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How-To Tutorials

7019 Articles
article-image-getting-started-multiplayer-game-programming
Packt
04 Jun 2015
33 min read
Save for later

Getting Started with Multiplayer Game Programming

Packt
04 Jun 2015
33 min read
In this article by Rodrigo Silveira author of the book Multiplayer gaming with HTML5 game development, if you're reading this, chances are pretty good that you are already a game developer. That being the case, then you already know just how exciting it is to program your own games, either professionally or as a highly gratifying hobby that is very time-consuming. Now you're ready to take your game programming skills to the next level—that is, you're ready to implement multiplayer functionality into your JavaScript-based games. (For more resources related to this topic, see here.) In case you have already set out to create multiplayer games for the Open Web Platform using HTML5 and JavaScript, then you may have already come to realize that a personal desktop computer, laptop, or a mobile device is not particularly the most appropriate device to share with another human player for games in which two or more players share the same game world at the same time. Therefore, what is needed in order to create exciting multiplayer games with JavaScript is some form of networking technology. We will be discussing the following principles and concepts: The basics of networking and network programming paradigms Socket programming with HTML5 Programming a game server and game clients Turn-based multiplayer games Understanding the basics of networking It is said that one cannot program games that make use of networking without first understanding all about the discipline of computer networking and network programming. Although having a deep understanding of any topic can be only beneficial to the person working on that topic, I don't believe that you must know everything there is to know about game networking in order to program some pretty fun and engaging multiplayer games. Saying that is the case is like saying that one needs to be a scholar of the Spanish language in order to cook a simple burrito. Thus, let us take a look at the most basic and fundamental concepts of networking. After you finish reading this article, you will know enough about computer networking to get started, and you will feel comfortable adding multiplayer aspects to your games. One thing to keep in mind is that, even though networked games are not nearly as old as single-player games, computer networking is actually a very old and well-studied subject. Some of the earliest computer network systems date back to the 1950s. Though some of the techniques have improved over the years, the basic idea remains the same: two or more computers are connected together to establish communication between the machines. By communication, I mean data exchange, such as sending messages back and forth between the machines, or one of the machines only sends the data and the other only receives it. With this brief introduction to the concept of networking, you are now grounded in the subject of networking, enough to know what is required to network your games—two or more computers that talk to each other as close to real time as possible. By now, it should be clear how this simple concept makes it possible for us to connect multiple players into the same game world. In essence, we need a way to share the global game data among all the players who are connected to the game session, then continue to update each player about every other player. There are several different techniques that are commonly used to achieve this, but the two most common approaches are peer-to-peer and client-server. Both techniques present different opportunities, including advantages and disadvantages. In general, neither is particularly better than the other, but different situations and use cases may be better suited for one or the other technique. Peer-to-peer networking A simple way to connect players into the same virtual game world is through the peer-to-peer architecture. Although the name might suggest that only two peers ("nodes") are involved, by definition a peer-to-peer network system is one in which two or more nodes are connected directly to each other without a centralized system orchestrating the connection or information exchange. On a typical peer-to-peer setup, each peer serves the same function as every other one—that is, they all consume the same data and share whatever data they produce so that others can stay synchronized. In the case of a peer-to-peer game, we can illustrate this architecture with a simple game of Tic-tac-toe. Once both the players have established a connection between themselves, whoever is starting the game makes a move by marking a cell on the game board. This information is relayed across the wire to the other peer, who is now aware of the decision made by his or her opponent, and can thus update their own game world. Once the second player receives the game's latest state that results from the first player's latest move, the second player is able to make a move of their own by checking some available space on the board. This information is then copied over to the first player who can update their own world and continue the process by making the next desired move. The process goes on until one of the peers disconnects or the game ends as some condition that is based on the game's own business logic is met. In the case of the game of Tic-tac-toe, the game would end once one of the players has marked three spaces on the board forming a straight line or if all nine cells are filled, but neither player managed to connect three cells in a straight path. Some of the benefits of peer-to-peer networked games are as follows: Fast data transmission: Here, the data goes directly to its intended target. In other architectures, the data could go to some centralized node first, then the central node (or the "server") contacts the other peer, sending the necessary updates. Simpler setup: You would only need to think about one instance of your game that, generally speaking, handles its own input, sends its input to other connected peers, and handles their output as input for its own system. This can be especially handy in turn-based games, for example, most board games such as Tic-tac-toe. More reliability: Here one peer that goes offline typically won't affect any of the other peers. However, in the simple case of a two-player game, if one of the players is unable to continue, the game will likely cease to be playable. Imagine, though, that the game in question has dozens or hundreds of connected peers. If a handful of them suddenly lose their Internet connection, the others can continue to play. However, if there is a server that is connecting all the nodes and the server goes down, then none of the other players will know how to talk to each other, and nobody will know what is going on. On the other hand, some of the more obvious drawbacks of peer-to-peer architecture are as follows: Incoming data cannot be trusted: Here, you don't know for sure whether or not the sender modified the data. The data that is input into a game server will also suffer from the same challenge, but once the data is validated and broadcasted to all the other peers, you can be more confident that the data received by each peer from the server will have at least been sanitized and verified, and will be more credible. Fault tolerance can be very low: If enough players share the game world, one or more crashes won't make the game unplayable to the rest of the peers. Now, if we consider the many cases where any of the players that suddenly crash out of the game negatively affect the rest of the players, we can see how a server could easily recover from the crash. Data duplication when broadcasting to other peers: Imagine that your game is a simple 2D side scroller, and many other players are sharing that game world with you. Every time one of the players moves to the right, you receive the new (x, y) coordinates from that player, and you're able to update your own game world. Now, imagine that you move your player to the right by a very few pixels; you would have to send that data out to all of the other nodes in the system. Overall, peer-to-peer is a very powerful networking architecture and is still widely used by many games in the industry. Since current peer-to-peer web technologies are still in their infancy, most JavaScript-powered games today do not make use of peer-to-peer networking. For this and other reasons that should become apparent soon, we will focus almost exclusively on the other popular networking paradigm, namely, the client-server architecture. Client-server networking The idea behind the client-server networking architecture is very simple. If you squint your eyes hard enough, you can almost see a peer-to-peer graph. The most obvious difference between them, is that, instead of every node being an equal peer, one of the nodes is special. That is, instead of every node connecting to every other node, every node (client) connects to a main centralized node called the server. While the concept of a client-server network seems clear enough, perhaps a simple metaphor might make it easier for you to understand the role of each type of node in this network format as well as differentiate it from peer-to-peer . In a peer-to-peer network, you can think of it as a group of friends (peers) having a conversation at a party. They all have access to all the other peers involved in the conversation and can talk to them directly. On the other hand, a client-server network can be viewed as a group of friends having dinner at a restaurant. If a client of the restaurant wishes to order a certain item from the menu, he or she must talk to the waiter, who is the only person in that group of people with access to the desired products and the ability to serve the products to the clients. In short, the server is in charge of providing data and services to one or more clients. In the context of game development, the most common scenario is when two or more clients connect to the same server; the server will keep track of the game as well as the distributed players. Thus, if two players are to exchange information that is only pertinent to the two of them, the communication will go from the first player to and through the server and will end up at the other end with the second player. Following the example of the two players involved in a game of Tic-tac-toe, we can see how similar the flow of events is on a client-server model. Again, the main difference is that players are unaware of each other and only know what the server tells them. While you can very easily mimic a peer-to-peer model by using a server to merely connect the two players, most often the server is used much more actively than that. There are two ways to engage the server in a networked game, namely in an authoritative and a non-authoritative way. That is to say, you can have the enforcement of the game's logic strictly in the server, or you can have the clients handle the game logic, input validation, and so on. Today, most games using the client-server architecture actually use a hybrid of the two (authoritative and non-authoritative servers). For all intents and purposes, however, the server's purpose in life is to receive input from each of the clients and distribute that input throughout the pool of connected clients. Now, regardless of whether you decide to go with an authoritative server instead of a non-authoritative one, you will notice that one of challenges with a client-server game is that you will need to program both ends of the stack. You will have to do this even if your clients do nothing more than take input from the user, forward it to the server, and render whatever data they receive from the server; if your game server does nothing more than forward the input that it receives from each client to every other client, you will still need to write a game client and a game server. We will discuss game clients and servers later. For now, all we really need to know is that these two components are what set this networking model apart from peer-to-peer. Some of the benefits of client-server networked games are as follows: Separation of concerns: If you know anything about software development, you know that this is something you should always aim for. That is, good, maintainable software is written as discrete components where each does one "thing", and it is done well. Writing individual specialized components lets you focus on performing one individual task at a time, making your game easier to design, code, test, reason, and maintain. Centralization: While this can be argued against as well as in favor of, having one central place through which all communication must flow makes it easier to manage such communication, enforce any required rules, control access, and so forth. Less work for the client: Instead of having a client (peer) in charge of taking input from the user as well as other peers, validating all the input, sharing data among other peers, rendering the game, and so on, the client can focus on only doing a few of these things, allowing the server to offload some of this work. This is particularly handy when we talk about mobile gaming, and how much subtle divisions of labor can impact the overall player experience. For example, imagine a game where 10 players are engaged in the same game world. In a peer-to-peer setup, every time one player takes an action, he or she would need to send that action to nine other players (in other words, there would need to be nine network calls, boiling down to more mobile data usage). On the other hand, on a client-server configuration, one player would only need to send his or her action to one of the peers, that is, the server, who would then be responsible for sending that data to the remaining nine players. Common drawbacks of client-server architectures, whether or not the server is authoritative, are as follows: Communication takes longer to propagate: In the very best possible scenario imaginable, every message sent from the first player to the second player would take twice as long to be delivered as compared to a peer-to-peer connection. That is, the message would be first sent from the first player to the server and then from the server to the second player. There are many techniques that are used today to solve the latency problem faced in this scenario, some of which we will discuss in much more depth later. However, the underlying dilemma will always be there. More complexity due to more moving parts: It doesn't really matter how you slice the pizza; the more code you need to write (and trust me, when you build two separate modules for a game, you will write more code), the greater your mental model will have to be. While much of your code can be reused between the client and the server (especially if you use well-established programming techniques, such as object-oriented programming), at the end of the day, you need to manage a greater level of complexity. Single point of failure and network congestion: Up until now, we have mostly discussed the case where only a handful of players participates in the same game. However, the more common case is that a handful of groups of players play different games at the same time. Using the same example of the two-player game of Tic-tac-toe, imagine that there are thousands of players facing each other in single games. In a peer-to-peer setup, once a couple of players have directly paired off, it is as though there are no other players enjoying that game. The only thing to keep these two players from continuing their game is their own connection with each other. On the other hand, if the same thousands of players are connected to each other through a server sitting between the two, then two singled out players might notice severe delays between messages because the server is so busy handling all of the messages from and to all of the other people playing isolated games. Worse yet, these two players now need to worry about maintaining their own connection with each other through the server, but they also hope that the server's connection between them and their opponent will remain active. All in all, many of the challenges involved in client-server networking are well studied and understood, and many of the problems you're likely to face during your multiplayer game development will already have been solved by someone else. Client-server is a very popular and powerful game networking model, and the required technology for it, which is available to us through HTML5 and JavaScript, is well developed and widely supported. Networking protocols – UDP and TCP By discussing some of the ways in which your players can talk to each other across some form of network, we have yet only skimmed over how that communication is actually done. Let us then describe what protocols are and how they apply to networking and, more importantly, multiplayer game development. The word protocol can be defined as a set of conventions or a detailed plan of a procedure [Citation [Def. 3,4]. (n.d.). In Merriam Webster Online, Retrieved February 12, 2015, from http://www.merriam-webster.com/dictionary/protocol]. In computer networking, a protocol describes to the receiver of a message how the data is organized so that it can be decoded. For example, imagine that you have a multiplayer beat 'em up game, and you want to tell the game server that your player just issued a kick command and moved 3 units to the left. What exactly do you send to the server? Do you send a string with a value of "kick", followed by the number 3? Otherwise, do you send the number first, followed by a capitalized letter "K", indicating that the action taken was a kick? The point I'm trying to make is that, without a well-understood and agreed-upon protocol, it is impossible to successfully and predictably communicate with another computer. The two networking protocols that we'll discuss in the section, and that are also the two most widely used protocols in multiplayer networked games, are the Transmission Control Protocol (TCP) and the User Datagram Protocol (UDP). Both protocols provide communication services between clients in a network system. In simple terms, they are protocols that allow us to send and receive packets of data in such a way that the data can be identified and interpreted in a predictable way. When data is sent through TCP, the application running in the source machine first establishes a connection with the destination machine. Once a connection has been established, data is transmitted in packets in such a way that the receiving application can then put the data back together in the appropriate order. TCP also provides built-in error checking mechanisms so that, if a packet is lost, the target application can notify the sender application, and any missing packets are sent again until the entire message is received. In short, TCP is a connection-based protocol that guarantees the delivery of the full data in the correct order. Use cases where this behavior is desirable are all around us. When you download a game from a web server, for example, you want to make sure that the data comes in correctly. You want to be sure that your game assets will be properly and completely downloaded before your users start playing your game. While this guarantee of delivery may sound very reassuring, it can also be thought of as a slow process, which, as we'll see briefly, may sometimes be more important than knowing that the data will arrive in full. In contrast, UDP transmits packets of data (called datagrams) without the use of a pre-established connection. The main goal of the protocol is to be a very fast and frictionless way of sending data towards some target application. In essence, you can think of UDP as the brave employees who dress up as their company's mascot and stand outside their store waving a large banner in the hope that at least some of the people driving by will see them and give them their business. While at first, UDP may seem like a reckless protocol, the use cases that make UDP so desirable and effective includes the many situations when you care more about speed than missing packets a few times, getting duplicate packets, or getting them out of order. You may also want to choose UDP over TCP when you don't care about the reply from the receiver. With TCP, whether or not you need some form of confirmation or reply from the receiver of your message, it will still take the time to reply back to you, at least acknowledging that the message was received. Sometimes, you may not care whether or not the server received the data. A more concrete example of a scenario where UDP is a far better choice over TCP is when you need a heartbeat from the client letting the server know if the player is still there. If you need to let your server know that the session is still active every so often, and you don't care if one of the heartbeats get lost every now and again, then it would be wise to use UDP. In short, for any data that is not mission-critical and you can afford to lose, UDP might be the best option. In closing, keep in mind that, just as peer-to-peer and client-server models can be built side by side, and in the same way your game server can be a hybrid of authoritative and non-authoritative, there is absolutely no reason why your multiplayer games should only use TCP or UDP. Use whichever protocol a particular situation calls for. Network sockets There is one other protocol that we'll cover very briefly, but only so that you can see the need for network sockets in game development. As a JavaScript programmer, you are doubtlessly familiar with Hypertext Transfer Protocol (HTTP). This is the protocol in the application layer that web browsers use to fetch your games from a Web server. While HTTP is a great protocol to reliably retrieve documents from web servers, it was not designed to be used in real-time games; therefore, it is not ideal for this purpose. The way HTTP works is very simple: a client sends a request to a server, which then returns a response back to the client. The response includes a completion status code, indicating to the client that the request is either in process, needs to be forwarded to another address, or is finished successfully or erroneously. There are a handful of things to note about HTTP that will make it clear that a better protocol is needed for real-time communication between the client and server. Firstly, after each response is received by the requester, the connection is closed. Thus, before making each and every request, a new connection must be established with the server. Most of the time, an HTTP request will be sent through TCP, which, as we've seen, can be slow, relatively speaking. Secondly, HTTP is by design a stateless protocol. This means that, every time you request a resource from a server, the server has no idea who you are and what is the context of the request. (It doesn't know whether this is your first request ever or if you're a frequent requester.) A common solution to this problem is to include a unique string with every HTTP request that the server keeps track of, and can thus provide information about each individual client on an ongoing basis. You may recognize this as a standard session. The major downside with this solution, at least with regard to real-time gaming, is that mapping a session cookie to the user's session takes additional time. Finally, the major factor that makes HTTP unsuitable for multiplayer game programming is that the communication is one way—only the client can connect to the server, and the server replies back through the same connection. In other words, the game client can tell the game server that a punch command has been entered by the user, but the game server cannot pass that information along to other clients. Think of it like a vending machine. As a client of the machine, we can request specific items that we wish to buy. We formalize this request by inserting money into the vending machine, and then we press the appropriate button. Under no circumstance will a vending machine issue commands to a person standing nearby. That would be like waiting for a vending machine to dispense food, expecting people to deposit the money inside it afterwards. The answer to this lack of functionality in HTTP is pretty straightforward. A network socket is an endpoint in a connection that allows for two-way communication between the client and the server. Think of it more like a telephone call, rather than a vending machine. During a telephone call, either party can say whatever they want at any given time. Most importantly, the connection between both parties remains open throughout the duration of the conversation, making the communication process highly efficient. WebSocket is a protocol built on top of TCP, allowing web-based applications to have two-way communication with a server. The way a WebSocket is created consists of several steps, including a protocol upgrade from HTTP to WebSocket. Thankfully, all of the heavy lifting is done behind the scenes by the browser and JavaScript. For now, the key takeaway here is that with a TCP socket (yes, there are other types of socket including UDP sockets), we can reliably communicate with a server, and the server can talk back to us as per the need. Socket programming in JavaScript Let's now bring the conversation about network connections, protocols, and sockets to a close by talking about the tools—JavaScript and WebSockets—that bring everything together, allowing us to program awesome multiplayer games in the language of the open Web. The WebSocket protocol Modern browsers and other JavaScript runtime environments have implemented the WebSocket protocol in JavaScript. Don't make the mistake of thinking that just because we can create WebSocket objects in JavaScript, WebSockets are part of JavaScript. The standard that defines the WebSocket protocol is language-agnostic and can be implemented in any programming language. Thus, before you start to deploy your JavaScript games that make use of WebSockets, ensure that the environment that will run your game uses an implementation of the ECMA standard that also implements WebSockets. In other words, not all browsers will know what to do when you ask for a WebSocket connection. For the most part, though, the latest versions, as of this writing, of the most popular browsers today (namely, Google Chrome, Safari, Mozilla Firefox, Opera, and Internet Explorer) implement the current latest revision of RFC 6455. Previous versions of WebSockets (such as protocol version - 76, 7, or 10) are slowly being deprecated and have been removed by some of the previously mentioned browsers. Probably the most confusing thing about the WebSocket protocol is the way each version of the protocol is named. The very first draft (which dates back to 2010), was named draft-hixie-thewebsocketprotocol-75. The next version was named draft-hixie-thewebsocketprotocol-76. Some people refer to these versions as 75 and 76, which can be quite confusing, especially since the fourth version of the protocol is named draft-ietf-hybi-thewebsocketprotocol-07, which is named in the draft as WebSocket Version 7. The current version of the protocol (RFC 6455) is 13. Let us take a quick look at the programming interface (API) that we'll use within our JavaScript code to interact with a WebSocket server. Keep in mind that we'll need to write both the JavaScript clients that use WebSockets to consume data as well as the WebSocket server, which uses WebSockets but plays the role of the server. The difference between the two will become apparent as we go over some examples. Creating a client-side WebSocket The following code snippet creates a new object of type WebSocket that connects the client to some backend server. The constructor takes two parameters; the first is required and represents the URL where the WebSocket server is running and expecting connections. The second URL, is an optional list of sub-protocols that the server may implement. var socket = new WebSocket('ws://www.game-domain.com'); Although this one line of code may seem simple and harmless enough, here are a few things to keep in mind: We are no longer in HTTP territory. The address to your WebSocket server now starts with ws:// instead of http://. Similarly, when we work with secure (encrypted) sockets, we would specify the server's URL as wss://, just like in https://. It may seem obvious to you, but a common pitfall that those getting started with WebSockets fall into is that, before you can establish a connection with the previous code, you need a WebSocket server running at that domain. WebSockets implement the same-origin security model. As you may have already seen with other HTML5 features, the same-origin policy states that you can only access a resource through JavaScript if both the client and the server are in the same domain. For those who are not familiar with the same-domain (also known as the same-origin) policy, the three things that constitute a domain, in this context, are the protocol, host, and port of the resource being accessed. In the previous example, the protocol, host, and port number were, respectively ws (and not wss, http, or ssh), www.game-domain.com (any sub-domain, such as game-domain.com or beta.game-domain.com would violate the same-origin policy), and 80 (by default, WebSocket connects to port 80, and port 443 when it uses wss). Since the server in the previous example binds to port 80, we don't need to explicitly specify the port number. However, had the server been configured to run on a different port, say 2667, then the URL string would need to include a colon followed by the port number that would need to be placed at the end of the host name, such as ws://www.game-domain.com:2667. As with everything else in JavaScript, WebSocket instances attempt to connect to the backend server asynchronously. Thus, you should not attempt to issue commands on your newly created socket until you're sure that the server has connected; otherwise, JavaScript will throw an error that may crash your entire game. This can be done by registering a callback function on the socket's onopen event as follows: var socket = new WebSocket('ws://www.game-domain.com'); socket.onopen = function(event) {    // socket ready to send and receive data }; Once the socket is ready to send and receive data, you can send messages to the server by calling the socket object's send method, which takes a string as the message to be sent. // Assuming a connection was previously established socket.send('Hello, WebSocket world!'); Most often, however, you will want to send more meaningful data to the server, such as objects, arrays, and other data structures that have more meaning on their own. In these cases, we can simply serialize our data as JSON strings. var player = {    nickname: 'Juju',    team: 'Blue' };   socket.send(JSON.stringify(player)); Now, the server can receive that message and work with it as the same object structure that the client sent it, by running it through the parse method of the JSON object. var player = JSON.parse(event.data); player.name === 'Juju'; // true player.team === 'Blue'; // true player.id === undefined; // true If you look at the previous example closely, you will notice that we extract the message that is sent through the socket from the data attribute of some event object. Where did that event object come from, you ask? Good question! The way we receive messages from the socket is the same on both the client and server sides of the socket. We must simply register a callback function on the socket's onmessage event, and the callback will be invoked whenever a new message is received. The argument passed into the callback function will contain an attribute named data, which will contain the raw string object with the message that was sent. socket.onmessage = function(event) {    event instanceof MessageEvent; // true      var msg = JSON.parse(event.data); }; Other events on the socket object on which you can register callbacks include onerror, which is triggered whenever an error related to the socket occurs, and onclose, which is triggered whenever the state of the socket changes to CLOSED; in other words, whenever the server closes the connection with the client for any reason or the connected client closes its connection. As mentioned previously, the socket object will also have a property called readyState, which behaves in a similar manner to the equally-named attribute in AJAX objects (or more appropriately, XMLHttpRequest objects). This attribute represents the current state of the connection and can have one of four values at any point in time. This value is an unsigned integer between 0 and 3, inclusive of both the numbers. For clarity, there are four accompanying constants on the WebSocket class that map to the four numerical values of the instance's readyState attribute. The constants are as follows: WebSocket.CONNECTING: This has a value of 0 and means that the connection between the client and the server has not yet been established. WebSocket.OPEN: This has a value of 1 and means that the connection between the client and the server is open and ready for use. Whenever the object's readyState attribute changes from CONNECTING to OPEN, which will only happen once in the object's life cycle, the onopen callback will be invoked. WebSocket.CLOSING: This has a value of 2 and means that the connection is being closed. WebSocket.CLOSED: This has a value of 3 and means that the connection is now closed (or could not be opened to begin with). Once the readyState has changed to a new value, it will never return to a previous state in the same instance of the socket object. Thus, if a socket object is CLOSING or has already become CLOSED, it will never OPEN again. In this case, you would need a new instance of WebSocket if you would like to continue to communicate with the server. To summarize, let us bring together the simple WebSocket API features that we discussed previously and create a convenient function that simplifies data serialization, error checking, and error handling when communicating with the game server: function sendMsg(socket, data) {    if (socket.readyState === WebSocket.OPEN) {      socket.send(JSON.stringify(data));        return true;    }      return false; }; Game clients Earlier, we talked about the architecture of a multiplayer game that was based on the client-server pattern. Since this is the approach we will take for the games that we'll be developing, let us define some of the main roles that the game client will fulfill. From a higher level, a game client will be the interface between the human player and the rest of the game universe (which includes the game server and other human players who are connected to it). Thus, the game client will be in charge of taking input from the player, communicating this to the server, receive any further instructions and information from the server, and then render the final output to the human player again. Depending on the type of game server used, the client can be more sophisticated than just an input application that renders static data received from the server. For example, the client could very well simulate what the game server will do and present the result of this simulation to the user while the server performs the real calculations and tells the results to the client. The biggest selling point of this technique is that the game would seem a lot more dynamic and real-time to the user since the client responds to input almost instantly. Game servers The game server is primarily responsible for connecting all the players to the same game world and keeping the communication going between them. However as you will soon realize, there may be cases where you will want the server to be more sophisticated than a routing application. For example, just because one of the players is telling the server to inform the other participants that the game is over, and the player sending the message is the winner, we may still want to confirm the information before deciding that the game is in fact over. With this idea in mind, we can label the game server as being of one of the two kinds: authoritative or non-authoritative. In an authoritative game server, the game's logic is actually running in memory (although it normally doesn't render any graphical output like the game clients certainly will) all the time. As each client reports the information to the server by sending messages through its corresponding socket, the server updates the current game state and sends the updates back to all of the players, including the original sender. This way we can be more certain that any data coming from the server has been verified and is accurate. In a non-authoritative server, the clients take on a much more involved part in the game logic enforcement, which gives the client a lot more trust. As suggested previously, what we can do is take the best of both worlds and create a mix of both the techniques. What we will do is, have a strictly authoritative server, but clients that are smart and can do some of the work on their own. Since the server has the ultimate say in the game, however, any messages received by clients from the server are considered as the ultimate truth and supersede any conclusions it came to on its own. Summary Overall, we discussed the basics of networking and network programming paradigms. We saw how WebSockets makes it possible to develop real-time, multiplayer games in HTML5. Finally, we implemented a simple game client and game server using widely supported web technologies and built a fun game of Tic-tac-toe. Resources for Article: Further resources on this subject: HTML5 Game Development – A Ball-shooting Machine with Physics Engine [article] Creating different font files and using web fonts [article] HTML5 Canvas [article]
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04 Jun 2015
25 min read
 In this article by Alex Libby, author of the book Mastering jQuery, we will examine some of the options available to help develop your skills even further. (For more resources related to this topic, see here.) Local or CDN, I wonder…? Which version…? Do I support old IE…? Installing jQuery is a thankless task that has to be done countless times by any developer—it is easy to imagine that person asking some of the questions. It is easy to imagine why most people go with the option of using a Content Delivery Network (CDN) link, but there is more to installing jQuery than taking the easy route! There are more options available, where we can be really specific about what we need to use—throughout this article, we will. We'll cover a number of topics, which include: Downloading and installing jQuery Customizing jQuery downloads Building from Git Using other sources to install jQuery Adding source map support Working with Modernizr as a fallback Intrigued? Let's get started. Downloading and installing jQuery As with all projects that require the use of jQuery, we must start somewhere—no doubt you've downloaded and installed jQuery a thousand times; let's just quickly recap to bring ourselves up to speed. If we browse to http://www.jquery.com/download, we can download jQuery using one of the two methods: downloading the compressed production version or the uncompressed development version. If we don't need to support old IE (IE6, 7, and 8), then we can choose the 2.x branch. If, however, you still have some diehards who can't (or don't want to) upgrade, then the 1.x branch must be used instead. To include jQuery, we just need to add this link to our page: <script src="http://code.jquery.com/jquery-X.X.X.js"></script> Here, X.X.X marks the version number of jQuery or the Migrate plugin that is being used in the page. Conventional wisdom states that the jQuery plugin (and this includes the Migrate plugin too) should be added to the <head> tag, although there are valid arguments to add it as the last statement before the closing <body> tag; placing it here may help speed up loading times to your site. This argument is not set in stone; there may be instances where placing it in the <head> tag is necessary and this choice should be left to the developer's requirements. My personal preference is to place it in the <head> tag as it provides a clean separation of the script (and the CSS) code from the main markup in the body of the page, particularly on lighter sites. I have even seen some developers argue that there is little perceived difference if jQuery is added at the top, rather than at the bottom; some systems, such as WordPress, include jQuery in the <head> section too, so either will work. The key here though is if you are perceiving slowness, then move your scripts to just before the <body> tag, which is considered a better practice. Using jQuery in a development capacity A useful point to note at this stage is that best practice recommends that CDN links should not be used within a development capacity; instead, the uncompressed files should be downloaded and referenced locally. Once the site is complete and is ready to be uploaded, then CDN links can be used. Adding the jQuery Migrate plugin If you've used any version of jQuery prior to 1.9, then it is worth adding the jQuery Migrate plugin to your pages. The jQuery Core team made some significant changes to jQuery from this version; the Migrate plugin will temporarily restore the functionality until such time that the old code can be updated or replaced. The plugin adds three properties and a method to the jQuery object, which we can use to control its behavior: Property or Method Comments jQuery.migrateWarnings This is an array of string warning messages that have been generated by the code on the page, in the order in which they were generated. Messages appear in the array only once even if the condition has occurred multiple times, unless jQuery.migrateReset() is called. jQuery.migrateMute Set this property to true in order to prevent console warnings from being generated in the debugging version. If this property is set, the jQuery.migrateWarnings array is still maintained, which allows programmatic inspection without console output. jQuery.migrateTrace Set this property to false if you want warnings but don't want traces to appear on the console. jQuery.migrateReset() This method clears the jQuery.migrateWarnings array and "forgets" the list of messages that have been seen already. Adding the plugin is equally simple—all you need to do is add a link similar to this, where X represents the version number of the plugin that is used: <script src="http://code.jquery.com/jquery-migrate- X.X.X.js"></script> If you want to learn more about the plugin and obtain the source code, then it is available for download from https://github.com/jquery/jquery-migrate. Using a CDN We can equally use a CDN link to provide our jQuery library—the principal link is provided by MaxCDN for the jQuery team, with the current version available at http://code.jquery.com. We can, of course, use CDN links from some alternative sources, if preferred—a reminder of these is as follows: Google (https://developers.google.com/speed/libraries/devguide#jquery) Microsoft (http://www.asp.net/ajaxlibrary/cdn.ashx#jQuery_Releases_on_the_CDN_0) CDNJS (http://cdnjs.com/libraries/jquery/) jsDelivr (http://www.jsdelivr.com/#%!jquery) Don't forget though that if you need, we can always save a copy of the file provided on CDN locally and reference this instead. The jQuery CDN will always have the latest version, although it may take a couple of days for updates to appear via the other links. Using other sources to install jQuery Right. Okay, let's move on and develop some code! "What's next?" I hear you ask. Aha! If you thought downloading and installing jQuery from the main site was the only way to do this, then you are wrong! After all, this is about mastering jQuery, so you didn't think I will only talk about something that I am sure you are already familiar with, right? Yes, there are more options available to us to install jQuery than simply using the CDN or main download page. Let's begin by taking a look at using Node. Each demo is based on Windows, as this is the author's preferred platform; alternatives are given, where possible, for other platforms. Using Node JS to install jQuery So far, we've seen how to download and reference jQuery, which is to use the download from the main jQuery site or via a CDN. The downside of this method is the manual work required to keep our versions of jQuery up to date! Instead, we can use a package manager to help manage our assets. Node.js is one such system. Let's take a look at the steps that need to be performed in order to get jQuery installed: We first need to install Node.js—head over to http://www.nodejs.org in order to download the package for your chosen platform; accept all the defaults when working through the wizard (for Mac and PC). Next, fire up a Node Command Prompt and then change to your project folder. In the prompt, enter this command: npm install jquery Node will fetch and install jQuery—it displays a confirmation message when the installation is complete: You can then reference jQuery by using this link: <name of drive>:websitenode_modulesjquerydistjquery.min.js. Node is now installed and ready for use—although we've installed it in a folder locally, in reality, we will most likely install it within a subfolder of our local web server. For example, if we're running WampServer, we can install it, then copy it into the /wamp/www/js folder, and reference it using http://localhost/js/jquery.min.js. If you want to take a look at the source of the jQuery Node Package Manager (NPM) package, then check out https://www.npmjs.org/package/jquery. Using Node to install jQuery makes our work simpler, but at a cost. Node.js (and its package manager, NPM) is primarily aimed at installing and managing JavaScript components and expects packages to follow the CommonJS standard. The downside of this is that there is no scope to manage any of the other assets that are often used within websites, such as fonts, images, CSS files, or even HTML pages. "Why will this be an issue?," I hear you ask. Simple, why make life hard for ourselves when we can manage all of these assets automatically and still use Node? Installing jQuery using Bower A relatively new addition to the library is the support for installation using Bower—based on Node, it's a package manager that takes care of the fetching and installing of packages from over the Internet. It is designed to be far more flexible about managing the handling of multiple types of assets (such as images, fonts, and CSS files) and does not interfere with how these components are used within a page (unlike Node). For the purpose of this demo, I will assume that you have already installed it; if not, you will need to revisit it before continuing with the following steps: Bring up the Node Command Prompt, change to the drive where you want to install jQuery, and enter this command: bower install jquery This will download and install the script, displaying the confirmation of the version installed when it has completed. The library is installed in the bower_components folder on your PC. It will look similar to this example, where I've navigated to the jquery subfolder underneath. By default, Bower will install jQuery in its bower_components folder. Within bower_components/jquery/dist/, we will find an uncompressed version, compressed release, and source map file. We can then reference jQuery in our script using this line: <script src="/bower_components/jquery/jquery.js"></script> We can take this further though. If we don't want to install the extra files that come with a Bower installation by default, we can simply enter this in a Command Prompt instead to just install the minified version 2.1 of jQuery: bower install http://code.jquery.com/jquery-2.1.0.min.js Now, we can be really clever at this point; as Bower uses Node's JSON files to control what should be installed, we can use this to be really selective and set Bower to install additional components at the same time. Let's take a look and see how this will work—in the following example, we'll use Bower to install jQuery 2.1 and 1.10 (the latter to provide support for IE6-8). In the Node Command Prompt, enter the following command: bower init This will prompt you for answers to a series of questions, at which point you can either fill out information or press Enter to accept the defaults. Look in the project folder; you should find a bower.json file within. Open it in your favorite text editor and then alter the code as shown here: {"ignore": [ "**/.*", "node_modules", "bower_components","test", "tests" ] ,"dependencies": {"jquery-legacy": "jquery#1.11.1","jquery-modern": "jquery#2.10"}} At this point, you have a bower.json file that is ready for use. Bower is built on top of Git, so in order to install jQuery using your file, you will normally need to publish it to the Bower repository. Instead, you can install an additional Bower package, which will allow you to install your custom package without the need to publish it to the Bower repository: In the Node Command Prompt window, enter the following at the prompt: npm install -g bower-installer When the installation is complete, change to your project folder and then enter this command line: bower-installer The bower-installer command will now download and install both the versions of jQuery. At this stage, you now have jQuery installed using Bower. You're free to upgrade or remove jQuery using the normal Bower process at some point in the future. If you want to learn more about how to use Bower, there are plenty of references online; https://www.openshift.com/blogs/day-1-bower-manage-your-client-side-dependencies is a good example of a tutorial that will help you get accustomed to using Bower. In addition, there is a useful article that discusses both Bower and Node, available at http://tech.pro/tutorial/1190/package-managers-an-introductory-guide-for-the-uninitiated-front-end-developer. Bower isn't the only way to install jQuery though—while we can use it to install multiple versions of jQuery, for example, we're still limited to installing the entire jQuery library. We can improve on this by referencing only the elements we need within the library. Thanks to some extensive work undertaken by the jQuery Core team, we can use the Asynchronous Module Definition (AMD) approach to reference only those modules that are needed within our website or online application. Using the AMD approach to load jQuery In most instances, when using jQuery, developers are likely to simply include a reference to the main library in their code. There is nothing wrong with it per se, but it loads a lot of extra code that is surplus to our requirements. A more efficient method, although one that takes a little effort in getting used to, is to use the AMD approach. In a nutshell, the jQuery team has made the library more modular; this allows you to use a loader such as require.js to load individual modules when needed. It's not suitable for every approach, particularly if you are a heavy user of different parts of the library. However, for those instances where you only need a limited number of modules, then this is a perfect route to take. Let's work through a simple example to see what it looks like in practice. Before we start, we need one additional item—the code uses the Fira Sans regular custom font, which is available from Font Squirrel at http://www.fontsquirrel.com/fonts/fira-sans. Let's make a start using the following steps: The Fira Sans font doesn't come with a web format by default, so we need to convert the font to use the web font format. Go ahead and upload the FiraSans-Regular.otf file to Font Squirrel's web font generator at http://www.fontsquirrel.com/tools/webfont-generator. When prompted, save the converted file to your project folder in a subfolder called fonts. We need to install jQuery and RequireJS into our project folder, so fire up a Node.js Command Prompt and change to the project folder. Next, enter these commands one by one, pressing Enter after each: bower install jquerybower install requirejs We need to extract a copy of the amd.html and amd.css files—it contains some simple markup along with a link to require.js; the amd.css file contains some basic styling that we will use in our demo. We now need to add in this code block, immediately below the link for require.js—this handles the calls to jQuery and RequireJS, where we're calling in both jQuery and Sizzle, the selector engine for jQuery: <script>require.config({paths: {"jquery": "bower_components/jquery/src","sizzle": "bower_components/jquery/src/sizzle/dist/sizzle"}});require(["js/app"]);</script> Now that jQuery has been defined, we need to call in the relevant modules. In a new file, go ahead and add the following code, saving it as app.js in a subfolder marked js within our project folder: define(["jquery/core/init", "jquery/attributes/classes"],function($) {$("div").addClass("decoration");}); We used app.js as the filename to tie in with the require(["js/app"]); reference in the code. If all went well, when previewing the results of our work in a browser. Although we've only worked with a simple example here, it's enough to demonstrate how easy it is to only call those modules we need to use in our code rather than call the entire jQuery library. True, we still have to provide a link to the library, but this is only to tell our code where to find it; our module code weighs in at 29 KB (10 KB when gzipped), against 242 KB for the uncompressed version of the full library! Now, there may be instances where simply referencing modules using this method isn't the right approach; this may apply if you need to reference lots of different modules regularly. A better alternative is to build a custom version of the jQuery library that only contains the modules that we need to use and the rest are removed during build. It's a little more involved but worth the effort—let's take a look at what is involved in the process. Customizing the downloads of jQuery from Git If we feel so inclined, we can really push the boat out and build a custom version of jQuery using the JavaScript task runner, Grunt. The process is relatively straightforward but involves a few steps; it will certainly help if you have some prior familiarity with Git! The demo assumes that you have already installed Node.js—if you haven't, then you will need to do this first before continuing with the exercise. Okay, let's make a start by performing the following steps: You first need to install Grunt if it isn't already present on your system—bring up the Node.js Command Prompt and enter this command: npm install -g grunt-cli Next, install Git—for this, browse to http://msysgit.github.io/ in order to download the package. Double-click on the setup file to launch the wizard, accepting all the defaults is sufficient for our needs. If you want more information on how to install Git, head over and take a look at https://github.com/msysgit/msysgit/wiki/InstallMSysGit for more details. Once Git is installed, change to the jquery folder from within the Command Prompt and enter this command to download and install the dependencies needed to build jQuery: npm install The final stage of the build process is to build the library into the file we all know and love; from the same Command Prompt, enter this command: grunt Browse to the jquery folder—within this will be a folder called dist, which contains our custom build of jQuery, ready for use. If there are modules within the library that we don't need, we can run a custom build. We can set the Grunt task to remove these when building the library, leaving in those that are needed for our project. For a complete list of all the modules that we can exclude, see https://github.com/jquery/jquery#modules. For example, to remove AJAX support from our build, we can run this command in place of step 5, as shown previously: grunt custom:-ajax This results in a file saving on the original raw version of 30 KB as shown in the following screenshot: The JavaScript and map files can now be incorporated into our projects in the usual way. For a detailed tutorial on the build process, this article by Dan Wellman is worth a read (https://www.packtpub.com/books/content/building-custom-version-jquery). Using a GUI as an alternative There is an online GUI available, which performs much the same tasks, without the need to install Git or Grunt. It's available at hhttp://projects.jga.me/jquery-builder/, although it is worth noting that it hasn't been updated for a while! Okay, so we have jQuery installed; let's take a look at one more useful function that will help in the event of debugging errors in our code. Support for source maps has been made available within jQuery since version 1.9. Let's take a look at how they work and see a simple example in action. Adding source map support Imagine a scenario, if you will, where you've created a killer site, which is running well, until you start getting complaints about problems with some of the jQuery-based functionality that is used on the site. Sounds familiar? Using an uncompressed version of jQuery on a production site is not an option; instead we can use source maps. Simply put, these map a compressed version of jQuery against the relevant line in the original source. Historically, source maps have given developers a lot of heartache when implementing, to the extent that the jQuery Team had to revert to disabling the automatic use of maps! For best effects, it is recommended that you use a local web server, such as WAMP (PC) or MAMP (Mac), to view this demo and that you use Chrome as your browser. Source maps are not difficult to implement; let's run through how you can implement them: Extract a copy of the sourcemap folder and save it to your project area locally. Press Ctrl + Shift + I to bring up the Developer Tools in Chrome. Click on Sources, then double-click on the sourcemap.html file—in the code window, and finally click on 17. Now, run the demo in Chrome—we will see it paused; revert back to the developer toolbar where line 17 is highlighted. The relevant calls to the jQuery library are shown on the right-hand side of the screen: If we double-click on the n.event.dispatch entry on the right, Chrome refreshes the toolbar and displays the original source line (highlighted) from the jQuery library, as shown here: It is well worth spending the time to get to know source maps—all the latest browsers support it, including IE11. Even though we've only used a simple example here, it doesn't matter as the principle is exactly the same, no matter how much code is used in the site. For a more in-depth tutorial that covers all the browsers, it is worth heading over to http://blogs.msdn.com/b/davrous/archive/2014/08/22/enhance-your-javascript-debugging-life-thanks-to-the-source-map-support-available-in-ie11-chrome-opera-amp-firefox.aspx—it is worth a read! Adding support for source maps We've just previewed the source map, source map support has already been added to the library. It is worth noting though that source maps are not included with the current versions of jQuery by default. If you need to download a more recent version or add support for the first time, then follow these steps: Source maps can be downloaded from the main site using http://code.jquery.com/jquery-X.X.X.min.map, where X represents the version number of jQuery being used. Open a copy of the minified version of the library and then add this line at the end of the file: //# sourceMappingURL=jquery.min.map Save it and then store it in the JavaScript folder of your project. Make sure you have copies of both the compressed and uncompressed versions of the library within the same folder. Let's move on and look at one more critical part of loading jQuery: if, for some unknown reason, jQuery becomes completely unavailable, then we can add a fallback position to our site that allows graceful degradation. It's a small but crucial part of any site and presents a better user experience than your site simply falling over! Working with Modernizr as a fallback A best practice when working with jQuery is to ensure that a fallback is provided for the library, should the primary version not be available. (Yes, it's irritating when it happens, but it can happen!) Typically, we might use a little JavaScript, such as the following example, in the best practice suggestions. This would work perfectly well but doesn't provide a graceful fallback. Instead, we can use Modernizr to perform the check for us and provide a graceful degradation if all fails. Modernizr is a feature detection library for HTML5/CSS3, which can be used to provide a standardized fallback mechanism in the event of a functionality not being available. You can learn more at http://www.modernizr.com. As an example, the code might look like this at the end of our website page. We first try to load jQuery using the CDN link, falling back to a local copy if that hasn't worked or an alternative if both fail: <body><script src="js/modernizr.js"></script><script type="text/javascript">Modernizr.load([{load: 'http://code.jquery.com/jquery-2.1.1.min.js',complete: function () {// Confirm if jQuery was loaded using CDN link// if not, fall back to local versionif ( !window.jQuery ) {Modernizr.load('js/jquery-latest.min.js');}}},// This script would wait until fallback is loaded, beforeloading{ load: 'jquery-example.js' }]);</script></body> In this way, we can ensure that jQuery either loads locally or from the CDN link—if all else fails, then we can at least make a graceful exit. Best practices for loading jQuery So far, we've examined several ways of loading jQuery into our pages, over and above the usual route of downloading the library locally or using a CDN link in our code. Now that we have it installed, it's a good opportunity to cover some of the best practices we should try to incorporate into our pages when loading jQuery: Always try to use a CDN to include jQuery on your production site. We can take advantage of the high availability and low latency offered by CDN services; the library may already be precached too, avoiding the need to download it again. Try to implement a fallback on your locally hosted library of the same version. If CDN links become unavailable (and they are not 100 percent infallible), then the local version will kick in automatically, until the CDN link becomes available again: <script type="text/javascript" src="//code.jquery.com/jquery-1.11.1.min.js"></script><script>window.jQuery || document.write('<scriptsrc="js/jquery-1.11.1.min.js"></script>')</script> Note that although this will work equally well as using Modernizr, it doesn't provide a graceful fallback if both the versions of jQuery should become unavailable. Although one hopes to never be in this position, at least we can use CSS to provide a graceful exit! Use protocol-relative/protocol-independent URLs; the browser will automatically determine which protocol to use. If HTTPS is not available, then it will fall back to HTTP. If you look carefully at the code in the previous point, it shows a perfect example of a protocol-independent URL, with the call to jQuery from the main jQuery Core site. If possible, keep all your JavaScript and jQuery inclusions at the bottom of your page—scripts block the rendering of the rest of the page until they have been fully rendered. Use the jQuery 2.x branch, unless you need to support IE6-8; in this case, use jQuery 1.x instead—do not load multiple jQuery versions. If you load jQuery using a CDN link, always specify the complete version number you want to load, such as jquery-1.11.1.min.js. If you are using other libraries, such as Prototype, MooTools, Zepto, and so on, that use the $ sign as well, try not to use $ to call jQuery functions and simply use jQuery instead. You can return the control of $ back to the other library with a call to the $.noConflict() function. For advanced browser feature detection, use Modernizr. It is worth noting that there may be instances where it isn't always possible to follow best practices; circumstances may dictate that we need to make allowances for requirements, where best practices can't be used. However, this should be kept to a minimum where possible; one might argue that there are flaws in our design if most of the code doesn't follow best practices! Summary If you thought that the only methods to include jQuery were via a manual download or using a CDN link, then hopefully this article has opened your eyes to some alternatives—let's take a moment to recap what we have learned. We kicked off with a customary look at how most developers are likely to include jQuery before quickly moving on to look at other sources. We started with a look at how to use Node, before turning our attention to using the Bower package manager. Next, we had a look at how we can reference individual modules within jQuery using the AMD approach. We then moved on and turned our attention to creating custom builds of the library using Git. We then covered how we can use source maps to debug our code, with a look at enabling support for them within Google's Chrome browser. To round out our journey of loading jQuery, we saw what might happen if we can't load jQuery at all and how we can get around this, by using Modernizr to allow our pages to degrade gracefully. We then finished the article with some of the best practices that we can follow when referencing jQuery. Resources for Article: Further resources on this subject: Using different jQuery event listeners for responsive interaction [Article] Building a Custom Version of jQuery [Article] Learning jQuery [Article]
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Packt
04 Jun 2015
19 min read
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Getting Started with Hyper-V Architecture and Components

