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  • Silverlight 4 Twitter Client &ndash; Part 3

    - by Max
    Finally Silverlight 4 RC is released and also that Windows 7 Phone Series will rely heavily on Silverlight platform for apps platform. its a really good news for Silverlight developers and designers. More information on this here. You can use SL 4 RC with VS 2010. SL 4 RC does not come with VS 2010, you need to download it separately and install it. So for the next part, be ready with VS 2010 and SL4 RC, we will start using them and not With this momentum, let us go to the next part of our twitter client tutorial. This tutorial will cover setting your status in Twitter and also retrieving your 1) As everything in Silverlight is asynchronous, we need to have some visual representation showing that something is going on in the background. So what I did was to create a progress bar with indeterminate animation. The XAML is here below. <ProgressBar Maximum="100" Width="300" Height="50" Margin="20" Visibility="Collapsed" IsIndeterminate="True" Name="progressBar1" VerticalAlignment="Center" HorizontalAlignment="Center" /> 2) I will be toggling this progress bar to show the background work. So I thought of writing this small method, which I use to toggle the visibility of this progress bar. Just pass a bool to this method and this will toggle it based on its current visibility status. public void toggleProgressBar(bool Option){ if (Option) { if (progressBar1.Visibility == System.Windows.Visibility.Collapsed) progressBar1.Visibility = System.Windows.Visibility.Visible; } else { if (progressBar1.Visibility == System.Windows.Visibility.Visible) progressBar1.Visibility = System.Windows.Visibility.Collapsed; }} 3) Now let us create a grid to hold a textbox and a update button. The XAML will look like something below <Grid HorizontalAlignment="Center"> <Grid.RowDefinitions> <RowDefinition Height="50"></RowDefinition> </Grid.RowDefinitions> <Grid.ColumnDefinitions> <ColumnDefinition Width="400"></ColumnDefinition> <ColumnDefinition Width="200"></ColumnDefinition> </Grid.ColumnDefinitions> <TextBox Name="TwitterStatus" Width="380" Height="50"></TextBox> <Button Name="UpdateStatus" Content="Update" Grid.Row="1" Grid.Column="2" Width="200" Height="50" Click="UpdateStatus_Click"></Button></Grid> 4) The click handler for this update button will be again using the Web Client to post values. Posting values using Web Client. The code is: private void UpdateStatus_Click(object sender, RoutedEventArgs e){ toggleProgressBar(true); string statusupdate = "status=" + TwitterStatus.Text; WebRequest.RegisterPrefix("https://", System.Net.Browser.WebRequestCreator.ClientHttp);  WebClient myService = new WebClient(); myService.AllowReadStreamBuffering = true; myService.UseDefaultCredentials = false; myService.Credentials = new NetworkCredential(GlobalVariable.getUserName(), GlobalVariable.getPassword());  myService.UploadStringCompleted += new UploadStringCompletedEventHandler(myService_UploadStringCompleted); myService.UploadStringAsync(new Uri("https://twitter.com/statuses/update.xml"), statusupdate);  this.Dispatcher.BeginInvoke(() => ClearTextBoxValue());} 5) In the above code, we have a event handler which will be fired on this request is completed – !! Remember SL is Asynch !! So in the myService_UploadStringCompleted, we will just toggle the progress bar and change some status text to say that its done. The code for this will be StatusMessage is just another textblock conveniently positioned in the page.  void myService_UploadStringCompleted(object sender, UploadStringCompletedEventArgs e){ if (e.Error != null) { StatusMessage.Text = "Status Update Failed: " + e.Error.Message.ToString(); } else { toggleProgressBar(false); TwitterCredentialsSubmit(); }} 6) Now let us look at fetching the friends updates of the logged in user and displaying it in a datagrid. So just define a data grid and set its autogenerate columns as true. 7) Let us first create a data structure for use with fetching the friends timeline. The code is something like below: namespace MaxTwitter.Classes{ public class Status { public Status() {} public string ID { get; set; } public string Text { get; set; } public string Source { get; set; } public string UserID { get; set; } public string UserName { get; set; } }} You can add as many fields as you want, for the list of fields, have a look at here. It will ask for your Twitter username and password, just provide them and this will display the xml file. Go through them pick and choose your desired fields and include in your Data Structure. 8) Now the web client request for this is similar to the one we saw in step 4. Just change the uri in the last but one step to https://twitter.com/statuses/friends_timeline.xml Be sure to change the event handler to something else and within that we will use XLINQ to fetch the required details for us. Now let us how this event handler fetches details. public void parseXML(string text){ XDocument xdoc; if(text.Length> 0) xdoc = XDocument.Parse(text); else xdoc = XDocument.Parse(@"I USED MY OWN LOCAL COPY OF XML FILE HERE FOR OFFLINE TESTING"); statusList = new List<Status>(); statusList = (from status in xdoc.Descendants("status") select new Status { ID = status.Element("id").Value, Text = status.Element("text").Value, Source = status.Element("source").Value, UserID = status.Element("user").Element("id").Value, UserName = status.Element("user").Element("screen_name").Value, }).ToList(); //MessageBox.Show(text); //this.Dispatcher.BeginInvoke(() => CallDatabindMethod(StatusCollection)); //MessageBox.Show(statusList.Count.ToString()); DataGridStatus.ItemsSource = statusList; StatusMessage.Text = "Datagrid refreshed."; toggleProgressBar(false);} in the event handler, we call this method with e.Result.ToString() Parsing XML files using LINQ is super cool, I love it.   I am stopping it here for  this post. Will post the completed files in next post, as I’ve worked on a few more features in this page and don’t want to confuse you. See you soon in my next post where will play with Twitter lists. Have a nice day! Technorati Tags: Silverlight,LINQ,XLINQ,Twitter API,Twitter,Network Credentials

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  • Two network interfaces and two IP addresses on the same subnet in Linux

    - by Scott Duckworth
    I recently ran into a situation where I needed two IP addresses on the same subnet assigned to one Linux host so that we could run two SSL/TLS sites. My first approach was to use IP aliasing, e.g. using eth0:0, eth0:1, etc, but our network admins have some fairly strict settings in place for security that squashed this idea: They use DHCP snooping and normally don't allow static IP addresses. Static addressing is accomplished by using static DHCP entries, so the same MAC address always gets the same IP assignment. This feature can be disabled per switchport if you ask and you have a reason for it (thankfully I have a good relationship with the network guys and this isn't hard to do). With the DHCP snooping disabled on the switchport, they had to put in a rule on the switch that said MAC address X is allowed to have IP address Y. Unfortunately this had the side effect of also saying that MAC address X is ONLY allowed to have IP address Y. IP aliasing required that MAC address X was assigned two IP addresses, so this didn't work. There may have been a way around these issues on the switch configuration, but in an attempt to preserve good relations with the network admins I tried to find another way. Having two network interfaces seemed like the next logical step. Thankfully this Linux system is a virtual machine, so I was able to easily add a second network interface (without rebooting, I might add - pretty cool). A few keystrokes later I had two network interfaces up and running and both pulled IP addresses from DHCP. But then the problem came in: the network admins could see (on the switch) the ARP entry for both interfaces, but only the first network interface that I brought up would respond to pings or any sort of TCP or UDP traffic. After lots of digging and poking, here's what I came up with. It seems to work, but it also seems to be a lot of work for something that seems like it should be simple. Any alternate ideas out there? Step 1: Enable ARP filtering on all interfaces: # sysctl -w net.ipv4.conf.all.arp_filter=1 # echo "net.ipv4.conf.all.arp_filter = 1" >> /etc/sysctl.conf From the file networking/ip-sysctl.txt in the Linux kernel docs: arp_filter - BOOLEAN 1 - Allows you to have multiple network interfaces on the same subnet, and have the ARPs for each interface be answered based on whether or not the kernel would route a packet from the ARP'd IP out that interface (therefore you must use source based routing for this to work). In other words it allows control of which cards (usually 1) will respond to an arp request. 0 - (default) The kernel can respond to arp requests with addresses from other interfaces. This may seem wrong but it usually makes sense, because it increases the chance of successful communication. IP addresses are owned by the complete host on Linux, not by particular interfaces. Only for more complex setups like load- balancing, does this behaviour cause problems. arp_filter for the interface will be enabled if at least one of conf/{all,interface}/arp_filter is set to TRUE, it will be disabled otherwise Step 2: Implement source-based routing I basically just followed directions from http://lartc.org/howto/lartc.rpdb.multiple-links.html, although that page was written with a different goal in mind (dealing with two ISPs). Assume that the subnet is 10.0.0.0/24, the gateway is 10.0.0.1, the IP address for eth0 is 10.0.0.100, and the IP address for eth1 is 10.0.0.101. Define two new routing tables named eth0 and eth1 in /etc/iproute2/rt_tables: ... top of file omitted ... 1 eth0 2 eth1 Define the routes for these two tables: # ip route add default via 10.0.0.1 table eth0 # ip route add default via 10.0.0.1 table eth1 # ip route add 10.0.0.0/24 dev eth0 src 10.0.0.100 table eth0 # ip route add 10.0.0.0/24 dev eth1 src 10.0.0.101 table eth1 Define the rules for when to use the new routing tables: # ip rule add from 10.0.0.100 table eth0 # ip rule add from 10.0.0.101 table eth1 The main routing table was already taken care of by DHCP (and it's not even clear that its strictly necessary in this case), but it basically equates to this: # ip route add default via 10.0.0.1 dev eth0 # ip route add 130.127.48.0/23 dev eth0 src 10.0.0.100 # ip route add 130.127.48.0/23 dev eth1 src 10.0.0.101 And voila! Everything seems to work just fine. Sending pings to both IP addresses works fine. Sending pings from this system to other systems and forcing the ping to use a specific interface works fine (ping -I eth0 10.0.0.1, ping -I eth1 10.0.0.1). And most importantly, all TCP and UDP traffic to/from either IP address works as expected. So again, my question is: is there a better way to do this? This seems like a lot of work for a seemingly simple problem.

