk. now before we convert this, lets look at the code we use to control the backend. public void onmessage(string data) { if (data.startswith("start")) { // before we start we cleanup anything left over cleanup(); currentcommand = "start"; } else if (data.startswith("stop")) { currentcommand = "stop"; } else if (data.startswith("clear")) { // just remove the current recording cleanup(); } else if (data.startswith("analyze")) { // convert to flac ... // send the request to the google speech to text service ... } } the previous method responded to binary websockets messages. the one shown above responds to string messages. we use this to control, from the browser, what the backend should do. let's look at the analyze command, since that is the interesting one. when this command is issued from the frontend the backend needs to convert the wav file to flac and send it to the google service. convert to flac for the conversion to flac we need an external library since java standard has no support for this. i used the javaflacencoder from here for this. // get an encoder flac_fileencoder flacencoder = new flac_fileencoder(); // point to the input file file inputfile = new file("out.wav"); file outputfile = new file("out2.flac"); // encode the file log.info("start encoding wav file to flac."); flacencoder.encode(inputfile, outputfile); log.info("finished encoding wav file to flac."); easy as that. now we got a flac file that we can send to google for analysis. send to google for analysis a couple of weeks ago i ran across an article that explained how someone analyzed chrome and found out about an undocumented google api you can use for speech to text. if you post a flac file to this url: https://www.google.com/speech-api/v1/recognize?xjerr=1&client=chromium&l... you receive a response like this: { "status": 0, "id": "ae466ffa24a1213f5611f32a17d5a42b-1", "hypotheses": [ { "utterance": "the quick brown fox", "confidence": 0.857393 }] } to do this from java code, using httpclient, you do the following: // send the request to the google speech to text service log.info("sending file to google for speech2text"); httpclient client = new defaulthttpclient(); httppost p = new httppost(url); p.addheader("content-type", "audio/x-flac; rate=44100"); p.setentity(new fileentity(outputfile, "audio/x-flac; rate=44100")); httpresponse response = client.execute(p); f (response.getstatusline().getstatuscode() == 200) { log.info("received valid response, sending back to browser."); string result = new string(ioutils.tobytearray(response.getentity().getcontent())); this.connection.sendmessage(result); } and that are all the steps that are needed.
October 5, 2012
by Jos Dirksen
· 20,657 Views · 1 Like
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SQL Query Optimization and Normalization
Explore SQL query optimization and normalization.
October 4, 2012
by Michael Georgiou
· 37,934 Views · 2 Likes
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Difference Between Mysql Replace and Insert on Duplicate Key Update
While me and my friend roshan recently working as a support developers at Australia famous e-commerce website. recently roshan as assign a new bug in this site it’s related to the product synchronize process in the ware house product table and the e-commerce site, his main task was check the quickly the site product table and check with ware house product table product if the either insert new data into a site database, or update an existing record on the site database, Of course, doing a lookup to see if the record exists already and then either updating or inserting would be an expensive process (existing items are defined either by a unique key or a primary key). Luckily, MySQL offers two functions to combat this (each with two very different approaches). 1. REPLACE = DELETE+INSERT 2. INSERT ON DUPLICATE KEY UPDATE = UPDATE + INSERT 1 . REPLACE This syntax is the same as the INSERT function. When dealing with a record with a unique or primary key, REPLACE will either do a DELETE and then an INSERT, or just an INSERT if use this this function will cause a record to be removed, and inserted at the end. It will cause the indexing to get broken apart, decreasing the efficiency of the table. If, however REPLACE INTO ds_product SET pID = 3112, catID = 231, uniCost = 232.50, salePrice = 250.23; 2. ON DUPLICATE KEY UPDATE ON DUPLICATE KEY UPDATE clause to the INSERT function. This one actively hunts down an existing record in the table which has the same UNIQUE or PRIMARY KEY as the one we’re trying to update. If it finds an existing one, you specify a clause for which column(s) you would like to UPDATE. Otherwise, it will do a normal INSERT. INSERT INTO ds_product SET pID = 3112, catID = 231, uniCost = 232.50, salePrice = 250.23, ON DUPLICATE KEY UPDATE uniCost = 232.50, salePrice = 250.23; This should be helpful when trying to create database queries that add and update information, without having to go through the extra step. Thanks Have a Nice Day
October 3, 2012
by Prathap Givantha Kalansuriya
· 14,478 Views
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Parsing a Connection String With 'Sprache' C# Parser