Packt
04 Jun 2015
19 min read
In this article by Vinícius R. Apolinário, author of the book Learning Hyper-V, we will cover the following topics: Hypervisor architecture Type 1 and 2 Hypervisors Microkernel and Monolithic Type 1 Hypervisors Hyper-V requirements and processor features Memory configuration Non-Uniform Memory Access (NUMA) architecture (For more resources related to this topic, see here.) Hypervisor architecture If you've used Microsoft Virtual Server or Virtual PC, and then moved to Hyper-V, I'm almost sure that your first impression was: "Wow, this is much faster than Virtual Server". You are right. And there is a reason why Hyper-V performance is much better than Virtual Server or Virtual PC. It's all about the architecture. There are two types of Hypervisor architectures. Hypervisor Type 1, like Hyper-V and ESXi from VMware, and Hypervisor Type 2, like Virtual Server, Virtual PC, VMware Workstation, and others. The objective of the Hypervisor is to execute, manage and control the operation of the VM on a given hardware. For that reason, the Hypervisor is also called Virtual Machine Monitor (VMM). The main difference between these Hypervisor types is the way they operate on the host machine and its operating systems. As Hyper-V is a Type 1 Hypervisor, we will cover Type 2 first, so we can detail Type 1 and its benefits later. Type 1 and Type 2 Hypervisors Hypervisor Type 2, also known as hosted, is an implementation of the Hypervisor over and above the OS installed on the host machine. With that, the OS will impose some limitations to the Hypervisor to operate, and these limitations are going to reflect on the performance of the VM. To understand that, let me explain how a process is placed on the processor: the processor has what we call Rings on which the processes are placed, based on prioritization. The main Rings are 0 and 3. Kernel processes are placed on Ring 0 as they are vital to the OS. Application processes are placed on Ring 3, and, as a result, they will have less priority when compared to Ring 0. The issue on Hypervisors Type 2 is that it will be considered an application, and will run on Ring 3. Let's have a look at it: As you can see from the preceding diagram, the hypervisor has an additional layer to access the hardware. Now, let's compare it with Hypervisor Type 1: The impact is immediate. As you can see, Hypervisor Type 1 has total control of the underlying hardware. In fact, when you enable Virtualization Assistance (hardware-assisted virtualization) at the server BIOS, you are enabling what we call Ring -1, or Ring decompression, on the processor and the Hypervisor will run on this Ring. The question you might have is "And what about the host OS?" If you install the Hyper-V role on a Windows Server for the first time, you may note that after installation, the server will restart. But, if you're really paying attention, you will note that the server will actually reboot twice. This behavior is expected, and the reason it will happen is because the OS is not only installing and enabling Hyper-V bits, but also changing its architecture to the Type 1 Hypervisor. In this mode, the host OS will operate in the same way a VM does, on top of the Hypervisor, but on what we call parent partition. The parent partition will play a key role as the boot partition and in supporting the child partitions, or guest OS, where the VMs are running. The main reason for this partition model is the key attribute of a Hypervisor: isolation. For Microsoft Hyper-V Server you don't have to install the Hyper-V role, as it will be installed when you install the OS, so you won't be able to see the server booting twice. With isolation, you can ensure that a given VM will never have access to another VM. That means that if you have a compromised VM, with isolation, the VM will never infect another VM or the host OS. The only way a VM can access another VM is through the network, like all other devices in your network. Actually, the same is true for the host OS. This is one of the reasons why you need an antivirus for the host and the VMs, but this will be discussed later. The major difference between Type 1 and Type 2 now is that kernel processes from both host OS and VM OS will run on Ring 0. Application processes from both host OS and VM OS will run on Ring 3. However, there is one piece left. The question now is "What about device drivers?" Microkernel and Monolithic Type 1 Hypervisors Have you tried to install Hyper-V on a laptop? What about an all-in-one device? A PC? A server? An x64 based tablet? They all worked, right? And they're supposed to work. As Hyper-V is a Microkernel Type 1 Hypervisor, all the device drivers are hosted on the parent partition. A Monolithic Type 1 Hypervisor hosts its drivers on the Hypervisor itself. VMware ESXi works this way. That's why you should never use a standard ESXi media to install an ESXi host. The hardware manufacturer will provide you with an appropriate media with the correct drivers for the specific hardware. The main advantage of the Monolithic Type 1 Hypervisor is that, as it always has the correct driver installed, you will never have a performance issue due to an incorrect driver. On the other hand, you won't be able to install this on any device. The Microkernel Type 1 Hypervisor, on the other hand, hosts its drivers on the parent partition. That means that if you installed the host OS on a device, and the drivers are working, the Hypervisor, and in this case Hyper-V, will work just fine. There are other hardware requirements. These will be discussed later in this article. The other side of this is that if you use a generic driver, or a wrong version of it, you may have performance issues, or even driver malfunction. What you have to keep in mind here is that Microsoft does not certify drivers for Hyper-V. Device drivers are always certified for Windows Server. If the driver is certified for Windows Server, it is also certified for Hyper-V. But you always have to ensure the use of correct driver for a given hardware. Let's take a better look at how Hyper-V works as a Microkernel Type 1 Hypervisor: As you can see from the preceding diagram, there are multiple components to ensure that the VM will run perfectly. However, the major component is the Integration Components (IC), also called Integration Services. The IC is a set of tools that you should install or upgrade on the VM, so that the VM OS will be able to detect the virtualization stack and run as a regular OS on a given hardware. To understand this more clearly, let's see how an application accesses the hardware and understand all the processes behind it. When the application tries to send a request to the hardware, the kernel is responsible for interpreting this call. As this OS is running on an Enlightened Child Partition (Means that IC is installed), the Kernel will send this call to the Virtual Service Client (VSC) that operates as a synthetic device driver. The VSC is responsible for communicating with the Virtual Service Provider (VSP) on the parent partition, through VMBus, so the VSC can use the hardware resource. The VMBus will then be able to communicate with the hardware for the VM. The VMBus, a channel-based communication, is actually responsible for communicating with the parent partition and hardware. For the VMBus to access the hardware, it will communicate directly with a component on the Hypervisor called hypercalls. These hypercalls are then redirected to the hardware. However, only the parent partition can actually access the physical processor and memory. The child partitions access a virtual view of these components that are translated on the guest and the host partitions. New processors have a feature called Second Level Address Translation (SLAT) or Nested Paging. This feature is extremely important on high performance VMs and hosts, as it helps reduce the overhead of the virtual to physical memory and processor translation. On Windows 8, SLAT is a requirement for Hyper-V. It is important to note that Enlightened Child Partitions, or partitions with IC, can be Windows or Linux OS. If the child partitions have a Linux OS, the name of the component is Linux Integration Services (LIS), but the operation is actually the same. Another important fact regarding ICs is that they are already present on Windows Server 2008 or later. But, if you are running a newer version of Hyper-V, you have to upgrade the IC version on the VM OS. For example, if you are running Hyper-V 2012 R2 on the host OS and the guest OS is running Windows Server 2012 R2, you probably don't have to worry about it. But if you are running Hyper-V 2012 R2 on the host OS and the guest OS is running Windows Server 2012, then you have to upgrade the IC on the VM to match the parent partition version. Running guest OS Windows Server 2012 R2 on a VM on top of Hyper-V 2012 is not recommended. For Linux guest OS, the process is the same. Linux kernel version 3 or later already have LIS installed. If you are running an old version of Linux, you should verify the correct LIS version of your OS. To confirm the Linux and LIS versions, you can refer to an article at http://technet.microsoft.com/library/dn531030.aspx. Another situation is when the guest OS does not support IC or LIS, or an Unenlightened Child Partition. In this case, the guest OS and its kernel will not be able to run as an Enlightened Child Partition. As the VMBus is not present in this case, the utilization of hardware will be made by emulation and performance will be degraded. This only happens with old versions of Windows and Linux, like Windows 2000 Server, Windows NT, and CentOS 5.8 or earlier, or in case that the guest OS does not have or support IC. Now that you understand how the Hyper-V architecture works, you may be thinking "Okay, so for all of this to work, what are the requirements?" Hyper-V requirements and processor features At this point, you can see that there is a lot of effort for putting all of this to work. In fact, this architecture is only possible because hardware and software companies worked together in the past. The main goal of both type of companies was to enable virtualization of operating systems without changing them. Intel and AMD created, each one with its own implementation, a processor feature called virtualization assistance so that the Hypervisor could run on Ring 0, as explained before. But this is just the first requirement. There are other requirement as well, which are as follows: Virtualization assistance (also known as Hardware-assisted virtualization): This feature was created to remove the necessity of changing the OS for virtualizing it. On Intel processors, it is known as Intel VT-x. All recent processor families support this feature, including Core i3, Core i5, and Core i7. The complete list of processors and features can be found at http://ark.intel.com/Products/VirtualizationTechnology. You can also use this tool to check if your processor meets this requirement which can be downloaded at: https://downloadcenter.intel.com/Detail_Desc.aspx?ProductID=1881&DwnldID=7838. On AMD Processors, this technology is known as AMD-V. Like Intel, all recent processor families support this feature. AMD provides a tool to check processor compatibility that can be downloaded at http://www.amd.com/en-us/innovations/software-technologies/server-solution/virtualization. Data Execution Prevention (DEP): This is a security feature that marks memory pages as either executable or nonexecutable. For Hyper-V to run, this option must be enabled on the System BIOS. For an Intel-based processor, this feature is called Execute Disable bit (Intel XD bit) and No Execute Bit (AMD NX bit). This configuration will vary from one System BIOS to another. Check with your hardware vendor how to enable it on System BIOS. x64 (64-bit) based processor: This processor feature uses a 64-bit memory address. Although you may find that all new processors are x64, you might want to check if this is true before starting your implementation. The compatibility checkers above, from Intel and AMD, will show you if your processor is x64. Second Level Address Translation (SLAT): As discussed before, SLAT is not a requirement for Hyper-V to work. This feature provides much more performance on the VMs as it removes the need for translating physical and virtual pages of memory. It is highly recommended to have the SLAT feature on the processor ait provides more performance on high performance systems. As also discussed before, SLAT is a requirement if you want to use Hyper-V on Windows 8 or 8.1. To check if your processor has the SLAT feature, use the Sysinternals tool—Coreinfo— that can be downloaded at http://technet.microsoft.com/en-us/sysinternals/cc835722.aspx. There are some specific processor features that are not used exclusively for virtualization. But when the VM is initiated, it will use these specific features from the processor. If the VM is initiated and these features are allocated on the guest OS, you can't simply remove them. This is a problem if you are going to Live Migrate this VM from a host to another host; if these specific features are not available, you won't be able to perform the operation. At this moment, you have to understand that Live Migration moves a powered-on VM from one host to another. If you try to Live Migrate a VM between hosts with different processor types, you may be presented with an error. Live Migration is only permitted between the same processor vendor: Intel-Intel or AMD-AMD. Intel-AMD Live Migration is not allowed under any circumstance. If the processor is the same on both hosts, Live Migration and Share Nothing Live Migration will work without problems. But even within the same vendor, there can be different processor families. In this case, you can remove these specific features from the Virtual Processor presented to the VM. To do that, open Hyper-V Manager | Settings... | Processor | Processor Compatibility. Mark the Migrate to a physical computer with a different processor version option. This option is only available if the VM is powered off. Keep in mind that enabling this option will remove processor-specific features for the VM. If you are going to run an application that requires these features, they will not be available and the application may not run. Now that you have checked all the requirements, you can start planning your server for virtualization with Hyper-V. This is true from the perspective that you understand how Hyper-V works and what are the requirements for it to work. But there is another important subject that you should pay attention to when planning your server: memory. Memory configuration I believe you have heard this one before "The application server is running under performance". In the virtualization world, there is an obvious answer to it: give more virtual hardware to the VM. Although it seems to be the logical solution, the real effect can be totally opposite. During the early days, when servers had just a few sockets, processors, and cores, a single channel made the communication between logical processors and memory. But server hardware has evolved, and today, we have servers with 256 logical processors and 4 TB of RAM. To provide better communication between these components, a new concept emerged. Modern servers with multiple logical processors and high amount of memory use a new design called Non-Uniform Memory Access (NUMA) architecture. Non-Uniform Memory Access (NUMA) architecture NUMA is a memory design that consists of allocating memory to a given node, or a cluster of memory and logical processors. Accessing memory from a processor inside the node is notably faster than accessing memory from another node. If a processor has to access memory from another node, the performance of the process performing the operation will be affected. Basically, to solve this equation you have to ensure that the process inside the guest VM is aware of the NUMA node and is able to use the best available option: When you create a virtual machine, you decide how many virtual processors and how much virtual RAM this VM will have. Usually, you assign the amount of RAM that the application will need to run and meet the expected performance. For example, you may ask a software vendor on the application requirements and this software vendor will say that the application would be using at least 8 GB of RAM. Suppose you have a server with 16 GB of RAM. What you don't know is that this server has four NUMA nodes. To be able to know how much memory each NUMA node has, you must divide the total amount of RAM installed on the server by the number of NUMA nodes on the system. The result will be the amount of RAM of each NUMA node. In this case, each NUMA node has a total of 4 GB of RAM. Following the instructions of the software vendor, you create a VM with 8 GB of RAM. The Hyper-V standard configuration is to allow NUMA spanning, so you will be able to create the VM and start it. Hyper-V will accommodate 4 GB of RAM on two NUMA nodes. This NUMA spanning configuration means that a processor can access the memory on another NUMA node. As mentioned earlier, this will have an impact on the performance if the application is not aware of it. On Hyper-V, prior to the 2012 version, the guest OS was not informed about the NUMA configuration. Basically, in this case, the guest OS would see one NUMA node with 8 GB of RAM, and the allocation of memory would be made without NUMA restrictions, impacting the final performance of the application. Hyper-V 2012 and 2012 R2 have the same feature—the guest OS will see the virtual NUMA (vNUMA) presented to the child partition. With this feature, the guest OS and/or the application can make a better choice on where to allocate memory for each process running on this VM. NUMA is not a virtualization technology. In fact, it has been used for a long time, and even applications like SQL Server 2005 already used NUMA to better allocate the memory that its processes are using. Prior to Hyper-V 2012, if you wanted to avoid this behavior, you had two choices: Create the VM and allocate the maximum vRAM of a single NUMA node for it, as Hyper-V will always try to allocate the memory inside of a single NUMA node. In the above case, the VM should not have more than 4 GB of vRAM. But for this configuration to really work, you should also follow the next choice. Disable NUMA Spanning on Hyper-V. With this configuration disabled, you will not be able to run a VM if the memory configuration exceeds a single NUMA node. To do this, you should clear the Allow virtual machines to span physical NUMA nodes checkbox on Hyper-V Manager | Hyper-V Settings... | NUMA Spanning. Keep in mind that disabling this option will prevent you from running a VM if no nodes are available. You should also remember that even with Hyper-V 2012, if you create a VM with 8 GB of RAM using two NUMA nodes, the application on top of the guest OS (and the guest OS) must understand the NUMA topology. If the application and/or guest OS are not NUMA aware, vNUMA will not have effect and the application can still have performance issues. At this point you are probably asking yourself "How do I know how many NUMA nodes I have on my server?" This was harder to find in the previous versions of Windows Server and Hyper-V Server. In versions prior to 2012, you should open the Performance Monitor and check the available counters in Hyper-V VM Vid NUMA Node. The number of instances represents the number of NUMA Nodes. In Hyper-V 2012, you can check the settings for any VM. Under the Processor tab, there is a new feature available for NUMA. Let's have a look at this screen to understand what it represents: In Configuration, you can easily confirm how many NUMA nodes the host running this VM has. In the case above, the server has only 1 NUMA node. This means that all memory will be allocated close to the processor. Multiple NUMA nodes are usually present on servers with high amount of logical processors and memory. In the NUMA topology section, you can ensure that this VM will always run with the informed configuration. This is presented to you because of a new Hyper-V 2012 feature called Share Nothing Live Migration, which will be explained in detail later. This feature allows you to move a VM from one host to another without turning the VM off, with no cluster and no shared storage. As you can move the VM turned on, you might want to force the processor and memory configuration, based on the hardware of your worst server, ensuring that your VM will always meet your performance expectations. The Use Hardware Topology button will apply the hardware topology in case you moved the VM to another host or in case you changed the configuration and you want to apply the default configuration again. To summarize, if you want to make sure that your VM will not have performance problems, you should check how many NUMA nodes your server has and divide the total amount of memory by it; the result is the total memory on each node. Creating a VM with more memory than a single node will make Hyper-V present a vNUMA to the guest OS. Ensuring that the guest OS and applications are NUMA aware is also important, so that the guest OS and application can use this information to allocate memory for a process on the correct node. NUMA is important to ensure that you will not have problems because of host configuration and misconfiguration on the VM. But, in some cases, even when planning the VM size, you will come to a moment when the VM memory is stressed. In these cases, Hyper-V can help with another feature called Dynamic Memory. Summary In this we learned about the Hypervisor architecture and different Hypervisor types. We explored in brief about Microkernel and Monolithic Type 1 Hypervisors. In addition to this, this article also explains the Hyper-V requirements and processor features, Memory configuration and the NUMA architecture. Resources for Article: Further resources on this subject: Planning a Compliance Program in Microsoft System Center 2012 [Article] So, what is Microsoft © Hyper-V server 2008 R2? [Article] Deploying Applications and Software Updates on Microsoft System Center 2012 Configuration Manager [Article]
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Packt
04 Jun 2015
11 min read
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Preparing Optimizations