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  • How do i route TCP connections via TOR? [on hold]

    - by acidzombie24
    I was reading about torchat which is essentially an anonymous chat program. It sounded cool so i wanted to experiment with making my own. First i wrote a test to grab a webpage using Http. Sicne .NET doesnt support SOCKS4A/SOCKS5 i used privoxy and my app worked. Then i switch to a TCP echo test and privoxy doesnt support TCP so i searched and installed 6+ proxy apps (freecap, socat, freeproxy, delegate are the ones i can remember from the top of my head, i also played with putty bc i know it supports tunnels and SOCK5) but i couldnt successfully get any of them to work let alone get it running with my http test that privoxy easily and painlessly did. What may i use to get TCP connections going through TOR? I spent more then 2 hours without success. I don't know if i am looking for a relay, tunnel, forwarder, proxy or a proxychain which all came up in my search. I use the config below for .NET. I need TCP working but i am first testing with http since i know i had it working using privoxy. What apps and configs do i use to get TCP going through tor? <?xml version="1.0" encoding="utf-8" ?> <configuration> <system.net> <defaultProxy enabled="true"> <proxy bypassonlocal="True" proxyaddress="http://127.0.0.1:8118"/> </defaultProxy> <settings> <httpWebRequest useUnsafeHeaderParsing="true"/> </settings> </system.net> </configuration> -edit- Thanks to Bernd i have a solution. Here is the code i ended up writing. It isn't amazing but its fair. static NetworkStream ConnectSocksProxy(string proxyDomain, short proxyPort, string host, short hostPort, TcpClient tc) { tc.Connect(proxyDomain, proxyPort); if (System.Text.RegularExpressions.Regex.IsMatch(host, @"[\:/\\]")) throw new Exception("Invalid Host name. Use FQDN such as www.google.com. Do not have http, a port or / in it"); NetworkStream ns = tc.GetStream(); var HostNameBuf = new ASCIIEncoding().GetBytes(host); var HostPortBuf = BitConverter.GetBytes(IPAddress.HostToNetworkOrder(hostPort)); if (true) //5 { var bufout = new byte[128]; var buflen = 0; ns.Write(new byte[] { 5, 1, 0 }, 0, 3); buflen = ns.Read(bufout, 0, bufout.Length); if (buflen != 2 || bufout[0] != 5 || bufout[1] != 0) throw new Exception(); var buf = new byte[] { 5, 1, 0, 3, (byte)HostNameBuf.Length }; var mem = new MemoryStream(); mem.Write(buf, 0, buf.Length); mem.Write(HostNameBuf, 0, HostNameBuf.Length); mem.Write(new byte[] { HostPortBuf[0], HostPortBuf[1] }, 0, 2); var memarr = mem.ToArray(); ns.Write(memarr, 0, memarr.Length); buflen = ns.Read(bufout, 0, bufout.Length); if (bufout[0] != 5 || bufout[1] != 0) throw new Exception(); } else //4a { var bufout = new byte[128]; var buflen = 0; var mem = new MemoryStream(); mem.WriteByte(4); mem.WriteByte(1); mem.Write(HostPortBuf, 0, 2); mem.Write(BitConverter.GetBytes(IPAddress.HostToNetworkOrder(1)), 0, 4); mem.WriteByte(0); mem.Write(HostNameBuf, 0, HostNameBuf.Length); mem.WriteByte(0); var memarr = mem.ToArray(); ns.Write(memarr, 0, memarr.Length); buflen = ns.Read(bufout, 0, bufout.Length); if (buflen != 8 || bufout[0] != 0 || bufout[1] != 90) throw new Exception(); } return ns; } Usage using (TcpClient client = new TcpClient()) using (var ns = ConnectSocksProxy("127.0.0.1", 9050, "website.com", 80, client)) {...}

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  • Android - Custom Adapter Problem

    - by Ryan
    Hello, I seem to be having a problem with my Custom Adapter view. When I display the list, it only displays a white screen. Here is how it works: 1.) I send a JSON request 2.) populate the ArrayList with the returned results 3.) create a custom adapter 4.) then bind the adapter. Here is steps 2-4 private void updateUI() { ListView myList = (ListView) findViewById(android.R.id.list); itemList = new ArrayList(); Iterator it = data.entrySet().iterator(); while (it.hasNext()) { //Get the key name and value for it Map.Entry pair = (Map.Entry)it.next(); String keyName = (String) pair.getKey(); String value = pair.getValue().toString(); if (value != null) { ListItem li = new ListItem(keyName, value, false); itemList.add(li); } } CustomAdapter mAdapter = new CustomAdapter( mContext, itemList); myList.setAdapter(mAdapter); //Bind the adapter to the list //Tell the dialog it's cool now. dismissDialog(0); //Show next screen flipper.setInAnimation(inFromRightAnimation()); flipper.setOutAnimation(outToRightAnimation()); flipper.showNext(); } And here is my CustomAdapter class: import java.util.List; import android.R.color; import android.content.Context; import android.view.View; import android.view.ViewGroup; import android.widget.BaseAdapter; import android.widget.ImageView; import android.widget.RelativeLayout; import android.widget.TextView; class MyAdapterView extends RelativeLayout { public MyAdapterView(Context c, ListItem li) { super( c ); RelativeLayout rL = new RelativeLayout(c); RelativeLayout.LayoutParams containerParams = new RelativeLayout.LayoutParams( ViewGroup.LayoutParams.FILL_PARENT, ViewGroup.LayoutParams.FILL_PARENT); rL.setLayoutParams(containerParams); rL.setBackgroundColor(color.white); ImageView img = new ImageView (c); img.setImageResource(li.getImage()); img.setPadding(5, 5, 10, 5); rL.addView(img, 48, 48); TextView top = new TextView(c); top.setText(li.getTopText()); top.setTextColor(color.black); top.setTextSize(20); top.setPadding(0, 20, 0, 0); rL.addView(top,ViewGroup.LayoutParams.FILL_PARENT, ViewGroup.LayoutParams.WRAP_CONTENT); TextView bot = new TextView( c ); bot.setText(li.getBottomText()); bot.setTextColor(color.black); bot.setTextSize(12); bot.setPadding(0, 0, 0, 10); bot.setAutoLinkMask(1); rL.addView(bot,ViewGroup.LayoutParams.FILL_PARENT, ViewGroup.LayoutParams.WRAP_CONTENT); } } public class CustomAdapter extends BaseAdapter { private Context context; private List itemList; public CustomAdapter(Context c, List itemL ) { this.context = c; this.itemList = itemL; } public int getCount() { return itemList.size(); } public Object getItem(int position) { return itemList.get(position); } public long getItemId(int position) { return position; } @Override public View getView(int position, View convertView, ViewGroup parent) { ListItem li = itemList.get(position); return new MyAdapterView(this.context, li); } } Does anyone have any idea why this displays a white screen upon completion?? Thanks in advance!