Sprache is a very cool lightweight parser library for C#. Today I was experimenting with parsing EasyNetQ connection strings, so I thought I’d have a go at getting Sprache to do it. An EasyNetQ connection string is a list of key-value pairs like this: key1=value1;key2=value2;key3=value3 The motivation for looking at something more sophisticated than simply chopping strings based on delimiters, is that I’m thinking of having more complex values that would themselves need parsing. But that’s for the future, today I’m just going to parse a simple connection string where the values can be strings or numbers (ushort to be exact). So, I want to parse a connection string that looks like this: virtualHost=Copa;username=Copa;host=192.168.1.1;password=abc_xyz;port=12345;requestedHeartbeat=3 … into a strongly typed structure like this: public class ConnectionConfiguration : IConnectionConfiguration { public string Host { get; set; } public ushort Port { get; set; } public string VirtualHost { get; set; } public string UserName { get; set; } public string Password { get; set; } public ushort RequestedHeartbeat { get; set; } } I want it to be as easy as possible to add new connection string items. First let’s define a name for a function that updates a ConnectionConfiguration. A uncommonly used version of the ‘using’ statement allows us to give a short name to a complex type: using UpdateConfiguration = Func; Now lets define a little function that creates a Sprache parser for a key value pair. We supply the key and a parser for the value and get back a parser that can update the ConnectionConfiguration. public static Parser BuildKeyValueParser( string keyName, Parser valueParser, Expression> getter) { return from key in Parse.String(keyName).Token() from separator in Parse.Char('=') from value in valueParser select (Func)(c => { CreateSetter(getter)(c, value); return c; }); } The CreateSetter is a little function that turns a property expression (like x => x.Name) into an Action. Next let’s define parsers for string and number values: public static Parser Text = Parse.CharExcept(';').Many().Text(); public static Parser Number = Parse.Number.Select(ushort.Parse); Now we can chain a series of BuildKeyValueParser invocations and Or them together so that we can parse any of our expected key-values: public static Parser Part = new List> { BuildKeyValueParser("host", Text, c => c.Host), BuildKeyValueParser("port", Number, c => c.Port), BuildKeyValueParser("virtualHost", Text, c => c.VirtualHost), BuildKeyValueParser("requestedHeartbeat", Number, c => c.RequestedHeartbeat), BuildKeyValueParser("username", Text, c => c.UserName), BuildKeyValueParser("password", Text, c => c.Password), }.Aggregate((a, b) => a.Or(b)); Each invocation of BuildKeyValueParser defines an expected key-value pair of our connection string. We just give the key name, the parser that understands the value, and the property on ConnectionConfiguration that we want to update. In effect we’ve defined a little DSL for connection strings. If I want to add a new connection string value, I simply add a new property to ConnectionConfiguration and a single line to the above code. Now lets define a parser for the entire string, by saying that we’ll parse any number of key-value parts: public static Parser> ConnectionStringBuilder = from first in Part from rest in Parse.Char(';').Then(_ => Part).Many() select Cons(first, rest); All we have to do now is parse the connection string and apply the chain of update functions to a ConnectionConfiguration instance: public IConnectionConfiguration Parse(string connectionString) { var updater = ConnectionStringGrammar.ConnectionStringBuilder.Parse(connectionString); return updater.Aggregate(new ConnectionConfiguration(), (current, updateFunction) => updateFunction(current)); } We get lots of nice things out of the box with Sprache, one of the best is the excellent error messages: Parsing failure: unexpected 'x'; expected host or port or virtualHost or requestedHeartbeat or username or password (Line 1, Column 1). Sprache is really nice for this kind of task. I’d recommend checking it out.
October 3, 2012
by Mike Hadlow
· 7,688 Views
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Customizing Spring Data JPA Repository
Spring Data is a very convenient library. However, as the project as quite new, it is not well featured. By default, Spring Data JPA will provide implementation of the DAO based on SimpleJpaRepository. In recent project, I have developed a customize repository base class so that I could add more features on it. You could add vendor specific features to this repository base class as you like. Configuration You have to add the following configuration to you spring beans configuration file. You have to specified a new repository factory class. We will develop the class later. extends SimpleJpaRepository implements GenericRepository , Serializable{ private static final long serialVersionUID = 1L; static Logger logger = Logger.getLogger(GenericRepositoryImpl.class); private final JpaEntityInformation entityInformation; private final EntityManager em; private final DefaultPersistenceProvider provider; private Class springDataRepositoryInterface; public Class getSpringDataRepositoryInterface() { return springDataRepositoryInterface; } public void setSpringDataRepositoryInterface( Class springDataRepositoryInterface) { this.springDataRepositoryInterface = springDataRepositoryInterface; } /** * Creates a new {@link SimpleJpaRepository} to manage objects of the given * {@link JpaEntityInformation}. * * @param entityInformation * @param entityManager */ public GenericRepositoryImpl (JpaEntityInformation entityInformation, EntityManager entityManager , Class springDataRepositoryInterface) { super(entityInformation, entityManager); this.entityInformation = entityInformation; this.em = entityManager; this.provider = DefaultPersistenceProvider.fromEntityManager(entityManager); this.springDataRepositoryInterface = springDataRepositoryInterface; } /** * Creates a new {@link SimpleJpaRepository} to manage objects of the given * domain type. * * @param domainClass * @param em */ public GenericRepositoryImpl(Class domainClass, EntityManager em) { this(JpaEntityInformationSupport.getMetadata(domainClass, em), em, null); } public S save(S entity) { if (this.entityInformation.isNew(entity)) { this.em.persist(entity); flush(); return entity; } entity = this.em.merge(entity); flush(); return entity; } public T saveWithoutFlush(T entity) { return super.save(entity); } public List saveWithoutFlush(Iterable entities) { List result = new ArrayList(); if (entities == null) { return result; } for (T entity : entities) { result.add(saveWithoutFlush(entity)); } return result; } } As a simple example here, I just override the default save method of the SimpleJPARepository. The default behaviour of the save method will not flush after persist. I modified to make it flush after persist. On the other hand, I add another method called saveWithoutFlush() to allow developer to call save the entity without flush. Define Custom repository factory bean The last step is to create a factory bean class and factory class to produce repository based on your customized base repository class. public class DefaultRepositoryFactoryBean , S, ID extends Serializable> extends JpaRepositoryFactoryBean { /** * Returns a {@link RepositoryFactorySupport}. * * @param entityManager * @return */ protected RepositoryFactorySupport createRepositoryFactory( EntityManager entityManager) { return new DefaultRepositoryFactory(entityManager); } } /** * * The purpose of this class is to override the default behaviour of the spring JpaRepositoryFactory class. * It will produce a GenericRepositoryImpl object instead of SimpleJpaRepository. * */ public class DefaultRepositoryFactory extends JpaRepositoryFactory{ private final EntityManager entityManager; private final QueryExtractor extractor; public DefaultRepositoryFactory(EntityManager entityManager) { super(entityManager); Assert.notNull(entityManager); this.entityManager = entityManager; this.extractor = DefaultPersistenceProvider.fromEntityManager(entityManager); } @SuppressWarnings({ "unchecked", "rawtypes" }) protected JpaRepository getTargetRepository( RepositoryMetadata metadata, EntityManager entityManager) { Class repositoryInterface = metadata.getRepositoryInterface(); JpaEntityInformation entityInformation = getEntityInformation(metadata.getDomainType()); if (isQueryDslExecutor(repositoryInterface)) { return new QueryDslJpaRepository(entityInformation, entityManager); } else { return new GenericRepositoryImpl(entityInformation, entityManager, repositoryInterface); //custom implementation } } @Override protected Class getRepositoryBaseClass(RepositoryMetadata metadata) { if (isQueryDslExecutor(metadata.getRepositoryInterface())) { return QueryDslJpaRepository.class; } else { return GenericRepositoryImpl.class; } } /** * Returns whether the given repository interface requires a QueryDsl * specific implementation to be chosen. * * @param repositoryInterface * @return */ private boolean isQueryDslExecutor(Class repositoryInterface) { return QUERY_DSL_PRESENT && QueryDslPredicateExecutor.class .isAssignableFrom(repositoryInterface); } } Conclusion You could now add more features to base repository class. In your program, you could now create your own repository interface extending GenericRepository instead of JpaRepository. public interface MyRepository extends GenericRepository { void someCustomMethod(ID id); } In next post, I will show you how to add hibernate filter features to this GenericRepository.
September 27, 2012
by Boris Lam
· 98,267 Views · 4 Likes
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Enabling JMX Monitoring for Hadoop & Hive
Hadoop’s NameNode and JobTracker expose interesting metrics and statistics over the JMX. Hive seems not to expose anything intersting but it still might be useful to monitor its JVM or do simpler profiling/sampling on it. Let’s see how to enable JMX and how to access it securely, over SSH. Background: We run NameNode, JobTracker and Hive on the same server. Monitoring og TaskTrackers and DataNodes isn’t that interesting but still might be useful to have. Configuration /etc/hadoop/hadoop-env.sh diff --git a/etc/hadoop/hadoop-env.sh b/etc/hadoop/hadoop-env.sh index 69a13b1..e8ca596 100644 --- a/etc/hadoop/hadoop-env.sh +++ b/etc/hadoop/hadoop-env.sh @@ -14,7 +14,8 @@ export HADOOP_CONF_DIR=${HADOOP_CONF_DIR:-"/etc/hadoop"} #export HADOOP_NAMENODE_INIT_HEAPSIZE="" # Extra Java runtime options. Empty by default. -export HADOOP_OPTS="-Djava.net.preferIPv4Stack=true $HADOOP_CLIENT_OPTS" +# Added $HIVE_OPTS that is set by hive-env.sh when starting hiveserver +export HADOOP_OPTS="-Djava.net.preferIPv4Stack=true $HADOOP_CLIENT_OPTS $HIVE_OPTS" # Command specific options appended to HADOOP_OPTS when specified export HADOOP_NAMENODE_OPTS="-Dhadoop.security.logger=INFO,DRFAS -Dhdfs.audit.logger=INFO,DRFAAUDIT $HADOOP_NAMENODE_OPTS" @@ -43,3 +44,16 @@ export HADOOP_SECURE_DN_PID_DIR=/var/run/hadoop # A string representing this instance of hadoop. $USER by default. export HADOOP_IDENT_STRING=$USER + +### JMX settings +export JMX_OPTS=" -Dcom.sun.management.jmxremote.authenticate=false \ + -Dcom.sun.management.jmxremote.ssl=false \ + -Dcom.sun.management.jmxremote.port" +# -Dcom.sun.management.jmxremote.password.file=$HADOOP_HOME/conf/jmxremote.password \ +# -Dcom.sun.management.jmxremote.access.file=$HADOOP_HOME/conf/jmxremote.access" +export HADOOP_NAMENODE_OPTS="$JMX_OPTS=8006 $HADOOP_NAMENODE_OPTS" +export HADOOP_SECONDARYNAMENODE_OPTS="$HADOOP_SECONDARYNAMENODE_OPTS" +export HADOOP_DATANODE_OPTS="$JMX_OPTS=8006 $HADOOP_DATANODE_OPTS" +export HADOOP_BALANCER_OPTS="$HADOOP_BALANCER_OPTS" +export HADOOP_JOBTRACKER_OPTS="$JMX_OPTS=8007 $HADOOP_JOBTRACKER_OPTS" +export HADOOP_TASKTRACKER_OPTS="$JMX_OPTS=8007 $HADOOP_TASKTRACKER_OPTS" The JMX setting is used for Hadoop’s daemons while the HIVE_OPTS was added for Hive. /conf/hive-env.sh Enable JMX when running the Hive thrift server (we don’t want it when running the command-line client etc. since it’s pointless and we wouldn’t need to make sure that each of them has a unique port): if [ "$SERVICE" = "hiveserver" ]; then JMX_OPTS="-Dcom.sun.management.jmxremote -Dcom.sun.management.jmxremote.authenticate=false -Dcom.sun.management.jmxremote.ssl=false -Dcom.sun.management.jmxremote.port=8008" export HIVE_OPTS="$HIVE_OPTS $JMX_OPTS" fi Pitfalls When you start Hive server via hive –service hiveserver then it actually executes “hadoop jar …” so to be able to pass options from hive-env.sh to the JVM we had to add $HIVE_OPTS in hadoop-env.sh. (I haven’t found a cleaner way to do it.) Effects When we now start Hive or any of the Hadoop daemons, they will expose their metrics at their respective ports (NameNode – 8006, JobTracker – 8007, Hive – 8008). (If you are running DataNode and/or TaskTracker on the same machine then you’ll need to change their ports to be unique.) Secure Connection Over SSH Read the post VisualVM: Monitoring Remote JVM Over SSH (JMX Or Not) to find out how to connect securely to the JMX ports over ssh, f.ex. with VisualVM (spolier: ssh -D 9696 hostname; use proxy at localhost:9696).
September 25, 2012
by Jakub Holý
· 15,348 Views
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Choosing Static vs. Dynamic Languages for Your Startup
Everyone is thinking why in the world would anyone pick static, when you can be dynamic? Usually the thought process is, "what language am I most proficient in, that can do the job." Totally not a bad way to go about it. Now does this choice affect anything else? Testing? Speed of development? Robustness? Dynamic vs. Static Dynamic languages are languages that don’t necessarily need variables to be declared before they are used. Examples of dynamic languages are Python, Ruby, and PHP. So in dynamic languages the following is possible: num = 10 We have successfully assigned a value to variable without declaring it before hand. Simple enough, try doing this in Java (you can’t). This can *increase* development speed, without having to write boilerplate code. This can somewhat be a double edge sword, since dynamic languages types are checked during runtime, there is no way to tell if there is a bug in code until it is run. I know you can test, but you can’t test for everything. You can’t test for everything. Here is an example albeit trivial. def get_first_problem(problems): for problem in problems: problam = problem + 1 return problam Now if you are raging to some serious dubstep, its easy enough to miss that small typo, you go screw it and do it live, and deploy to production. Python will simply create the new variable and not a single thing will be said. Only you can stop bugs in production! Static languages are languages that variables need to be declared before use and type checking is done at compile time. Examples of static languages include Java, C, and C++. So in static languages the following is enforced static int awesomeNumber; awesomeNumber = 10; Many argue this increases robustness as well as decrease chances of Runtime Errors. Since the compiler will catch those horrible horrible mistakes you made throughout your code. Your methods contracts are tighter, downside to this is crap ton of boilerplate code. Weak and Strong Typing can be often be confused with dynamic and static languages. Weak typed languages can lead to philosophical questions like what does the number 2 added to the word ‘two’ give you? Things like this are possible with a weak typed language. a = 2 b = "2" concatenate(a, b) // Returns "22" add(a, b) // Returns 4 Traditionally languages may place restriction on what transaction may occur for example in a strong typed language adding a string and integer will result in a type error as shown below. >>> a = 10 >>> b = 'ten' >>> a + b Traceback (most recent call last): File "", line 1, in TypeError: unsupported operand type(s) for +: 'int' and 'str' >>> Conclusion Regardless of where you land on this discussion, claiming one is better than the other would lead to flame war, but there are places where each is strong. Dynamic languages are good for fast quick development cycles and prototyping, while static languages are better suited to longer development cycles where trivial bugs could be extremely costly (telecommunication systems, air traffic control). For example if some giant company called Moo Corp. spent millions of dollars on QA and Testing and a bug somehow gets into the field, to fix it would mean another round of testing. When sitting in that chair the choice is clear static languages FTW, its a hard job but someone has to milk the cows. Test, test, and test. Just a little food for thought, for when you are starting your next project. You never know what limitations you maybe placing on yourself and your team. What do you do consider when selecting a programming language for a project?