Packt
04 Jun 2015
11 min read
In this article by Mayur Pandey and Suyog Sarda, authors of LLVM Cookbook, we will look into the following recipes: Various levels of optimization Writing your own LLVM pass Running your own pass with the opt tool Using another pass in a new pass (For more resources related to this topic, see here.) Once the source code transformation completes, the output is in the LLVM IR form. This IR serves as a common platform for converting into assembly code, depending on the backend. However, before converting into an assembly code, the IR can be optimized to produce more effective code. The IR is in the SSA form, where every new assignment to a variable is a new variable itself—a classic case of an SSA representation. In the LLVM infrastructure, a pass serves the purpose of optimizing LLVM IR. A pass runs over the LLVM IR, processes the IR, analyzes it, identifies the optimization opportunities, and modifies the IR to produce optimized code. The command-line interface opt is used to run optimization passes on LLVM IR. Various levels of optimization There are various levels of optimization, starting at 0 and going up to 3 (there is also s for space optimization). The code gets more and more optimized as the optimization level increases. Let's try to explore the various optimization levels. Getting ready... Various optimization levels can be understood by running the opt command-line interface on LLVM IR. For this, an example C program can first be converted to IR using the Clang frontend. Open an example.c file and write the following code in it: $ vi example.c int main(int argc, char **argv) { int i, j, k, t = 0; for(i = 0; i < 10; i++) {    for(j = 0; j < 10; j++) {      for(k = 0; k < 10; k++) {        t++;      }    }    for(j = 0; j < 10; j++) {      t++;    } } for(i = 0; i < 20; i++) {    for(j = 0; j < 20; j++) {      t++;    }    for(j = 0; j < 20; j++) {      t++;    } } return t; } Now convert this into LLVM IR using the clang command, as shown here: $ clang –S –O0 –emit-llvm example.c A new file, example.ll, will be generated, containing LLVM IR. This file will be used to demonstrate the various optimization levels available. How to do it… Do the following steps: The opt command-line tool can be run on the IR-generated example.ll file: $ opt –O0 –S example.ll The –O0 syntax specifies the least optimization level. Similarly, you can run other optimization levels: $ opt –O1 –S example.ll $ opt –O2 –S example.ll $ opt –O3 –S example.ll How it works… The opt command-line interface takes the example.ll file as the input and runs the series of passes specified in each optimization level. It can repeat some passes in the same optimization level. To see which passes are being used in each optimization level, you have to add the --debug-pass=Structure command-line option with the previous opt commands. See Also To know more on various other options that can be used with the opt tool, refer to http://llvm.org/docs/CommandGuide/opt.html Writing your own LLVM pass All LLVM passes are subclasses of the pass class, and they implement functionality by overriding the virtual methods inherited from pass. LLVM applies a chain of analyses and transformations on the target program. A pass is an instance of the Pass LLVM class. Getting ready Let's see how to write a pass. Let's name the pass function block counter; once done, it will simply display the name of the function and count the basic blocks in that function when run. First, a Makefile needs to be written for the pass. Follow the given steps to write a Makefile: Open a Makefile in the llvm lib/Transform folder: $ vi Makefile Specify the path to the LLVM root folder and the library name, and make this pass a loadable module by specifying it in Makefile, as follows: LEVEL = ../../.. LIBRARYNAME = FuncBlockCount LOADABLE_MODULE = 1 include $(LEVEL)/Makefile.common This Makefile specifies that all the .cpp files in the current directory are to be compiled and linked together in a shared object. How to do it… Do the following steps: Create a new .cpp file called FuncBlockCount.cpp: $ vi FuncBlockCount.cpp In this file, include some header files from LLVM: #include "llvm/Pass.h" #include "llvm/IR/Function.h" #include "llvm/Support/raw_ostream.h" Include the llvm namespace to enable access to LLVM functions: using namespace llvm; Then start with an anonymous namespace: namespace { Next declare the pass: struct FuncBlockCount : public FunctionPass { Then declare the pass identifier, which will be used by LLVM to identify the pass: static char ID; FuncBlockCount() : FunctionPass(ID) {} This step is one of the most important steps in writing a pass—writing a run function. Since this pass inherits FunctionPass and runs on a function, a runOnFunction is defined to be run on a function: bool runOnFunction(Function &F) override {      errs() << "Function " << F.getName() << 'n';      return false;    } }; } This function prints the name of the function that is being processed. The next step is to initialize the pass ID: char FuncBlockCount::ID = 0; Finally, the pass needs to be registered, with a command-line argument and a name: static RegisterPass<FuncBlockCount> X("funcblockcount", "Function Block Count", false, false); Putting everything together, the entire code looks like this: #include "llvm/Pass.h" #include "llvm/IR/Function.h" #include "llvm/Support/raw_ostream.h" using namespace llvm; namespace { struct FuncBlockCount : public FunctionPass { static char ID; FuncBlockCount() : FunctionPass(ID) {} bool runOnFunction(Function &F) override {    errs() << "Function " << F.getName() << 'n';    return false; }            };        }        char FuncBlockCount::ID = 0;        static RegisterPass<FuncBlockCount> X("funcblockcount", "Function Block Count", false, false); How it works A simple gmake command compiles the file, so a new file FuncBlockCount.so is generated at the LLVM root directory. This shared object file can be dynamically loaded to the opt tool to run it on a piece of LLVM IR code. How to load and run it will be demonstrated in the next section. See also To know more on how a pass can be built from scratch, visit http://llvm.org/docs/WritingAnLLVMPass.html Running your own pass with the opt tool The pass written in the previous recipe, Writing your own LLVM pass, is ready to be run on the LLVM IR. This pass needs to be loaded dynamically for the opt tool to recognize and execute it. How to do it… Do the following steps: Write the C test code in the sample.c file, which we will convert into an .ll file in the next step: $ vi sample.c   int foo(int n, int m) { int sum = 0; int c0; for (c0 = n; c0 > 0; c0--) {    int c1 = m;  for (; c1 > 0; c1--) {      sum += c0 > c1 ? 1 : 0;    } } return sum; } Convert the C test code into LLVM IR using the following command: $ clang –O0 –S –emit-llvm sample.c –o sample.ll This will generate a sample.ll file. Run the new pass with the opt tool, as follows: $ opt -load (path_to_.so_file)/FuncBlockCount.so -funcblockcount sample.ll The output will look something like this: Function foo How it works… As seen in the preceding code, the shared object loads dynamically into the opt command-line tool and runs the pass. It goes over the function and displays its name. It does not modify the IR. Further enhancement in the new pass is demonstrated in the next recipe. See also To know more about the various types of the Pass class, visit http://llvm.org/docs/WritingAnLLVMPass.html#pass-classes-and-requirements Using another pass in a new pass A pass may require another pass to get some analysis data, heuristics, or any such information to decide on a further course of action. The pass may just require some analysis such as memory dependencies, or it may require the altered IR as well. The new pass that you just saw simply prints the name of the function. Let's see how to enhance it to count the basic blocks in a loop, which also demonstrates how to use other pass results. Getting ready The code used in the previous recipe remains the same. Some modifications are required, however, to enhance it—as demonstrated in next section—so that it counts the number of basic blocks in the IR. How to do it… The getAnalysis function is used to specify which other pass will be used: Since the new pass will be counting the number of basic blocks, it requires loop information. This is specified using the getAnalysis loop function: LoopInfo *LI = &getAnalysis<LoopInfoWrapperPass>().getLoopInfo(); This will call the LoopInfo pass to get information on the loop. Iterating through this object gives the basic block information: unsigned num_Blocks = 0; Loop::block_iterator bb; for(bb = L->block_begin(); bb != L->block_end();++bb)    num_Blocks++; errs() << "Loop level " << nest << " has " << num_Blocks << " blocksn"; This will go over the loop to count the basic blocks inside it. However, it counts only the basic blocks in the outermost loop. To get information on the innermost loop, recursive calling of the getSubLoops function will help. Putting the logic in a separate function and calling it recursively makes more sense: void countBlocksInLoop(Loop *L, unsigned nest) { unsigned num_Blocks = 0; Loop::block_iterator bb; for(bb = L->block_begin(); bb != L->block_end();++bb)    num_Blocks++; errs() << "Loop level " << nest << " has " << num_Blocks << " blocksn"; std::vector<Loop*> subLoops = L->getSubLoops(); Loop::iterator j, f; for (j = subLoops.begin(), f = subLoops.end(); j != f; ++j)    countBlocksInLoop(*j, nest + 1); } virtual bool runOnFunction(Function &F) { LoopInfo *LI = &getAnalysis<LoopInfoWrapperPass>().getLoopInfo(); errs() << "Function " << F.getName() + "n"; for (Loop *L : *LI)    countBlocksInLoop(L, 0); return false; } How it works… The newly modified pass now needs to run on a sample program. Follow the given steps to modify and run the sample program: Open the sample.c file and replace its content with the following program: int main(int argc, char **argv) { int i, j, k, t = 0; for(i = 0; i < 10; i++) {    for(j = 0; j < 10; j++) {      for(k = 0; k < 10; k++) {        t++;      }    }    for(j = 0; j < 10; j++) {      t++;    } } for(i = 0; i < 20; i++) {    for(j = 0; j < 20; j++) {      t++;    }    for(j = 0; j < 20; j++) {      t++;    } } return t; } Convert it into a .ll file using Clang: $ clang –O0 –S –emit-llvm sample.c –o sample.ll Run the new pass on the previous sample program: $ opt -load (path_to_.so_file)/FuncBlockCount.so - funcblockcount sample.ll The output will look something like this: Function main Loop level 0 has 11 blocks Loop level 1 has 3 blocks Loop level 1 has 3 blocks Loop level 0 has 15 blocks Loop level 1 has 7 blocks Loop level 2 has 3 blocks Loop level 1 has 3 blocks There's more… The LLVM's pass manager provides a debug pass option that gives us the chance to see which passes interact with our analyses and optimizations, as follows: $ opt -load (path_to_.so_file)/FuncBlockCount.so - funcblockcount sample.ll –disable-output –debug-pass=Structure Summary In this article you have explored various optimization levels, and the optimization techniques kicking at each level. We also saw the step-by-step approach to writing our own LLVM pass. Resources for Article: Further resources on this subject: Integrating a D3.js visualization into a simple AngularJS application [article] Getting Up and Running with Cassandra [article] Cassandra Architecture [article]
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Packt
04 Jun 2015
8 min read
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Deploying New Hosts with vCenter

Packt
04 Jun 2015
8 min read
In this article by Konstantin Kuminsky author of the book, VMware vCenter Cookbook, we will review some options and features available in vCenter to improve an administrator's efficiency. (For more resources related to this topic, see here.) Deploying new hosts faster with scripted installation Scripted installation is an alternative way to deploy ESXi hosts. It can be used when several hosts need to be deployed or upgraded. The installation script contains ESXi settings and can be accessed by a host during the ESXi boot from the following locations: FTP HTTP or HTTPS NFS USB flash drive or CD-ROM How to do it... The following sections describe the process of creating an installation script and using it to boot the ESXi host. Creating an installation script An installation script contains installation options for ESXi. It's a text file with the .cfg extension. The best way to create an installation script is to use the default script supplied with the ESXi installer and modify it. The default script is located in the /etc/vmware/weasel/ folder location and is called ks.cfg. Commands that can be modified include, but are not limited to: The install, installorupgrade, or upgrade commands define the ESXi disk—location, where the installation or upgrade will be installed. The available options are: --disk: This option is the disk name which can be specified as path (/vmfs/devices/disks/vmhbaX:X:X), VML name (vml.xxxxxxxx) or as LUN UID (vmkLUM_UID) –overwritevmfs: This option wipes the existing datastore. --preservevmfs: This option keeps the existing datastore. --novmfsondisk: This option prevents a new partition from being created. The Network command, which specifies the network settings. Most of the available options are self-explanatory: --bootproto=[dhcp|static] --device: MAC address of NIC to use --ip --gateway --nameserver --netmask --hostname --vlanid A full list of installation and upgrade commands can be found in the vSphere5 documentation on the VMware website at https://www.vmware.com/support/pubs/. Use the installation script to configure ESXi In order to use the installation script, you will need to use additional ESXi boot options. Boot a host from the ESXi installation disk. When the ESXi installer screen appears, press Shift + O to provide additional boot options. In the command prompt, type the following: ks=<location of the script> <additional boot options> The valid locations are as follows: ks=cdrom:/path ks=file://path ks=protocol://path ks=usb:/path The additional options available are as follows: gateway: This option is the default gateway ip: This option is the IP address nameserver: This option is the DNS server netmask: This option is the subnet mask vlanid: This option is the VLAN ID netdevice: This option is the MAC address of NIC to use bootif: This option is the MAC address of NIC to use in PXELINUX format For example, for the HTTP location, the command will look like this: ks=http://XX.XX.XX.XX/scripts/ks-v1.cfg nameserver=XX.XX.XX.XX ip=XX.XX.XX.XX netmask=255.255.255.0 gateway=XX.XX.XX.XX Deploying new hosts faster with auto deploy vSphere Auto Deploy is VMware's solution to simplify the deployment of large numbers of ESXi hosts. It is one of the available options for ESXi deployment along with an interactive and scripted installation. The main difference of Auto Deploy compared to other deployment options is that the ESXi configuration is not stored on the host's disk. Instead, it's managed with image and host profiles by the Auto Deploy server. Getting ready Before using Auto Deploy, confirm the following: The Auto Deploy server is installed and registered with vCenter. It can be installed as a standalone server or as part of the vCenter installation. The DHCP server exists in the environment. The DHCP server is configured to point to the TFTP server for PXE boot (option 66) with the boot filename undionly.kpxe.vmw-hardwired. The TFTP server that will be used for PXE boot exists and is configured properly. The machine where Auto Deploy cmdlets will run has the following installed: Microsoft .NET 2.0 or later PowerShell 2.0 or later PowerCLI including Auto Deploy cmdlets New hosts that will be provisioned with Auto Deploy must: Meet the hardware requirements for ESXi 5 Have network connectivity to vCenter, preferably 1 Gbps or higher Have PXE boot enabled How to do it... Once prerequisites are met, the following steps are required to start deploying hosts. Configuring the TFTP server In order to configure the TFTP server with the correct boot image for ESXi, execute the following steps: In vCenter, go to Home | Auto Deploy. Switch to the Administration tab. From the Auto Deploy page, click on Download TFTP Boot ZIP. Download the file and unzip it to the appropriate folder on the TFTP server. Creating an image profile Image profies are created using Image Builder PowerCLI cmdlets. Image Builder requires PowerCLI and can be installed on a machine that's used to run administrative tasks. It doesn't have to be a vCenter server or Auto Deploy server and the only requirement for this machine is that it must have access to the software depot—a file server that stores image profiles. Image profiles can be created from scratch or by cloning an existing profile. The following steps outline the process of creating an image profile by cloning. The steps assume that: The Image Builder has been installed. The appropriate software depot has been downloaded from the VMware website by going to http://www.vmware.com/downloads and searching for the software depot. Cloning an existing profile included in the depot is the easiest way to create a new profile. The steps to do so are as follows: Add a depot with the image profile to be cloned: Add-EsxSoftwareDepot -DepotUrl <Path to softwaredepot> Find the name of the profile to be cloned using Get-ESXImageProfile. Clone the profile: New-EsxImageProfile -CloneProfile <Existing profile name> - Name <New profile name> Add a software package to the new image profile: Add-EsxSoftwarePackage -ImageProfile <New profile name> - SoftwarePackage <Package> At this point, the software package will be validated and in case of errors, or if there are any dependencies that need to be resolved, an appropriate message will be displayed. Assigning an image profile to hosts To create a rule that assigns an image profile to a host, execute the following steps: Connect to vCenter with PowerCLI: Connect-VIServer <vCenter IP or FQDN> Add the software depot with the correct image profile to the PowerCLI session: Add-EsxSoftwareDepot <depot URL> Locate the image profile using the Get-EsxImageProfile cmdlet. Define a rule that assigns hosts with certain attributes to an image profile. For example, for hosts with IP addresses for a range, run the following command: New-DeployRule -Name <Rule name> -Item <Profile name> -Pattern "ipv4=192.168.1.10-192.168.1.20" Add-DeployRule <Rule name> Assigning a host profile to hosts Optionally, the existing host profile can be assigned to hosts. To accomplish this, execute the following steps: Connect to vCenter with PowerCLI: Connect-VIServer <vCenter IP or FQDN> Locate the host profile name using the Get-VMhostProfile command. Define a rule that assigns hosts with certain attributes to a host profile. For example, for hosts with IP addresses for a range, run the following command: New-DeployRule -Name <Rule name> -Item <Profile name> -Pattern "ipv4=192.168.1.10-192.168.1.20" Add-DeployRule <Rule name> Assigning a host to a folder or cluster in vCenter To make sure a host is placed in a certain folder or cluster once it boots, do the following: Connect to vCenter with PowerCLI: Connect-VIServer <vCenter IP or FQDN> Define a rule that assigns hosts with certain attributes to a folder or cluster. For example, for hosts with IP addresses for a range, run the following command: New-DeployRule -Name <Rule name> -Item <Folder name> -Pattern "ipv4=192.168.1.10-192.168.1.20" Add-DeployRule <Rule name> If a host is assigned to a cluster it inherits that cluster's host profile. How it works... Auto Deploy utilizes the PXE boot to connect to the Auto Deploy server and get an image profile, vCenter location, and optionally, host profiles. The detailed process is as follows: The host gets gPXE executable and gPXE configuration files from the PXE TFTP server. As gPXE executes, it uses instructions from the configuration file to query the Auto Deploy server for specific information. The Auto Deploy server returns the requested information specified in the image and host profiles. The host boots using this information. Auto Deploy adds a host to the specified vCenter server. The host is placed in maintenance mode when additional information such as IP address is required from the administrator. To exit maintenance mode, the administrator will need to provide this information and reapply the host profile. When a new host boots for the first time, vCenter creates a new object and stores it together with the host and image profiles in the database. For any subsequent reboots, the existing object is used to get the correct host profile and any changes that have been made. More details can be found in the vSphere 5 documentation on the VMware website at https://www.vmware.com/support/pubs/. Summary In this article we learnt how new hosts can be deployed with scripted installation and auto deploy techniques. Resources for Article: Further resources on this subject: VMware vRealize Operations Performance and Capacity Management [Article] Backups in the VMware View Infrastructure [Article] Application Packaging in VMware ThinApp 4.7 Essentials [Article]
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Packt
04 Jun 2015
17 min read
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Data Analysis Using R

Packt
04 Jun 2015
17 min read
In this article by Viswa Viswanathan and Shanthi Viswanathan, the authors of the book R Data Analysis Cookbook, we discover how R can be used in various ways such as comparison, classification, applying different functions, and so on. We will cover the following recipes: Creating charts that facilitate comparisons Building, plotting, and evaluating – classification trees Using time series objects Applying functions to subsets of a vector (For more resources related to this topic, see here.) Creating charts that facilitate comparisons In large datasets, we often gain good insights by examining how different segments behave. The similarities and differences can reveal interesting patterns. This recipe shows how to create graphs that enable such comparisons. Getting ready If you have not already done so, download the code files and save the daily-bike-rentals.csv file in your R working directory. Read the data into R using the following command: > bike <- read.csv("daily-bike-rentals.csv") > bike$season <- factor(bike$season, levels = c(1,2,3,4),   labels = c("Spring", "Summer", "Fall", "Winter")) > attach(bike) How to do it... We base this recipe on the task of generating histograms to facilitate the comparison of bike rentals by season. Using base plotting system We first look at how to generate histograms of the count of daily bike rentals by season using R's base plotting system: Set up a 2 X 2 grid for plotting histograms for the four seasons: > par(mfrow = c(2,2)) Extract data for the seasons: > spring <- subset(bike, season == "Spring")$cnt > summer <- subset(bike, season == "Summer")$cnt > fall <- subset(bike, season == "Fall")$cnt > winter <- subset(bike, season == "Winter")$cnt Plot the histogram and density for each season: > hist(spring, prob=TRUE,   xlab = "Spring daily rentals", main = "") > lines(density(spring)) >  > hist(summer, prob=TRUE,   xlab = "Summer daily rentals", main = "") > lines(density(summer)) >  > hist(fall, prob=TRUE,   xlab = "Fall daily rentals", main = "") > lines(density(fall)) >  > hist(winter, prob=TRUE,   xlab = "Winter daily rentals", main = "") > lines(density(winter)) You get the following output that facilitates comparisons across the seasons: Using ggplot2 We can achieve much of the preceding results in a single command: > qplot(cnt, data = bike) + facet_wrap(~ season, nrow=2) +   geom_histogram(fill = "blue") You can also combine all four into a single histogram and show the seasonal differences through coloring: > qplot(cnt, data = bike, fill = season) How it works... When you plot a single variable with qplot, you get a histogram by default. Adding facet enables you to generate one histogram per level of the chosen facet. By default, the four histograms will be arranged in a single row. Use facet_wrap to change this. There's more... You can use ggplot2 to generate comparative boxplots as well. Creating boxplots with ggplot2 Instead of the default histogram, you can get a boxplot with either of the following two approaches: > qplot(season, cnt, data = bike, geom = c("boxplot"), fill = season) >  > ggplot(bike, aes(x = season, y = cnt)) + geom_boxplot() The preceding code produces the following output: The second line of the preceding code produces the following plot: Building, plotting, and evaluating – classification trees You can use a couple of R packages to build classification trees. Under the hood, they all do the same thing. Getting ready If you do not already have the rpart, rpart.plot, and caret packages, install them now. Download the data files and place the banknote-authentication.csv file in your R working directory. How to do it... This recipe shows you how you can use the rpart package to build classification trees and the rpart.plot package to generate nice-looking tree diagrams: Load the rpart, rpart.plot, and caret packages: > library(rpart) > library(rpart.plot) > library(caret) Read the data: > bn <- read.csv("banknote-authentication.csv") Create data partitions. We need two partitions—training and validation. Rather than copying the data into the partitions, we will just keep the indices of the cases that represent the training cases and subset as and when needed: > set.seed(1000) > train.idx <- createDataPartition(bn$class, p = 0.7, list = FALSE) Build the tree: > mod <- rpart(class ~ ., data = bn[train.idx, ], method = "class", control = rpart.control(minsplit = 20, cp = 0.01)) View the text output (your result could differ if you did not set the random seed as in step 3): > mod n= 961   node), split, n, loss, yval, (yprob)      * denotes terminal node   1) root 961 423 0 (0.55983351 0.44016649)    2) variance>=0.321235 511 52 0 (0.89823875 0.10176125)      4) curtosis>=-4.3856 482 29 0 (0.93983402 0.06016598)        8) variance>=0.92009 413 10 0 (0.97578692 0.02421308) *        9) variance< 0.92009 69 19 0 (0.72463768 0.27536232)        18) entropy< -0.167685 52   6 0 (0.88461538 0.11538462) *        19) entropy>=-0.167685 17   4 1 (0.23529412 0.76470588) *      5) curtosis< -4.3856 29   6 1 (0.20689655 0.79310345)      10) variance>=2.3098 7   1 0 (0.85714286 0.14285714) *      11) variance< 2.3098 22   0 1 (0.00000000 1.00000000) *    3) variance< 0.321235 450 79 1 (0.17555556 0.82444444)      6) skew>=6.83375 76 18 0 (0.76315789 0.23684211)      12) variance>=-3.4449 57   0 0 (1.00000000 0.00000000) *      13) variance< -3.4449 19   1 1 (0.05263158 0.94736842) *      7) skew< 6.83375 374 21 1 (0.05614973 0.94385027)      14) curtosis>=6.21865 106 16 1 (0.15094340 0.84905660)        28) skew>=-3.16705 16   0 0 (1.00000000 0.00000000) *       29) skew< -3.16705 90   0 1 (0.00000000 1.00000000) *      15) curtosis< 6.21865 268   5 1 (0.01865672 0.98134328) * Generate a diagram of the tree (your tree might differ if you did not set the random seed as in step 3): > prp(mod, type = 2, extra = 104, nn = TRUE, fallen.leaves = TRUE, faclen = 4, varlen = 8, shadow.col = "gray") The following output is obtained as a result of the preceding command: Prune the tree: > # First see the cptable > # !!Note!!: Your table can be different because of the > # random aspect in cross-validation > mod$cptable            CP nsplit rel error   xerror       xstd 1 0.69030733     0 1.00000000 1.0000000 0.03637971 2 0.09456265     1 0.30969267 0.3262411 0.02570025 3 0.04018913     2 0.21513002 0.2387707 0.02247542 4 0.01891253     4 0.13475177 0.1607565 0.01879222 5 0.01182033     6 0.09692671 0.1347518 0.01731090 6 0.01063830     7 0.08510638 0.1323877 0.01716786 7 0.01000000     9 0.06382979 0.1276596 0.01687712   > # Choose CP value as the highest value whose > # xerror is not greater than minimum xerror + xstd > # With the above data that happens to be > # the fifth one, 0.01182033 > # Your values could be different because of random > # sampling > mod.pruned = prune(mod, mod$cptable[5, "CP"]) View the pruned tree (your tree will look different): > prp(mod.pruned, type = 2, extra = 104, nn = TRUE, fallen.leaves = TRUE, faclen = 4, varlen = 8, shadow.col = "gray") Use the pruned model to predict for a validation partition (note the minus sign before train.idx to consider the cases in the validation partition): > pred.pruned <- predict(mod, bn[-train.idx,], type = "class") Generate the error/classification-confusion matrix: > table(bn[-train.idx,]$class, pred.pruned, dnn = c("Actual", "Predicted"))      Predicted Actual   0   1      0 213 11      1 11 176 How it works... Steps 1 to 3 load the packages, read the data, and identify the cases in the training partition, respectively. In step 3, we set the random seed so that your results should match those that we display. Step 4 builds the classification tree model: > mod <- rpart(class ~ ., data = bn[train.idx, ], method = "class", control = rpart.control(minsplit = 20, cp = 0.01)) The rpart() function builds the tree model based on the following:   Formula specifying the dependent and independent variables   Dataset to use   A specification through method="class" that we want to build a classification tree (as opposed to a regression tree)   Control parameters specified through the control = rpart.control() setting; here we have indicated that the tree should only consider nodes with at least 20 cases for splitting and use the complexity parameter value of 0.01—these two values represent the defaults and we have included these just for illustration Step 5 produces a textual display of the results. Step 6 uses the prp() function of the rpart.plot package to produce a nice-looking plot of the tree: > prp(mod, type = 2, extra = 104, nn = TRUE, fallen.leaves = TRUE, faclen = 4, varlen = 8, shadow.col = "gray")   use type=2 to get a plot with every node labeled and with the split label below the node   use extra=4 to display the probability of each class in the node (conditioned on the node and hence summing to 1); add 100 (hence extra=104) to display the number of cases in the node as a percentage of the total number of cases   use nn = TRUE to display the node numbers; the root node is node number 1 and node n has child nodes numbered 2n and 2n+1   use fallen.leaves=TRUE to display all leaf nodes at the bottom of the graph   use faclen to abbreviate class names in the nodes to a specific maximum length   use varlen to abbreviate variable names   use shadow.col to specify the color of the shadow that each node casts Step 7 prunes the tree to reduce the chance that the model too closely models the training data—that is, to reduce overfitting. Within this step, we first look at the complexity table generated through cross-validation. We then use the table to determine the cutoff complexity level as the largest xerror (cross-validation error) value that is not greater than one standard deviation above the minimum cross-validation error. Steps 8 through 10 display the pruned tree; use the pruned tree to predict the class for the validation partition and then generate the error matrix for the validation partition. There's more... We discuss in the following an important variation on predictions using classification trees. Computing raw probabilities We can generate probabilities in place of classifications by specifying type="prob": > pred.pruned <- predict(mod, bn[-train.idx,], type = "prob") Create the ROC Chart Using the preceding raw probabilities and the class labels, we can generate a ROC chart: > pred <- prediction(pred.pruned[,2], bn[-train.idx,"class"]) > perf <- performance(pred, "tpr", "fpr") > plot(perf) Using time series objects In this recipe, we look at various features to create and plot time-series objects. We will consider data with both a single and multiple time series. Getting ready If you have not already downloaded the data files, do it now and ensure that the files are in your R working directory. How to do it... Read the data. The file has 100 rows and a single column named sales: > s <- read.csv("ts-example.csv") Convert the data to a simplistic time series object without any explicit notion of time: > s.ts <- ts(s) > class(s.ts) [1] "ts" Plot the time series: > plot(s.ts) Create a proper time series object with proper time points: > s.ts.a <- ts(s, start = 2002) > s.ts.a Time Series: Start = 2002 End = 2101 Frequency = 1        sales [1,]   51 [2,]   56 [3,]   37 [4,]   101 [5,]   66 (output truncated) > plot(s.ts.a) > # results show that R treated this as an annual > # time series with 2002 as the starting year The result of the preceding commands is seen in the following graph: To create a monthly time series run the following command: > # Create a monthly time series > s.ts.m <- ts(s, start = c(2002,1), frequency = 12) > s.ts.m        Jan Feb Mar Apr May Jun Jul Aug Sep Oct Nov Dec 2002 51 56 37 101 66 63 45 68 70 107 86 102 2003 90 102 79 95 95 101 128 109 139 119 124 116 2004 106 100 114 133 119 114 125 167 149 165 135 152 2005 155 167 169 192 170 180 175 207 164 204 180 203 2006 215 222 205 202 203 209 200 199 218 221 225 212 2007 250 219 242 241 267 249 253 242 251 279 298 260 2008 269 257 279 273 275 314 288 286 290 288 304 291 2009 314 290 312 319 334 307 315 321 339 348 323 342 2010 340 348 354 291 > plot(s.ts.m) # note x axis on plot The following plot can be seen as a result of the preceding commands: > # Specify frequency = 4 for quarterly data > s.ts.q <- ts(s, start = 2002, frequency = 4) > s.ts.q        Qtr1 Qtr2 Qtr3 Qtr4 2002   51   56   37 101 2003   66   63   45   68 2004   70 107   86 102 2005   90 102   79   95 2006   95 101 128 109 (output truncated) > plot(s.ts.q) Query time series objects (we use the s.ts.m object we created in the previous step): > # When does the series start? > start(s.ts.m) [1] 2002   1 > # When does it end? > end(s.ts.m) [1] 2010   4 > # What is the frequency? > frequency(s.ts.m) [1] 12 Create a time series object with multiple time series. This data file contains US monthly consumer prices for white flour and unleaded gas for the years 1980 through 2014 (downloaded from the website of the US Bureau of Labor Statistics): > prices <- read.csv("prices.csv") > prices.ts <- ts(prices, start=c(1980,1), frequency = 12) Plot a time series object with multiple time series: > plot(prices.ts) The plot in two separate panels appears as follows: > # Plot both series in one panel with suitable legend > plot(prices.ts, plot.type = "single", col = 1:2) > legend("topleft", colnames(prices.ts), col = 1:2, lty = 1) Two series plotted in one panel appear as follow: How it works... Step 1 reads the data. Step 2 uses the ts function to generate a time series object based on the raw data. Step 3 uses the plot function to generate a line plot of the time series. We see that the time axis does not provide much information. Time series objects can represent time in more friendly terms. Step 4 shows how to create time series objects with a better notion of time. It shows how we can treat a data series as an annual, monthly, or quarterly time series. The start and frequency parameters help us to control these data series. Although the time series we provide is just a list of sequential values, in reality our data can have an implicit notion of time attached to it. For example, the data can be annual numbers, monthly numbers, or quarterly ones (or something else, such as 10-second observations of something). Given just the raw numbers (as in our data file, ts-example.csv), the ts function cannot figure out the time aspect and by default assumes no secondary time interval at all. We can use the frequency parameter to tell ts how to interpret the time aspect of the data. The frequency parameter controls how many secondary time intervals there are in one major time interval. If we do not explicitly specify it, by default frequency takes on a value of 1. Thus, the following code treats the data as an annual sequence, starting in 2002: > s.ts.a <- ts(s, start = 2002) The following code, on the other hand, treats the data as a monthly time series, starting in January 2002. If we specify the start parameter as a number, then R treats it as starting at the first subperiod, if any, of the specified start period. When we specify frequency as different from 1, then the start parameter can be a vector such as c(2002,1) to specify the series, the major period, and the subperiod where the series starts. c(2002,1) represent January 2002: > s.ts.m <- ts(s, start = c(2002,1), frequency = 12) Similarly, the following code treats the data as a quarterly sequence, starting in the first quarter of 2002: > s.ts.q <- ts(s, start = 2002, frequency = 4) The frequency values of 12 and 4 have a special meaning—they represent monthly and quarterly time sequences. We can supply start and end, just one of them, or none. If we do not specify either, then R treats the start as 1 and figures out end based on the number of data points. If we supply one, then R figures out the other based on the number of data points. While start and end do not play a role in computations, frequency plays a big role in determining seasonality, which captures periodic fluctuations. If we have some other specialized time series, we can specify the frequency parameter appropriately. Here are two examples:   With measurements taken every 10 minutes and seasonality pegged to the hour, we should specify frequency as 6   With measurements taken every 10 minutes and seasonality pegged to the day, use frequency = 24*6 (6 measurements per hour times 24 hours per day) Step 5 shows the use of the functions start, end, and frequency to query time series objects. Steps 6 and 7 show that R can handle data files that contain multiple time series. Applying functions to subsets of a vector The tapply function applies a function to each partition of the dataset. Hence, when we need to evaluate a function over subsets of a vector defined by a factor, tapply comes in handy. Getting ready Download the files and store the auto-mpg.csv file in your R working directory. Read the data and create factors for the cylinders variable: > auto <- read.csv("auto-mpg.csv", stringsAsFactors=FALSE) > auto$cylinders <- factor(auto$cylinders, levels = c(3,4,5,6,8),   labels = c("3cyl", "4cyl", "5cyl", "6cyl", "8cyl")) How to do it... To apply functions to subsets of a vector, follow these steps: Calculate mean mpg for each cylinder type: > tapply(auto$mpg,auto$cylinders,mean)      3cyl     4cyl     5cyl     6cyl     8cyl 20.55000 29.28676 27.36667 19.98571 14.96311 We can even specify multiple factors as a list. The following example shows only one factor since the out file has only one, but it serves as a template that you can adapt: > tapply(auto$mpg,list(cyl=auto$cylinders),mean)   cyl    3cyl     4cyl     5cyl     6cyl     8cyl 20.55000 29.28676 27.36667 19.98571 14.96311 How it works... In step 1 the mean function is applied to the auto$mpg vector grouped according to the auto$cylinders vector. The grouping factor should be of the same length as the input vector so that each element of the first vector can be associated with a group. The tapply function creates groups of the first argument based on each element's group affiliation as defined by the second argument and passes each group to the user-specified function. Step 2 shows that we can actually group by several factors specified as a list. In this case, tapply applies the function to each unique combination of the specified factors. There's more... The by function is similar to tapply and applies the function to a group of rows in a dataset, but by passing in the entire data frame. The following examples clarify this. Applying a function on groups from a data frame In the following example, we find the correlation between mpg and weight for each cylinder type: > by(auto, auto$cylinders, function(x) cor(x$mpg, x$weight)) auto$cylinders: 3cyl [1] 0.6191685 --------------------------------------------------- auto$cylinders: 4cyl [1] -0.5430774 --------------------------------------------------- auto$cylinders: 5cyl [1] -0.04750808 --------------------------------------------------- auto$cylinders: 6cyl [1] -0.4634435 --------------------------------------------------- auto$cylinders: 8cyl [1] -0.5569099 Summary Being an extensible system, R's functionality is divided across numerous packages with each one exposing large numbers of functions. Even experienced users cannot expect to remember all the details off the top of their head. In this article, we went through a few techniques using which R helps analyze data and visualize the results. Resources for Article: Further resources on this subject: Combining Vector and Raster Datasets [article] Factor variables in R [article] Big Data Analysis (R and Hadoop) [article]
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article-image-events-notifications-and-reporting
Packt
04 Jun 2015
55 min read
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Events, Notifications, and Reporting