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  • Streaming a webcam from Silverlight 4 (Beta)

    - by Ken Smith
    The new webcam stuff in Silverlight 4 is darned cool. By exposing it as a brush, it allows scenarios that are way beyond anything that Flash has. At the same time, accessing the webcam locally seems like it's only half the story. Nobody buys a webcam so they can take pictures of themselves and make funny faces out of them. They buy a webcam because they want other people to see the resulting video stream, i.e., they want to stream that video out to the Internet, a lay Skype or any of the dozens of other video chat sites/applications. And so far, I haven't figured out how to do that with It turns out that it's pretty simple to get a hold of the raw (Format32bppArgb formatted) bytestream, as demonstrated here. But unless we want to transmit that raw bytestream to a server (which would chew up way too much bandwidth), we need to encode that in some fashion. And that's more complicated. MS has implemented several codecs in Silverlight, but so far as I can tell, they're all focused on decoding a video stream, not encoding it in the first place. And that's apart from the fact that I can't figure out how to get direct access to, say, the H.264 codec in the first place. There are a ton of open-source codecs (for instance, in the ffmpeg project here), but they're all written in C, and they don't look easy to port to C#. Unless translating 10000+ lines of code that look like this is your idea of fun :-) const int b_xy= h->mb2b_xy[left_xy[i]] + 3; const int b8_xy= h->mb2b8_xy[left_xy[i]] + 1; *(uint32_t*)h->mv_cache[list][cache_idx ]= *(uint32_t*)s->current_picture.motion_val[list][b_xy + h->b_stride*left_block[0+i*2]]; *(uint32_t*)h->mv_cache[list][cache_idx+8]= *(uint32_t*)s->current_picture.motion_val[list][b_xy + h->b_stride*left_block[1+i*2]]; h->ref_cache[list][cache_idx ]= s->current_picture.ref_index[list][b8_xy + h->b8_stride*(left_block[0+i*2]>>1)]; h->ref_cache[list][cache_idx+8]= s->current_picture.ref_index[list][b8_xy + h->b8_stride*(left_block[1+i*2]>>1)]; The mooncodecs folder within the Mono project (here) has several audio codecs in C# (ADPCM and Ogg Vorbis), and one video codec (Dirac), but they all seem to implement just the decode portion of their respective formats, as do the java implementations from which they were ported. I found a C# codec for Ogg Theora (csTheora, http://www.wreckedgames.com/forum/index.php?topic=1053.0), but again, it's decode only, as is the jheora codec on which it's based. Of course, it would presumably be easier to port a codec from Java than from C or C++, but the only java video codecs that I found were decode-only (such as jheora, or jirac). So I'm kinda back at square one. It looks like our options for hooking up a webcam (or microphone) through Silverlight to the Internet are: (1) Wait for Microsoft to provide some guidance on this; (2) Spend the brain cycles porting one of the C or C++ codecs over to Silverlight-compatible C#; (3) Send the raw, uncompressed bytestream up to a server (or perhaps compressed slightly with something like zlib), and then encode it server-side; or (4) Wait for someone smarter than me to figure this out and provide a solution. Does anybody else have any better guidance? Have I missed something that's just blindingly obvious to everyone else? (For instance, does Silverlight 4 somewhere have some classes I've missed that take care of this?)

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  • .NET 4.0 Dynamic object used statically?

    - by Kevin Won
    I've gotten quite sick of XML configuration files in .NET and want to replace them with a format that is more sane. Therefore, I'm writing a config file parser for C# applications that will take a custom config file format, parse it, and create a Python source string that I can then execute in C# and use as a static object (yes that's right--I want a static (not the static type dyanamic) object in the end). Here's an example of what my config file looks like: // my custom config file format GlobalName: ExampleApp Properties { ExternalServiceTimeout: "120" } Python { // this allows for straight python code to be added to handle custom config def MyCustomPython: return "cool" } Using ANTLR I've created a Lexer/Parser that will convert this format to a Python script. So assume I have that all right and can take the .config above and run my Lexer/Parser on it to get a Python script out the back (this has the added benefit of giving me a validation tool for my config). By running the resultant script in C# // simplified example of getting the dynamic python object in C# // (not how I really do it) ScriptRuntime py = Python.CreateRuntime(); dynamic conf = py.UseFile("conftest.py"); dynamic t = conf.GetConfTest("test"); I can get a dynamic object that has my configuration settings. I can now get my config file settings in C# by invoking a dynamic method on that object: //C# calling a method on the dynamic python object var timeout = t.GetProperty("ExternalServiceTimeout"); //the config also allows for straight Python scripting (via the Python block) var special = t.MyCustonPython(); of course, I have no type safety here and no intellisense support. I have a dynamic representation of my config file, but I want a static one. I know what my Python object's type is--it is actually newing up in instance of a C# class. But since it's happening in python, it's type is not the C# type, but dynamic instead. What I want to do is then cast the object back to the C# type that I know the object is: // doesn't work--can't cast a dynamic to a static type (nulls out) IConfigSettings staticTypeConfig = t as IConfigSettings Is there any way to figure out how to cast the object to the static type? I'm rather doubtful that there is... so doubtful that I took another approach of which I'm not entirely sure about. I'm wondering if someone has a better way... So here's my current tactic: since I know the type of the python object, I am creating a C# wrapper class: public class ConfigSettings : IConfigSettings that takes in a dynamic object in the ctor: public ConfigSettings(dynamic settings) { this.DynamicProxy = settings; } public dynamic DynamicProxy { get; private set; } Now I have a reference to the Python dynamic object of which I know the type. So I can then just put wrappers around the Python methods that I know are there: // wrapper access to the underlying dynamic object // this makes my dynamic object appear 'static' public string GetSetting(string key) { return this.DynamicProxy.GetProperty(key).ToString(); } Now the dynamic object is accessed through this static proxy and thus can obviously be passed around in the static C# world via interface, etc: // dependency inject the dynamic object around IBusinessLogic logic = new BusinessLogic(IConfigSettings config); This solution has the benefits of all the static typing stuff we know and love while at the same time giving me the option of 'bailing out' to dynamic too: // the DynamicProxy property give direct access to the dynamic object var result = config.DynamicProxy.MyCustomPython(); but, man, this seems rather convoluted way of getting to an object that is a static type in the first place! Since the whole dynamic/static interaction world is new to me, I'm really questioning if my solution is optimal or if I'm missing something (i.e. some way of casting that dynamic object to a known static type) about how to bridge the chasm between these two universes.

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  • Core Data Model Design Question - Changing "Live" Objects also Changes Saved Objects

    - by mwt
    I'm working on my first Core Data project (on iPhone) and am really liking it. Core Data is cool stuff. I am, however, running into a design difficulty that I'm not sure how to solve, although I imagine it's a fairly common situation. It concerns the data model. For the sake of clarity, I'll use an imaginary football game app as an example to illustrate my question. Say that there are NSMO's called Downs and Plays. Plays function like templates to be used by Downs. The user creates Plays (for example, Bootleg, Button Hook, Slant Route, Sweep, etc.) and fills in the various properties. Plays have a to-many relationship with Downs. For each Down, the user decides which Play to use. When the Down is executed, it uses the Play as its template. After each down is run, it is stored in history. The program remembers all the Downs ever played. So far, so good. This is all working fine. The question I have concerns what happens when the user wants to change the details of a Play. Let's say it originally involved a pass to the left, but the user now wants it to be a pass to the right. Making that change, however, not only affects all the future executions of that Play, but also changes the details of the Plays stored in history. The record of Downs gets "polluted," in effect, because the Play template has been changed. I have been rolling around several possible fixes to this situation, but I imagine the geniuses of SO know much more about how to handle this than I do. Still, the potential fixes I've come up with are: 1) "Versioning" of Plays. Each change to a Play template actually creates a new, separate Play object with the same name (as far as the user can tell). Underneath the hood, however, it is actually a different Play. This would work, AFAICT, but seems like it could potentially lead to a wild proliferation of Play objects, esp. if the user keeps switching back and forth between several versions of the same Play (creating object after object each time the user switches). Yes, the app could check for pre-existing, identical Plays, but... it just seems like a mess. 2) Have Downs, upon saving, record the details of the Play they used, but not as a Play object. This just seems ridiculous, given that the Play object is there to hold those just those details. 3) Recognize that Play objects are actually fulfilling 2 functions: one to be a template for a Down, and the other to record what template was used. These 2 functions have a different relationship with a Down. The first (template) has a to-many relationship. But the second (record) has a one-to-one relationship. This would mean creating a second object, something like "Play-Template" which would retain the to-many relationship with Downs. Play objects would get reconfigured to have a one-to-one relationship with Downs. A Down would use a Play-Template object for execution, but use the new kind of Play object to store what template was used. It is this change from a to-many relationship to a one-to-one relationship that represents the crux of the problem. Even writing this question out has helped me get clearer. I think something like solution 3 is the answer. However if anyone has a better idea or even just a confirmation that I'm on the right track, that would be helpful. (Remember, I'm not really making a football game, it's just faster/easier to use a metaphor everyone understands.) Thanks.