September 25, 2012
by Mahdi Yusuf
· 25,145 Views
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Introducing the New Date and Time API for JDK 8
Date and time handling in Java is a somewhat tricky part when you are new to the language. Time can be accessed via the static method System.currentTimeMillis() which returns the current time in milliseconds from January 1st 1970. If you prefer to work with Objects instead you can use java.util.Date, a class whose methods are mostly deprecated in recent versions of Java. To work with time offsets, say add one month to a date, there is java.util.GregorianCalendar. All in all, those methods described here are not very convenient to work with. Java 7 and below are lacking a good date and time API. The Joda Time library is a common drop-in if you need to work with date/time. With JSR 310 (Java Specification Request) this is about to change. JSR 310 adds a new date, time and calendar API to Java 8. The ThreeTen project provides a reference implementation to this new API and can already be utilized in current Java projects (I however recommend not to do this for production). As the README states: The API is currently considered usable and accurate, yet incomplete and subject to change. If you use this API you must be able to handle incompatible changes in later versions. Building ThreeTen Building the ThreeTen project is relatively easy. It requires both Git and Ant to be installed on your system. git clone git://github.com/ThreeTen/threeten.git cd threeten ant This will first fetch the most recent version of ThreeTen and then start the build process using ant. Note that building the library also requires either OpenJDK 1.6 or Oracle JDK 1.6. JSR 310 The new API specifies a number of new classes which are divided into the categories of continuous and human time. Continuous time is based on Unix time and is represented as a single incrementing number. Class Description Instant A point in time in nanoseconds from January 1st 1970 Duration An amount of time measured in nanoseconds Human time is based on fields that we use in our daily lifes such as day, hour, minute and second. It is represented by a group of classes, some of which we will discuss in this article. Class Description LocalDate a date, without time of day, offset or zone LocalTime the time of day, without date, offset or zone LocalDateTime the date and time, without offset or zone OffsetDate a date with an offset such as +02:00, without time of day or zone OffsetTime the time of day with an offset such as +02:00, without date or zone OffsetDateTime the date and time with an offset such as +02:00, without a zone ZonedDateTime the date and time with a time zone and offset YearMonth a year and month MonthDay month and day Year/MonthOfDay/DayOfWeek/... classes for the important fields DateTimeFields stores a map of field-value pairs which may be invalid Calendrical access to the low-level API Period a descriptive amount of time, such as "2 months and 3 days" In addition to the above classes three support classes have been implemented. The Clock class wraps the current time and date, ZoneOffset is a time offset from UTC and ZoneId defines a time zone such as 'Australia/Brisbane'. Using the API Getting the current time The current time is represented by the Clock class. The class is abstract, so you can not create instances of it. The systemUTC() static method will return the current time based on your system clock and set to UTC. import javax.time.Clock; Clock clock = Clock.systemUTC(); To use the default time zone on your system there also is systemDefaultZone(). Clock clock = Clock.systemDefaultZone(); The millis() method can then be used to access the current time in milliseconds from January 1st, 1970. This shows, that the Clock class and all subclasses are wrapped around System.currentTimeMillis(). Clock clock = Clock.systemDefaultZone(); long time = clock.millis(); Working with time zones To work with time zones you need to import the ZoneId class. The class provides a method to get the default system time zone: import javax.time.ZoneId; import javax.time.Clock; ZoneId zone = ZoneId.systemDefault(); Clock clock = Clock.system(zone); As seen above, the ZoneId can then be used to get an instance of a Clock with that time zone. Other time zones can be accessed by their name, e.g.: ZoneId zone = ZoneId.of("Europe/Berlin"); Clock clock = Clock.system(zone); Getting human date and time Working with a time represented in a single long variable is not what we wanted. We want to work with objects that represent human readable time. The LocalDate, LocalTime and LocalDateTime classes do just that. import javax.time.LocalDate; // The now() method returns the current DateTime LocalDate date = LocalDate.now(); System.out.printf("%s-%s-%s", date.getYear(), date.getMonthValue(), date.getDayOfMonth() ); Using LocalDate to print the current date Doing calculations with times and dates One of the most important functionalities of JSR-310 is that you can do calculations with dates and times. The API makes it very easy to do that. import javax.time.LocalTime; import javax.time.Period; import static javax.time.calendrical.LocalPeriodUnit.HOURS; Period p = Period.of(5, HOURS); LocalTime time = LocalTime.now(); LocalTime newTime; newTime = time.plus(5, HOURS); // or newTime = time.plusHours(5); // or newTime = time.plus(p); Three ways of adding 5 hours to the current time Each class that represents human time implements the AdjustableDateTime interface. The interface requires the plus and the minus method that take a value and a PeriodUnit as argument. Conclusion This article gave a (very) brief introduction into the new date and time API that will ship with Java 8. The API seems to be very consistent and well thought through and provides many ways to interact with dates and times. Upon release of Java 8 the API will be moved from the javax.time package over to java.time, so there will be no conflict if you start using the current implementation.