Packt
04 Jun 2015
55 min read
In this article by Martin Wood, the author of the book, Mastering ServiceNow, has discussed about communication which is a key part of any business application. Not only does the boss need to have an updated report by Monday, but your customers and users also want to be kept informed. ServiceNow helps users who want to know what's going on. In this article, we'll explore the functionality available. The platform can notify and provide information to people in a variety of ways: Registering events and creating Scheduled Jobs to automate functionality Sending out informational e-mails when something happens Live dashboards and homepages showing the latest reports and statistics Scheduled reports that help with handover between shifts Capturing information with metrics Presenting a single set of consolidated data with database views (For more resources related to this topic, see here.) Dealing with events Firing an event is a way to tell the platform that something happened. Since ServiceNow is a data-driven system, in many cases, this means that a record has been updated in some way. For instance, maybe a guest has been made a VIP, or has stayed for 20 nights. Several parts of the system may be listening for an event to happen. When it does, they perform an action. One of these actions may be sending an e-mail to thank our guest for their continued business. These days, e-mail notifications don't need to be triggered by events. However, it is an excellent example. When you fire an event, you pass through a GlideRecord object and up to two string parameters. The item receiving this data can then use it as necessary, so if we wanted to send an e-mail confirming a hotel booking, we have those details to hand during processing. Registering events Before an event can be fired, it must be known to the system. We do this by adding it to Event Registry [sysevent_register], which can be accessed by navigating to System Policy > Events > Registry. It's a good idea to check whether there isn't one you can use before you add a new one. An event registration record consists of several fields, but most importantly a string name. An event can be called anything, but by convention it is in a dotted namespace style format. Often, it is prefixed by the application or table name and then by the activity that occurred. Since a GlideRecord object accompanies an event, the table that the record will come from should also be selected. It is also a good idea to describe your event and what will cause it in the Description and Fired by fields. Finally, there is a field that is often left empty, called Queue. This gives us the functionality to categorize events and process them in a specific order or frequency. Firing an event Most often, a script in a Business Rule will notice that something happens and will add an event to the Event [sysevent] queue. This table stores all of the events that have been fired, if it has been processed, and what page the user was on when it happened. As the events come in, the platform deals with them in a first in, first out order by default. It finds everything that is listening for this specific event and executes them. That may be an e-mail notification or a script. By navigating to System Policy > Events > Event Log, you can view the state of an event, when it was added to the queue, and when it was processed. To add an event to the queue, use the eventQueue function of GlideSystem. It accepts four parameters: the name of the event, a GlideRecord object, and two run time parameters. These can be any text strings, but most often are related to the user that caused the event. Sending an e-mail for new reservations Let's create an event that will fire when a Maintenance task has been assigned to one of our teams. Navigate to System Policy > Events > Registry. Click on New and set the following fields:     Event name: maintenance.assigned     Table: Maintenance [u_maintenance] Next, we need to add the event to the Event Queue. This is easily done with a simple Business Rule:     Name: Maintenance assignment events     Table: Maintenance [u_maintenance]     Advanced: <ticked>     When: after Make sure to always fire events after the record has been written to the database. This stops the possibility of firing an event even though another script has aborted the action. Insert: <ticked> Update: <ticked> Filter Conditions: Assignment group – changes Assignment group – is not empty Assigned to – is empty This filter represents when a task is sent to a new group but someone hasn't yet been identified to own the work. Script: gs.eventQueue('maintenance.assigned', current, gs.getUserID(), gs.getUserName()); This script follows the standard convention when firing events—passing the event name, current, which contains the GlideRecord object the Business Rule is working with, and some details about the user who is logged in. We'll pick this event up later and send an e-mail whenever it is fired. There are several events, such as <table_name>.view, that are fired automatically. A very useful one is the login event. Take a look at the Event Log to see what is happening. Scheduling jobs You may be wondering how the platform processes the event queue. What picks them up? How often are they processed? In order to make things happen automatically, ServiceNow has a System Scheduler. Processing the event queue is one job that is done on a repeated basis. ServiceNow can provide extra worker nodes that only process events. These shift the processing of things such as e-mails onto another system, enabling the other application nodes to better service user interactions. To see what is going on, navigate to System Scheduler > Scheduled Jobs > Today's Scheduled Jobs. This is a link to the Schedule Item [sys_trigger] table, a list of everything the system is doing in the background. You will see a job that collects database statistics, another that upgrades the instance (if appropriate), and others that send and receive e-mails or SMS messages. You should also spot one called events process, which deals with the event queue. A Schedule Item has a Next action date and time field. This is when the platform will next run the job. Exactly what will happen is specified through the Job ID field. This is a reference to the Java class in the platform that will actually do the work. The majority of the time, this is RunScriptJob, which will execute some JavaScript code. The Trigger type field specifies how often the job will repeat. Most jobs are run repetitively, with events process set to run every 30 seconds. Others run when the instance is started—perhaps to preload the cache. Another job that is run on a periodic basis is SMTP Sender. Once an e-mail has been generated and placed in the sys_email table, the SMTP Sender job performs the same function as many desktop e-mail clients: it connects to an e-mail server and asks it to deliver the message. It runs every minute by default. This schedule has a direct impact on how quickly our e-mail will be sent out. There may be a delay of up to 30 seconds in generating the e-mail from an event, and a further delay of up to a minute before the e-mail is actually sent. Other jobs may process a particular event queue differently. Events placed into the metric queue will be worked with after 5 seconds. Adding your own jobs The sys_trigger table is a backend data store. It is possible to add your own jobs and edit what is already there, but I don't recommend it. Instead, there is a more appropriate frontend: the Scheduled Job [sysauto] table. The sysauto table is designed to be extended. There are many things that can be automated in ServiceNow, including data imports, sending reports, and creating records, and they each have a table extended from sysauto. Once you create an entry in the sysauto table, the platform creates the appropriate record in the sys_trigger table. This is done through a call in the automation synchronizer Business Rule. Each table extended from sysauto contains fields that are relevant to its automation. For example, a Scheduled Email of Report [sysauto_report] requires e-mail addresses and reports to be specified. Creating events every day Navigate to System Definition > Scheduled Jobs. Unfortunately, the sys_trigger and sysauto tables have very similar module names. Be sure to pick the right one. When you click on New, an interceptor will fire, asking you to choose what you want to automate. Let's write a simple script that will create a maintenance task at the end of a hotel stay, so choose Automatically run a script of your choosing. Our aim is to fire an event for each room that needs cleaning. We'll keep this for midday to give our guests plenty of time to check out. Set the following fields: Name: Clean on end of reservation Time: 12:00:00 Run this script: var res = new GlideRecord('u_reservation'); res.addQuery('u_departure', gs.now()); res.addNotNullQuery('u_room'); res.query(); while (res.next()) { gs.eventQueue('room.reservation_end', res.u_room.getRefRecord()); } Remember to enclose scripts in a function if they could cause other scripts to run. Most often, this is when records are updated, but it is not the case here. Our reliable friend, GlideRecord, is employed to get reservation records. The first filter ensures that only reservations that are ending today will be returned, while the second filter ignores reservations that don't have a room. Once the database has been queried, the records are looped round. For each one, the eventQueue function of GlideSystem is used to add in an event into the event queue. The record that is being passed into the event queue is actually the Room record. The getRefRecord function of GlideElement dot-walks through a reference field and returns a newly initialized GlideRecord object rather than more GlideElement objects. Once the Scheduled Job has been saved, it'll generate the events at midday. But for testing, there is a handy Execute Now UI action. Ensure there is test data that fits the code and click on the button. Navigate to System Policy > Events > Event Log to see the entries. There is a Conditional checkbox with a separate Condition script field. However, I don't often use this; instead, I provide any conditions inline in the script that I'm writing, just like we did here. For anything more than a few lines, a Script Include should be used for modularity and efficiency. Running scripts on events The ServiceNow platform has several items that listen for events. Email Notifications are one, which we'll explore soon. Another is Script Actions. Script Actions is server-side code that is associated with a table and runs against a record, just like a Business Rule. But instead of being triggered by a database action, a Script Action is started with an event. There are many similarities between a Script Action and an asynchronous Business Rule. They both run server-side, asynchronous code. Unless there is a particular reason, stick to Business Rules for ease and familiarity. Just like a Business Rule, the GlideRecord variable called current is available. This is the same record that was passed into the second parameter when gs.eventQueue was called. Additionally, another GlideRecord variable called event is provided. It is initialized against the appropriate Event record on the sysevent table. This gives you access to the other parameters (event.param1 and event.param2) as well as who created the event, when, and more. Creating tasks automatically When creating a Script Action, the first step is to register or identify the event it will be associated with. Create another entry in Event Registry. Event name: room.reservation_end Table: Room [u_room] In order to make the functionality more data driven, let's create another template. Either navigate to System Definition > Templates or create a new Maintenance task and use the Save as Template option in the context menu. Regardless, set the following fields: Name: End of reservation room cleaning Table: Maintenance [u_maintenance] Template: Assignment group: Housekeeping Short description: End of reservation room cleaning Description: Please perform the standard cleaning for the room listed above. To create the Script Action, go to System Policy > Events > Script Actions and use the following details: Name: Produce maintenance tasks Event name: room.reservation_end Active: <ticked> Script: var tsk = new GlideRecord('u_maintenance'); tsk.newRecord(); tsk.u_room = current.sys_id; tsk.applyTemplate('End of reservation room cleaning'); tsk.insert(); This script is quite straightforward. It creates a new GlideRecord object that represents a record in the Maintenance table. The fields are initialized through newRecord, and the Room field is populated with the sys_id of current—which is the Room record that the event is associated with. The applyTemplate function is given the name of the template. It would be better to use a property here instead of hardcoding a template name. Now, the following items should occur every day: At midday, a Scheduled Job looks for any reservations that are ending today For each one, the room.reservation_end event is fired A Script Action will be called, which creates a new Maintenance task The Maintenance task is assigned, through a template, to the Housekeeping group. But how does Housekeeping know that this task has been created? Let's send them an e-mail! Sending e-mail notifications E-mail is ubiquitous. It is often the primary form of communication in business, so it is important that ServiceNow has good support. It is easy to configure ServiceNow to send out communications to whoever needs to know. There are a few general use cases for e-mail notifications: Action: Asking the receiver to do some work Informational: Giving the receiver an update or some data Approval: Asking for a decision While this is similar enough to an action e-mail, it is a common enough scenario to make it independent. We'll work through these scenarios in order to understand how ServiceNow can help. There are obviously a lot more ways you can use e-mails. One of them is for a machine-to-machine integration, such as e-bonding. It is possible to do this in ServiceNow, but it is not the best solution. Setting e-mail properties A ServiceNow instance uses standard protocols to send and receive e-mail. E-mails are sent by connecting to an SMTP server with a username and password, just like Outlook or any other e-mail client. When an instance is provisioned, it also gets an e-mail account. If your instance is available at instance.service-now.com through the Web, it has an e-mail address of instance@service-now.com. This e-mail account is not unusual. It is accessible via POP to receive mail, and uses SMTP to send it. Indeed, any standard e-mail account can be used with an instance. Navigate to System Properties > Email to investigate the settings. The properties are unusually laid out in two columns, for sending and receiving for the SMTP and POP connections. When you reach the page, the settings will be tested, so you can immediately see if the platform is capable of sending or receiving e-mails. Before you spend time configuring Email Notifications, make sure the basics work! ServiceNow will only use one e-mail account to send out e-mails, and by default, will only check for new e-mails in one account too. Tracking sent e-mails in the Activity Log One important feature of Email Notifications is that they can show up in the Activity Log if configured. This means that all e-mails associated with a ticket are associated and kept together. This is useful when tracking correspondence with a Requester. To configure the Activity Log, navigate to a Maintenance record. Right-click on the field and choose Personalize Activities. At the bottom of the Available list is Sent/Received Emails. Add it to the Selected list and click on Save. Once an e-mail has been sent out, check back to the Activity Formatter to see the results. Assigning work Our Housekeeping team is equipped with the most modern technology. Not only are they users of ServiceNow, but they have mobile phones that will send and receive e-mails. They have better things to do than constantly refresh the web interface, so let's ensure that ServiceNow will come to them. One of the most common e-mail notifications is for ServiceNow to inform people when they have been assigned a task. It usually gives an overview and a link to view more details. This e-mail tells them that something needs to happen and that ServiceNow should be updated with the result. Sending an e-mail notification on assignment When our Maintenance tasks have the Assignment group field populated, we need the appropriate team members to be aware. We are going to achieve this by sending an e-mail to everyone in that group. At Gardiner Hotels, we empower our staff: they know that one member of the team should pick the task up and own it by setting the Assigned to field to themselves and then get it done. Navigate to System Policy > Email > Notifications. You will see several examples that are useful to understand the basic configuration, but we'll create our own. Click on New. The Email Notifications form is split into three main sections: When to send, Who will receive, and What it will contain. Some options are hidden in a different view, so click on Advanced view to see them all. Start off by giving the basic details: Name: Group assignment Table: Maintenance [u_maintenance] Now, let's see each of the sections of Email Notifications form in detail, in the following sections. When to send This section gives you a choice of either using an event to determine which record should be worked with or for the e-mail notification system to monitor the table directly. Either way, Conditions and Advanced conditions lets you provide a filter or a script to ensure you only send e-mails at the right time. If you are using an event, the event must be fired and the condition fields satisfied for the e-mail to be sent. The Weight field is often overlooked. A single event or record update may satisfy the condition of multiple Email Notifications. For example, a common scenario is to send an e-mail to the Assignment group when it is populated and to send an e-mail to the Assigned to person when that is populated. But what if they both happen at the same time? You probably don't want the Assignment group being told to pick up a task if it has already been assigned. One way is to give the Assignment group e-mail a higher weight: if two e-mails are being generated, only the lower weight will be sent. The other will be marked as skipped. Another way to achieve this scenario is through conditions. Only send the Assignment group e-mail if the Assigned to field is empty. Since we've already created an event, let's use it. And because of the careful use of conditions in the Business Rule, it only sends out the event in the appropriate circumstances. That means no condition is necessary in this Email Notification. Send when: Event is fired Event name: maintenance.assigned Who will receive Once we've determined when an e-mail should be sent, we need to know who it will go to. The majority of the time, it'll be driven by data on the record. This scenario is exactly that: the people who will receive the e-mail are those in the Assignment group field on the Maintenance task. Of course, it is possible to hardcode recipients and the system can also deliver e-mails to Users and Groups that have been sent as a parameter when creating the event. Users/Groups in fields: Assignment group You can also use scripts to specify the From, To, CC, and BCC of an e-mail. The wiki here contains more information: http://wiki.servicenow.com/?title=Scripting_for_Email_Notifications Send to event creator When someone comes to me and says: "Martin, I've set up the e-mail notification, but it isn't working. Do you know why?", I like to put money on the reason. I very often win, and you can too. Just answer: "Ensure Send to event creator is ticked and try again". The Send to event creator field is only visible on the Advanced view, but is the cause of this problem. So tick Send to event creator. Make sure this field is ticked, at least for now. If you do not, when you test your e-mail notifications, you will not receive your e-mail. Why? By default, the system will not send confirmation e-mails. If you were the person to update a record and it causes e-mails to be sent, and it turns out that you are one of the recipients, it'll go to everyone other than you. The reasoning is straightforward: you carried out the action so why do you need to be informed that it happened? This cuts down on unnecessary e-mails and so is a good thing. But it confuses everyone who first comes across it. If there is one tip I can give to you in this article, it is this – tick the Send to event creator field when testing e-mails. Better still, test realistically! What it will contain The last section is probably the simplest to understand, but the one that takes most time: deciding what to send. The standard view contains just a few fields: a space to enter your message, a subject line, and an SMS alternate field that is used for text messages. Additionally, there is an Email template field that isn't often used but is useful if you want to deliver the same content in multiple e-mail messages. View them by navigating to System Policy > Email > Templates. These fields all support variable substitution. This is a special syntax that instructs the instance to insert data from the record that the e-mail is triggered for. This Maintenance e-mail can easily contain data from the Maintenance record. This lets you create data-driven e-mails. I like to compare it to a mail-merge system; you have some fixed text, some placeholders, and some data, and the platform puts them all together to produce a personalized e-mail. By default, the message will be delivered as HTML. This means you can make your messages look more styled by using image tags and font controls, among other options. Using variable substitution The format for substitution is ${variable}. All of the fields on the record are available as variables, so to include the Short description field in an e-mail, use ${short_description}. Additionally, you can dot-walk. So by having ${assigned_to.email} in the message, you insert the e-mail address of the user that the task is assigned to. Populate the fields with the following information and save: Subject: Maintenance task assigned to your group Message HTML: Hello ${assignment_group}. Maintenance task ${number} has been assigned to your group, for room: ${u_room}. Description: ${description} Please assign to a team member here: ${URI} Thanks! To make this easier, there is a Select variables section on the Message HTML and SMS alternate fields that will create the syntax in a single click. But don't forget that variable substitution is available for the Subject field too. In addition to adding the value of fields, variable substitution like the following ones also makes it easy to add HTML links. ${<reference field>.URI} will create an HTML link to the reference field, with the text LINK ${<reference field>.URI_REF} will create an HTML link, but with the display value of the record as the text Linking to CMS sites is possible through ${CMS_URI+<site>/<page>} Running scripts in e-mail messages If the variables aren't giving you enough control, like everywhere else in ServiceNow, you can add a script. To do so, create a new entry in the Email Scripts [sys_script_email] table, which is available under System Policy > Email > Notification Email Scripts. Typical server-side capability is present, including the current GlideRecord variable. To output text, use the print function of the template object. For example: template.print('Hello, world!'); Like a Script Include, the Name field is important. Call the script by placing ${mail_script:<name>} in the Message HTML field in the e-mail. An object called email is also available. This gives much more control with the resulting e-mail, giving functions such as setImportance, addAddress, and setReplyTo. This wiki has more details: http://wiki.servicenow.com/?title=Scripting_for_Email_Notifications. Controlling the watermark Every outbound mail contains a reference number embedded into the body of the message, in the format Ref:MSG0000100. This is very important for the inbound processing of e-mails, as discussed in a later section. Some options are available to hide or remove the watermark, but this may affect how the platform treats a reply. Navigating to System Mailboxes > Administration > Watermarks shows a full list of every watermark and the associated record and e-mail. Including attachments and other options There are several other options to control how an e-mail is processed: Include Attachments: It will copy any attachments from the record into the e-mail. There is no selection available: it simply duplicates each one every time. You probably wouldn't want this option ticked on many e-mails, since otherwise you will fill up the recipient's inboxes quickly! The attach_links Email Script is a good alternative—it gives HTML links that will let an interested recipient download the file from the instance. Importance: This allows a Low or High priority flag to be set on an e-mail From and Reply-To fields: They'll let you configure who the e-mail purports to be from, on a per–e-mail basis. It is important to realize that this is e-mail spoofing: while the e-mail protocols accept this, it is often used by spam to forge a false address. Sending informational updates Many people rely on e-mails to know what is going on. In addition to telling users when they need to do work, ServiceNow can keep everyone informed as to the current situation. This often takes the form of one of these scenarios: Automatic e-mails, often based on a change of the State field Completely freeform text, with or without a template A combination of the preceding two: a textual update given by a person, but in a structured template Sending a custom e-mail Sometimes, you need to send an e-mail that doesn't fit into a template. Perhaps you need to attach a file, copy in additional people, or want more control over formatting. In many cases, you would turn to the e-mail client on your desktop, such as Outlook or perhaps even Lotus Notes. But the big disadvantage is that the association between the e-mail and the record is lost. Of course, you could save the e-mail and upload it as an attachment, but that isn't as good as it being part of the audit history. ServiceNow comes with a basic e-mail client built in. In fact, it is just shortcutting the process. When you use the e-mail client, you are doing exactly the same as the Email Notifications engine would, by generating an entry in the sys_email table. Enabling the e-mail client The Email Client is accessed by a little icon in the form header of a record. In order to show it, a property must be set in the Dictionary Entry of the table. Navigate to System Definition > Dictionary and find the entry for the u_maintenance table that does not have an entry in the Column name field. The value for the filter is Table - is - u_maintenance and Column name - is – empty. Click on Advanced view. Ensure the Attributes field contains email_client. Navigate to an existing Maintenance record, and next to the attachments icon is the envelope icon. Click on it to open the e-mail client window. The Email Client is a simple window, and the fields should be obvious. Simply fill them out and click on Send to deliver the mail. You may have noticed that some of the fields were prepopulated. You can control what each field initially contains by creating an Email Client Template. Navigate to System Policy > Email > Client Templates, click on New, and save a template for the appropriate table. You can use the variable substitution syntax to place the contents of fields in the e-mail. There is a Conditions field you can add to the form to have the right template used. Quick Messages are a way to let the e-mail user populate Message Text, similar to a record template. Navigate to System Policy > Email > Quick Messages and define some text. These are then available in a dropdown selection field at the top of the e-mail client. The e-mail client is often seized upon by customers who send a lot of e-mail. However, it is a simple solution and does not have a whole host of functionality that is often expected. I've found that this gap can be frustrating. For example, there isn't an easy way to include attachments from the parent record. Instead, often a more automated way to send custom text is useful. Sending e-mails with Additional comments and Work notes The journal fields on the task table are useful enough, allowing you to record results that are then displayed on the Activity log in a who, what, when fashion. But sending out the contents via e-mail makes them especially helpful. This lets you combine two actions in one: documenting information against the ticket and also giving an update to interested parties. The Task table has two fields that let you specify who those people are: the Watch list and the Work notes list. An e-mail notification can then use this information in a structured manner to send out the work note. It can include the contents of the work notes as well as images, styled text, and background information. Sending out Work notes The Work notes field should already be on the Maintenance form. Use Form Design to include the Work notes list field too, placing it somewhere appropriate, such as underneath the Assignment group field. Both the Watch list and the Work notes list are List fields (often referred to as Glide Lists). These are reference fields that contain more than one sys_id from the sys_user table. This makes it is easy to add a requester or fulfiller who is interested in updates to the ticket. What is special about List fields is that although they point towards the sys_user table and store sys_id references, they also store e-mail addresses in the same database field. The e-mail notification system knows all about this. It will run through the following logic: If it is a sys_id, the user record is looked up. The e-mail address in the user record is used. If it is an e-mail address, the user record is searched for. If one is found, any notification settings they have are respected. A user may turn off e-mails, for example, by setting the Notification field to Disabled in their user record. If a user record is not found, the e-mail is sent directly to the e-mail address. Now create a new Email Notification and fill out the following fields: Name: Work notes update Table: Maintenance [u_maintenance] Inserted: <ticked> Updated: <ticked> Conditions: Work notes - changes Users/Groups in fields: Work notes list Subject: New work notes update on ${number} Send to event creator: <ticked> Message: ${number} - ${short_description} has a new work note added.   ${work_notes} This simple message would normally be expanded and made to fit into the corporate style guidelines—use appropriate colors and styles. By default, the last three entries in the Work notes field would be included. If this wasn't appropriate, the global property could be updated or a mail script could use getJournalEntry(1) to grab the last one. Refer to this wiki article for more information: http://wiki.servicenow.com/?title=Using_Journal_Fields#Restrict_the_Number_of_Entries_Sent_in_a_Notification. To test, add an e-mail address or a user into the Work notes list, enter something into the Work notes field, and save. Don't forget about Send to event creator! This is a typical example of how, normally, the person doing the action wouldn't need to receive the e-mail update, since they were the one doing it. But set it so it'll work with your own updates. Approving via e-mail Graphical Workflow generates records that someone will need to evaluate and make a decision on. Most often, approvers will want to receive an e-mail notification to alert them to the situation. There are two approaches to sending out an e-mail when an approval is needed. An e-mail is associated with a particular record; and with approvals, there are two records to choose from: The Approval record, asking for your decision. The response will be processed by the Graphical Workflow. The system will send out one e-mail to each person that is requested to approve it. The Task record that generated the Approval request. The system will send out one e-mail in total. Attaching notifications to the task is sometimes helpful, since it gives you access to all the fields on the record without dot-walking. This section deals with how the Approval record itself uses e-mail notifications. Using the Approval table An e-mail that is sent out from the Approval table often contains the same elements: Some text describing what needs approving: perhaps the Short description or Priority. This is often achieved by dot-walking to the data through the Approval for reference field. A link to view the task that needs approval. A link to the approval record. Two mailto links that allow the user to approve or reject through e-mail. This style is captured in the Email Template named change.itil.approve.role and is used in an Email Notification called Approval Request that is against the Approval [sys_approver] table. The mailto links are generated through a special syntax: ${mailto:mailto.approval} and ${mailto:mailto.rejection}. These actually refer to Email Templates themselves (navigate to System Policy > Email > Templates and find the template called mailto.approval). Altogether, these generate HTML code in the e-mail message that looks something like this: <a href="mailto:<instance>@service-now.com.com?subject=Re:MAI0001001 - approve&body=Ref:MSG0000001">Click here to approve MAI0001001</a> Normally, this URL would be encoded, but I've removed the characters for clarity. When this link is clicked on in the receiver's e-mail client, it creates a new e-mail message addressed to the instance, with Re:MAI0001001 - approve in the subject line and Ref:MSG0000001 in the body. If this e-mail was sent, the instance would process it and approve the approval record. A later section, on processing inbound e-mails, shows in detail how this happens. Testing the default approval e-mail In the baseline system, there is an Email Notification called Approval Request. It is sent when an approval event is fired, which happens in a Business Rule on the Approval table. It uses the e-mail template mentioned earlier, giving the recipient information and an opportunity to approve it either in their web browser, or using their e-mail client. If Howard Johnson was set as the manager of the Maintenance group, he will be receiving any approval requests generated when the Send to External button is clicked on. Try changing the e-mail address in Howard's user account to your own, but ensure the Notification field is set to Enable. Then try creating some approval requests. Specifying Notification Preferences Every user that has access to the standard web interface can configure their own e-mail preferences through the Subscription Notification functionality. Navigate to Self-Service > My profile and click on Notification Preferences to explore what is available. It represents the Notification Messages [cmn_notif_message] table in a straightforward user interface. The Notification Preferences screen shows all the notifications that the user has received, such as the Approval Request and Work notes update configured earlier. They are organized by device. By default, every user has a primary e-mail device. To never receive a notification again, just choose the Off selection and save. This is useful if you are bombarded by e-mails and would rather use the web interface to see updates! If you want to ensure a user cannot unsubscribe, check the Mandatory field in the Email Notification definition record. You may need to add it to the form. This disables the choice, as per the Work notes update notification in the screenshot. Subscribing to Email Notifications The Email Notifications table has a field labeled Subscribable. If this is checked, then users can choose to receive a message every time the Email Notification record's conditions are met. This offers a different way of working: someone can decide if they want more information, rather than the administrator deciding. Edit the Work notes update Email Notification. Switch to the Advanced view, and using Form Design, add the Subscribable field to the Who will receive section on the form. Now make the following changes. Once done, use Insert and Stay to make a copy.     Name: Work notes update (Subscribable)     Users/Groups in fields: <blank>     Subscribable: <ticked> Go to Notification Preferences and click on To subscribe to a new notification click here. The new notification can be selected from the list. Now, every time a Work note is added to any Maintenance record, a notification will be sent to the subscriber. It is important to clear Users/Groups in field if Subscribable is ticked. Otherwise, everyone in the Work notes list will then become subscribed and receive every single subsequent notification for every record! The user can also choose to only receive a subset of the messages. The Schedule field lets them choose when to receive notifications: perhaps only during working hours. The filter lets you define conditions, such as only receiving notifications for important issues. In this instance, a Notification Filter could be created for the Maintenance table, based upon the Priority field. Then, only Work notes for high-priority Maintenance tasks would be sent out. Creating a new device The Notification Devices [cmn_notif_device] table stores e-mail addresses for users. It allows every user to have multiple e-mail addresses, or even register mobile phones for text messages. When a User record is created, a Business Rule named Create primary email device inserts a record in the Notification Devices table. The value in the Email field on the User table is just copied to this table by another Business Rule named Update Email Devices. A new device can be added from the Notification Preferences page, or a Related List can be added to the User form. Navigate to User Administration > Users and create a new user. Once saved, you should receive a message saying Primary email device created for user (the username is displayed in place of user). Then add the Notification Device > User Related List to the form where the e-mail address record should be visible. Click on New. The Notification Device form allows you to enter the details of your e-mail- or SMS-capable device. Service provider is a reference field to the Notification Service Provider table, which specifies how an SMS message should be sent. If you have an account with one of the providers listed, enter your details. There are many hundreds of inactive providers in the Notification Service Provider [cmn_notif_service_provider] table. You may want to try enabling some, though many do not work for the reasons discussed soon. Once a device has been added, they can be set up to receive messages through Notification Preferences. For example, a user can choose to receive approval requests via a text message by adding the Approval Request Notification Message and associating their SMS device. Alternatively, they could have two e-mail addresses, with one for an assistant. If a Notification is sent to a SMS device, the contents of the SMS alternate field are used. Remember that a text message can only be 160 characters at maximum. The Notification Device table has a field called Primary Email. This determines which device is used for a notification that has not been sent to this user before. Despite the name, Primary Email can be ticked for an SMS device. Sending text messages Many mobile phone networks in the US supply e-mail-to-SMS gateways. AT&T gives every subscriber an e-mail address in the form of 5551234567@txt.att.net. This allows the ServiceNow instance to actually send an e-mail and have the gateway convert it into an SMS. The Notification Service Provider form gives several options to construct the appropriate e-mail address. In this scheme, the recipient pays for the text message, so the sending of text messages is free. Many European providers do not provide such functionality, since the sender is responsible for paying. Therefore, it is more common to use the Web to deliver the message to the gateway: perhaps using REST or SOAP. This gives an authenticated method of communication, which allows charging. The Notifications Service Provider table also provides an Advanced notification checkbox that enables a script field. The code is run whenever the instance needs to send out an e-mail. This is a great place to call a Script Include that does the actual work, providing it with the appropriate parameters. Some global variables are present: email.SMSText contains the SMS alternate text and device is the GlideRecord of the Notification Device. This means device.phone_number and device.user are very useful values to access. Delivering an e-mail There are a great many steps that the instance goes through to send an e-mail. Some may be skipped or delivered as a shortcut, depending on the situation, but there are usually a great many steps that are processed. An e-mail may not be sent if any one of these steps goes wrong! A record is updated: Most notifications are triggered when a task changes state or a comment is added. Use debugging techniques to determine what is changing. These next two steps may not be used if the Notification does not use events. An event is fired: A Business Rule may fire an event. Look under System Policy > Events > Event Log to see if it was fired. The event is processed: A Scheduled Job will process each event in turn. Look in the Event Log and ensure that all events have their state changed to Processed. An Email Notification is processed: The event is associated with an Email Notification or the Email Notification uses the Inserted and Updated checkboxes to monitor a table directly. Conditions are evaluated: The platform checks the associated record and ensures the conditions are met. If not, no further processing occurs. The receivers are evaluated: The recipients are determined from the logic in the Email Notification. The use of Send to event creator makes a big impact on this step. The Notification Device is determined: The Notification Messages table is queried. The appropriate Notification Device is then found. If the Notification Device is set to inactive, the recipient is dropped. The Notification field on the User record will control the Active flag of the Notification Devices. Any Notification Device filters are applied: Any further conditions set in the Notification Preferences interface are evaluated, such as Schedule and Filter. An e-mail record is generated: Variable substitution takes place on the Message Text and a record is saved into the sys_email table, with details of the messages in the Outbox. The Email Client starts at this point. The weight is evaluated: If an Email Notification with a lower weight has already been generated for the same event, the e-mail has the Mailbox field set to Skipped. The email is sent: The SMTP Sender Scheduled Job runs every minute. It picks up all messages in the Outbox, generates the message ID, and connects to the SMTP server specified in Email properties. This only occurs if Mail sending is enabled in the properties. Errors will be visible under System Mailboxes > Outbound > Failed. The generated e-mails can be monitored in the System Mailboxes Application Menu, or through System Logs > Emails. They are categorized into Mailboxes, just like an e-mail client. This should be considered a backend table, though some customers who want more control over e-mail notifications make this more accessible. Knowing who the e-mail is from ServiceNow uses one account when sending e-mails. This account is usually the one provided by ServiceNow, but it can be anything that supports SMTP: Exchange, Sendmail, NetMail, or even Gmail. The SMTP protocol lets the sender specify who the mail is from. By default, no checks are done to ensure that the sender is allowed to send from that address. Every e-mail client lets you specify who the e-mail address is from, so I could change the settings in Outlook to say my e-mail address is president@whitehouse.gov or primeminister@number10.gov.uk. Spammers and virus writers have taken advantage of this situation to fill our mailboxes with unwanted e-mails. Therefore, e-mail systems are doing more authentication and checking of addresses when the message is received. You may have seen some e-mails from your client saying an e-mail has been delivered on behalf of another when this validation fails, or it even falling into the spam directly. ServiceNow uses SPF to specify which IP addresses can deliver service-now.com e-mails. Spam filters often use this to check if a sender is authorized. If you spoof the e-mail address, you may need to make an exception for ServiceNow. Read up more about it at: http://en.wikipedia.org/wiki/Sender_Policy_Framework. You may want to change the e-mail addresses on the instance to be your corporate domain. That means that your ServiceNow instance will send the message but will pretend that it is coming from another source. This runs the real risk of the e-mails being marked as spam. Instead, think about only changing the From display (not the e-mail address) or use your own e-mail account. Receiving e-mails Many systems can send e-mails. But isn't it annoying when they are broadcast only? When I get sent a message, I want to be able to reply to it. E-mail should be a conversation, not a fire-and-forget distribution mechanism. So what happens when you reply to a ServiceNow e-mail? It gets categorized, and then processed according to the settings in Inbound Email Actions. Lots of information is available on the wiki: http://wiki.servicenow.com/?title=Inbound_Email_Actions. Determining what an inbound e-mail is Every two minutes, the platform runs the POP Reader scheduled job. It connects to the e-mail account specified in the properties and pulls them all into the Email table, setting the Mailbox to be Inbox. Despite the name, the POP Reader job also supports IMAP accounts. This fires an event called email.read, which in turn starts the classification of the e-mail. It uses a series of logic decisions to determine how it should respond. The concept is that an inbound e-mail can be a reply to something that the platform has already sent out, is an e-mail that someone forwarded, or is part of an e-mail chain that the platform has not seen before; that is, it is a new e-mail. Each of these are handled differently, with different assumptions. As the first step in processing the e-mail, the platform attempts to find the sender in the User table. It takes the address that the e-mail was sent from as the key to search for. If it cannot find a User, it either creates a new User record (if the property is set), or uses the Guest account. Should this e-mail be processed at all? If either of the following conditions match, then the e-mail has the Mailbox set to skipped and no further processing takes place:     Does the subject line start with recognized text such as "out of office autoreply"?     Is the User account locked out? Is this a forward? Both of the following conditions must match, else the e-mail will be checked as a reply:     Does the subject line start with a recognized prefix (such as FW)?     Does the string "From" appear anywhere in the body? Is this a reply? One of the following conditions must match, else the e-mail will be processed as new:     Is there a valid, appropriate watermark that matches an existing record?     Is there an In-Reply-To header in the e-mail that references an e-mail sent by the instance?     Does the subject line start with a recognized prefix (such as RE) and contain a number prefix (such as MAI000100)? If none of these are affirmative, the e-mail is treated as a new e-mail. The prefixes and recognized text are controlled with properties available under System Properties > Email. This order of processing and logic cannot be changed. It is hardcoded into the platform. However, clever manipulation of the properties and prefixes allows great control over what will happen. One common request is to treat forwarded e-mails just like replies. To accomplish this, a nonsensical string should be added into the forward_subject_prefix, and the standard values added to the reply_subject prefix. property. For example, the following values could be used: Forward prefix: xxxxxxxxxxx Reply prefix: re:, aw:, r:, fw:, fwd:… This will ensure that a match with the forwarding prefixes is very unlikely, while the reply logic checks will be met. Creating Inbound Email Actions Once an e-mail has been categorized, it will run through the appropriate Inbound Email Action. The main purpose of an Inbound Email Action is to run JavaScript code that manipulates a target record in some way. The target record depends upon what the e-mail has been classified as: A forwarded or new e-mail will create a new record A reply will update an existing record Every Inbound Email Action is associated with a table and a condition, just like Business Rules. Since a reply must be associated with an existing record (usually found using the watermark), the platform will only look for Inbound Email Actions that are against the same table. The platform initializes the GlideRecord object current as the existing record. An e-mail classified as Reply must have an associated record, found via the watermark, the In-Reply-To header, or by running a search for a prefix stored in the sys_number table, or else it will not proceed. Forwarded and new e-mails will create new records. They will use the first Inbound Email Action that meets the condition, regardless of the table. It will then initialize a new GlideRecord object called current, expecting it to be inserted into the table. Accessing the e-mail information In order to make the scripting easier, the platform parses the e-mail and populates the properties of an object called email. Some of the more helpful properties are listed here: email.to is a comma-separated list of e-mail addresses that the e-mail was sent to and was CC'ed to. email.body_text contains the full text of the e-mail, but does not include the previous entries in the e-mail message chain. This behavior is controlled by a property. For example, anything that appears underneath two empty lines plus -----Original Message----- is ignored. email.subject is the subject line of the e-mail. email.from contains the e-mail address of the User record that the platform thinks sent the e-mail. email.origemail uses the e-mail headers to get the e-mail address of the original sender. email.body contains the body of the e-mail, separated into name:value pairs. For instance, if a line of the body was hello:world, it would be equivalent to email.body.hello = 'world'. Approving e-mails using Inbound Email Actions The previous section looked at how the platform can generate mailto links, ready for a user to select. They generate an e-mail that has the word approve or reject in the subject line and watermark in the body. This is a great example of how e-mail can be used to automate steps in ServiceNow. Approving via e-mail is often much quicker than logging in to the instance, especially if you are working remotely and are on the road. It means approvals happen faster, which in turn provides better service to the requesters and reduces the effort for our approvers. Win win! The Update Approval Request Inbound Email Action uses the information in the inbound e-mail to update the Approval record appropriately. Navigate to System Policy > Email > Inbound Actions to see what it does. We'll inspect a few lines of the code to get a feel for what is possible when automating actions with incoming e-mails. Understanding the code in Update Approval Request One of the first steps within the function, validUser, performs a check to ensure the sender is allowed to update this Approval. They must either be a delegate or the user themselves. Some companies prefer to use an e-Signature method to perform approval, where a password must be entered. This check is not up to that level, but does go some way to helping. E-mail addresses (and From strings) can be spoofed in an e-mail client. Assuming the validation is passed, the Comments field of the Approval record is updated with the body of the e-mail. current.comments = "reply from: " + email.from + "nn" + email.body_text; In order to set the State field, and thus make the decision on the Approval request, the script simply runs a search for the existence of approve or reject within the subject line of the e-mail using the standard indexOf string function. If it is found, the state is set. if (email.subject.indexOf("approve") >= 0) current.state = "approved"; if (email.subject.indexOf("reject") >= 0) current.state = "rejected"; Once the fields have been updated, it saves the record. This triggers the standard Business Rules and will run the Workflow as though this was done in the web interface. Updating the Work notes of a Maintenance task Most often, a reply to an e-mail is to add Additional comments or Work notes to a task. Using scripting, you could differentiate between the two scenarios by seeing who has sent the e-mail: a requester would provide Additional comments and a fulfiller may give either, but it is safer to assume Work notes. Let's make a simple Inbound Email Action to process e-mails and populate the Work notes field. Navigate to System Policy > Email > Inbound Actions and click on New. Use these details: Name: Work notes for Maintenance task Target table: Maintenance [u_maintenance] Active: <ticked> Type: Reply Script: current.work_notes = "Reply from: " + email.origemail + "nn" + email.body_text; current.update(); This script is very simple: it just updates our task record after setting the Work notes field with the e-mail address of the sender and the text they sent. It is separated out with a few new lines. The platform impersonates the sender, so the Activity Log will show the update as though it was done in the web interface. Once the record has been saved, the Business Rules run as normal. This includes ServiceNow sending out e-mails. Anyone who is in the Work notes list will receive the e-mail. If Send to event creator is ticked, it means the person who sent the e-mail may receive another in return, telling them they updated the task! Having multiple incoming e-mail addresses Many customers want to have logic based upon inbound e-mail addresses. For example, sending a new e-mail to invoices@gardiner-hotels.com would create a task for the Finance team, while wifi@gardiner-hotels.com creates a ticket for the Networking group. These are easy to remember and work with, and implementing ServiceNow should not mean that this simplicity should be removed. ServiceNow provides a single e-mail account that is in the format instance@service-now.com and is not able to provide multiple or custom e-mail addresses. There are two broad options for meeting this requirement: Checking multiple accounts Redirecting e-mails Using the Email Accounts plugin While ServiceNow only provides a single e-mail address, it has the ability to pull in e-mails from multiple e-mail accounts through the Email Accounts plugin. The wiki has more information here: http://wiki.servicenow.com/?title=Email_Accounts. Once the plugin has been activated, it converts the standard account information into a new Email Account [sys_email_account] record. There can be multiple Email Accounts for a particular instance, and the POP Reader job is repurposed to check each one. Once the e-mails have been brought into ServiceNow, they are treated as normal. Since ServiceNow does not provide multiple e-mail accounts, it is the customer's responsibility to create, maintain, and configure the instance with the details, including the username and passwords. The instance will need to connect to the e-mail account, which is often hosted within the customer's datacenter. This means that firewall rules or other security methods may need to be considered. Redirecting e-mails Instead of having the instance check multiple e-mail accounts, it is often preferable to continue to work with a single e-mail address. The additional e-mail addresses can be redirected to the one that ServiceNow provides. The majority of e-mail platforms, such as Microsoft Exchange, make it possible to redirect e-mail accounts. When an e-mail is received by the e-mail system, it is resent to the ServiceNow account. This process differs from e-mail forwarding: Forwarding involves adding the FW: prefix to the subject line, altering the message body, and changing the From address. Redirection sends the message unaltered, with the original To address, to the new address. There is little indication that the message has not come directly from the original sender. Redirection is often an easier method to work with than having multiple e-mail accounts. It gives more flexibility to the customer's IT team, since they do not need to provide account details to the instance, and enables them to change the redirection details easily. If a new e-mail address has to be added or an existing one decommissioned, only the e-mail platform needs to be involved. It also reduces the configuration on the ServiceNow instance; nothing needs to change. Processing multiple e-mail address Once the e-mails have been brought into ServiceNow, the platform will need to examine who the e-mail was sent to and make some decisions. This will allow the e-mails sent to wifi@gardiner-hotels.com to be routed as tasks to the networking team. There are several methods available for achieving this: A look-up table can be created, containing a list of e-mail addresses and a matching Group reference. The Inbound Email Script would use a GlideRecord query to find the right entry and populate the Assignment group on the new task. The e-mail address could be copied over into a new field on the task. Standard routing techniques, such as Assignment Rules and Data Lookup, could be used to examine the new field and populate the Assignment group. The Inbound Email Action could contain the addresses hardcoded in the script. While this is not a scalable or maintainable solution, it may be appropriate for a simple deployment. Recording Metrics ServiceNow provides several ways to monitor the progress of a task. These are often reported and e-mailed to the stakeholders, thus providing insight into the effectiveness of processes. Metrics are a way to record information. It allows the analysis and improvement of a process by measuring statistics, based upon particular defined criteria. Most often, these are time based. One of the most common metrics is how long it takes to complete a task: from when the record was created to the moment the Active flag became false. The duration can then be averaged out and compared over time, helping to answer questions such as Are we getting quicker at completing tasks? Metrics provide a great alternative to creating lots of extra fields and Business Rules on a table. Other metrics are more complex and may involve getting more than one result per task. How long does each Assignment group take to deal with the ticket? How long does an SLA get paused for? How many times does the incident get reassigned? The difference between Metrics and SLAs At first glance, a Metric appears to be very similar to an SLA, since they both record time. However, there are some key differences between Metrics and SLAs: There is no target or aim defined in a Metric. It cannot be breached; the duration is simply recorded. A Metric cannot be paused or made to work to a schedule. There is no Workflow associated with a Metric. In general, a Metric is a more straightforward measurement, designed for collecting statistics rather than being in the forefront when processing a task. Running Metrics Every time the Task table gets updated, the metrics events Business Rule fires an event called metric.update. A Script Action named Metric Update is associated with the event and calls the appropriate Metric Definitions. If you define a metric on a non-task-based table, make sure you fire the metric.update event through a Business Rule. The Metric Definition [metric_definition] table specifies how a metric should be recorded, while the Metric Instance [metric_instance] table records the results. As ever, each Metric Definition is applied to a specific table. The Type field of a Metric Definition refers to two situations: Field value duration is associated with a field on the table. Each time the field changes value, the platform creates a new Metric Instance. The duration for which that value was present is recorded. No code is required, but if some is given, it is used as a condition. Script calculation uses JavaScript to determine what the Metric Instance contains. Scripting a Metric Definition There are several predefined variables available to a Metric Definition: current refers to the GlideRecord under examination and definition is a GlideRecord of the Metric Definition. The MetricInstance Script Include provides some helpful functions, including startDuration and endDuration, but it is really only relevant for time-based metrics. Metrics can be used to calculate many statistics (like the number of times a task is reopened), but code must be written to accomplish this. Monitoring the duration of Maintenance tasks Navigate to Metrics > Definitions and click on New. Set the following fields: Name: Maintenance states Table: Maintenance [u_maintenance] Field: State Timeline: <ticked> Once saved, test it out by changing the State field on a Maintenance record to several different values. Make sure to wait 30 seconds or so between each State change, so that the Scheduled Job has time to fire. Right-click on the Form header and choose Metrics Timeline to visualize the changes in the State field. Adding the Metrics Related List to the Maintenance form will display all the captured data. Another Related List is available on the Maintenance Definition form. Summary This article showed how to deal with all the data collected in ServiceNow. The key to this is the automated processing of information. We started with exploring events. When things happen in ServiceNow, the platform can notice and set a flag for processing later. This keeps the system responsive for the user, while ensuring all the work that needs to get done, does get done. Scheduled Jobs is the background for a variety of functions: scheduled reports, scripts, or even task generation. They run on a periodic basis, such as every day or every hour. They are often used for the automatic closure of tasks if the requester hasn't responded recently. Email Notifications are a critical part of any business application. We explored how e-mails are used to let requesters know when they've got work to do, to give requesters a useful update, or when an approver must make a decision. We even saw how approvers can make that decision using only e-mail. Every user has a great deal of control over how they receive these notifications. The Notification Preferences interface lets them add multiple devices, including mobile phones to receive text messages. The Email Client in ServiceNow gives a simple, straightforward interface to send out e-mails, but the Additional comments and Work notes fields are often better and quicker to use. Every e-mail can include the contents of fields and even the output of scripts. Every two minutes, ServiceNow checks for e-mails sent to its account. If it finds any, the e-mail is categorized into being a reply, forward, or new and runs Inbound Email Actions to update or create new records.
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04 Jun 2015
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Predicting Hospital Readmission Expense Using Cascading