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  • Calculate a set of concatenated sets of n sets

    - by Andras Zoltan
    Okay - I'm not even sure that the term is right - and I'm sure there is bound to be a term for this - but I'll do my best to explain. This is not quite a cross product here, and the order of the results are absolutely crucial. Given: IEnumerable<IEnumerable<string>> sets = new[] { /* a */ new[] { "a", "b", "c" }, /* b */ new[] { "1", "2", "3" }, /* c */ new[] { "x", "y", "z" } }; Where each inner enumerable represents an instruction to produce a set of concatenations as follows (the order here is important): set a* = new string[] { "abc", "ab", "a" }; set b* = new string[] { "123", "12", "1" }; set c* = new string[] { "xyz", "xy", "x" }; I want to produce set ordered concatenations as follows: set final = new string { a*[0] + b*[0] + c*[0], /* abc123xyz */ a*[0] + b*[0] + c*[1], /* abc123xy */ a*[0] + b*[0] + c*[2], /* abc123x */ a*[0] + b*[0], /* abc123 */ a*[0] + b*[1] + c*[0], /* abc12xyz */ a*[0] + b*[1] + c*[1], /* abc12xy */ a*[0] + b*[1] + c*[2], /* abc12x */ a*[0] + b*[1], /* abc12 */ a*[0] + b*[2] + c*[0], /* abc1xyz */ a*[0] + b*[2] + c*[1], /* abc1xy */ a*[0] + b*[2] + c*[2], /* abc1x */ a*[0] + b*[2], /* abc1 */ a*[0], /* abc */ a*[1] + b*[0] + c*[0], /* ab123xyz */ /* and so on for a*[1] */ /* ... */ a*[2] + b*[0] + c*[0], /* a123xyz */ /* and so on for a*[2] */ /* ... */ /* now lop off a[*] and start with b + c */ b*[0] + c*[0], /* 123xyz */ /* rest of the combinations of b + c with b on its own as well */ /* then finally */ c[0], c[1], c[2]}; So clearly, there are going to be a lot of combinations! I can see similarities with Numeric bases (since the order is important as well), and I'm sure there are permutations/combinations lurking in here too. The question is - how to write an algorithm like this that'll cope with any number of sets of strings? Linq, non-Linq; I'm not fussed. Why am I doing this? Indeed, why!? In Asp.Net MVC - I want to have partial views that can be redefined for a given combination of back-end/front-end culture and language. The most basic of these would be, for a given base view View, we could have View-en-GB, View-en, View-GB, and View, in that order of precedence (recognising of course that the language/culture codes could be the same, so some combinations might be the same - a Distinct() will solve that). But I also have other views that, in themselves, have other possible combinations before culture is even taken into account (too long to go into - but the fact is, this algo will enable a whole bunch of really cool that I want to offer my developers!). I want to produce a search list of all the acceptable view names, iterate through the whole lot until the most specific match is found (governed by the order that this algo will produce these concatenations in) then serve up the resolved Partial View. The result of the search can later be cached to avoid the expense of running the algorithm all the time. I already have a really basic version of this working that just has one enumerable of strings. But this is a whole different kettle of seafood! Any help greatly appreciated.

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  • Can't access a map member from a pointer

    - by fjfnaranjo
    Hi. That's my first question :) I'm storing the configuration of my program in a Group->Key->Value form, like the old INIs. I'm storing the information in a pair of structures. First one, I'm using a std::map with string+ptr for the groups info (the group name in the string key). The second std::map value is a pointer to the sencond structure, a std::list of std::maps, with the finish Key->Value pairs. The Key-Value pairs structure is created dynamically, so the config structure is: std::map< std::string , std::list< std::map<std::string,std::string> >* > lv1; Well, I'm trying to implement two methods to check the existence of data in the internal config. The first one, check the existence of a group in the structure: bool isConfigLv1(std::string); bool ConfigManager::isConfigLv1(std::string s) { return !(lv1.find(s)==lv1.end()); } The second method, is making me crazy... It check the existence for a key inside a group. bool isConfigLv2(std::string,std::string); bool ConfigManager::isConfigLv2(std::string s,std::string d) { if(!isConfigLv1(s)) return false; std::map< std::string , std::list< std::map<std::string,std::string> >* >::iterator it; std::list< std::map<std::string,std::string> >* keyValue; std::list< std::map<std::string,std::string> >::iterator keyValueIt; it = lv1.find(s); keyValue = (*it).second; for ( keyValueIt = keyValue->begin() ; keyValueIt != keyValue->end() ; keyValueIt++ ) if(!((*keyValueIt).second.find(d)==(*keyValueIt).second.end())) return true; return false; } I don't understand what is wrong. The compiler says: ConfigManager.cpp||In member function ‘bool ConfigManager::isConfigLv2(std::string, std::string)’:| ConfigManager.cpp|(line over return true)|error: ‘class std::map<std::basic_string<char, std::char_traits<char>, std::allocator<char> >, std::basic_string<char, std::char_traits<char>, std::allocator<char> >, std::less<std::basic_string<char, std::char_traits<char>, std::allocator<char> > >, std::allocator<std::pair<const std::basic_string<char, std::char_traits<char>, std::allocator<char> >, std::basic_string<char, std::char_traits<char>, std::allocator<char> > > > >’ has no member named ‘second’| But it has to have the second member, because it's a map iterator... Any suggestion about what's happening? Sorry for my English :P, and consider I'm doing it as a exercise, I know there are a lot of cool configuration managers.

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  • Python: Improving long cumulative sum

    - by Bo102010
    I have a program that operates on a large set of experimental data. The data is stored as a list of objects that are instances of a class with the following attributes: time_point - the time of the sample cluster - the name of the cluster of nodes from which the sample was taken code - the name of the node from which the sample was taken qty1 = the value of the sample for the first quantity qty2 = the value of the sample for the second quantity I need to derive some values from the data set, grouped in three ways - once for the sample as a whole, once for each cluster of nodes, and once for each node. The values I need to derive depend on the (time sorted) cumulative sums of qty1 and qty2: the maximum value of the element-wise sum of the cumulative sums of qty1 and qty2, the time point at which that maximum value occurred, and the values of qty1 and qty2 at that time point. I came up with the following solution: dataset.sort(key=operator.attrgetter('time_point')) # For the whole set sys_qty1 = 0 sys_qty2 = 0 sys_combo = 0 sys_max = 0 # For the cluster grouping cluster_qty1 = defaultdict(int) cluster_qty2 = defaultdict(int) cluster_combo = defaultdict(int) cluster_max = defaultdict(int) cluster_peak = defaultdict(int) # For the node grouping node_qty1 = defaultdict(int) node_qty2 = defaultdict(int) node_combo = defaultdict(int) node_max = defaultdict(int) node_peak = defaultdict(int) for t in dataset: # For the whole system ###################################################### sys_qty1 += t.qty1 sys_qty2 += t.qty2 sys_combo = sys_qty1 + sys_qty2 if sys_combo > sys_max: sys_max = sys_combo # The Peak class is to record the time point and the cumulative quantities system_peak = Peak(time_point=t.time_point, qty1=sys_qty1, qty2=sys_qty2) # For the cluster grouping ################################################## cluster_qty1[t.cluster] += t.qty1 cluster_qty2[t.cluster] += t.qty2 cluster_combo[t.cluster] = cluster_qty1[t.cluster] + cluster_qty2[t.cluster] if cluster_combo[t.cluster] > cluster_max[t.cluster]: cluster_max[t.cluster] = cluster_combo[t.cluster] cluster_peak[t.cluster] = Peak(time_point=t.time_point, qty1=cluster_qty1[t.cluster], qty2=cluster_qty2[t.cluster]) # For the node grouping ##################################################### node_qty1[t.node] += t.qty1 node_qty2[t.node] += t.qty2 node_combo[t.node] = node_qty1[t.node] + node_qty2[t.node] if node_combo[t.node] > node_max[t.node]: node_max[t.node] = node_combo[t.node] node_peak[t.node] = Peak(time_point=t.time_point, qty1=node_qty1[t.node], qty2=node_qty2[t.node]) This produces the correct output, but I'm wondering if it can be made more readable/Pythonic, and/or faster/more scalable. The above is attractive in that it only loops through the (large) dataset once, but unattractive in that I've essentially copied/pasted three copies of the same algorithm. To avoid the copy/paste issues of the above, I tried this also: def find_peaks(level, dataset): def grouping(object, attr_name): if attr_name == 'system': return attr_name else: return object.__dict__[attrname] cuml_qty1 = defaultdict(int) cuml_qty2 = defaultdict(int) cuml_combo = defaultdict(int) level_max = defaultdict(int) level_peak = defaultdict(int) for t in dataset: cuml_qty1[grouping(t, level)] += t.qty1 cuml_qty2[grouping(t, level)] += t.qty2 cuml_combo[grouping(t, level)] = (cuml_qty1[grouping(t, level)] + cuml_qty2[grouping(t, level)]) if cuml_combo[grouping(t, level)] > level_max[grouping(t, level)]: level_max[grouping(t, level)] = cuml_combo[grouping(t, level)] level_peak[grouping(t, level)] = Peak(time_point=t.time_point, qty1=node_qty1[grouping(t, level)], qty2=node_qty2[grouping(t, level)]) return level_peak system_peak = find_peaks('system', dataset) cluster_peak = find_peaks('cluster', dataset) node_peak = find_peaks('node', dataset) For the (non-grouped) system-level calculations, I also came up with this, which is pretty: dataset.sort(key=operator.attrgetter('time_point')) def cuml_sum(seq): rseq = [] t = 0 for i in seq: t += i rseq.append(t) return rseq time_get = operator.attrgetter('time_point') q1_get = operator.attrgetter('qty1') q2_get = operator.attrgetter('qty2') timeline = [time_get(t) for t in dataset] cuml_qty1 = cuml_sum([q1_get(t) for t in dataset]) cuml_qty2 = cuml_sum([q2_get(t) for t in dataset]) cuml_combo = [q1 + q2 for q1, q2 in zip(cuml_qty1, cuml_qty2)] combo_max = max(cuml_combo) time_max = timeline.index(combo_max) q1_at_max = cuml_qty1.index(time_max) q2_at_max = cuml_qty2.index(time_max) However, despite this version's cool use of list comprehensions and zip(), it loops through the dataset three times just for the system-level calculations, and I can't think of a good way to do the cluster-level and node-level calaculations without doing something slow like: timeline = defaultdict(int) cuml_qty1 = defaultdict(int) #...etc. for c in cluster_list: timeline[c] = [time_get(t) for t in dataset if t.cluster == c] cuml_qty1[c] = [q1_get(t) for t in dataset if t.cluster == c] #...etc. Does anyone here at Stack Overflow have suggestions for improvements? The first snippet above runs well for my initial dataset (on the order of a million records), but later datasets will have more records and clusters/nodes, so scalability is a concern. This is my first non-trivial use of Python, and I want to make sure I'm taking proper advantage of the language (this is replacing a very convoluted set of SQL queries, and earlier versions of the Python version were essentially very ineffecient straight transalations of what that did). I don't normally do much programming, so I may be missing something elementary. Many thanks!