September 25, 2012
by Fabian Becker
· 78,738 Views
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Nested Data Structures, and non-1NF design in PostgreSQL
This has been adapted from an ongoing series currently running on my blog. It has been adapted to be more self-contained, and rely less on other blog entries. For more see http://ledgersmbdev.blogspot.com PostgreSQL provides a very advanced set of tools for doing data modelling in ways which drift back and forth across a relational and non-relational divide. While it is generally a good idea to make the database relational first, and add objects later, the principles of object-relational database design allow you to do a lot more with PostgreSQL than you can on many other database platforms. This article will discuss the use of non-first-normal-form designs, in particular the storage of arrays of tuples in columns to simulate a nested table. The possible uses and problems of such a design will be discussed in detail. One of the promises of object-relational modelling is the ability to address information modelling on complex and nested data structures. Nested data structures bring considerable richness to the database, which is lost in a pure, flat, relational model. Nested data structures can be used to model tuple constraints in ways that are impossible to do when looking at flat data structures, at least as long as those constraints are limited to the information in a single tuple. At the same time there are cases where they simplify things and cases where they complicate things. This is true both in the case of using these for storage and for interfacing with stored procedures. PostgreSQL allows for nested tuples to be stored in a database, and for arrays of tuples. Other ORDBMS's allow something similar (Informix, DB2, and Oracle all support nested tables). Nested tables in PostgreSQL provide a number of gotchas, and additionally exposing the data in them to relational queries takes some extra work. In this post we will look at modelling general ledger transactions using a nested table approach, and both the benefits and limitations of this approach. In general this trades one set of problems for another and it is important to recognize the problems going in. The storage example came out of a brainstorming session I had with Marc Balmer of Micro Systems, though it is worth noting that this is not the solution they use in their products, nor is it the approach currently used by LedgerSMB. Basic Table Structure: The basic data schema will end up looking like this: CREATE TABLE journal_type ( id serial not null unique, label text primary key ); CREATE TABLE account ( id serial not null unique, control_code text primary key, -- account number description text ); CREATE TYPE journal_line_type AS ( account_id int, amount numeric ); CREATE TABLE journal_entry ( id serial not null unique, journal_type int references journal_type(id), source_document_id text,-- for example invoice number date_posted date not null, description text, line_items journal_line_type[], PRIMARY KEY (journal_type, source_document_id) ); This schema has a number of obvious gotchas and cannot, by itself, guarantee the sorts of things we want to do. However, using object-relational modelling we can fix these in ways that cannot do in a purely relational schema. The main problems are: First, since this is a double entry model, we need a constraint that says that the sum of the amounts of the lines must always equal zero. However, if we just add a sum() aggregate, we will end up with it summing every record in the db every time we do an insert, which is not what we want. We also want to make sure that no account_id's are null and no amounts are null. Additionally it is not possible in the schema above to easily expose the journal line information to purely relational tools. However we can use a VIEW to do this, though this produces yet more problems. Finally referential integrity enforcement between the account lines and accounts cannot be done declaratively. We will have to create TRIGGERs to enforce this manually. These problems are traded off against the fact that the relational model does not allow for the first problem to be solved at all so we trade off the fact that we have some solutions which are a bit of a pain for the fact that we have some solutions at all. Nested Table Constraints If we simply had a tuple as a column, we could look inside the tuple with check constraints. Something like check((column).subcolumn is not null). However in this case we cannot do that because we need to aggregate on a set of tuples attached to the row. To do this instead we create a set of table methods for managing the constraints: CREATE OR REPLACE FUNCTION is_balanced(journal_entry) RETURNS BOOL LANGUAGE SQL AS $$ SELECT sum(amount) = 0 FROM unnest($1.line_items); $$; CREATE OR REPLACE FUNCTION has_no_null_account_ids(journal_entry) RETURNS BOOL LANGUAGE SQL AS $$ SELECT bool_and(account_id is not null) FROM unnest($1.line_items); $$; CREATE OR REPLACE FUNCTION has_no_null_amounts(journal_entry) RETURNS BOOL LANGUAGE SQL AS $$ select bool_and(amount is not null) from unnest($1.line_items); $$; We can then create our constraints. Note that because we have to create the methods first, we have to add our constraints after the functions are defined, and these are added after the table is constructed. I have gone ahead and given these friendly names so that errors are easier for people (and machines) to process and handle. ALTER TABLE journal_entry ADD CONSTRAINT is_balanced CHECK ((journal_entry).is_balanced); ALTER TABLE journal_entry ADD CONSTRAINT has_no_null_account_ids CHECK ((journal_entry).has_no_null_account_ids); ALTER TABLE journal_entry ADD CONSTRAINT has_no_null_amounts CHECK ((journal_entry).has_no_null_amounts); Now we have integrity constraints reaching into our nested data. So let's test this out. insert into journal_type (label) values ('General'); We will re-use the account data from the previous post: or_examples=# select * from account; id | control_code | description ----+--------------+------------- 1 | 1500 | Inventory 2 | 4500 | Sales 3 | 5500 | Purchase (3 rows) Let's try inserting a few meaningless transactions, some of which violate our constraints: insert into journal_entry (journal_type, source_document_id, date_posted, description, line_items) values (1, 'ref-10001', now()::date, 'This is a test', ARRAY[row(1, 100)::journal_line_type]); ERROR: new row for relation "journal_entry" violates check constraint "is_balanced" So far so good. insert into journal_entry (journal_type, source_document_id, date_posted, description, line_items) values (1, 'ref-10001', now()::date, 'This is a test', ARRAY[row(1, 100)::journal_line_type, row(null, -100)::journal_line_type]); ERROR: new row for relation "journal_entry" violates check constraint "has_no_null_account_ids" Still good. insert into journal_entry (journal_type, source_document_id, date_posted, description, line_items) values (1, 'ref-10001', now()::date, 'This is a test', ARRAY[row(1, 100)::journal_line_type, row(2, -100)::journal_line_type, row(3, NULL)::journal_line_type]) ERROR: new row for relation "journal_entry" violates check constraint "has_no_null_amounts" Great. All constraints working properly. Let's try inserting a valid row: insert into journal_entry (journal_type, source_document_id, date_posted, description, line_items) values (1, 'ref-10001', now()::date, 'This is a test', ARRAY[row(1, 100)::journal_line_type, row(2, -100)::journal_line_type]); And it works! or_examples=# select * from journal_entry; id | journal_type | source_document_id | date_posted | description | li ne_items ----+--------------+--------------------+-------------+----------------+------------------------ 5 | 1 | ref-10001 | 2012-08-23 | This is a test | {"(1,100)","(2,-100)"} (1 row) Break-Out Views A second major problem that we will be facing with this schema is that if someone wants to create a report using a reporting tool that only really supports relational data very well, then the financial data will be opaque and not available. This scenario is one of the reasons why I think it is important generally to push the relational model to its breaking point before looking at object-relational functions. Consequently I think when doing nested tables it is important to ensure that the data in them is available through a relational interface, in this case, a view. In this case, we may want to model debits and credits in a way which is re-usable, so we will start by creating two type methods: CREATE OR REPLACE FUNCTION debits(journal_line_type) RETURNS NUMERIC LANGUAGE SQL AS $$ SELECT CASE WHEN $1.amount < 0 THEN $1.amount * -1 ELSE NULL END $$; CREATE OR REPLACE FUNCTION credits(journal_line_type) RETURNS NUMERIC LANGUAGE SQL AS $$ SELECT CASE WHEN $1.amount > 0 THEN $1.amount ELSE NULL END $$; Now we can use these as virtual columns anywhere a journal_line_type is used. The view definition itself is rather convoluted and this may impact performance. I am waiting for the LATERAL construct to become available which will make this easier. CREATE VIEW journal_line_items AS SELECT id AS journal_entry_id, (li).*, (li).debits, (li).credits FROM (SELECT je.*, unnest(line_items) li FROM journal_entry je) j; Remember li.debits and li.credits gets turned by the parser into debits(li) and credits(li), allowing for class.method notation here. Testing this out: SELECT * FROM journal_line_items; gives us journal_entry_id | account_id | amount | debits | credits ------------------+------------+--------+--------+--------- 5 | 1 | 100 | | 100 5 | 2 | -100 | 100 | 6 | 1 | 200 | | 200 6 | 3 | -200 | 200 | As you can see, this works. Now people with purely relational tools can access the information in the nested table. In general it is almost always worth creating break-out views of this sort where nested data is stored. However it is important to note that with larger data sets this is insufficient because indexing considerations makes it hard to look up specific information on a row level. This may or may not be the end of the world depending on data set size. Referential Integrity Controls The final problem is that relational integrity is not a well defined concept for nested data. For this reason, if we value relational integrity and foreign keys are involved, we must find ways of enforcing these. The simplest solution is a trigger which runs on insert, update, or delete, and manages another relation which can be used as a proxy for relational integrity checks. For example, we could: CREATE TABLE je_account ( je_id int references journal_entry (id), account_id int references account(id), primary key (je_id, account_id) ); This will be a very narrow table and so should be quick to search. It may also be useful in determining which accounts to look at for transactions if we need to do that. This table could then be used to optimize queries. To maintain the table we need to recognize that never ever will a journal entry's line items be updated or deleted. This is due to the need to maintain clear audit controls and trails. We may add other flags to the table to indicate transactions but we can handle insert, update, and delete conditions with a trigger, namely: CREATE FUNCTION je_ri_management() RETURNS TRIGGER LANGUAGE PLPGSQL AS $$ DECLARE accounts int[]; BEGIN IF TG_OP ILIKE 'INSERT' THEN INSERT INTO je_account (je_id, account_id) SELECT NEW.id, account_id FROM unnest(NEW.line_items) GROUP BY account_id; RETURN NEW; ELSIF TG_OP ILIKE 'UPDATE' THEN IF NEW.line_items <> OLD.line_items THEN RAISE EXCEPTION 'Cannot journal entry line items!'; ELSE RETURN NEW; END IF; ELSIF TG_OP ILIKE 'DELETE' THEN RAISE EXCEPTION 'Cannot delete journal entries!'; ELSE RAISE EXCEPTION 'Invalid TG_OP in trigger'; END IF; END; $$; Then we add the trigger with: CREATE TRIGGER je_breakout_for_ri AFTER INSERT OR UPDATE OR DELETE ON journal_entry FOR EACH ROW EXECUTE PROCEDURE je_ri_management(); The final invalid TG_OP could be omitted but this is not a bad check to have. Let's try this out: insert into journal_entry (journal_type, source_document_id, date_posted, description, line_items) values (1, 'ref-10003', now()::date, 'This is a test', ARRAY[row(1, 200)::journal_line_type, row(3, -200)::journal_line_type]); or_examples=# select * from je_account; je_id | account_id -------+------------ 10 | 3 10 | 1 (2 rows) In this way referential integrity can be enforced. Solution 2.0: Refactoring the above to eliminate the view. The above solution will work great for small businesses but for larger businesses, querying this data will become slow for certain kinds of reports. Storage here is tied to a specific criteria, and indexing is somewhat problematic. There are ways we can address this, but they are not always optimal. At the same time our work is simplified because the actual accounting details are append-only. One solution to this is to refactor the above solution. Instead of: Main table Relational view Materialized view for referential integrity checking we can have: Main table, with tweaked storage for line items Materialized view for RI checking and relational access Unfortunately this sort of refactoring after the fact