Packt
04 Jun 2015
10 min read
In this article by Michael Covert, author of the book Learning Cascading, we will look at a system that allows for health care providers to create complex predictive models that can assess who is most at risk for such readmission using Cascading. (For more resources related to this topic, see here.) Overview Hospital readmission is an event that health care providers are attempting to reduce, and it is the primary target of new regulations of the Affordable Care Act, passed by the US government. A readmission is defined as any reentry to a hospital 30 days or less from a prior discharge. The financial impact of this is that US Medicare and Medicaid will either not pay or will reduce the payment made to hospitals for expenses incurred. By the end of 2014, over 2600 hospitals will incur these losses from a Medicare and Medicaid tab that is thought to exceed $24 billion annually. Hospitals are seeking to find ways to predict when a patient is susceptible to readmission so that actions can be taken to fully treat the patient before discharge. Many of them are using big data and machine learning-based predictive analytics. One such predictive engine is MedPredict from Analytics Inside, a company based in Westerville, Ohio. MedPredict is the predictive modeling component of the MedMiner suite of health care products. These products use Concurrent Cascading products to perform nightly rescoring of inpatients using a highly customizable calculation known as LACE, which stands for the following: Length of stay: This refers to the number of days a patient been in hospital Acute admissions through emergency department: This refers to whether a patient has arrived through the ER Comorbidities: A comorbidity refers to the presence of a two or more individual conditions in a patient. Each condition is designated by a diagnosis code. Diagnosis codes can also indicate complications and severity of a condition. In LACE, certain conditions are associated with the probability of readmission through statistical analysis. For instance, a diagnosis of AIDS, COPD, diabetes, and so on will each increase the probability of readmission. So, each diagnosis code is assigned points, with other points indicating "seriousness" of the condition. Diagnosis codes: These refer to the International Classification of Disease codes. Version 9 (ICD-9) and now version 10 (ICD-10) standards are available as well. Emergency visits: This refers to the number of emergency room visits the patient has made in a particular window of time. The LACE engine looks at a patient's history and computes a score that is a predictor of readmissions. In order to compute the comorbidity score, the Charlson Comorbidity Index (CCI) calculation is used. It is a statistical calculation that factors in the age and complexity of the patient's condition. Using Cascading to control predictive modeling The full data workflow to compute the probability of readmissions is as follows: Read all hospital records and reformat them into patient records, diagnosis records, and discharge records. Read all data related to patient diagnosis and diagnosis records, that is, ICD-9/10, date of diagnosis, complications, and so on. Read all tracked diagnosis records and join them with patient data to produce a diagnosis (comorbidity) score by summing up comorbidity "points". Read all data related to patient admissions, that is, records associated with admission and discharge, length of stay, hospital, admittance location, stay type, and so on. Read patient profile record, that is, age, race, gender, ethnicity, eye color, body mass indicator, and so on. Compute all intermediate scores for age, emergency visits, and comorbidities. Calculate the LACE score (refer to Figure 2). Assign a date and time to it. Take all the patient information, as mentioned in the preceding points, and run it through MedPredict to produce these variety of metrics: Expected length of stay Expected expense Expected outcome Probability of readmission Figure 1 – The data workflow The Cascading LACE engine The calculational aspects of computing LACE scores makes it ideal for Cascading as a series of reusable subassemblies. Firstly, the extraction, transformation, and loading (ETL) of patient data is complex and costly. Secondly, the calculations are data-intensive. The CCI alone has to examine a patient's medical history and must find all matching diagnosis codes (such as ICD-9 or ICD-10) to assign a score. This score must be augmented by the patient's age, and lastly, a patient's inpatient discharge records must be examined for admittance to the ER as well as emergency room visits. Also, many hospitals desire to customize these calculations. The LACE engine supports and facilitates this since scores are adjustable at the diagnosis code level, and MedPredict automatically produces metrics about how significant an individual feature is to the resulting score. Medical data is quite complex too. For instance, the particular diagnosis codes that represent cancer are many, and their meanings are quite nuanced. In some cases, metastasis (spreading of cancer to other locations in the body) may have occurred, and this is treated as a more severe situation. In other situations, measured values may be "bucketed", so this implies that we track the number of emergency room visits over 1 year, 6 months, 90 days, and 30 days. The Cascading LACE engine performs these calculations easily. It is customized through a set of hospital supplied parameters, and it has the capability to perform full calculations nightly due to its usage of Hadoop. Using this capability, a patient's record can track the full history of the LACE index over time. Additionally, different sets of LACE indices can be computed simultaneously, maybe one used for diabetes, the other for Chronic Obstructive Pulmonary Disorder (COPD), and so on. Figure 2 – The LACE subassembly MedPredict tracking The Lace engine metrics feed into MedPredict along with many other variables cited previously. These records are rescored nightly and the patient history is updated. This patient history is then used to analyze trends and generate alerts when the patient is showing an increased likelihood of variance to the desired metric values. What Cascading does for us We chose Cascading to help reduce the complexity of our development efforts. MapReduce provided us with the scalability that we desired, but we found that we were developing massive amounts of code to do so. Reusability was difficult, and the Java code library was becoming large. By shifting to Cascading, we found that we could encapsulate our code better and achieve significantly greater reusability. Additionally, we reduced complexity as well. The Cascading API provides simplification and understandability, which accelerates our development velocity metrics and also reduces bugs and maintenance cycles. We allow Cascading to control the end-to-end workflow of these nightly calculations. It handles preprocessing and formatting of data. Then, it handles running these calculations in parallel, allowing high speed hash joins to be performed, and also for each leg of the calculation to be split into a parallel pipe. Next, all these calculations are merged and the final score is produced. The last step is to analyze the patient trends and generate alerts where potential problems are likely to occur. Cascading has allowed us to produce a reusable assembly that is highly parameterized, thereby allowing hospitals to customize their usage. Not only can thresholds, scores, and bucket sizes be varied, but if it's desired, additional information could be included for things, such as medical procedures performed on the patient. The local mode of Cascading allows for easy testing, and it also provides a scaled down version that can be run against a small number of patients. However, by using Cascading in the Hadoop mode, massive scalability can be achieved against very large patient populations and ICD-9/10 code sets. Concurrent also provides an excellent framework for predictive modeling using machine learning through its Pattern component. MedPredict uses this to integrate its predictive engine, which is written using Cascading, MapReduce, and Mahout. Pattern provides an interface for the integration of other external analysis products through the exchange of Predictive Model Markup Language (PMML), an XML dialect that allows many of the MedPredict proprietary machine learning algorithms to be directly incorporated into the full Cascading LACE workflow. MedPredict then produces a variety of predictive metrics in a single pass of the data. The LACE scores (current and historical trends) are used as features for these predictions. Additionally, Concurrent provides a product called Driven that greatly reduces the development cycle time for such large, complex applications. Their lingual product provides seamless integration with relational databases, which is also key to enterprise integration. Results Numerous studies have now been performed using LACE risk estimates. Many hospitals have shown the ability to reduce readmission rates by 5-10 percent due to early intervention and specific guidance given to a patient as a result of an elevated LACE score. Other studies are examining the efficacy of additional metrics, and of segmentation of the patients into better identifying groups, such as heart failure, cancer, diabetes, and so on. Additional effort is being put in to study the ability of modifying the values of the comorbidity scores, taking into account combinations and complications. In some cases, even more dramatic improvements have taken place using these techniques. For up-to-date information, search for LACE readmissions, which will provide current information about implementations and results. Analytics Inside LLC Analytics Inside is based in Westerville, Ohio. It was founded in 2005 and specializes in advanced analytical solutions and services. Analytics Inside produces the RelMiner family of relationship mining systems. These systems are based on machine learning, big data, graph theories, data visualizations, and Natural Language Processing (NLP). For further information, visit our website at http://www.AnalyticsInside.us, or e-mail us at info@AnalyticsInside.us. MedMiner Advanced Analytics for Health Care is an integrated software system designed to help an organization or patient care team in the following ways: Predicting the outcomes of patient cases and tracking these predictions over time Generating alerts based on patient case trends that will help direct remediation Complying better with ARRA value-based purchasing and meaningful use guidelines Providing management dashboards that can be used to set guidelines and track performance Tracking performance of drug usage, interactions, potentials for drug diversion, and pharmaceutical fraud Extracting medical information contained within text documents Designating data security is a key design point PHI can be hidden through external linkages, so data exchange is not required If PHI is required, it is kept safe through heavy encryption, virus scanning, and data isolation Using both cloud-based and on premise capabilities to meet client needs Concurrent Inc. Concurrent Inc. is the leader in big data application infrastructure, delivering products that help enterprises create, deploy, run, and manage data applications at scale. The company's flagship enterprise solution, Driven, was designed to accelerate the development and management of enterprise data applications. Concurrent is the team behind Cascading, the most widely deployed technology for data applications with more than 175,000 user downloads a month. Used by thousands of businesses, including eBay, Etsy, The Climate Corporation, and Twitter, Cascading is the defacto standard in open source application infrastructure technology. Concurrent is headquartered in San Francisco and can be found online at http://concurrentinc.com. Summary Hospital readmission is an event that health care providers are attempting to reduce, and it is a primary target of new regulation from the Affordable Care Act, passed by the US government. This article describes a system that allows for health care providers to create complex predictive models that can assess who is most at risk for such readmission using Cascading. Resources for Article: Further resources on this subject: Hadoop Monitoring and its aspects [article] Introduction to Hadoop [article] YARN and Hadoop [article]
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04 Jun 2015
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The Splunk Web Framework