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  • My google map android app keeps crashing

    - by Manny264
    I have followed about two tutorials from vogella and some other tutorial that looked similar...very similar but to no avail. I load the app on my nexus 7 and it just crashes "Unfortunately MyMapView has stopped working" on launch.This is the manifest: ` <uses-sdk android:minSdkVersion="17" android:targetSdkVersion="17" /> <uses-feature android:glEsVersion="0x00020000" android:required="true"/> <uses-permission android:name="android.permission.INTERNET" /> <uses-permission android:name="android.permission.ACCESS_NETWORK_STATE" /> <uses-permission android:name="android.permission.ACCESS_COARSE_LOCATION" /> <uses-permission android:name="android.permission.ACCESS_FINE_LOCATION" /> <uses-permission android:name="android.permission.WRITE_EXTERNAL_STORAGE"/> <uses-permission android:name="com.google.android.providers.gsf.permission.READ_GSERVICES"/> <application android:allowBackup="true" android:icon="@drawable/ic_launcher" android:label="@string/app_name" android:theme="@style/AppTheme" > <uses-library android:name="com.google.android.maps" /> <activity android:name="com.macmozart.mymapview.MainActivity" android:label="@string/app_name" > <intent-filter> <action android:name="android.intent.action.MAIN" /> <category android:name="android.intent.category.LAUNCHER" /> </intent-filter> </activity> <meta-data android:name="com.google.android.maps.v2.API_KEY" android:value="AIzaSyBZ1Bt7rjB863Jy-B05zls6k8XZsBGQ6-4" /> </application> ` Followed by my main layout: <fragment android:id="@+id/map" android:layout_width="match_parent" android:layout_height="match_parent" class="com.google.android.gms.maps.MapFragment" /> and finally my java class: package com.macmozart.mymapview; import android.app.Activity; import android.os.Bundle; import com.google.android.gms.maps.CameraUpdateFactory; import com.google.android.gms.maps.GoogleMap; import com.google.android.gms.maps.MapFragment; import com.google.android.gms.maps.model.BitmapDescriptorFactory; import com.google.android.gms.maps.model.LatLng; import com.google.android.gms.maps.model.Marker; import com.google.android.gms.maps.model.MarkerOptions; import com.google.android.maps.*; public class MainActivity extends Activity { static final LatLng HAMBURG = new LatLng(53.558, 9.927); static final LatLng KIEL = new LatLng(53.551, 9.993); private GoogleMap map; @Override protected void onCreate(Bundle savedInstanceState) { super.onCreate(savedInstanceState); setContentView(R.layout.activity_main); map = ((MapFragment) getFragmentManager().findFragmentById(R.id.map)) .getMap(); if (map != null) { Marker hamburg = map.addMarker(new MarkerOptions() .position(HAMBURG).title("Hamburg")); Marker kiel = map.addMarker(new MarkerOptions() .position(KIEL) .title("Kiel") .snippet("Kiel is cool") .icon(BitmapDescriptorFactory .fromResource(R.drawable.ic_launcher))); map.moveCamera(CameraUpdateFactory.newLatLngZoom(HAMBURG, 15)); // Zoom in, animating the camera. map.animateCamera(CameraUpdateFactory.zoomTo(10), 2000, null); } } } Any idea what im doing wrong I really need this to work

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  • Writing a code example

    - by Stefano Borini
    I would like to have your feedback regarding code examples. One of the most frustrating experiences I sometimes have when learning a new technology is finding useless examples. I think an example as the most precious thing that comes with a new library, language, or technology. It must be a starting point, a wise and unadulterated explanation on how to achieve a given result. A perfect example must have the following characteristics: Self contained: it should be small enough to be compiled or executed as a single program, without dependencies or complex makefiles. An example is also a strong functional test if you correctly installed the new technology. The more issues could arise, the more likely is that something goes wrong, and the more difficult is to debug and solve the situation. Pertinent: it should demonstrate one, and only one, specific feature of your software/library, involving the minimal additional behavior from external libraries. Helpful: the code should bring you forward, step by step, using comments or self-documenting code. Extensible: the example code should be a small “framework” or blueprint for additional tinkering. A learner can start by adding features to this blueprint. Recyclable: it should be possible to extract parts of the example to use in your own code Easy: An example code is not the place to show your code-fu skillz. Keep it easy. helpful acronym: SPHERE. Prototypical examples of violations of those rules are the following: Violation of self-containedness: an example spanning multiple files without any real need for it. If your example is a python program, keep everything into a single module file. Don’t sub-modularize it. In Java, try to keep everything into a single class, unless you really must partition some entity into a meaningful object you need to pass around (and java mandates one class per file, if I remember correctly). Violation of Pertinency: When showing how many different shapes you can draw, adding radio buttons and complex controls with all the possible choices for point shapes is a bad idea. You de-focalize your example code, introducing code for event handling, controls initialization etc., and this is not part the feature you want to demonstrate, they are unnecessary noise in the understanding of the crucial mechanisms providing the feature. Violation of Helpfulness: code containing dubious naming, wrong comments, hacks, and functions longer than one page of code. Violation of Extensibility: badly factored code that have everything into a single function, with potentially swappable entities embedded within the code. Example: if an example reads data from a file and displays it, create a method getData() returning a useful entity, instead of opening the file raw and plotting the stuff. This way, if the user of the library needs to read data from a HTTP server instead, he just has to modify the getData() module and use the example almost as-is. Another violation of Extensibility comes if the example code is not under a fully liberal (e.g. MIT or BSD) license. Violation of Recyclability: when the code layout is so intermingled that is difficult to easily copy and paste parts of it and recycle them into another program. Again, licensing is also a factor. Violation of Easiness: Yes, you are a functional-programming nerd and want to show how cool you are by doing everything on a single line of map, filter and so on, but that could not be helpful to someone else, who is already under pressure to understand your library, and now has to understand your code as well. And in general, the final rule: if it takes more than 10 minutes to do the following: compile the code, run it, read the source, and understand it fully, it means that the example is not a good one. Please let me know your opinion, either positive or negative, or experience on this regard.

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  • Qt, MSVC, and /Zc:wchar_t- == I want to blow up the world

    - by Noah Roberts
    So Qt is compiled with /Zc:wchar_t- on windows. What this means is that instead of wchar_t being a typedef for some internal type (__wchar_t I think) it becomes a typedef for unsigned short. The really cool thing about this is that the default for MSVC is the opposite, which of course means that the libraries you're using are likely compiled with wchar_t being a different type than Qt's wchar_t. This doesn't become an issue of course until you try to use something like std::wstring in your code; especially when one or more libraries have functions that accept it as parameters. What effectively happens is that your code happily compiles but then fails to link because it's looking for definitions using std::wstring<unsigned short...> but they only contain definitions expecting std::wstring<__wchar_t...> (or whatever). So I did some web searching and ran into this link: http://bugreports.qt.nokia.com/browse/QTBUG-6345 Based on the statement by Thiago Macieira, "Sorry, we will not support building Qt like this," I've been worried that fixing Qt to work like everything else might cause some problem and have been trying to avoid it. We recompiled all of our support libraries with the /Zc:wchar_t- flag and have been fairly content with that until a couple days ago when we started trying to port over (we're in the process of switching from Wx to Qt) some serialization code. Because of how win32 works, and because Wx just wraps win32, we've been using std::wstring to represent string data with the intent of making our product as i18n ready as possible. We did some testing and Wx did not work with multibyte characters when trying to print special stuff (even not so special stuff like the degree symbol was an issue). I'm not so sure that Qt has this problem since QString isn't just a wrapper to the underlying _TCHAR type but is a Unicode monster of some sort. At any rate, the serialization library in boost has compiled parts. We've attempted to recompile boost with /Zc:wchar_t- but so far our attempts to tell bjam to do this have gone unheeded. We're at an impasse. From where I'm sitting I have three options: Recompile Qt and hope it works with /Zc:wchar_t. There's some evidence around the web that others have done this but I have no way of predicting what will happen. All attempts to ask Qt people on forums and such have gone unanswered. Hell, even in that very bug report someone asks why and it just sat there for a year. Keep fighting with bjam until it listens. Right now I've got someone under me doing that and I have more experience fighting with things to get what I want but I do have to admit to getting rather tired of it. I'm also concerned that I'll KEEP running into this issue just because Qt wants to be a c**t. Stop using wchar_t for anything. Unfortunately my i18n experience is pretty much 0 but it seems to me that I just need to find the right to/from function in QString (it has a BUNCH) to encode the Unicode into 8-bytes and visa-versa. UTF8 functions look promising but I really want to be sure that no data will be lost if someone from Zimbabfuckegypt starts writing in their own language and the documentation in QString frightens me a little into thinking that could happen. Of course, I could always run into some library that insists I use wchar_t and then I'm back to 1 or 2 but I rather doubt that would happen. So, what's my question... Which of these options is my best bet? Is Qt going to eventually cause me to gouge out my own eyes because I decided to compile it with /Zc:wchar_t anyway? What's the magic incantation to get boost to build with /Zc:wchar_t- and will THAT cause permanent mental damage? Can I get away with just using the standard 8-bit (well, 'common' anyway) character classes and be i18n compliant/ready? How do other Qt developers deal with this mess?