isn't simple. Typically you want to convert the journal_line_type type to a journal_line_type table, and inherit this in your materialized view table. You cannot simply drop and recreate since the column you are storing the data in is dependent on the structure. The solution is to rename the type, create a new one in its place. This must be done manually and there is no current capability to copy a composite type's structure into a table. You will then need to create a cast and a cast function. Then, when you can afford the downtime, you will want to convert the table to the new type. It is quite possible that the downtime will be delayed and you will have an extended time period where you are half-way through migrating the structure of your database. You can, however, decide to create a cast between the table and the type, perhaps an implicit one (though this is not inherited) and use this to centralize your logic. Unfortunately this leads to duplication-related complexity and in an ideal world would be avoided. However, assuming that the downtime ends up being tolerable, the resulting structures will end up such that they can be more readily optimized for a variety of workloads. In this regard you would have a main table, most likely with line_items moved to extended storage, whose function is to model journal entries as journal entries and apply relevant constraints, and a second table which models journal entry lines as independent lines. This also simplifies some of the constraint issues on the first table, and makes the modelling easier because we only have to look into the nested storage where we are looking at subset constraints. This section then provides a warning regarding the use of advanced ORDBMS functionality, namely that it is easy to get tunnel vision and create problems for the future. The complexity cost here is so high, that the primary model should generally remain relational, with things like nested storage primarily used to create constraints that cannot be effectively modelled otherwise. However, this becomes a great deal more complicated where values may be update or deleted. Here, however, we have a relatively simple case regarding data writes combined with complex constraints that cannot be effectively expressed in normalized, relational SQL. Therefore the standard maintenance concerns that counsel against duplicating information may give way to the fact that such duplication allows for richer constraints. Now, if we had been aware of the problems going in we would have chosen this structure all along. Our design would have been: CREATE TYPE journal_line AS ( entry_id bigserial primary key, --only possible key je_id int not null, account_id int, amount numeric ); After creating the journal entry table we'd: ALTER TABLE journal_line ADD FOREIGN KEY (je_id) REFERENCES journal_entry(id); If we have to handle purging old data we can make that key ON DELETE CASCADE. And the lines would have been of this type instead. We can then get rid of all constraints and their supporting functions other than the is_balanced one. Our debit and credit functions then also reference this type. Our trigger then looks like: CREATE FUNCTION je_ri_management() RETURNS TRIGGER LANGUAGE PLPGSQL AS $$ DECLARE accounts int[]; BEGIN IF TG_OP ILIKE 'INSERT' THEN INSERT INTO journal_line (je_id, account_id, amount) SELECT NEW.id, account_id, amount FROM unnest(NEW.line_items); RETURN NEW; ELSIF TG_OP ILIKE 'UPDATE' THEN RAISE EXCEPTION 'Cannot journal entry line items!'; ELSIF TG_OP ILIKE 'DELETE' THEN RAISE EXCEPTION 'Cannot delete journal entries!'; ELSE RAISE EXCEPTION 'Invalid TG_OP in trigger'; END IF; END; $$; Approval workflows can be handled with a separate status table with its own constraints. Deletions of old information (up to a specific snapshot) can be handled by a stored procedure which is unit tested and disables this trigger before purging data. This system has the advantage of having several small components which are all complete and easily understood, and it is made possible because the data is exclusively append-only. As you can see from the above examples, nested data structures greatly complicate the data model and create problems with relational math that must be addressed if data logic will remain meaningful. This is a complex field, and it adds a lot of complexity to storage. In general, these are best avoided in actual data storage except where this approach makes formerly insurmountable problems manageable. Moreover, they add complexity to optimization once data gets large. Thus while non-atomic fields in this regard make sense as an initial point of entry in some narrow cases, as a point of actual query, they are very rarely the right approaches. It is possible that, at some point, nested storage will be able to have its own indexes, foreign keys, etc. but I cannot imagine this being a high priority and so it isn't clear that this will ever happen. In general, it usually makes the most sense to simply store the data in a pseudo-normalized way, with any non-1NF designs being the initial point of entry in a linear write model. Nested Data Structures as Interfaces Nested data structures as interfaces to stored procedures are a little more manageable. The main difficulties are in application-side data construction and output parsing. Some languages handle this more easily than others. Upper-level construction and handling of these structures is relatively straight-forward on the database-side and poses none of these problems. However, they do cause additional complexity and this must be managed carefully. The biggest issue when interfacing with an application is that ROW types are not usually automatically constructed by application-level frameworks even if they have arrays. This leaves the programmer to choose between unstructured text arrays which are fundamentally non-discoverable (and thus brittle), and arrays of tuples which are discoverable but require a lot of additional application code to handle. At the same time as a chicken and egg problem, frameworks will not add handling for this sort of problem unless people are already trying to do it. So my general recommendation is to use nested data types everywhere in the database sparingly, only where the benefits clearly outweigh the complexity costs. Complexity costs are certainly lower in the interface level and there are many more cases where it these techniques are net wins there, but that does not mean that they should be routinely used even there.
September 25, 2012
by Chris Travers
· 21,036 Views
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