Packt
04 Jun 2015
10 min read
In this article by the author, Kyle Smith, of the book, Splunk Developer's Guide, we learn about search-related and view-related modules. We will be covering the following topics: Search-related modules View-related modules (For more resources related to this topic, see here.) Search-related modules Let's talk JavaScript modules. For each module, we will review their primary purpose, their module path, the default variable used in an HTML dashboard, and the JavaScript instantiation of the module. We will also cover which attributes are required and which are optional. SearchManager The SearchManager is a primary driver of any dashboard. This module contains an entire search job, including the query, properties, and the actual dispatch of the job. Let's instantiate an object, and dissect the options from this sample code: Module Path: splunkjs/mvc/searchmanager Default Variable: SearchManager JavaScript Object instantiation    Var mySearchManager = new SearchManager({        id: "search1",        earliest_time: "-24h@h",        latest_time: "now",        preview: true,        cache: false,        search: "index=_internal | stats count by sourcetype"    }, {tokens: true, tokenNamespace: "submitted"}); The only required property is the id property. This is a reference ID that will be used to access this object from other instantiated objects later in the development of the page. It is best to name it something concise, yet descriptive with no spaces. The search property is optional, and contains the SPL query that will be dispatched from the module. Make sure to escape any quotes properly, if not, you may cause a JavaScript exception. earliest_time and latest_time are time modifiers that restrict the range of the events. At the end of the options object, notice the second object with token references. This is what automatically executes the search. Without these options, you would have to trigger the search manually. There are a few other properties shown, but you can refer to the actual documentation at the main documentation page http://docs.splunk.com/DocumentationStatic/WebFramework/1.1/compref_searchmanager.html. SearchManagers are set to autostart on page load. To prevent this, set autostart to false in the options. SavedSearchManager The SavedSearchManager is very similar in operation to the SearchManager, but works with a saved report, instead of an ad hoc query. The advantage to using a SavedSearchManager is in performance. If the report is scheduled, you can configure the SavedSearchManager to use the previously run jobs to load the data. If any other user runs the report within Splunk, the SavedSearchManager can reuse that user's results in the manager to boost performance. Let's take a look at a few sections of our code: Module Path: splunkjs/mvc/savedsearchmanager Default Variable: SavedSearchManager JavaScript Object instantiation        Var mySavedSearchManager = new SavedSearchManager({            id: "savedsearch1",        searchname: "Saved Report 1"            "dispatch.earliest_time": "-24h@h",            "dispatch.latest_time": "now",            preview: true,            cache: true        }); The only two required properties are id and searchname. Both of those must be present in order for this manager to run correctly. The other options are very similar to the SearchManager, except for the dispatch options. The SearchManager has the option "earliest_time", whereas the SavedSearchManager uses the option "dispatch.earliest_time". They both have the same restriction but are named differently. The additional options are listed in the main documentation page available at http://docs.splunk.com/DocumentationStatic/WebFramework/1.1/compref_savedsearchmanager.html. PostProcessManager The PostProcessManager does just that, post processes the results of a main search. This works in the same way as the post processing done in SimpleXML; a main search to load the event set, and a secondary search to perform an additional analysis and transformation. Using this manager has its own performance considerations as well. By loading a single job first, and then performing additional commands on those results, you avoid having concurrent searches for the same information. Your usage of CPU and RAM will be less, as you only store one copy of the results, instead of multiple. Module Path: splunkjs/mvc/postprocessmanager Default Variable: PostProcessManager JavaScript Object instantiation        Var mysecondarySearch = new PostProcessManager({            id: "after_search1",        search: "stats count by sourcetype",    managerid: "search1"        }); The property id is the only required property. The module won't do anything when instantiated with only an id property, but you can set it up to populate later. The other options are similar to the SearchManager, the major difference being that the search property in this case is appended to the search property of the manager listed in the managerid property. For example, if the manager search is search index=_internal source=*splunkd.log, and the post process manager search is stats count by host, then the entire search for the post process manager is search index=_internal source=*splunkd.log | stats count by host. The additional options are listed at the main documentation page http://docs.splunk.com/DocumentationStatic/WebFramework/1.1/compref_postprocessmanager.html. View-related modules These modules are related to the views and data visualizations that are native to Splunk. They range in use from charts that display data, to control groups, such as radio groups or dropdowns. These are also included with Splunk and are included by default in the RequireJS declaration. ChartView The ChartView displays a series of data in the formats in the list as follows. Item number one shows an example of how each different chart is described and presented. Each ChartView is instantiated in the same way, the only difference is in what searches are used with which chart. Module Path: splunkjs/mvc/chartview Default Variable: ChartView JavaScript Object instantiation        Var myBarChart = new ChartView({            id: "myBarChart",             managerid: "searchManagerId",            type: "bar",            el: $("#mybarchart")        }); The only required property is the id property. This assigns the object an id that can be later referenced as needed. The el option refers to the HTML element in the page that this view will be assigned and created within. The managerid relates to an existing search, saved search, or post process manager object. The results are passed from the manager into the chart view and displayed as indicated. Each chart view can be customized extensively using the charting.* properties. For example, charting.chart.overlayFields, when set to a comma separated list of field names, will overlay those fields over the chart of other data, making it possible to display SLA times over the top of Customer Service Metrics. The full list of configurable options can be found at the following link: http://docs.splunk.com/Documentation/Splunk/latest/Viz/ChartConfigurationReference. The different types of ChartView Now that we've introduced the ChartView module, let's look at the different types of charts that are built-in. This section has been presented in the following format: Name of Chart Short description of the chart type Type property for use in the JavaScript configuration Example chart command that can be displayed with this chart type Example image of the chart The different ChartView types we will cover in this section include the following: Area The area chart is similar to the line chart, and compares quantitative data. The graph is filled with color to show volume. This is commonly used to show statistics of data over time. An example of an area chart is as follows: timechart span=1h max(results.collection1{}.meh_clicks) as MehClicks max(results.collection1{}.visitors) as Visits Bar The bar chart is similar to the column chart, except that the x and y axes have been switched, and the bars run horizontally and not vertically. The bar chart is used to compare different categories. An example of a bar chart is as follows: stats max(results.collection1{}.visitors) as Visits max(results.collection1{}.meh_clicks) as MehClicks by results.collection1{}.title.text Column The column chart is similar to the bar chart, but the bars are displayed vertically. An example of a column chart is as follows: timechart span=1h avg(DPS) as "Difference in Products Sold" Filler gauge The filler gauge is a Splunk-provided visualization. It is intended for single values, normally as a percentage, but can be adjusted to use discrete values as well. The gauge uses different colors for different ranges of values, by default using green, yellow, and red, in that order. These colors can also be changed using the charting.* properties. One of the differences between this gauge and the other single value gauges is that it shows both the color and value close together, whereas the others do not. An example of a filler gauge chart is as follows: eval diff = results.collection1{}.meh_clicks / results.collection1{}.visitors * 100 | stats latest(diff) as D Line The line chart is similar to the area chart but does not fill the area under the line. This chart can be used to display discrete measurements over time. An example of a line chart is as follows: timechart span=1h max(results.collection1{}.meh_clicks) as MehClicks max(results.collection1{}.visitors) as Visits Marker gauge The marker gauge is a Splunk native visualization intended for use with a single value. Normally this will be a percentage of a value, but can be adjusted as needed. The gauge uses different colors for different ranges of values, by default using green, yellow, and red, in that order. These colors can also be changed using the charting.* properties. An example of a marker gauge chart is as follows: eval diff = results.collection1{}.meh_clicks / results.collection1{}.visitors * 100 | stats latest(diff) as D Pie Chart A pie chart is useful for displaying percentages. It gives you the ability to quickly see which part of the "pie" is disproportionate to the others. Actual measurements may not be relevant. An example of a pie chart is as follows: top op_action Radial gauge The radial gauge is another single value chart provided by Splunk. It is normally used to show percentages, but can be adjusted to show discrete values. The gauge uses different colors for different ranges of values, by default using green, yellow, and red, in that order. These colors can also be changed using the charting.* properties. An example of a radial gauge is as follows: eval diff = MC / V * 100 | stats latest(diff) as D Scatter The scatter plot can plot two sets of data on an x and y axis chart (Cartesian coordinates). This chart is primarily time independent, and is useful for finding correlations (but not necessarily causation) in data. An example of a scatter plot is as follows: table MehClicks Visitors Summary We covered some deeper elements of Splunk applications and visualizations. We reviewed each of the SplunkJS modules, how to instantiate them, and gave an example of each search-related modules and view-related modules. Resources for Article: Further resources on this subject: Introducing Splunk [article] Lookups [article] Loading data, creating an app, and adding dashboards and reports in Splunk [article]
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03 Jun 2015
6 min read
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Running Cucumber

Packt
03 Jun 2015
6 min read
In this article by Shankar Garg, author of the book Cucumber Cookbook, we will cover the following topics: Integrating Cucumber with Maven Running Cucumber from the Terminal Overriding options from the Terminal (For more resources related to this topic, see here.) Integrating Cucumber with Maven Maven has a lot of advantages over other build tools, such as dependency management, lots of plugins and the convenience of running integration tests. So let's also integrate our framework with Maven. Maven will allow our test cases to be run in different flavors, such as from the Terminal, integrating with Jenkins, and parallel execution. So how do we integrate with Maven? Let's find out in the next section. Getting ready I am assuming that we know the basics of Maven (the basics of Maven are out of the scope of this book). Follow the upcoming instructions to install Maven on your system and to create a sample Maven project. We need to install Maven on our system first. So, follow instructions mentioned on the following blogs: For Windows: http://www.mkyong.com/maven/how-to-install-maven-in-windows/ For Mac: http://www.mkyong.com/maven/install-maven-on-mac-osx/ We can also install the Maven Eclipse plugin by following the instructions mentioned on this blog: http://theopentutorials.com/tutorials/eclipse/installing-m2eclipse-maven-plugin-for-eclipse/. To import a Maven project into Eclipse, follow the instructions on this blog: http://www.tutorialspoint.com/maven/maven_eclispe_ide.htm. How to do it… Since it is a Maven project, we are going to change the pom.xml file to add the Cucumber dependencies. First we are going to declare some custom properties which will be used by us in managing the dependency version: <properties>    <junit.version>4.11</junit.version>    <cucumber.version>1.2.2</cucumber.version>    <selenium.version>2.45.0</selenium.version>    <maven.compiler.version>2.3.2</maven.compiler.version> </properties> Now, we are going to add the dependency for Cucumber-JVM with scope as test: <!—- Cucumber-java--> <dependency>    <groupId>info.cukes</groupId>    <artifactId>cucumber-java</artifactId>    <version>${cucumber.version}</version>    <scope>test</scope> </dependency> Now we need to add the dependency for Cucumber-JUnit with scope as test. <!-— Cucumber-JUnit --> <dependency>    <groupId>info.cukes</groupId>    <artifactId>cucumber-junit</artifactId>    <version>${cucumber.version}</version>    <scope>test</scope> </dependency> That's it! We have integrated Cucumber and Maven. How it works… By following these Steps, we have created a Maven project and added the Cucumber-Java dependency. At the moment, this project only has a pom.xml file, but this project can be used for adding different modules such as Feature files and Step Definitions. The advantage of using properties is that we are making sure that the dependency version is declared at one place in the pom.xml file. Otherwise, we declare a dependency at multiple places and may end up with a discrepancy in the dependency version. The Cucumber-Java dependency is the main dependency necessary for the different building blocks of Cucumber. The Cucumber-JUnit dependency is for Cucumber JUnit Runner, which we use in running Cucumber test cases. Running Cucumber from the Terminal Now we have integrated Cucumber with Maven, running Cucumber from the Terminal will not be a problem. Running any test framework from the Terminal has its own advantages, such as overriding the run configurations mentioned in the code. So how do we run Cucumber test cases from the Terminal? Let's find out in our next section. How to do it… Open the command prompt and cd until the project root directory. First, let's run all the Cucumber Scenarios from the command prompt. Since it's a Maven project and we have added Cucumber in test scope dependency and all features are also added in test packages, run the following command in the command prompt: mvn test This is the output:     The previous command runs everything as mentioned in the JUnit Runner class. However, if we want to override the configurations mentioned in the Runner, then we need to use following command: mvn test –DCucumber.options="<<OPTIONS>>" If you need help on these Cucumber options, then enter the following command in the command prompt and look at the output: mvn test -Dcucumber.options="--help" This is the output: How it works… mvn test runs Cucumber Features using Cucumber's JUnit Runner. The @RunWith (Cucumber.class) annotation on the RunCukesTest class tells JUnit to kick off Cucumber. The Cucumber runtime parses the command-line options to know what Feature to run, where the Glue Code lives, what plugins to use, and so on. When you use the JUnit Runner, these options are generated from the @CucumberOptions annotation on your test. Overriding Options from the Terminal When it is necessary to override the options mentioned in the JUnit Runner, then we need Dcucumber.options from the Terminal. Let's look at some of the practical examples. How to do it… If we want to run a Scenario by specifying the filesystem path, run the following command and look at the output: mvn test -Dcucumber.options= "src/test/java/com/features/sample.feature:5"   In the preceding code, "5" is the Feature file line number where a Scenario starts. If we want to run the test cases using Tags, then we run the following command and notice the output: mvn test -Dcucumber.options="--tags @sanity" The following is the output of the preceding command: If we want to generate a different report, then we can use the following command and see the JUnit report generate at the location mentioned: mvn test -Dcucumber.options= "--plugin junit:target/cucumber-junit-report.xml" How it works… When you override the options with -Dcucumber.options, you will completely override whatever options are hardcoded in your @CucumberOptions. There is one exception to this rule, and that is the --plugin option. This will not override, but instead, it will add a plugin. Summary In this article we learned that for successful implementation of any testing framework, it is mandatory that test cases can be run in multiple ways so that people with different competency levels can use it how they need to. In this article, we also covered advanced topics of running Cucumber test cases in parallel by a combination of Cucumber and Maven. Resources for Article: Further resources on this subject: Signing an application in Android using Maven [article] Apache Maven and m2eclipse [article] Understanding Maven [article]
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03 Jun 2015
25 min read
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SceneKit

Packt
03 Jun 2015
25 min read
So, this is it! Finally, we move from the 2D world to 3D. With SceneKit, we can make 3D games quite easily, especially since the syntax for SceneKit is quite similar to SpriteKit. When we say 3D games, we don't mean that you get to put on your 3D glasses to make the game. In 2D games, we mostly work in the x and y coordinates. In 3D games, we deal with all three axes x, y, and z. Additionally, in 3D games, we have different types of lights that we can use. Also, SceneKit has an inbuilt physics engine that will take care of forces such as gravity and will also aid collision detection. We can also use SpriteKit in SceneKit for GUI and buttons so that we can add scores and interactivity to the game. So, there is a lot to cover in this article. Let's get started. The topics covered in this article by Siddharth Shekar, the author of Learning iOS 8 Game Development Using Swift, are as follows: Creating a scene with SCNScene Adding objects to a scene Importing scenes from external 3D applications Adding physics to the scene Adding an enemy (For more resources related to this topic, see here.) Creating a scene with SCNScene First, we create a new SceneKit project. It is very similar to creating other projects. Only this time, make sure you select SceneKit from the Game Technology drop-down list. Don't forget to select Swift for the language field. Choose iPad as the device and click on Next to create the project in the selected directory, as shown in the following screenshot: Once the project is created, open it. Click on the GameViewController class, and delete all the contents in the viewDidLoad function, delete the handleTap function, as we will be creating a separate class, and add touch behavior. Create a new class called GameSCNScene and import the following headers. Inherit from the SCNScene class and add an init function that takes in a parameter called view of type SCNView: import Foundation import UIKit import SceneKit   class GameSCNScene: SCNScene{      let scnView: SCNView!    let _size:CGSize!    var scene: SCNScene!           required init(coder aDecoder: NSCoder) {        fatalError("init(coder:) has not been implemented")    }       init(currentview view: SCNView) {               super.init()    } } Also, create two new constants scnView and _size of type SCNView and CGSize, respectively. Also, add a variable called scene of type SCNScene. Since we are making a SceneKit game, we have to get the current view, which is the type SCNView, similar to how we got the view in SpriteKit where we typecasted the current view in SpriteKit to SKView. We create a _size constant to get the current size of the view. We then create a new variable scene of type SCNScene. SCNScene is the class used to make scenes in SceneKit, similar to how we would use SKScene to create scenes in SpriteKit. Swift would automatically ask to create the required init function, so we might as well include it in the class. Now, move to the GameViewController class and create a global variable called gameSCNScene of type GameSCNScene and assign it in the viewDidLoad function, as follows: class GameViewController: UIViewController { var gameSCNScene:GameSCNScene!    override func viewDidLoad() {      super.viewDidLoad()      let scnView = view as SCNView      gameSCNScene = GameSCNScene(currentview: scnView)    } }// UIViewController Class Great! Now we can add objects in the GameSCNScene class. It is better to move all the code to a single class so that we can keep the GameSceneController class clean. In the init function of GameSCNScene, add the following after the super.init function: scnView = view _size = scnView.bounds.size                         // retrieve the SCNView scene = SCNScene() scnView.scene = scene scnView.allowsCameraControl = true scnView.showsStatistics = true scnView.backgroundColor = UIColor.yellowColor() Here, we first assign the current view to the scnView constant. Next, we set the _size constant to the dimensions of the current view. Next we initialize the scene variable. Then, assign the scene to the scene of scnView. Next, enable allowCameraControls and showStatistics. This will enable us to control the camera and move it around to have a better look at the scene. Also, with statistics enabled, we will see the performance of the game to make sure that the FPS is maintained. The backgroundColor property of scnView enables us to set the color of the view. I have set it to yellow so that objects are easily visible in the scene, as shown in the following screenshot. With all this set we can run the scene. Well, it is not all that awesome yet. One thing to notice is that we have still not added a camera or a light, but we still see the yellow scene. This is because while we have not added anything to the scene yet, SceneKit automatically provides a default light and camera for the scene created. Adding objects to a scene Let us next add geometry to the scene. We can create some basic geometry such as spheres, boxes, cones, tori, and so on in SceneKit with ease. Let us create a sphere first and add it to the scene. Adding a sphere to the scene Create a function called addGeometryNode in the class and add the following code in it: func addGeometryNode(){      let sphereGeometry = SCNSphere(radius: 1.0)    sphereGeometry.firstMaterial?.diffuse.contents = UIColor.orangeColor()           let sphereNode = SCNNode(geometry: sphereGeometry)    sphereNode.position = SCNVector3Make(0.0, 0.0, 0.0)    scene.rootNode.addChildNode(sphereNode)       } For creating geometry, we use the SCNSphere class to create a sphere shape. We can also call SCNBox, SCNCone, SCNTorus, and so on to create box, cone, or torus shapes respectively. While creating the sphere, we have to provide the radius as a parameter, which will determine the size of the sphere. Although to place the shape, we have to attach it to a node so that we can place and add it to the scene. So, create a new constant called sphereNode of type SCNNode and pass in the sphere geometry as a parameter. For positioning the node, we have to use the SCNvector3Make property to place our object in 3D space by providing the values for x, y, and z. Finally, to add the node to the scene, we have to call scene.rootNode to add the sphereNode to scene, unlike SpriteKit where we would simply use addChild to add objects to the scene. With the sphere added, let us run the scene. Don't forget to add self.addGeometryNode() in the init function. We did add a sphere, so why are we getting a circle (shown in the following screenshot)? Well, the basic light source used by SceneKit just enables to us to see objects in the scene. If we want to see the actual sphere, we have to improve the light source of the scene. Adding light sources Let us create a new function called addLightSourceNode as follows so that we can add custom lights to our scene: func addLightSourceNode(){           let lightNode = SCNNode()    lightNode.light = SCNLight()    lightNode.light!.type = SCNLightTypeOmni    lightNode.position = SCNVector3(x: 10, y: 10, z: 10)    scene.rootNode.addChildNode(lightNode)         let ambientLightNode = SCNNode()    ambientLightNode.light = SCNLight()    ambientLightNode.light!.type = SCNLightTypeAmbient    ambientLightNode.light!.color = UIColor.darkGrayColor()    scene.rootNode.addChildNode(ambientLightNode) } We can add some light sources to see some depth in our sphere object. Here we add two types of light source. The first is an omni light. Omni lights start at a point and then the light is scattered equally in all directions. We also add an ambient light source. An ambient light is the light that is reflected by other objects, such as moonlight. There are two more types of light sources: directional and spotlight. Spotlight is easy to understand, and we usually use it if a certain object needs to be brought to attention like a singer on a stage. Directional lights are used if you want light to go in a single direction, such as sunlight. The Sun is so far from the Earth that the light rays are almost parallel to each other when we see them. For creating a light source, we create a node called lightNode of type SCNNode. We then assign SCNLight to the light property of lightNode. We assign the omni light type to be the type of the light. We assign position of the light source to be at 10 in all three x, y, and z coordinates. Then, we add it to the rootnode of the scene. Next we add an ambient light to the scene. The first two steps of the process are the same as for creating any light source: For the type of light we have to assign SCNLightTypeAmbient to assign an ambient type light source. Since we don't want the light source to be very strong, as it is reflected, we assign a darkGrayColor to the color. Finally, we add the light source to the scene. There is no need to add the ambient light source to the scene but it will make the scene have softer shadows. You can remove the ambient light source to see the difference. Call the addLightSourceNode function in the init function. Now, build and run the scene to see an actual sphere with proper lighting, as shown in the following screenshot: You can place a finger on the screen and move it to rotate the cameras as we have enabled camera control. You can use two fingers to pan the camera and you can double tap to reset the camera to its original position and direction. Adding a camera to the scene Next let us add a camera to the scene, as the default camera is very close. Create a new function called addCameraNode to the class and add the following code in it: func addCameraNode(){        let cameraNode = SCNNode()    cameraNode.camera = SCNCamera()    cameraNode.position = SCNVector3(x: 0, y: 0, z: 15)    scene.rootNode.addChildNode(cameraNode)       } Here, again we create an empty node called cameraNode. We assign SCNCamera to the camera property of cameraNode. Next we position the camera such that we keep the x and y values at zero and move the camera back in the z direction by 15 units. Then we add the camera to the rootnode of the scene. Call the addCameraNode at the bottom of the init function. In this scene, the origin is at the center of the scene, unlike SpriteKit where the origin of a scene is always at bottom right of the scene. Here the positive x and y are to the right and up from the center. The positive z direction is toward you. We didn't move the sphere back or reduce its size here. This is purely because we brought the camera backward in the scene. Let us next create a floor so that we can have a better understanding of the depth in the scene. Also, in this way, we will learn how to create floors in the scene. Adding a floor In the class, add a new function called addFloorNode and add the following code: func addFloorNode(){                   var floorNode = SCNNode()      floorNode.geometry = SCNFloor()      floorNode.position.y = -1.0      scene.rootNode.addChildNode(floorNode) } For creating a floor, we create a variable called floorNode of type SCNNode. We then assign SCNFloor to the geometry property of floorNode. For the position, we assign the y value to -1 as we want the sphere to appear above the floor. At the end, as usual, we assign the floorNode to the root node of the scene. In the following screenshot, I have rotated the camera to show the scene in full action. Here we can see the floor is gray in color and the sphere is casting its reflection on the floor, and we can also see the bright omni light at the top left of the sphere. Importing scenes from external 3D applications Although we can add objects, cameras, and lights through code, it will become very tedious and confusing when we have a lot of objects added to the scene. In SceneKit, this problem can be easily overcome by importing scenes prebuilt in other 3D applications. All 3D applications such as 3D StudioMax, Maya, Cheetah 3D, and Blender have the ability to export scenes in Collada (.dae) and Alembic (.abc) format. We can import these scenes with lighting, camera, and textured objects into SceneKit directly, without the need for setting up the scene. In this section, we will import a Collada file into the scene. Drag this file into the current project. Along with the DAE file, also add the monster.png file to the project, otherwise you will see only the untextured monster mesh in the scene. Click on the monsterScene.DAE file. If the textured monster is not automatically loaded, drag the monster.png file from the project into the monster mesh in the preview window. Release the mouse button once you see a (+) sign while over the monster mesh. Now you will be able to see the monster properly textured. The panel on the left shows the entities in the scene. Below the entities, the scene graph is shown and the view on the right is the preview pane. Entities show all the objects in the scene and the scene graph shows the relation between these entities. If you have certain objects that are children to other objects, the scene graph will show them as a tree. For example, if you open the triangle next to CATRigHub001, you will see all the child objects under it. You can use the scene graph to move and rotate objects in the scene to fine-tune your scene. You can also add nodes, which can be accessed by code. You can see that we already have a camera and a spotlight in the scene. You can select each object and move it around using the arrow at the pivot point of the object. You can also rotate the scene to get a better view by clicking and dragging the left mouse button on the preview scene. For zooming, scroll your mouse wheel up and down. To pan, hold the Alt button on the keyboard and left-click and drag on the preview pane. One thing to note is that rotating, zooming, and panning in the preview pane won't actually move your camera. The camera is still at the same position and angle. To view from the camera, again select the Camera001 option from the drop-down list in the preview pane and the view will reset to the camera view. At the bottom of the preview window, we can either choose to see the view through the camera or spotlight, or click-and-drag to rotate the free camera. If you have more than one camera in your scene, then you will have Camera002, Camera003, and so on in the drop-down list. Below the view selection dropdown in the preview panel you also have a play button. If you click on the play button, you can look at the default animation of the monster getting played in the preview window. The preview panel is just that; it is just to aid you in having a better understanding of the objects in the scene. In no way is it a replacement for a regular 3D package such as 3DSMax, Maya, or Blender. You can create cameras, lights, and empty nodes in the scene graph, but you can't add geometry such as boxes and spheres. You can add an empty node and position it in the scene graph, and then add geometry in code and attach it to the node. Now that we have an understanding of the scene graph, let us see how we can run this scene in SceneKit. In the init function, delete the line where we initialized the scene and add the following line instead. Also delete the objects, light, and camera we added earlier. init(currentview view:SCNView){    super.init()    scnView = view    _size = scnView.bounds.size       //retrieve the SCNView    //scene = SCNScene()    scene = SCNScene(named: "monsterScene.DAE")       scnView.scene = scene    scnView.allowsCameraControl = true    scnView.showsStatistics = true    scnView.backgroundColor = UIColor.yellowColor()    //   self.addGeometryNode() //   self.addLightSourceNode() //   self.addCameraNode() //   self.addFloorNode() //   } Build and run the game to see the following screenshot: You will see the monster running and the yellow background that we initially assigned to the scene. While exporting the scene, if you export the animations as well, once the scene loads in SceneKit the animation starts playing automatically. Also, you will notice that we have deleted the camera and light in the scene. So, how come the default camera and the light aren't loaded in the scene? What is happening here is that while I exported the file, I inserted a camera in the scene and also added a spotlight. So, when we imported the file into the scene, SceneKit automatically understood that there is a camera already present, so it will use the camera as its default camera. Similarly, a spotlight is already added in the scene, which is taken as the default light source, and lighting is calculated accordingly. Adding objects and physics to the scene Let us now see how we can access each of the objects in the scene graph and add gravity to the monster. Accessing the hero object and adding a physics body So, create a new function called addColladaObjects and call an addHero function in it. Create a global variable called heroNode of type SCNNode. We will use this node to access the hero object in the scene. In the addHero function, add the following code: init(currentview view:SCNView){    super.init()    scnView = view    _size = scnView.bounds.size       //retrieve the SCNView    //scene = SCNScene()    scene = SCNScene(named: "monster.scnassets/monsterScene.DAE")       scnView.scene = scene    scnView.allowsCameraControl = true    scnView.showsStatistics = true    scnView.backgroundColor = UIColor.yellowColor()       self.addColladaObjects()    //   self.addGeometryNode() //   self.addLightSourceNode() //   self.addCameraNode() //   self.addFloorNode()    }   func addHero(){      heroNode = SCNNode()        var monsterNode = scene.rootNode.childNodeWithName( "CATRigHub001", recursively: false)    heroNode.addChildNode(monsterNode!) heroNode.position = SCNVector3Make(0, 0, 0)                     let collisionBox = SCNBox(width: 10.0, height: 10.0,            length: 10.0, chamferRadius: 0)      heroNode.physicsBody?.physicsShape = SCNPhysicsShape(geometry: collisionBox, options: nil)    heroNode.physicsBody = SCNPhysicsBody.dynamicBody()      heroNode.physicsBody?.mass = 20    heroNode.physicsBody?.angularVelocityFactor = SCNVector3Zero heroNode.name = "hero"           scene.rootNode.addChildNode(heroNode) } First, we call the addColladaObjects function in the init function, as highlighted. Then we create the addHero function. In it we initiate the heroNode. Then, to actually move the monster, we need access to the CatRibHub001 node to move the monster. We gain access to it through the ChildWithName property of scene.rootNode. For each object that we wish to gain access to through code, we will have to use the ChildWithName property of the rootNode of the scene and pass in the name of the object. If recursively is set to true, to get said object, SceneKit will go through all the child nodes to get access to the specific node. Since the node that we are looking for is right on top, we said false to save processing time. We create a temporary variable called monsterNode. In the next step, we add the monsterNode variable to heroNode. We then set the position of the hero node to the origin. For heroNode to interact with other physics bodies in the scene, we have to assign a shape to the physics body of heroNode. We could use the mesh of the monster, but the shape might not be calculated properly and a box is a much simpler shape than the mesh of the monster. For creating a box collider, we create a new box geometry roughly the width, height, and depth of the monster. Then, using the physicsBody.physicsShape property of the heroNode, we assign the shape of the collisionBox we created for it. Since we want the body to be affected by gravity, we assign the physics body type to be dynamic. Later we will see other body types. Since we want the body to be highly responsive to gravity, we assign a value of 20 to the mass of the body. In the next step, we set the angularVelocityFactor to 0 in all three directions, as we want the body to move straight up and down when a vertical force is applied. If we don't do this, the body will flip-flop around. We also assign the name hero to the monster to check if the collided object is the hero or not. This will come in handy when we check for collision with other objects. Finally, we add heroNode to the scene. Add the addColladaObjects to the init function and comment or delete the self.addGeometryNode, self.addLightSourceNode, self.addCameraNode, and self.addFloorNode functions if you haven't already. Then, run the game to see the monster slowly falling through. We will create a small patch of ground right underneath the monster so that it doesn't fall down. Adding ground Create a new function called addGround and add the following: func addGround(){           let groundBox = SCNBox(width: 10, height: 2,                            length: 10, chamferRadius: 0)      let groundNode = SCNNode(geometry: groundBox)           groundNode.position = SCNVector3Make(0, -1.01, 0)    groundNode.physicsBody = SCNPhysicsBody.staticBody()    groundNode.physicsBody?.restitution = 0.0      scene.rootNode.addChildNode(groundNode) } We create a new constant called groundBox of type SCNBox, with a width and length of 10, and height of 2. Chamfer is the rounding of the edges of the box. Since we didn't want any rounding of the corners, it is set to 0. Next we create a SCNNode called groundNode and assign groundBox to it. We place it slightly below the origin. Since the height of the box is 2, we place it at –1.01 so that heroNode will be (0, 0, 0) when the monster rests on the ground. Next we assign the physics body of type static body. Also, since we don't want the hero to bounce off the ground when he falls on it, we set the restitution to 0. Finally, we then add the ground to the scene's rootnode. The reason we made this body static instead of dynamic is because a dynamic body gets affected by gravity and other forces but a static one doesn't. So, in this scene, even though gravity is acting downward, the hero will fall but groundBox won't as it is a static body. You will see that the physics syntax is very similar to SpriteKit with static bodies and dynamic bodies, gravity, and so on. And once again, similar to SpriteKit, the physics simulation is automatically turned on when we run the scene. Add the addGround function in the addColladaObjects functions and run the game to see the monster getting affected by gravity and stopping after coming in touch with the ground. Adding an enemy node To check collision in SceneKit, we can check for collision between the hero and the ground. But let us make it a little more interesting and also learn a new kind of body type: the kinematic body. For this, we will create a new box called enemy and make it move and collide with the hero. Create a new global SCNNode called enemyNode as follows: let scnView: SCNView! let _size:CGSize! var scene: SCNScene! var heroNode:SCNNode! var enemyNode:SCNNode! Also, create a new function called addEnemy to the class and add the following in it: func addEnemy(){           let geo = SCNBox(width: 4.0, height: 4.0, length: 4.0, chamferRadius: 0.0)           geo.firstMaterial?.diffuse.contents = UIColor.yellowColor()           enemyNode = SCNNode(geometry: geo)    enemyNode.position = SCNVector3Make(0, 20.0 , 60.0)    enemyNode.physicsBody = SCNPhysicsBody.kinematicBody()    scene.rootNode.addChildNode(enemyNode)           enemyNode.name = "enemy" } Nothing too fancy here! Just as when adding the groundNode, we have created a cube with all its sides four units long. We have also added a yellow color to its material. We then initialize enemyNode in the function. We position the node along the x, y, and z axes. Assign the body type as kinematic instead of static or dynamic. Then we add the body to the scene and finally name the enemyNode as enemy, which we will be needing while checking for collision. Before we forget, call the addEnemy function in the addColladaObjects function after where we called the addHero function. The difference between the kinematic body and other body types is that, like static, external forces cannot act on the body, but we can apply a force to a kinematic body to move it. In the case of a static body, we saw that it is not affected by gravity and even if we apply a force to it, the body just won't move. Here we won't be applying any force to move the enemy block but will simply move the object like we moved the enemy in the SpriteKit game. So, it is like making the same game, but in 3D instead of 2D, so that you can see that although we have a third dimension, the same principles of game development can be applied to both. For moving the enemy, we need an update function for the enemy. So, let us add it to the scene by creating an updateEnemy function and adding the following to it: func updateEnemy(){         enemyNode.position.z += -0.9             if((enemyNode.position.z - 5.0) < -40){                   var factor = arc4random_uniform(2) + 1                   if( factor == 1 ){            enemyNode.position = SCNVector3Make(0, 2.0 , 60.0)        }else{            enemyNode.position = SCNVector3Make(0, 15.0 , 60.0)        }    } } In the update function, similar to how we moved the enemy in the SpriteKit game, we increment the Z position of the enemy node by 0.9. The difference being that we are moving the z direction. Once the enemy has gone beyond –40 in the z direction, we reset the position of the enemy. To create an additional challenge to the player, when the enemy resets, a random number is chosen between 1 and 2. If it is 1, then the enemy is placed closer to the ground, otherwise it is placed at 15 units from the ground. Later, we will add a jump mechanic to the hero. So, when the enemy is closer to the ground, the hero has to jump over the enemy box, but when the enemy is spawned at a height, the hero shouldn't jump. If he jumps and hits the enemy box, then it is game over. Later we will also add a scoring mechanism to keep score. For updating the enemy, we actually need an update function to add the enemyUpdate function to so that the enemy moves and his position resets. So, create a function called update in the class and call the updateEnemy function in it as follows:    func update(){           updateEnemy()    } Summary This article has given insight on how to create a scene with SCNScene, how to add objects to a scene, how to import scenes from external 3D applications, how to adding physics to the scene, and how to add an enemy. Resources for Article: Further resources on this subject: Creating a Brick Breaking Game [article] iOS Security Overview [article] Code Sharing Between iOS and Android [article]
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03 Jun 2015
11 min read
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Reactive Data Streams