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  • How do you stop scripters from slamming your website hundreds of times a second?

    - by davebug
    [update] I've accepted an answer, as lc deserves the bounty due to the well thought-out answer, but sadly, I believe we're stuck with our original worst case scenario: CAPTCHA everyone on purchase attempts of the crap. Short explanation: caching / web farms make it impossible for us to actually track hits, and any workaround (sending a non-cached web-beacon, writing to a unified table, etc.) slows the site down worse than the bots would. There is likely some pricey bit of hardware from Cisco or the like that can help at a high level, but it's hard to justify the cost if CAPTCHAing everyone is an alternative. I'll attempt to do a more full explanation in here later, as well as cleaning this up for future searchers (though others are welcome to try, as it's community wiki). I've added bounty to this question and attempted to explain why the current answers don't fit our needs. First, though, thanks to all of you who have thought about this, it's amazing to have this collective intelligence to help work through seemingly impossible problems. I'll be a little more clear than I was before: This is about the bag o' crap sales on woot.com. I'm the president of Woot Workshop, the subsidiary of Woot that does the design, writes the product descriptions, podcasts, blog posts, and moderates the forums. I work in the css/html world and am only barely familiar with the rest of the developer world. I work closely with the developers and have talked through all of the answers here (and many other ideas we've had). Usability of the site is a massive part of my job, and making the site exciting and fun is most of the rest of it. That's where the three goals below derive. CAPTCHA harms usability, and bots steal the fun and excitement out of our crap sales. To set up the scenario a little more, bots are slamming our front page tens of times a second screenscraping (and/or scanning our rss) for the Random Crap sale. The moment they see that, it triggers a second stage of the program that logs in, clicks I want One, fills out the form, and buys the crap. In current (2/6/2009) order of votes: lc: On stackoverflow and other sites that use this method, they're almost always dealing with authenticated (logged in) users, because the task being attempted requires that. On Woot, anonymous (non-logged) users can view our home page. In other words, the slamming bots can be non-authenticated (and essentially non-trackable except by IP address). So we're back to scanning for IPs, which a) is fairly useless in this age of cloud networking and spambot zombies and b) catches too many innocents given the number of businesses that come from one IP address (not to mention the issues with non-static IP ISPs and potential performance hits to trying to track this). Oh, and having people call us would be the worst possible scenario. Can we have them call you? BradC Ned Batchelder's methods look pretty cool, but they're pretty firmly designed to defeat bots built for a network of sites. Our problem is bots are built specifically to defeat our site. Some of these methods could likely work for a short time until the scripters evolved their bots to ignore the honeypot, screenscrape for nearby label names instead of form ids, and use a javascript-capable browser control. lc again "Unless, of course, the hype is part of you

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  • C++ MySQL++ Delete query statement brain killer question

    - by shauny
    Hello all, I'm relatively new to the MySQL++ connector in C++, and have an really annoying issue with it already! I've managed to get stored procedures working, however i'm having issues with the delete statements. I've looked high and low and have found no documentation with examples. First I thought maybe the code needs to free the query/connection results after calling the stored procedure, but of course MySQL++ doesn't have a free_result method... or does it? Anyways, here's what I've got: #include <iostream> #include <stdio.h> #include <queue> #include <deque> #include <sys/stat.h> #include <mysql++/mysql++.h> #include <boost/thread/thread.hpp> #include "RepositoryQueue.h" using namespace boost; using namespace mysqlpp; class RepositoryChecker { private: bool _isRunning; Connection _con; public: RepositoryChecker() { try { this->_con = Connection(false); this->_con.set_option(new MultiStatementsOption(true)); this->_con.set_option(new ReconnectOption(true)); this->_con.connect("**", "***", "***", "***"); this->ChangeRunningState(true); } catch(const Exception& e) { this->ChangeRunningState(false); } } /** * Thread method which runs and creates the repositories */ void CheckRepositoryQueues() { //while(this->IsRunning()) //{ std::queue<RepositoryQueue> queues = this->GetQueue(); if(queues.size() > 0) { while(!queues.empty()) { RepositoryQueue &q = queues.front(); char cmd[256]; sprintf(cmd, "svnadmin create /home/svn/%s/%s/%s", q.GetPublicStatus().c_str(), q.GetUsername().c_str(), q.GetRepositoryName().c_str()); if(this->DeleteQueuedRepository(q.GetQueueId())) { printf("query deleted?\n"); } printf("Repository created!\n"); queues.pop(); } } boost::this_thread::sleep(boost::posix_time::milliseconds(500)); //} } protected: /** * Gets the latest queue of repositories from the database * and returns them inside a cool queue defined with the * RepositoryQueue class. */ std::queue<RepositoryQueue> GetQueue() { std::queue<RepositoryQueue> queues; Query query = this->_con.query("CALL sp_GetRepositoryQueue();"); StoreQueryResult result = query.store(); RepositoryQueue rQ; if(result.num_rows() > 0) { for(unsigned int i = 0;i < result.num_rows(); ++i) { rQ = RepositoryQueue((unsigned int)result[i][0], (unsigned int)result[i][1], (String)result[i][2], (String)result[i][3], (String)result[i][4], (bool)result[i][5]); queues.push(rQ); } } return queues; } /** * Allows the thread to be shut off. */ void ChangeRunningState(bool isRunning) { this->_isRunning = isRunning; } /** * Returns the running value of the active thread. */ bool IsRunning() { return this->_isRunning; } /** * Deletes the repository from the mysql queue table. This is * only called once it has been created. */ bool DeleteQueuedRepository(unsigned int id) { char cmd[256]; sprintf(cmd, "DELETE FROM RepositoryQueue WHERE Id = %d LIMIT 1;", id); Query query = this->_con.query(cmd); return (query.exec()); } }; I've removed all the other methods as they're not needed... Basically it's the DeleteQueuedRepository method which isn't working, the GetQueue works fine. PS: This is on a Linux OS (Ubuntu server) Many thanks, Shaun

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  • Change the content of a <style> element through JavaScript

    - by paercebal
    The Problem I have the following code: <html> <head> <style id="ID_Style"> .myStyle { color : #FF0000 ; } </style> </head> <body> <p class="myStyle"> Hello World ! </p> </body> </html> And I want to modify the contents of <style> through JavaScript. The Expected Solution The first solution was to use the innerHTML property of the style element (retrieved through its id), but while it works on Firefox, it fails on Internet Explorer 7. So, I used pure DOM methods, that is, creating an element called style, a text node with the desired content, and append the text node as a child of the style node, etc. It fails, too. According to MSDN, the <style> element has an innerHTML property, and according to W3C, the <style> element is a HTMLStyleElement, which derives from HTMLElement, deriving from Element deriving from Node, which has the appendChild method. It seems to behave as if the content of a <style> element was readonly on Internet Explorer. The Question So the question is: Is there a way to modify the content of a <style> element on Internet Explorer? While the current problem is with IE7, a cross-browser solution would be cool, if possible. Appendix Sources: Style Element (MSDN): http://msdn.microsoft.com/en-us/library/ms535898.aspx HTMLStyleElement (W3C): http://www.w3.org/TR/2003/REC-DOM-Level-2-HTML-20030109/html.html#ID-16428977 Complete Test Code You can use this test code if you want to reproduce your problem: <html> <head> <style id="ID_Style"> .myStyle { color : #FF0000 ; } </style> <script> function replaceStyleViaDOM(p_strContent) { var oOld = document.getElementById("ID_Style") ; var oParent = oOld.parentNode ; oParent.removeChild(oOld) ; var oNew = document.createElement("style") ; oParent.appendChild(oNew) ; oNew.setAttribute("id", "ID_Style") ; var oText = document.createTextNode(p_strContent) ; oNew.appendChild(oText) ; } function replaceStyleViaInnerHTML(p_strContent) { document.getElementById("ID_Style").innerHTML = p_strContent ; } </script> <script> function setRedViaDOM() { replaceStyleViaDOM("\n.myStyle { color : #FF0000 ; }\n") } function setRedViaInnerHTML() { replaceStyleViaInnerHTML("\n.myStyle { color : #FF0000 ; }\n") } function setBlueViaDOM() { replaceStyleViaDOM("\n.myStyle { color : #0000FF ; }\n") } function setBlueViaInnerHTML() { replaceStyleViaInnerHTML("\n.myStyle { color : #0000FF ; }\n") } function alertStyle() { alert("*******************\n" + document.getElementById("ID_Style").innerHTML + "\n*******************") ; } </script> </head> <body> <div> <button type="button" onclick="alertStyle()">alert Style</button> <br /> <button type="button" onclick="setRedViaDOM()">set Red via DOM</button> <button type="button" onclick="setRedViaDOM()">set Red via InnerHTML</button> <br /> <button type="button" onclick="setBlueViaDOM()">set Blue via DOM</button> <button type="button" onclick="setBlueViaInnerHTML()">set Blue via InnerHTML</button> </div> <p class="myStyle"> Hello World ! </p> </body> </html> Thanks !