Packt
03 Jun 2015
11 min read
In this article by Shiti Saxena, author of the book Mastering Play Framework for Scala, we will discuss the Iteratee approach used to handle such situations. This article also covers the basics of handling data streams with a brief explanation of the following topics: Iteratees Enumerators Enumeratees (For more resources related to this topic, see here.) Iteratee Iteratee is defined as a trait, Iteratee[E, +A], where E is the input type and A is the result type. The state of an Iteratee is represented by an instance of Step, which is defined as follows: sealed trait Step[E, +A] {def it: Iteratee[E, A] = this match {case Step.Done(a, e) => Done(a, e)case Step.Cont(k) => Cont(k)case Step.Error(msg, e) => Error(msg, e)}}object Step {//done state of an iterateecase class Done[+A, E](a: A, remaining: Input[E]) extends Step[E, A]//continuing state of an iteratee.case class Cont[E, +A](k: Input[E] => Iteratee[E, A]) extendsStep[E, A]//error state of an iterateecase class Error[E](msg: String, input: Input[E]) extends Step[E,Nothing]} The input used here represents an element of the data stream, which can be empty, an element, or an end of file indicator. Therefore, Input is defined as follows: sealed trait Input[+E] {def map[U](f: (E => U)): Input[U] = this match {case Input.El(e) => Input.El(f(e))case Input.Empty => Input.Emptycase Input.EOF => Input.EOF}}object Input {//An input elementcase class El[+E](e: E) extends Input[E]// An empty inputcase object Empty extends Input[Nothing]// An end of file inputcase object EOF extends Input[Nothing]} An Iteratee is an immutable data type and each result of processing an input is a new Iteratee with a new state. To handle the possible states of an Iteratee, there is a predefined helper object for each state. They are: Cont Done Error Let's see the definition of the readLine method, which utilizes these objects: def readLine(line: List[Array[Byte]] = Nil): Iteratee[Array[Byte],String] = Cont {case Input.El(data) => {val s = data.takeWhile(_ != 'n')if (s.length == data.length) {readLine(s :: line)} else {Done(new String(Array.concat((s :: line).reverse: _*),"UTF-8").trim(), elOrEmpty(data.drop(s.length + 1)))}}case Input.EOF => {Error("EOF found while reading line", Input.Empty)}case Input.Empty => readLine(line)} The readLine method is responsible for reading a line and returning an Iteratee. As long as there are more bytes to be read, the readLine method is called recursively. On completing the process, an Iteratee with a completed state (Done) is returned, else an Iteratee with state continuous (Cont) is returned. In case the method encounters EOF, an Iteratee with state Error is returned. In addition to these, Play Framework exposes a companion Iteratee object, which has helper methods to deal with Iteratees. The API exposed through the Iteratee object is documented at https://www.playframework.com/documentation/2.3.x/api/scala/index.html#play.api.libs.iteratee.Iteratee$. The Iteratee object is also used internally within the framework to provide some key features. For example, consider the request body parsers. The apply method of the BodyParser object is defined as follows: def apply[T](debugName: String)(f: RequestHeader =>Iteratee[Array[Byte], Either[Result, T]]): BodyParser[T] = newBodyParser[T] {def apply(rh: RequestHeader) = f(rh)override def toString = "BodyParser(" + debugName + ")"} So, to define BodyParser[T], we need to define a method that accepts RequestHeader and returns an Iteratee whose input is an Array[Byte] and results in Either[Result,T]. Let's look at some of the existing implementations to understand how this works. The RawBuffer parser is defined as follows: def raw(memoryThreshold: Int): BodyParser[RawBuffer] =BodyParser("raw, memoryThreshold=" + memoryThreshold) { request =>import play.core.Execution.Implicits.internalContextval buffer = RawBuffer(memoryThreshold)Iteratee.foreach[Array[Byte]](bytes => buffer.push(bytes)).map {_ =>buffer.close()Right(buffer)}} The RawBuffer parser uses Iteratee.forEach method and pushes the input received into a buffer. The file parser is defined as follows: def file(to: File): BodyParser[File] = BodyParser("file, to=" +to) { request =>import play.core.Execution.Implicits.internalContextIteratee.fold[Array[Byte], FileOutputStream](newFileOutputStream(to)) {(os, data) =>os.write(data)os}.map { os =>os.close()Right(to)}} The file parser uses the Iteratee.fold method to create FileOutputStream of the incoming data. Now, let's see the implementation of Enumerator and how these two pieces fit together. Enumerator Similar to the Iteratee, an Enumerator is also defined through a trait and backed by an object of the same name: trait Enumerator[E] {parent =>def apply[A](i: Iteratee[E, A]): Future[Iteratee[E, A]]...}object Enumerator{def apply[E](in: E*): Enumerator[E] = in.length match {case 0 => Enumerator.emptycase 1 => new Enumerator[E] {def apply[A](i: Iteratee[E, A]): Future[Iteratee[E, A]] =i.pureFoldNoEC {case Step.Cont(k) => k(Input.El(in.head))case _ => i}}case _ => new Enumerator[E] {def apply[A](i: Iteratee[E, A]): Future[Iteratee[E, A]] =enumerateSeq(in, i)}}...} Observe that the apply method of the trait and its companion object are different. The apply method of the trait accepts Iteratee[E, A] and returns Future[Iteratee[E, A]], while that of the companion object accepts a sequence of type E and returns an Enumerator[E]. Now, let's define a simple data flow using the companion object's apply method; first, get the character count in a given (Seq[String]) line: val line: String = "What we need is not the will to believe, butthe wish to find out."val words: Seq[String] = line.split(" ")val src: Enumerator[String] = Enumerator(words: _*)val sink: Iteratee[String, Int] = Iteratee.fold[String,Int](0)((x, y) => x + y.length)val flow: Future[Iteratee[String, Int]] = src(sink)val result: Future[Int] = flow.flatMap(_.run) The variable result has the Future[Int] type. We can now process this to get the actual count. In the preceding code snippet, we got the result by following these steps: Building an Enumerator using the companion object's apply method: val src: Enumerator[String] = Enumerator(words: _*) Getting Future[Iteratee[String, Int]] by binding the Enumerator to an Iteratee: val flow: Future[Iteratee[String, Int]] = src(sink) Flattening Future[Iteratee[String,Int]] and processing it: val result: Future[Int] = flow.flatMap(_.run) Fetching the result from Future[Int]: Thankfully, Play provides a shortcut method by merging steps 2 and 3 so that we don't have to repeat the same process every time. The method is represented by the |>>> symbol. Using the shortcut method, our code is reduced to this: val src: Enumerator[String] = Enumerator(words: _*)val sink: Iteratee[String, Int] = Iteratee.fold[String, Int](0)((x, y)=> x + y.length)val result: Future[Int] = src |>>> sink Why use this when we can simply use the methods of the data type? In this case, do we use the length method of String to get the same value (by ignoring whitespaces)? In this example, we are getting the data as a single String but this will not be the only scenario. We need ways to process continuous data, such as a file upload, or feed data from various networking sites, and so on. For example, suppose our application receives heartbeats at a fixed interval from all the devices (such as cameras, thermometers, and so on) connected to it. We can simulate a data stream using the Enumerator.generateM method: val dataStream: Enumerator[String] = Enumerator.generateM {Promise.timeout(Some("alive"), 100 millis)} In the preceding snippet, the "alive" String is produced every 100 milliseconds. The function passed to the generateM method is called whenever the Iteratee bound to the Enumerator is in the Cont state. This method is used internally to build enumerators and can come in handy when we want to analyze the processing for an expected data stream. An Enumerator can be created from a file, InputStream, or OutputStream. Enumerators can be concatenated or interleaved. The Enumerator API is documented at https://www.playframework.com/documentation/2.3.x/api/scala/index.html#play.api.libs.iteratee.Enumerator$. Using the Concurrent object The Concurrent object is a helper that provides utilities for using Iteratees, enumerators, and Enumeratees concurrently. Two of its important methods are: Unicast: It is useful when sending data to a single iterate. Broadcast: It facilitates sending the same data to multiple Iteratees concurrently. Unicast For example, the character count example in the previous section can be implemented as follows: val unicastSrc = Concurrent.unicast[String](channel =>channel.push(line))val unicastResult: Future[Int] = unicastSrc |>>> sink The unicast method accepts the onStart, onError, and onComplete handlers. In the preceding code snippet, we have provided the onStart method, which is mandatory. The signature of unicast is this: def unicast[E](onStart: (Channel[E]) ⇒ Unit,onComplete: ⇒ Unit = (),onError: (String, Input[E]) ⇒ Unit = (_: String, _: Input[E])=> ())(implicit ec: ExecutionContext): Enumerator[E] {…} So, to add a log for errors, we can define the onError handler as follows: val unicastSrc2 = Concurrent.unicast[String](channel => channel.push(line),onError = { (msg, str) => Logger.error(s"encountered $msg for$str")}) Now, let's see how broadcast works. Broadcast The broadcast[E] method creates an enumerator and a channel and returns a (Enumerator[E], Channel[E]) tuple. The enumerator and channel thus obtained can be used to broadcast data to multiple Iteratees: val (broadcastSrc: Enumerator[String], channel:Concurrent.Channel[String]) = Concurrent.broadcast[String]private val vowels: Seq[Char] = Seq('a', 'e', 'i', 'o', 'u')def getVowels(str: String): String = {val result = str.filter(c => vowels.contains(c))result}def getConsonants(str: String): String = {val result = str.filterNot(c => vowels.contains(c))result}val vowelCount: Iteratee[String, Int] = Iteratee.fold[String,Int](0)((x, y) => x + getVowels(y).length)val consonantCount: Iteratee[String, Int] =Iteratee.fold[String, Int](0)((x, y) => x +getConsonants(y).length)val vowelInfo: Future[Int] = broadcastSrc |>>> vowelCountval consonantInfo: Future[Int] = broadcastSrc |>>>consonantCountwords.foreach(w => channel.push(w))channel.end()vowelInfo onSuccess { case count => println(s"vowels:$count")}consonantInfo onSuccess { case count =>println(s"consonants:$count")} Enumeratee Enumeratee is also defined using a trait and its companion object with the same Enumeratee name. It is defined as follows: trait Enumeratee[From, To] {...def applyOn[A](inner: Iteratee[To, A]): Iteratee[From,Iteratee[To, A]]def apply[A](inner: Iteratee[To, A]): Iteratee[From, Iteratee[To,A]] = applyOn[A](inner)...} An Enumeratee transforms the Iteratee given to it as input and returns a new Iteratee. Let's look at a method that defines an Enumeratee by implementing the applyOn method. An Enumeratee's flatten method accepts Future[Enumeratee] and returns an another Enumeratee, which is defined as follows: def flatten[From, To](futureOfEnumeratee:Future[Enumeratee[From, To]]) = new Enumeratee[From, To] {def applyOn[A](it: Iteratee[To, A]): Iteratee[From,Iteratee[To, A]] =Iteratee.flatten(futureOfEnumeratee.map(_.applyOn[A](it))(dec))} In the preceding snippet, applyOn is called on the Enumeratee whose future is passed and dec is defaultExecutionContext. Defining an Enumeratee using the companion object is a lot simpler. The companion object has a lot of methods to deal with enumeratees, such as map, transform, collect, take, filter, and so on. The API is documented at https://www.playframework.com/documentation/2.3.x/api/scala/index.html#play.api.libs.iteratee.Enumeratee$. Let's define an Enumeratee by working through a problem. The example we used in the previous section to find the count of vowels and consonants will not work correctly if a vowel is capitalized in a sentence, that is, the result of src |>>> vowelCount will be incorrect when the line variable is defined as follows: val line: String = "What we need is not the will to believe, but the wish to find out.".toUpperCase To fix this, let's alter the case of all the characters in the data stream to lowercase. We can use an Enumeratee to update the input provided to the Iteratee. Now, let's define an Enumeratee to return a given string in lowercase: val toSmallCase: Enumeratee[String, String] =Enumeratee.map[String] {s => s.toLowerCase} There are two ways to add an Enumeratee to the dataflow. It can be bound to the following: Enumerators Iteratees Summary In this article, we discussed the concept of Iteratees, Enumerators, and Enumeratees. We also saw how they were implemented in Play Framework and used internally. Resources for Article: Further resources on this subject: Play Framework: Data Validation Using Controllers [Article] Play Framework: Introduction to Writing Modules [Article] Integrating with other Frameworks [Article]
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Packt
03 Jun 2015
14 min read
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Let's Build with AngularJS and Bootstrap

Packt
03 Jun 2015
14 min read
In this article by Stephen Radford, author of the book Learning Web Development with Bootstrap and AngularJS, we're going to use Bootstrap and AngularJS. We'll look at building a maintainable code base as well as exploring the full potential of both frameworks. (For more resources related to this topic, see here.) Working with directives Something we've been using already without knowing it is what Angular calls directives. These are essentially powerful functions that can be called from an attribute or even its own element, and Angular is full of them. Whether we want to loop data, handle clicks, or submit forms, Angular will speed everything up for us. We first used a directive to initialize Angular on the page using ng-app, and all of the directives we're going to look at in this article are used in the same way—by adding an attribute to an element. Before we take a look at some more of the built-in directives, we need to quickly make a controller. Create a new file and call it controller.js. Save this to your js directory within your project and open it up in your editor. Controllers are just standard JS constructor functions that we can inject Angular's services such as $scope into. These functions are instantiated when Angular detects the ng-controller attribute. As such, we can have multiple instances of the same controller within our application, allowing us to reuse a lot of code. This familiar function declaration is all we need for our controller. function AppCtrl(){} To let the framework know this is the controller we want to use, we need to include this on the page after Angular is loaded and also attach the ng-controller directive to our opening <html> tag: <html ng-controller="AppCtl">…<script type="text/javascript"src="assets/js/controller.js"></script> ng-click and ng-mouseover One of the most basic things you'll have ever done with JavaScript is listened for a click event. This could have been using the onclick attribute on an element, using jQuery, or even with an event listener. In Angular, we use a directive. To demonstrate this, we'll create a button that will launch an alert box—simple stuff. First, let's add the button to our content area we created earlier: <div class="col-sm-8"><button>Click Me</button></div> If you open this up in your browser, you'll see a standard HTML button created—no surprises there. Before we attach the directive to this element, we need to create a handler in our controller. This is just a function within our controller that is attached to the scope. It's very important we attach our function to the scope or we won't be able to access it from our view at all: function AppCtl($scope){$scope.clickHandler = function(){window.alert('Clicked!');};} As we already know, we can have multiple scopes on a page and these are just objects that Angular allows the view and the controller to have access to. In order for the controller to have access, we've injected the $scope service into our controller. This service provides us with the scope Angular creates on the element we added the ng-controller attribute to. Angular relies heavily on dependency injection, which you may or may not be familiar with. As we've seen, Angular is split into modules and services. Each of these modules and services depend upon one another and dependency injection provides referential transparency. When unit testing, we can also mock objects that will be injected to confirm our test results. DI allows us to tell Angular what services our controller depends upon, and the framework will resolve these for us. An in-depth explanation of AngularJS' dependency injection can be found in the official documentation at https://docs.angularjs.org/guide/di. Okay, so our handler is set up; now we just need to add our directive to the button. Just like before, we need to add it as an additional attribute. This time, we're going to pass through the name of the function we're looking to execute, which in this case is clickHandler. Angular will evaluate anything we put within our directive as an AngularJS expression, so we need to be sure to include two parentheses indicating that this is a function we're calling: <button ng-click="clickHandler()">Click Me</button> If you load this up in your browser, you'll be presented with an alert box when you click the button. You'll also notice that we don't need to include the $scope variable when calling the function in our view. Functions and variables that can be accessed from the view live within the current scope or any ancestor scope.   Should we wish to display our alert box on hover instead of click, it's just a case of changing the name of the directive to ng-mouseover, as they both function in the exact same way. ng-init The ng-init directive is designed to evaluate an expression on the current scope and can be used on its own or in conjunction with other directives. It's executed at a higher priority than other directives to ensure the expression is evaluated in time. Here's a basic example of ng-init in action: <div ng-init="test = 'Hello, World'"></div>{{test}} This will display Hello, World onscreen when the application is loaded in your browser. Above, we've set the value of the test model and then used the double curly-brace syntax to display it. ng-show and ng-hide There will be times when you'll need to control whether an element is displayed programmatically. Both ng-show and ng-hide can be controlled by the value returned from a function or a model. We can extend upon our clickHandler function we created to demonstrate the ng-click directive to toggle the visibility of our element. We'll do this by creating a new model and toggling the value between true or false. First of all, let's create the element we're going to be showing or hiding. Pop this below your button: <div ng-hide="isHidden">Click the button above to toggle.</div> The value within the ng-hide attribute is our model. Because this is within our scope, we can easily modify it within our controller: $scope.clickHandler = function(){$scope.isHidden = !$scope.isHidden;}; Here we're just reversing the value of our model, which in turn toggles the visibility of our <div>. If you open up your browser, you'll notice that the element isn't hidden by default. There are a few ways we could tackle this. Firstly, we could set the value of $scope.hidden to true within our controller. We could also set the value of hidden to true using the ng-init directive. Alternatively, we could switch to the ng-show directive, which functions in reverse to ng-hide and will only make an element visible if a model's value is set to true. Ensure Angular is loaded within your header or ng-hide and ng-show won't function correctly. This is because Angular uses its own classes to hide elements and these need to be loaded on page render. ng-if Angular also includes an ng-if directive that works in a similar fashion to ng-show and ng-hide. However, ng-if actually removes the element from the DOM whereas ng-show and ng-hide just toggles the elements' visibility. Let's take a quick look at how we'd use ng-if with the preceding code: <div ng-if="isHidden">Click the button above to toggle.</div> If we wanted to reverse the statement's meaning, we'd simply just need to add an exclamation point before our expression: <div ng-if="!isHidden">Click the button above to toggle.</div> ng-repeat Something you'll come across very quickly when building a web app is the need to render an array of items. For example, in our contacts manager, this would be a list of contacts, but it could be anything. Angular allows us to do this with the ng-repeat directive. Here's an example of some data we may come across. It's array of objects with multiple properties within it. To display the data, we're going to need to be able to access each of the properties. Thankfully, ng-repeat allows us to do just that. Here's our controller with an array of contact objects assigned to the contacts model: function AppCtrl($scope){$scope.contacts = [{name: 'John Doe',phone: '01234567890',email: 'john@example.com'},{name: 'Karan Bromwich',phone: '09876543210',email: 'karan@email.com'}];} We have just a couple of contacts here, but as you can imagine, this could be hundreds of contacts served from an API that just wouldn't be feasible to work with without ng-repeat. First, add an array of contacts to your controller and assign it to $scope.contacts. Next, open up your index.html file and create a <ul> tag. We're going to be repeating a list item within this unordered list so this is the element we need to add our directive to: <ul><li ng-repeat="contact in contacts"></li></ul> If you're familiar with how loops work in PHP or Ruby, then you'll feel right at home here. We create a variable that we can access within the current element being looped. The variable after the in keyword references the model we created on $scope within our controller. This now gives us the ability to access any of the properties set on that object with each iteration or item repeated gaining a new scope. We can display these on the page using Angular's double curly-brace syntax. <ul><li ng-repeat="contact in contacts">{{contact.name}}</li></ul> You'll notice that this outputs the name within our list item as expected, and we can easily access any property on our contact object by referencing it using the standard dot syntax. ng-class Often there are times where you'll want to change or add a class to an element programmatically. We can use the ng-class directive to achieve this. It will let us define a class to add or remove based on the value of a model. There are a couple of ways we can utilize ng-class. In its most simple form, Angular will apply the value of the model as a CSS class to the element: <div ng-class="exampleClass"></div> Should the model referenced be undefined or false, Angular won't apply a class. This is great for single classes, but what if you want a little more control or want to apply multiple classes to a single element? Try this: <div ng-class="{className: model, class2: model2}"></div> Here, the expression is a little different. We've got a map of class names with the model we wish to check against. If the model returns true, then the class will be added to the element. Let's take a look at this in action. We'll use checkboxes with the ng-model attribute, to apply some classes to a paragraph: <p ng-class="{'text-center': center, 'text-danger': error}">Lorem ipsum dolor sit amet</p> I've added two Bootstrap classes: text-center and text-danger. These observe a couple of models, which we can quickly change with some checkboxes: The single quotations around the class names within the expression are only required when using hyphens, or an error will be thrown by Angular. <label><input type="checkbox" ng-model="center"> textcenter</label><label><input type="checkbox" ng-model="error"> textdanger</label> When these checkboxes are checked, the relevant classes will be applied to our element. ng-style In a similar way to ng-class, this directive is designed to allow us to dynamically style an element with Angular. To demonstrate this, we'll create a third checkbox that will apply some additional styles to our paragraph element. The ng-style directive uses a standard JavaScript object, with the keys being the property we wish to change (for example, color and background). This can be applied from a model or a value returned from a function. Let's take a look at hooking it up to a function that will check a model. We can then add this to our checkbox to turn the styles off and on. First, open up your controller.js file and create a new function attached to the scope. I'm calling mine styleDemo: $scope.styleDemo = function(){if(!$scope.styler){return;}return {background: 'red',fontWeight: 'bold'};}; Inside the function, we need to check the value of a model; in this example, it's called styler. If it's false, we don't need to return anything, otherwise we're returning an object with our CSS properties. You'll notice that we used fontWeight rather than font-weight in our returned object. Either is fine, and Angular will automatically switch the CamelCase over to the correct CSS property. Just remember than when using hyphens in JavaScript object keys, you'll need to wrap them in quotation marks. This model is going to be attached to a checkbox, just like we did with ng-class: <label><input type="checkbox" ng-model="styler"> ng-style</label> The last thing we need to do is add the ng-style directive to our paragraph element: <p .. ng-style="styleDemo()">Lorem ipsum dolor sit amet</p> Angular is clever enough to recall this function every time the scope changes. This means that as soon as our model's value changes from false to true, our styles will be applied and vice versa. ng-cloak The final directive we're going to look at is ng-cloak. When using Angular's templates within our HTML page, the double curly braces are temporarily displayed before AngularJS has finished loading and compiling everything on our page. To get around this, we need to temporarily hide our template before it's finished rendering. Angular allows us to do this with the ng-cloak directive. This sets an additional style on our element whilst it's being loaded: display: none !important;. To ensure there's no flashing while content is being loaded, it's important that Angular is loaded in the head section of our HTML page. Summary We've covered a lot in this article, let's recap it all. Bootstrap allowed us to quickly create a responsive navigation. We needed to include the JavaScript file included with our Bootstrap download to enable the toggle on the mobile navigation. We also looked at the powerful responsive grid system included with Bootstrap and created a simple two-column layout. While we were doing this, we learnt about the four different column class prefixes as well as nesting our grid. To adapt our layout, we discovered some of the helper classes included with the framework to allow us to float, center, and hide elements. In this article, we saw in detail Angular's built-in directives: functions Angular allows us to use from within our view. Before we could look at them, we needed to create a controller, which is just a function that we can pass Angular's services into using dependency injection. Directives such as ng-click and ng-mouseover are essentially just new ways of handling events that you will have no doubt done using either jQuery or vanilla JavaScript. However, directives such as ng-repeat will probably be a completely new way of working. It brings some logic directly within our view to loop through data and display it on the page. We also looked at directives that observe models on our scope and perform different actions based on their values. Directives like ng-show and ng-hide will show or hide an element based on a model's value. We also saw this in action in ng-class, which allowed us to add some classes to our elements based on our models' values. Resources for Article: Further resources on this subject: AngularJS Performance [Article] AngularJS Web Application Development Cookbook [Article] Role of AngularJS [Article]
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Packt
03 Jun 2015
14 min read
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Pointers and references

Packt
03 Jun 2015
14 min read
In this article by Ivo Balbaert, author of the book, Rust Essentials, we will go through the pointers and memory safety. (For more resources related to this topic, see here.) The stack and the heap When a program starts, by default a 2 MB chunk of memory called the stack is granted to it. The program will use its stack to store all its local variables and function parameters; for example, an i32 variable takes 4 bytes of the stack. When our program calls a function, a new stack frame is allocated to it. Through this mechanism, the stack knows the order in which the functions are called so that the functions return correctly to the calling code and possibly return values as well. Dynamically sized types, such as strings or arrays, can't be stored on the stack. For these values, a program can request memory space on its heap, so this is a potentially much bigger piece of memory than the stack. When possible, stack allocation is preferred over heap allocation because accessing the stack is a lot more efficient. Lifetimes All variables in a Rust code have a lifetime. Suppose we declare an n variable with the let n = 42u32; binding. Such a value is valid from where it is declared to when it is no longer referenced, which is called the lifetime of the variable. This is illustrated in the following code snippet: fn main() { let n = 42u32; let n2 = n; // a copy of the value from n to n2 life(n); println!("{}", m); // error: unresolved name `m`. println!("{}", o); // error: unresolved name `o`. }   fn life(m: u32) -> u32 {    let o = m;    o } The lifetime of n ends when main() ends; in general, the start and end of a lifetime happen in the same scope. The words lifetime and scope are synonymous, but we generally use the word lifetime to refer to the extent of a reference. As in other languages, local variables or parameters declared in a function do not exist anymore after the function has finished executing; in Rust, we say that their lifetime has ended. This is the case for the m and o variables in the preceding code snippet, which are only known in the life function. Likewise, the lifetime of a variable declared in a nested block is restricted to that block, like phi in the following example: {    let phi = 1.618; } println!("The value of phi is {}", phi); // is error Trying to use phi when its lifetime is over results in an error: unresolved name 'phi'. The lifetime of a value can be indicated in the code by an annotation, for example 'a, which reads as lifetime where a is simply an indicator; it could also be written as 'b, 'n, or 'life. It's common to see single letters being used to represent lifetimes. In the preceding example, an explicit lifetime indication was not necessary since there were no references involved. All values tagged with the same lifetime have the same maximum lifetime. In the following example, we have a transform function that explicitly declares the lifetime of its s parameter to be 'a: fn transform<'a>(s: &'a str) { /* ... */ } Note the <'a> indication after the name of the function. In nearly all cases, this explicit indication is not needed because the compiler is smart enough to deduce the lifetimes, so we can simply write this: fn transform_without_lifetime(s: &str) { /* ... */ } Here is an example where even when we indicate a lifetime specifier 'a, the compiler does not allow our code. Let's suppose that we define a Magician struct as follows: struct Magician { name: &'static str, power: u32 } We will get an error message if we try to construct the following function: fn return_magician<'a>() -> &'a Magician { let mag = Magician { name: "Gandalf", power: 4625}; &mag } The error message is error: 'mag' does not live long enough. Why does this happen? The lifetime of the mag value ends when the return_magician function ends, but this function nevertheless tries to return a reference to the Magician value, which no longer exists. Such an invalid reference is known as a dangling pointer. This is a situation that would clearly lead to errors and cannot be allowed. The lifespan of a pointer must always be shorter than or equal to than that of the value which it points to, thus avoiding dangling (or null) references. In some situations, the decision to determine whether the lifetime of an object has ended is complicated, but in almost all cases, the borrow checker does this for us automatically by inserting lifetime annotations in the intermediate code; so, we don't have to do it. This is known as lifetime elision. For example, when working with structs, we can safely assume that the struct instance and its fields have the same lifetime. Only when the borrow checker is not completely sure, we need to indicate the lifetime explicitly; however, this happens only on rare occasions, mostly when references are returned. One example is when we have a struct with fields that are references. The following code snippet explains this: struct MagicNumbers { magn1: &u32, magn2: &u32 } This won't compile and will give us the following error: missing lifetime specifier [E0106]. Therefore, we have to change the code as follows: struct MagicNumbers<'a> { magn1: &'a u32, magn2: &'a u32 } This specifies that both the struct and the fields have the lifetime as 'a. Perform the following exercise: Explain why the following code won't compile: fn main() {    let m: &u32 = {        let n = &5u32;        &*n    };    let o = *m; } Answer the same question for this code snippet as well: let mut x = &3; { let mut y = 4; x = &y; } Copying values and the Copy trait In the code that we discussed in earlier section the value of n is copied to a new location each time n is assigned via a new let binding or passed as a function argument: let n = 42u32; // no move, only a copy of the value: let n2 = n; life(n); fn life(m: u32) -> u32 {    let o = m;    o } At a certain moment in the program's execution, we would have four memory locations that contain the copied value 42, which we can visualize as follows: Each value disappears (and its memory location is freed) when the lifetime of its corresponding variable ends, which happens at the end of the function or code block in which it is defined. Nothing much can go wrong with this Copy behavior, in which the value (its bits) is simply copied to another location on the stack. Many built-in types, such as u32 and i64, work similar to this, and this copy-value behavior is defined in Rust as the Copy trait, which u32 and i64 implement. You can also implement the Copy trait for your own type, provided all of its fields or items implement Copy. For example, the MagicNumber struct, which contains a field of the u64 type, can have the same behavior. There are two ways to indicate this: One way is to explicitly name the Copy implementation as follows: struct MagicNumber {    value: u64 } impl Copy for MagicNumber {} Otherwise, we can annotate it with a Copy attribute: #[derive(Copy)] struct MagicNumber {    value: u64 } This now means that we can create two different copies, mag and mag2, of a MagicNumber by assignment: let mag = MagicNumber {value: 42}; let mag2 = mag; They are copies because they have different memory addresses (the values shown will differ at each execution): println!("{:?}", &mag as *const MagicNumber); // address is 0x23fa88 println!("{:?}", &mag2 as *const MagicNumber); // address is 0x23fa80 The *const function is a so-called raw pointer. A type that does not implement the Copy trait is called non-copyable. Another way to accomplish this is by letting MagicNumber implement the Clone trait: #[derive(Clone)] struct MagicNumber {    value: u64 } Then, we can use clone() mag into a different object called mag3, effectively making a copy as follows: let mag3 = mag.clone(); println!("{:?}", &mag3 as *const MagicNumber); // address is 0x23fa78 mag3 is a new pointer referencing a new copy of the value of mag. Pointers The n variable in the let n = 42i32; binding is stored on the stack. Values on the stack or the heap can be accessed by pointers. A pointer is a variable that contains the memory address of some value. To access the value it points to, dereference the pointer with *. This happens automatically in simple cases such as in println! or when a pointer is given as a parameter to a method. For example, in the following code, m is a pointer containing the address of n: let m = &n; println!("The address of n is {:p}", m); println!("The value of n is {}", *m); println!("The value of n is {}", m); This prints out the following output, which differs for each program run: The address of n is 0x23fb34 The value of n is 42 The value of n is 42 So, why do we need pointers? When we work with dynamically allocated values, such as a String, that can change in size, the memory address of that value is not known at compile time. Due to this, the memory address needs to be calculated at runtime. So, to be able to keep track of it, we need a pointer for it whose value will change when the location of String in memory changes. The compiler automatically takes care of the memory allocation of pointers and the freeing up of memory when their lifetime ends. You don't have to do this yourself like in C/C++, where you could mess up by freeing memory at the wrong moment or at multiple times. The incorrect use of pointers in languages such as C++ leads to all kinds of problems. However, Rust enforces a strong set of rules at compile time called the borrow checker, so we are protected against them. We have already seen them in action, but from here onwards, we'll explain the logic behind their rules. Pointers can also be passed as arguments to functions, and they can be returned from functions, but the compiler severely restricts their usage. When passing a pointer value to a function, it is always better to use the reference-dereference &* mechanism, as shown in this example: let q = &42; println!("{}", square(q)); // 1764 fn square(k: &i32) -> i32 {    *k * *k } References In our previous example, m, which had the &n value, is the simplest form of pointer, and it is called a reference (or borrowed pointer); m is a reference to the stack-allocated n variable and has the &i32 type because it points to a value of the i32 type. In general, when n is a value of the T type, then the &n reference is of the &T type. Here, n is immutable, so m is also immutable; for example, if you try to change the value of n through m with *m = 7; you will get a cannot assign to immutable borrowed content '*m' error. Contrary to C, Rust does not let you change an immutable variable via its pointer. Since there is no danger of changing the value of n through a reference, multiple references to an immutable value are allowed; they can only be used to read the value, for example: let o = &n; println!("The address of n is {:p}", o); println!("The value of n is {}", *o); It prints out as described earlier: The address of n is 0x23fb34 The value of n is 42 We could represent this situation in the memory as follows: It is clear that working with pointers such as this or in much more complex situations necessitates much stricter rules than the Copy behavior. For example, the memory can only be freed when there are no variables or pointers associated with it anymore. And when the value is mutable, can it be changed through any of its pointers? Mutable references do exist, and they are declared as let m = &mut n. However, n also has to be a mutable value. When n is immutable, the compiler rejects the m mutable reference binding with the error, cannot borrow immutable local variable 'n' as mutable. This makes sense since immutable variables cannot be changed even when you know their memory location. To reiterate, in order to change a value through a reference, both the variable and its reference have to be mutable, as shown in the following code snippet: let mut u = 3.14f64; let v = &mut u; *v = 3.15; println!("The value of u is now {}", *v); This will print: The value of u is now 3.15. Now, the value at the memory location of u is changed to 3.15. However, note that we now cannot change (or even print) that value anymore by using the u: u = u * 2.0; variable gives us a compiler error: cannot assign to 'u' because it is borrowed. We say that borrowing a variable (by making a reference to it) freezes that variable; the original u variable is frozen (and no longer usable) until the reference goes out of scope. In addition, we can only have one mutable reference: let w = &mut u; which results in the error: cannot borrow 'u' as mutable more than once at a time. The compiler even adds the following note to the previous code line with: let v = &mut u; note: previous borrow of 'u' occurs here; the mutable borrow prevents subsequent moves, borrows, or modification of `u` until the borrow ends. This is logical; the compiler is (rightfully) concerned that a change to the value of u through one reference might change its memory location because u might change in size, so it will not fit anymore within its previous location and would have to be relocated to another address. This would render all other references to u as invalid, and even dangerous, because through them we might inadvertently change another variable that has taken up the previous location of u! A mutable value can also be changed by passing its address as a mutable reference to a function, as shown in this example: let mut m = 7; add_three_to_magic(&mut m); println!("{}", m); // prints out 10 With the function add_three_to_magic declared as follows: fn add_three_to_magic(num: &mut i32) {    *num += 3; // value is changed in place through += } To summarize, when n is a mutable value of the T type, then only one mutable reference to it (of the &mut T type) can exist at any time. Through this reference, the value can be changed. Using ref in a match If you want to get a reference to a matched variable inside a match function, use the ref keyword, as shown in the following example: fn main() { let n = 42; match n {      ref r => println!("Got a reference to {}", r), } let mut m = 42; match m {      ref mut mr => {        println!("Got a mutable reference to {}", mr);        *mr = 43;      }, } println!("m has changed to {}!", m); } Which prints out: Got a reference to 42 Got a mutable reference to 42 m has changed to 43! The r variable inside the match has the &i32 type. In other words, the ref keyword creates a reference for use in the pattern. If you need a mutable reference, use ref mut. We can also use ref to get a reference to a field of a struct or tuple in a destructuring via a let binding. For example, while reusing the Magician struct, we can extract the name of mag by using ref and then return it from the match: let mag = Magician { name: "Gandalf", power: 4625}; let name = {    let Magician { name: ref ref_to_name, power: _ } = mag;    *ref_to_name }; println!("The magician's name is {}", name); Which prints: The magician's name is Gandalf. References are the most common pointer type and have the most possibilities; other pointer types should only be applied in very specific use cases. Summary In this article, we learned the intelligence behind the Rust compiler, which is embodied in the principles of ownership, moving values, and borrowing. Resources for Article: Further resources on this subject: Getting Started with NW.js [article] Creating Random Insults [article] Creating Man-made Materials in Blender 2.5 [article]
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Packt
03 Jun 2015
34 min read
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Getting Started with LiveCode Mobile