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  • Proxy Issues with Javascript Cross Domain RSS Feed Parsing

    - by Amir
    This is my Javascript function which grabs an rss feed via the proxy script and then spits out the 5 latest rss items from the feed along with a link to my stylesheet: function getWidget (feed,limit) { if (window.XMLHttpRequest) { xhttp=new XMLHttpRequest() } else { xhttp=new ActiveXObject("Microsoft.XMLHTTP") } xhttp.open("GET","http://MYSITE/proxy.php?url="+feed,false); xhttp.send(""); xmlDoc=xhttp.responseXML; var x = 1; var div = document.getElementById("div"); srdiv.innerHTML = '<link type="text/css" href="http://MYSITE/css/widget.css" rel="stylesheet" /><div id="rss-title"></div></h3><div id="items"></div><br /><br /><a href="http://MYSITE">Powered by MYSITE</a>'; document.body.appendChild(div); content=xmlDoc.getElementsByTagName("title"); thelink=xmlDoc.getElementsByTagName("link"); document.getElementByTagName("rss-title").innerHTML += content[0].childNodes[0].nodeValue; for (x=1;x<=limit;srx++) { y=x; y--; var shout = '<div class="item"><a href="'+thelink[y].childNodes[0].nodeValue+'">'+content[x].childNodes[0].nodeValue+'</a></div>'; document.getElementById("items").innerHTML += shout; } } Here is the the code from proxy.php: $session = curl_init($_GET['url']); // Open the Curl session curl_setopt($session, CURLOPT_HEADER, false); // Don't return HTTP headers curl_setopt($session, CURLOPT_RETURNTRANSFER, true); // Do return the contents of the call $xml = curl_exec($session); // Make the call header("Content-Type: text/xml"); // Set the content type appropriately echo $xml; // Spit out the xml curl_close($session); // And close the session Now when I try to load this on any domain that's not my site nothing loads. I get no JS errors, but I in the Console tab in firebug I get "407 Proxy Authentication Required" So I'm not really sure how to make this work. The goal is to be able to grab the RSS feed, parse it to grab the titles and links and spit it out into some HTML on any website on the web. I"m basically making a simple RSS widget for my site's various RSS feeds. My Javascript is wack Also, I'm really a beginner with Javascript. I know jQuery pretty well, but I wasn't able to use it in this case, because this script will be embeded on any site and I can't really rely on the jQuery library. So I was decided to write some basic Javascript relying on the default XML parsing options available. Any suggestions here would be cool. Thanks! What's with the x and y They way my site creates RSS feeds is that the first title is actually the RSS feed title. The second title is the title of the first item. The first link is the link to the first item. So when using the javascript to get the title, I had to first grab the first title (which is the RSS title) and then start with the second title that being the first title of the item. Sorry for the confusion, but I don't think this is related to my issue. Just wanted to clarify my code.

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  • Simple Detached pThread does not cancel! (cout blocks and interleaves even if mutexed)

    - by Gabriel
    I have a hard problem here, which I can not solve and do not find the right answer on the net: I have created a detached thread with a clean up routing, the problem is that on my Imac and Ubuntu 9.1 (Dual Core). I am not able to correctly cancel the detached thread in the fallowing code: #include <iostream> #include <pthread.h> #include <sched.h> #include <signal.h> #include <time.h> pthread_mutex_t mutex_t; using namespace std; static void cleanup(void *arg){ pthread_mutex_lock(&mutex_t); cout << " doing clean up"<<endl; pthread_mutex_unlock(&mutex_t); } static void *thread(void *aArgument) { pthread_setcancelstate(PTHREAD_CANCEL_ENABLE,NULL); pthread_setcanceltype(PTHREAD_CANCEL_DEFERRED,NULL); pthread_cleanup_push(&cleanup,NULL); int n=0; while(1){ pthread_testcancel(); sched_yield(); n++; pthread_mutex_lock(&mutex_t); cout << " Thread 2: "<< n<<endl; pthread_mutex_unlock(&mutex_t); } pthread_cleanup_pop(0); return NULL; } int main() { pthread_t thread_id; pthread_attr_t attr; pthread_attr_init(&attr); pthread_attr_setdetachstate(&attr,PTHREAD_CREATE_DETACHED); int error; if (pthread_mutex_init(&mutex_t,NULL) != 0) return 1; if (pthread_create(&thread_id, &attr, &(thread) , NULL) != 0) return 1; pthread_mutex_lock(&mutex_t); cout << "waiting 1s for thread...\n" <<endl; pthread_mutex_unlock(&mutex_t); int n =0; while(n<1E3){ pthread_testcancel(); sched_yield(); n++; pthread_mutex_lock(&mutex_t); cout << " Thread 1: "<< n<<endl; pthread_mutex_unlock(&mutex_t); } pthread_mutex_lock(&mutex_t); cout << "canceling thread...\n" <<endl; pthread_mutex_unlock(&mutex_t); if (pthread_cancel(thread_id) == 0) { //This doesn't wait for the thread to exit pthread_mutex_lock(&mutex_t); cout << "detaching thread...\n"<<endl; pthread_mutex_unlock(&mutex_t); pthread_detach(thread_id); while (pthread_kill(thread_id,0)==0) { sched_yield(); } pthread_mutex_lock(&mutex_t); cout << "thread is canceled"; pthread_mutex_unlock(&mutex_t); } pthread_mutex_lock(&mutex_t); cout << "exit"<<endl; pthread_mutex_unlock(&mutex_t); return 0; } When I replace the Cout with printf() i workes to the end "exit" , but with the cout (even locked) the executable hangs after outputting "detaching thread... It would be very cool to know from a Pro, what the problem here is?. Why does this not work even when cout is locked by a mutex!? Thanks a lot for your support!!

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  • CABasicAnimation not scrolling with rest of View

    - by morgman
    All, I'm working on reproducing the pulsing Blue circle effect that the map uses to show your own location... I layered a UIView over a MKMapView. The UIView contains the pulsing animation. I ended up using the following code gleaned from numerous answers here on stackoverflow: CABasicAnimation* pulseAnimation = [CABasicAnimation animationWithKeyPath:@"opacity"]; pulseAnimation.toValue = [NSNumber numberWithFloat: 0.0]; pulseAnimation.duration = 1.0; pulseAnimation.repeatCount = HUGE_VALF; pulseAnimation.autoreverses = YES; pulseAnimation.timingFunction = [CAMediaTimingFunction functionWithName:kCAMediaTimingFunctionEaseInEaseOut]; [self.layer addAnimation:pulseAnimation forKey:@"pulseAnimation"]; CABasicAnimation* resizeAnimation = [CABasicAnimation animationWithKeyPath:@"bounds.size"]; resizeAnimation.toValue = [NSValue valueWithCGSize:CGSizeMake(0.0f, 0.0f)]; resizeAnimation.fillMode = kCAFillModeBoth; resizeAnimation.duration = 1.0; resizeAnimation.repeatCount = HUGE_VALF; resizeAnimation.autoreverses = YES; resizeAnimation.timingFunction = [CAMediaTimingFunction functionWithName:kCAMediaTimingFunctionEaseInEaseOut]; [self.layer addAnimation:resizeAnimation forKey:@"resizeAnimation"]; This does an acceptable job of pulsing/fading a circle I drew in the UIView using: CGContextRef context = UIGraphicsGetCurrentContext(); CGContextSetRGBStrokeColor(context, 1.0, 0.0, 0.0, 0.4); // And draw with a blue fill color CGContextSetRGBFillColor(context, 1.0, 0.0, 0.0, 0.1); // Draw them with a 2.0 stroke width so they are a bit more visible. CGContextSetLineWidth(context, 2.0); CGContextAddEllipseInRect(context, CGRectMake(x-30, y-30, 60.0, 60.0)); CGContextDrawPath(context, kCGPathFillStroke); But I soon found that while this appears to work, as soon as I drag the underlying map to another position the animation screws up... While the position I want highlighted is centered on the screen the animation works ok, once I've dragged the position so it's no longer centered the animation now pulses starting at the center of the screen scaling up to center on the position dragged off center, and back again... A humorous and cool effect, but not what I was looking for. I realize I may have been lucky on several fronts. I'm not sure what I misunderstood. I think the animation is scaling the entire layer which just happens to have my circle drawn in the middle of it. So it works when centered but not when off center. I tried the following gleaned from one of the questions suggested by stackoverflow when I started this question: CABasicAnimation* translateAnimation = [CABasicAnimation animationWithKeyPath:@"position"]; translateAnimation.fromValue = [NSValue valueWithCGPoint:CGPointMake(oldx, oldy )]; translateAnimation.toValue = [NSValue valueWithCGPoint:CGPointMake(x, y )]; // translateAnimation.fillMode = kCAFillModeBoth; translateAnimation.duration = 1.0; translateAnimation.timingFunction = [CAMediaTimingFunction functionWithName:kCAMediaTimingFunctionEaseInEaseOut]; [self.layer addAnimation:translateAnimation forKey:@"translateAnimation"]; But it doesn't help much, when I play with it sometimes it looks like once time it animates in the correct spot after I've moved it offcenter, but then it switches back to the animating from the center point to the new location. Sooo, any suggestions or do I need to provide additional information.