Packt
03 Jun 2015
34 min read
In this article written by Joel Gerdeen, author of the book LiveCode Mobile Development: Beginner's Guide - Second Edition we will learn the following topics: Sign up for Google Play Sign up for Amazon Appstore Download and install the Android SDK Configure LiveCode so that it knows where to look for the Android SDK Become an iOS developer with Apple Download and install Xcode Configure LiveCode so that it knows where to look for iOS SDKs Set up simulators and physical devices Test a stack in a simulator and physical device (For more resources related to this topic, see here.) Disclaimer This article references many Internet pages that are not under our control. Here, we do show screenshots or URLs, so remember that the content may have changed since we wrote this. The suppliers may also have changed some of the details, but in general, our description of procedures should still work the way we have described them. Here we go... iOS, Android, or both? It could be that you only have interest in iOS or Android. You should be able to easily skip to the sections you're interested in unless you're intrigued about how the other half works! If, like me, you're a capitalist, then you should be interested in both the operating systems. Far fewer steps are needed to get the Android SDK than the iOS developer tools because for iOS, we have to sign up as a developer with Apple. However, the configuration for Android is more involved. We'll go through all the steps for Android and then the ones for iOS. If you're an iOS-only kind of person, skip the next few pages and start up again at the Becoming an iOS Developer section. Becoming an Android developer It is possible to develop Android OS apps without signing up for anything. We'll try to be optimistic and assume that within the next 12 months, you will find time to make an awesome app that will make you rich! To that end, we'll go over everything that is involved in the process of signing up to publish your apps in both Google Play (formally known as Android Market) and Amazon Appstore. Google Play The starting location to open Google Play is http://developer.android.com/: We will come back to this page again, shortly to download the Android SDK, but for now, click on the Distribute link in the menu bar and then on the Developer Console button on the following screen. Since Google changes these pages occasionally, you can use the URL https://play.google.com/apps/publish/ or search for "Google Play Developer Console". The screens you will progress through are not shown here since they tend to change with time. There will be a sign-in page; sign in using your usual Google details. Which e-mail address to use? Some Google services are easier to sign up for if you have a Gmail account. Creating a Google+ account, or signing up for some of their cloud services, requires a Gmail address (or so it seemed to me at the time!). If you have previously set up Google Wallet as part of your account, some of the steps in signing up become simpler. So, use your Gmail address and if you don't have one, create one! Google charges you a $25 fee to sign up for Google Play. At least now, you know about this! Enter the developer name, e-mail address, website URL (if you have one), and your phone number. The payment of $25 will be done through Google Wallet, which will save you from entering the billing details yet again. Now, you're all signed up and ready to make your fortune! Amazon Appstore Although the rules and costs for Google Play are fairly relaxed, Amazon has a more Apple-like approach, both in the amount they charge you to register and in the review process to accept app submissions. The URL to open Amazon Appstore is http://developer.amazon.com/public: Follow these steps to start with Amazon Appstore: When you select Get Started, you need to sign in to your Amazon account. Which email address to use? This feels like déjà vu! There is no real advantage of using your Google e-mail address when signing up for the Amazon Appstore Developer Program, but if you happen to have an account with Amazon, sign in with that one. It will simplify the payment stage, and your developer account and the general Amazon account will be associated with each other. You are then asked to agree to the Appstore Distribution Agreement terms before learning about the costs. These costs are $99 per year, but the first year is free. So that's good! Unlike the Google Android Market, Amazon asks for your bank details up front, ready to send you lots of money later, we hope! That's it, you're ready to make another fortune to go along with the one that Google sent you! Pop quiz – when is something too much? You're at the end of developing your mega app, it's 49.5 MB in size, and you just need to add title screen music. Why would you not add the two-minute epic tune you have lined up? It would take too long to load. People tend to skip the title screen soon anyway. The file size is going to be over 50 MB. Heavy metal might not be appropriate for a children's storybook app! Answer: 3 The other answers are valid too, though you could play the music as an external sound to reduce loading time, but if your file size goes over 50 MB, you would then cut out potential sales from people who are connected by cellular and not wireless networks. At the time of writing this aticle, all the stores require that you be connected to the site via a wireless network if you intend to download apps that are over 50 MB. Downloading the Android SDK Head back to http://developer.android.com/ and click on the Get the SDK link or go straight to http://developer.android.com/sdk/index.html. This link defaults to the OS that you are running on. Click on the Other Download Options link to see the full set of options for other systems, as shown here: In this article, we're only going to cover Windows and Mac OS X (Intel) and only as much as is needed to make LiveCode work with the Android and iOS SDKs. If you intend to make native Java-based applications, you may be interested in reading through all the steps that are described in the web page http://developer.android.com/sdk/installing.html. Click on the SDK download link for your platform. Note that you don't need the ADT Bundle unless you plan to develop outside the LiveCode IDE. The steps you'll have to go through are different for Mac and Windows. Let's start with Mac. Installing the Android SDK on Mac OS X (Intel) LiveCode itself doesn't require Intel Mac; you can develop stacks using a PowerPC-based Mac, but both the Android SDK and some of the iOS tools require an Intel-based Mac, which sadly means that if you're reading this as you sit next to your Mac G4 or G5, you're not going to get too far! The Android SDK requires the Java Runtime Environment (JRE). Since Apple stopped including the JRE in more recent OS X systems, you should check whether you have it in your system by typing java –version in a Terminal window. The terminal will display the version of Java installed. If not, you may get a message like the following: Click on the More Info button and follow the instructions to install the JRE and verify its installation. At the time of writing this article, JRE 8 doesn't work with OS X 10.10 and I had to use the JRE 6 obtained from http://support.apple.com/kb/DL1572. The file that you just downloaded will automatically expand to show a folder named android-sdk-macosx. It may be in your downloads folder right now, but a more natural place for it would be in your Documents folder, so move it there before performing the next steps. There is an SDK readme text file that lists the steps you need to follow during the installation. If these steps are different to what we have here, then follow the steps in the readme file in case they have been updated since the procedure here was written. Open the Terminal application, which is in Applications/Utilities. You need to change the default directories present in the android-sdk-macosx folder. One handy trick, using Terminal, is that you can drag items into the Terminal window to get the file path to that item. Using this trick, you can type cd and a space in the Terminal window and then drag the android-sdk-macosx folder after the space character. You'll end up with this line if your username is Fred: new-host-3:~ fred$ cd /Users/fred/Documents/android-sdk-macosx Of course, the first part of the line and the user folder will match yours, not Fred's! Whatever your name is, press the Return or Enter key after entering the preceding line. The location line now changes to look like this: new-host-3:android-sdk-macosx colin$ Either carefully type or copy and paste the following line from the readme file: tools/android update sdk --no-ui Press Return or Enter again. How long the file takes to get downloaded depends on your Internet connection. Even with a very fast Internet connection, it could still take over an hour. If you care to follow the update progress, you can just run the android file in the tools directory. This will open the Android SDK Manager, which is similar to the Windows version shown a couple of pages further on in this article. Installing the Android SDK on Windows The downloads page recommends that you use the .exe download link, as it gives extra services to you, such as checking whether you have the Java Development Kit (JDK) installed. When you click on the link, either use the Run or Save options, as you would with any download of a Windows installer. Here, we've opted to use Run; if you use Save, then you need to open the file after it has been saved to your hard drive. In the following case, as the JDK wasn't installed, a dialog box appears saying go to Oracle's site to get the JDK: If you see this screen too, you can leave the dialog box open and click on the Visit java.oracle.com button. On the Oracle page, click on a checkbox to agree to their terms and then on the download link that corresponds with your platform. Choose the 64-bit option if you are running a 64-bit version of Windows or the x86 option if you are running a 32-bit version of Windows. Either way, you're greeted with another installer that you can Run or Save as you prefer. Naturally, it takes a while for the installer to do its thing too! When the installation is complete, you will see a JDK registration page and it's up to you, to register or not. Back at the Android SDK installer dialog box, you can click on the Back button and then the Next button to get back to the JDK checking stage; only now, it sees that you have the JDK installed. Complete the remaining steps of the SDK installer as you would with any Windows installer. One important thing to note is that the last screen of the installer offers to open the SDK Manager. You should do that, so resist the temptation to uncheck that box! Click on Finish and you'll be greeted with a command-line window for a few moments, as shown in the following screenshot, and then, the Android SDK Manager will appear and do its thing: As with the Mac version, it takes a very long time for all these add-ons to download. Pointing LiveCode to the Android SDK After all the installation and command-line work, it's a refreshing change to get back to LiveCode! Open the LiveCode Preferences and choose Mobile Support: We will set the two iOS entries after we get iOS going (but these options will be grayed out in Windows). For now, click on the … button next to the Android development SDK root field and navigate to where the SDK is installed. If you've followed the earlier steps correctly, then the SDK will be in the Documents folder on Mac or you can navigate to C:Program Files (x86)Android to find it on Windows (or somewhere else, if you choose to use a custom location). Depending on the APIs that were loaded in the SDK Manager, you may get a message that the path does not include support for Android 2.2 (API 8). If so, use the Android SDK Manager to install it. LiveCode seems to want API 8 even though at this time Android 5.0 uses API 21. Phew! Now, let's do the same for iOS… Pop quiz – tasty code names An Android OS uses some curious code names for each version. At the time of writing this article, we were on Android OS 5, which had a code name of Lollipop. Version 4.1 was Jelly Bean and version 4.4 was KitKat. Which of these is most likely to be the code name for the next Android OS? Lemon Cheesecake Munchies Noodle Marshmallow Answer: 4 The pattern, if it isn't obvious, is that the code name takes on the next letter of the alphabet, is a kind of food, but more specifically, it's a dessert. "Munchies" almost works for Android OS 6, but "Marshmallow" or "Meringue Pie" would be a better choices! Becoming an iOS developer Creating iOS LiveCode applications requires that LiveCode must have access to the iOS SDK. This is installed as part of the Xcode developer tools and is a Mac-only program. Also, when you upload an app to the iOS App Store, the application used is Mac only and is part of the Xcode installation. If you are a Windows-based developer and wish to develop and publish for iOS, you need either an actual Mac based system or a virtual machine that can run the Mac OS. We can even use VirtualBox for running a Mac based virtual machine, but performance will be an issue. Refer to http://apple.stackexchange.com/questions/63147/is-mac-os-x-in-a-virtualbox-vm-suitable-for-ios-development for more information. The biggest difference between becoming an Android developer and becoming an iOS developer is that you have to sign up with Apple for their developer program even if you never produce an app for the iOS App Store, but no such signing up is required when becoming an Android developer. If things go well and you end up making an app for various stores, then this isn't such a big deal. It will cost you $25 to submit an app to the Android Market, $99 a year (with the first year free) to submit an app to the Amazon Appstore, and $99 a year (including the first year) to be an iOS developer with Apple. Just try to sell more than 300 copies of your amazing $0.99 app and you'll find that it has paid for itself! Note that there is a free iOS App Store and app licensing included, with LiveCode Membership, which also costs $99 per year. As a LiveCode member, you can submit your free non-commercial app to RunRev who will provide a license that will allow you to submit your app as "closed source" to iOS App Store. This service is exclusively available for LiveCode members. The first submission each year is free; after that, there is a $25 administration fee per submission. Refer to http://livecode.com/membership/ for more information. You can enroll yourself in the iOS Developer Program for iOS at http://developer.apple.com/programs/ios/: While signing up to be an iOS developer, there are a number of possibilities when it comes to your current status. If you already have an Apple ID, which you use with your iTunes or Apple online store purchases, you could choose the I already have an Apple ID… option. In order to illustrate all the steps to sign up, we will start as a brand new user, as shown in the following screenshot: You can choose whether you want to sign up as an individual or as a company. We will choose Individual, as shown in the following screenshot: With any such sign up process, you need to enter your personal details, set a security question, and enter your postal address: Most Apple software and services have their own legal agreement for you to sign. The one shown in the following screenshot is the general Registered Apple Developer Agreement: In order to verify the e-mail address you have used, a verification code is sent to you with a link in the e-mail, you can click this, or enter the code manually. Once you have completed the verification code step, you can then enter your billing details. It could be that you might go on to make LiveCode applications for the Mac App Store, in which case, you will need to add the Mac Developer Program product. For our purpose, we only need to sign up for the iOS Developer Program, as shown in the following screenshot: Each product that you sign up for has its own agreement. Lots of small print to read! The actual purchasing of the iOS developer account is handled through the Apple Store of your own region, shown as follows: As you can see in the next screenshot, it is going to cost you $99 per year or $198 per year if you also sign up for the Mac Developer account. Most LiveCode users won't need to sign up for the Mac Developer account unless their plan is to submit desktop apps to the Mac App Store. After submitting the order, you are rewarded with a message that tells you that you are now registered as an Apple developer! Sadly, you won't get an instant approval, as was the case with Android Market or Amazon Appstore. You have to wait for the approval for five days. In the early iPhone Developer days, the approval could take a month or more, so 24 hours is an improvement! Pop quiz – iOS code names You had it easy with the pop quiz about Android OS code names! Not so with iOS. Which of these names is more likely to be a code name for a future version of iOS? Las Vegas Laguna Beach Hunter Mountain Death Valley Answer: 3 Although not publicized, Apple does use code names for each version of iOS. Previous examples included Big Bear, Apex, Kirkwood, and Telluride. These, and all the others are apparently ski resorts. Hunter Mountain is a relatively small mountain (3,200 feet), so if it does get used, perhaps it would be a minor update! Installing Xcode Once you receive confirmation of becoming an iOS developer, you will be able to log in to the iOS Dev Center at https://developer.apple.com/devcenter/ios/index.action. This same page is used by iOS developers who are not using LiveCode and is full of support documents that can help you create native applications using Xcode and Objective-C. We don't need all the support documents, but we do need to download Xcode's support documents. In the downloads area of the iOS Dev Center page, you will see a link to the current version of Xcode and a link to get to the older versions as well. The current version is delivered via Mac App Store; when you try the given link, you will see a button that takes you to the App Store application. Installing Xcode from Mac App Store is very straightforward. It's just like buying any other app from the store, except that it's free! It does require you to use the latest version of Mac OS X. Xcode will show up in your Applications folder. If you are using an older system, then you need to download one of the older versions from the developer page. The older Xcode installation process is much like the installation process of any other Mac application: The older version of Xcode takes a long time to get installed, but in the end, you should have the Developer folder or a new Xcode application ready for LiveCode. Coping with newer and older devices In early 2012, Apple brought to the market a new version of iPad. The main selling point of this one compared to iPad 2 is that it has a Retina display. The original iPads have a resolution of 1024 x 768 and the Retina version has a resolution of 2048 x 1536. If you wish to build applications to take advantage of this, you must get the current version of Xcode from Mac App Store and not one of the older versions from the developer page. The new version of Xcode demands that you work on Mac OS 10.10 or its later versions. So, to fully support the latest devices, you may have to update your system software more than you were expecting! But wait, there's more… By taking a later version of Xcode, you are missing the iOS SDK versions needed to support older iOS devices, such as the original iPhone and iPhone 3G. Fortunately, you can go to Preferences in Xcode where there is a Downloads tab where you can get these older SDKs downloaded in the new version of Xcode. Typically, Apple only allows you to download one version older than the one that is currently provided in Xcode. There are older versions available, but are not accepted by Apple for App Store submission. Pointing LiveCode to the iOS SDKs Open the LiveCode Preferences and choose Mobile Support: Click on the Add Entry button in the upper-right section of the window to see a dialog box that asks whether you are using Xcode 4.2 or 4.3 or a later version. If you choose 4.2, then go on to select the folder named Developer at the root of your hard drive. For 4.3 or later versions, choose the Xcode application itself in your Applications folder. LiveCode knows where to find the SDKs for iOS. Before we make our first mobile app… Now that the required SDKs are installed and LiveCode knows where they are, we can make a stack and test it in a simulator or on a physical device. We do, however, have to get the simulators and physical devices warmed up… Getting ready for test development on an Android device Simulating on iOS is easier than it is on Android, and testing on a physical device is easier on Android than on iOS, but the setting up of physical Android devices can be horrendous! Time for action – starting an Android Virtual Device You will have to dig a little deep in the Android SDK folders to find the Android Virtual Device setup program. You might as well provide a shortcut or an alias to it for quicker access. The following steps will help you setup and start an Android virtual device: Navigate to the Android SDK tools folder located at C:Program Files (x86)Androidandroid-sdk on Windows and navigate to your Documents/android-sdk-macosx/tools folder on Mac. Open AVD Manager on Windows or android on Mac (these look like a Unix executable file; just double-click on it and the application will open via a command-line window). If you're on Mac, select Manage AVDs… from the Tools menu. Select Tablet from the list of devices if there is one. If not, you can add your own custom devices as described in the following section. Click on the Start button. Sit patiently while the virtual device starts up! Open LiveCode, create a new Mainstack, and click on Save to save the stack to your hard drive. Navigate to File | Standalone Application Settings…. Click on the Android icon and click on the Build for Android checkbox to select it. Close the settings dialog box and take a look at the Development menu. If the virtual machine is up and running, you should see it listed in the Test Target submenu. Creating an Android Virtual Device If there are no devices listed when you open the Android Virtual Device (AVD) Manager, you may If you wish to create a device, so click on the Create button. The following screenshot will appear when you do so. Further explanation of the various fields can be found at https://developer.android.com/tools/devices/index.html. After you have created a device, you can click on Start to start the virtual device and change some of the Launch Options. You should typically select Scale display to real size unless it is too big for your development screen. Then, click on Launch to fire up the emulator. Further information on how to run the emulator can be found at http://developer.android.com/tools/help/emulator.html. What just happened? Now that you've opened an Android virtual device, LiveCode will be able to test stacks using this device. Once it has finished loading, that is! Connecting a physical Android device Connecting a physical Android device can be extremely straightforward: Connect your device to the system by USB. Select your device from the Development | Test Target submenu. Select Test from the Development menu or click on the Test button in the Tool Bar. There can be problem cases though, and Google Search will become your best friend before you are done solving these problems! We should look at an example problem case, so that you get an idea of how to solve similar situations that you may encounter. Using Kindle Fire When it comes to finding Android devices, the Android SDK recognizes a lot of them automatically. Some devices are not recognized and you have to do something to help Android Debug Bridge (ADB) find these devices. Android Debug Bridge (ADB) is part of the Android SDK that acts as an intermediary between your device and any software that needs to access the device. In some cases, you will need to go to the Android system on the device to tell it to allow access for development purposes. For example, on an Android 3 (Honeycomb) device, you need to go to the Settings | Applications | Development menu and you need to activate the USB debugging mode. Before ADB connects to a Kindle Fire device, that device must first be configured, so that it allows connection. This is enabled by default on the first generation Kindle Fire device. On all other Kindle Fire models, go to the device settings screen, select Security, and set Enable ADB to On. The original Kindle Fire model comes with USB debugging already enabled, but the ADB system doesn't know about the device at all. You can fix this! Time for action – adding Kindle Fire to ADB It only takes one line of text to add Kindle Fire to the list of devices that ADB knows about. The hard part is tracking down the text file to edit and getting ADB to restart after making the required changes. Things are more involved when using Windows than with Mac because you also have to configure the USB driver, so the two systems are shown here as separate steps. The steps to be followed for adding a Kindle Fire to ADB for a Windows OS are as follows: In Windows Explorer, navigate to C:Usersyourusername.android where the adv_usb.ini file is located. Open the adv_usb.ini text file in a text editor. The file has no visible line breaks, so it is better to use WordPad than NotePad. On the line after the three instruction lines, type 0x1949. Make sure that there are no blank lines; the last character in the text file would be 9 at the end of 0x1949. Now, save the file. Navigate to C:Program Files (x86)Androidandroid-sdkextrasgoogleusb_driver where android_winusb.inf is located. Right-click on the file and in Properties, Security, select Users from the list and click on Edit to set the permissions, so that you are allowed to write the file. Open the android_winusb.inf file in NotePad. Add the following three lines to the [Google.NTx86] and [Google.NTamd64] sections and save the file: ;Kindle Fire %SingleAdbInterface% = USB_Install, USBVID_1949&PID_0006 %CompositeAdbInterface% = USB_Install, USBVID_1949&PID_0006&MI_01 You need to set the Kindle so that it uses the Google USB driver that you just edited. In the Windows control panel, navigate to Device Manager and find the Kindle entry in the list that is under USB. Right-click on the Kindle entry and choose Update Driver Software…. Choose the option that lets you find the driver on your local drive, navigate to the googleusb_driver folder, and then select it to be the new driver. When the driver is updated, open a command window (a handy trick to open a command window is to use Shift-right-click on the desktop and to choose "Open command window here"). Change the directories to where the ADB tool is located by typing: cd C:Program Files (x86)Androidandroid-sdkplatform-tools Type the following three line of code and press Enter after each line: adb kill-server adb start-server adb devices You should see the Kindle Fire listed (as an obscure looking number) as well as the virtual device if you still have that running. The steps to be followed for a Mac (MUCH simpler!) system are as follows: Navigate to where the adv_usb.ini file is located. On Mac, in Finder, select the menu by navigating to Go | Go to Folder… and type ~/.android/in. Open the adv_usb.ini file in a text editor. On the line after the three instruction lines, type 0x1949. Make sure that there are no blank lines; the last character in the text file would be 9 at the end of 0x1949. Save the adv_usb.ini file. Navigate to Utilities | Terminal. You can let OS X know how to find ADB from anywhere by typing the following line (replace yourusername with your actual username and also change the path if you've installed the Android SDK to some other location): export PATH=$PATH:/Users/yourusername/Documents/android-sdk-macosx/platform-tools Now, try the same three lines as we did with Windows: adb kill-server adb start-server adb devices Again, you should see the Kindle Fire listed here. What just happened? I suspect that you're going to have nightmares about all these steps! It took a lot of research on the Web to find out some of these obscure hacks. The general case with Android devices on Windows is that you have to modify the USB driver for the device to be handled using the Google USB driver, and you may have to modify the adb_usb.ini file (on Mac too) for the device to be considered as an ADB compatible device. Getting ready for test development on an iOS device If you carefully went through all these Android steps, especially on Windows, you will hopefully be amused by the brevity of this section! There is a catch though; you can't really test on an iOS device from LiveCode. We'll look at what you have to do instead in a moment, but first, we'll look at the steps required to test an app in the iOS simulator. Time for action – using the iOS simulator The initial steps are much like what we did for Android apps, but the process becomes a lot quicker in later steps. Remember, this only applies to a Mac OS; you can only do these things on Windows if you are using a Mac OS in a virtual machine, which may have performance issues. This is most likely not covered by the Mac OS's user agreement! In other words, get a Mac OS if you intend to develop for iOS. The following steps will help you achieve that: Open LiveCode and create a new Mainstack and save the stack to your hard drive. Select File and then Standalone Application Settings…. Click on the iOS icon to select the Build for iOS checkbox. Close the settings dialog box and take a look at the Test Target menu under Development. You will see a list of simulator options for iPhone and iPad and different versions of iOS. To start the iOS simulator, select an option and click on the Test button. What just happened? This was all it took for us to get the testing done using the iOS simulators! To test on a physical iOS device, we need to create an application file first. Let's do that. Appiness at last! At this point, you should be able to create a new Mainstack, save it, select either iOS or Android in the Standalone Settings dialog box, and be able to see simulators or virtual devices in the Development/Test menu item. In the case of an Android app, you will also see your device listed if it is connected via USB at the time. Time for action – testing a simple stack in the simulators Feel free to make things that are more elaborate than the ones we have made through these steps! The following instructions make an assumption that you know how to find things by yourself in the object inspector palette: Open LiveCode, create a new Mainstack, and save it someplace where it is easy to find in a moment from now. Set the card window to the size 480 x 320 and uncheck the Resizable checkbox. Drag a label field to the top-left corner of the card window and set its contents to something appropriate. Hello World might do. If you're developing on Windows, skip to step 11. Open the Standalone Application Settings dialog box, click on the iOS icon, and click on the Build for iOS checkbox. Under Orientation Options, set the iPhone Initial Orientation to Landscape Left. Close the dialog box. Navigate to the Development | Test Target submenu and choose an iPhone Simulator. Select Test from the Development menu. You should now be able to see your test stack running in the iOS simulator! As discussed earlier, launch the Android virtual device. Open the Standalone Application Settings dialog box, click on the Android icon, and click on the Build for Android checkbox. Under User Interface Options, set the Initial Orientation to Landscape. Close the dialog box. If the virtual device is running by now, do whatever it takes to get past the locked home screen, if that's what it is showing. From the Development/Test Target submenu, choose the Android emulator. Select Test from the Development menu. You should now see your test stack running in the Android emulator! What just happened? All being well, you just made and ran your first mobile app on both Android and iOS! For an encore, we should try this on physical devices only to give Android a chance to show how easy it can be done. There is a whole can of worms we didn't open yet that has to do with getting an iOS device configured, so that it can be used for testing. You could visit the iOS Provisioning Portal at https://developer.apple.com/ios/manage/overview/index.action and look at the How To tab in each of the different sections. Time for action – testing a simple stack on devices Now, let's try running our tests on physical devices. Get your USB cables ready and connect the devices to your computer. Lets go through the steps for an Android device first: You should still have Android selected in Standalone Application Settings. Get your device to its home screen past the initial Lock screen if there is one. Choose Development/Test Target and select your Android device. It may well say "Android" and a very long number. Choose Development/Test. The stack should now be running on your Android device. Now, we'll go through the steps to test a simple stack on an iOS device: Change the Standalone Application Settings back to iOS. Under Basic Application Settings of the iOS settings is a Profile drop-down menu of the provisioning files that you have installed. Choose one that is configured for the device you are going to test. Close the dialog box and choose Save as Standalone Application… from the File menu. In Finder, locate the folder that was just created and open it to reveal the app file itself. As we didn't give the stack a sensible name, it will be named Untitled 1. Open Xcode, which is in the Developer folder you installed earlier, in the Applications subfolder. In the Xcode folder, choose Devices from the Window menu if it isn't already selected. You should see your device listed. Select it and if you see a button labeled Use for Development, click on that button. Drag the app file straight from the Finder menu to your device in the Devices window. You should see a green circle with a + sign. You can also click on the + sign below Installed Apps and locate your app file in the Finder window. You can also replace or delete an installed app from this window. You can now open the app on your iOS device! What just happened? In addition to getting a test stack to work on real devices, we also saw how easy it is, once it's all configured, to test a stack, straight on an Android device. If you are developing an app that is to be deployed on both Android and iOS, you may find that the fastest way to work is to test with the iOS Simulator for iOS tests, but for this, you need to test directly on an Android device instead of using the Android SDK virtual devices. Have a go hero – Nook Until recently, the Android support for the Nook Color from Barnes & Noble wasn't good enough to install LiveCode apps. It seems to have improved though and could well be another worthwhile app store for you to target. Investigate about the sign up process, download their SDK, and so on. With any luck, some of the processes that you've learned while signing up for the other stores will also apply to the Nook store. You can start the signing up process at https://nookdeveloper.barnesandnoble.com. Further reading The SDK providers, Google and Apple, have extensive pages of information on how to set up development environments, create certificates and provisioning files, and so on. The information covers a lot of topics that don't apply to LiveCode, so try not to get lost! These URLs would be good starting points if you want to read further: http://developer.android.com/ http://developer.apple.com/ios/ Summary Signing up for programs, downloading files, using command lines all over the place, and patiently waiting for the Android emulator to launch. Fortunately, you only have to go through it once. In this article, we worked through a number of tasks that you have to do before you create a mobile app in LiveCode. We had to sign up as an iOS developer before we could download and install Xcode and iOS SDKs. We then downloaded and installed the Android SDK and configured LiveCode for devices and simulators. We also covered some topics that will be useful once you are ready to upload a finished app. We showed you how to sign up for the Android Market and Amazon Appstore. There will be a few more mundane things that we have to cover at the end of the article, but not for a while! Next up, we will start to play with some of the special abilities of mobile devices. Resources for Article: Further resources on this subject: LiveCode: Loops and Timers [article] Creating Quizzes [article] Getting Started with LiveCode for Mobile [article]
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