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  • Slow Javascript touch events on Android

    - by oneself
    I'm trying to write a simple html based drawing application (standalone simplified code attached bellow). I've tested this on the following devices: iPad 1 and 2: Works great ASUS T101 running Windows: Works great Samsung Galaxy Tab: Extremely slow and patchy -- unusable. Lenovo IdeaPad K1: Extremely slow and patchy -- unusable. Asus Transformer Prime: Noticeable lag compare with the iPad -- close to usable. The Asus tablet is running ICS, the other android tablets are running 3.1 and 3.2. I tested using the stock Android browser. I also tried the Android Chrome Beta, but that was even worse. My questions is why are the Android tablets so slow? Am I doing something wrong or is it an inherit problem with Android OS or browser, or is there anything I can do about it in my code? multi.html: <html> <body> <style media="screen"> canvas { border: 1px solid #CCC; } </style> <canvas style="" id="draw" height="450" width="922"></canvas> <script class="jsbin" src="jquery.js"></script> <script src="multi.js"></script> </body> </html> multi.js: var CanvasDrawr = function(options) { // grab canvas element var canvas = document.getElementById(options.id), ctxt = canvas.getContext("2d"); canvas.style.width = '100%' canvas.width = canvas.offsetWidth; canvas.style.width = ''; // set props from options, but the defaults are for the cool kids ctxt.lineWidth = options.size || Math.ceil(Math.random() * 35); ctxt.lineCap = options.lineCap || "round"; ctxt.pX = undefined; ctxt.pY = undefined; var lines = [,,]; var offset = $(canvas).offset(); var eventCount = 0; var self = { // Bind click events init: function() { // Set pX and pY from first click canvas.addEventListener('touchstart', self.preDraw, false); canvas.addEventListener('touchmove', self.draw, false); }, preDraw: function(event) { $.each(event.touches, function(i, touch) { var id = touch.identifier; lines[id] = { x : this.pageX - offset.left, y : this.pageY - offset.top, color : 'black' }; }); event.preventDefault(); }, draw: function(event) { var e = event, hmm = {}; eventCount += 1; $.each(event.touches, function(i, touch) { var id = touch.identifier, moveX = this.pageX - offset.left - lines[id].x, moveY = this.pageY - offset.top - lines[id].y; var ret = self.move(id, moveX, moveY); lines[id].x = ret.x; lines[id].y = ret.y; }); event.preventDefault(); }, move: function(i, changeX, changeY) { ctxt.strokeStyle = lines[i].color; ctxt.beginPath(); ctxt.moveTo(lines[i].x, lines[i].y); ctxt.lineTo(lines[i].x + changeX, lines[i].y + changeY); ctxt.stroke(); ctxt.closePath(); return { x: lines[i].x + changeX, y: lines[i].y + changeY }; }, }; return self.init(); }; $(function(){ var drawr = new CanvasDrawr({ id: "draw", size: 5 }); });

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  • Facelet components layouts and javascript

    - by Java Drinker
    Hi all, I have a question regarding the placement of javascript within facelet components. This is more regarding best practice/style than a programming issue, but I feel like all the solutions I have thought of have been hacks at best. Ok here is my scenario: I have a facelet template like so (my faces, and apache Trinidad)... <ui:composition> <f:view locale="#{myLocale}"> <ui:insert name="messageBundles" /><!--Here we add load bundle tags--> <tr:document mode="strict" styleClass="coolStyleDoc"> <f:facet name="metaContainer"> <!--This trinidad defined facet is added to HTML head--> <tr:group> <!-- blah bal my own styles and js common to all --> <ui:insert name="metaData" /> </tr:group> </f:facet> <tr:form usesUpload="#{empty usesUpload ? 'false' : usesUpload}"> <div id="formTemplateHeader"> <ui:insert name="contentHeader" /> </div> <div id="formTemplateContentContainer"> <div id="formTemplateContent"> <ui:insert name="contentBody" /> </div> </div> <div id="formTemplateFooter"> <ui:insert name="contentFooter"> </ui:insert> </div> </tr:form> <!-- etc...---> Now, a facelet that wants to use this template would look like the following: <ui:composition template="/path/to/my/template.jspx"> <ui:define name="bundles"> <custom:loadBundle basename="messagesStuff" var="bundle" /> </ui:define> <ui:define name="metaData"> <script> <!-- cool javascript stuff goes here--> </script> </ui:define> <ui:define name="contentHeader"> <!-- MY HEADING!--> </ui:define> <ui:define name="contentBody"> <!-- MY Body!--> </ui:define> <ui:define name="contentFooter"> <!-- Copyright/footer stuff!--> </ui:define> </ui:composition> All this works quite well, but the problem I have is when I want to use a facelet component inside this page. If the facelet component has its own javascript code (jQuery stuff etc), how can I make it so that that javascript code is included in the header section of the generated html? Any help would be appreciated. Please let me know if this is not clear or something... thanks in advance

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  • Calculate the number of ways to roll a certain number

    - by helloworld
    I'm a high school Computer Science student, and today I was given a problem to: Program Description: There is a belief among dice players that in throwing three dice a ten is easier to get than a nine. Can you write a program that proves or disproves this belief? Have the computer compute all the possible ways three dice can be thrown: 1 + 1 + 1, 1 + 1 + 2, 1 + 1 + 3, etc. Add up each of these possibilities and see how many give nine as the result and how many give ten. If more give ten, then the belief is proven. I quickly worked out a brute force solution, as such int sum,tens,nines; tens=nines=0; for(int i=1;i<=6;i++){ for(int j=1;j<=6;j++){ for(int k=1;k<=6;k++){ sum=i+j+k; //Ternary operators are fun! tens+=((sum==10)?1:0); nines+=((sum==9)?1:0); } } } System.out.println("There are "+tens+" ways to roll a 10"); System.out.println("There are "+nines+" ways to roll a 9"); Which works just fine, and a brute force solution is what the teacher wanted us to do. However, it doesn't scale, and I am trying to find a way to make an algorithm that can calculate the number of ways to roll n dice to get a specific number. Therefore, I started generating the number of ways to get each sum with n dice. With 1 die, there is obviously 1 solution for each. I then calculated, through brute force, the combinations with 2 and 3 dice. These are for two: There are 1 ways to roll a 2 There are 2 ways to roll a 3 There are 3 ways to roll a 4 There are 4 ways to roll a 5 There are 5 ways to roll a 6 There are 6 ways to roll a 7 There are 5 ways to roll a 8 There are 4 ways to roll a 9 There are 3 ways to roll a 10 There are 2 ways to roll a 11 There are 1 ways to roll a 12 Which looks straightforward enough; it can be calculated with a simple linear absolute value function. But then things start getting trickier. With 3: There are 1 ways to roll a 3 There are 3 ways to roll a 4 There are 6 ways to roll a 5 There are 10 ways to roll a 6 There are 15 ways to roll a 7 There are 21 ways to roll a 8 There are 25 ways to roll a 9 There are 27 ways to roll a 10 There are 27 ways to roll a 11 There are 25 ways to roll a 12 There are 21 ways to roll a 13 There are 15 ways to roll a 14 There are 10 ways to roll a 15 There are 6 ways to roll a 16 There are 3 ways to roll a 17 There are 1 ways to roll a 18 So I look at that, and I think: Cool, Triangular numbers! However, then I notice those pesky 25s and 27s. So it's obviously not triangular numbers, but still some polynomial expansion, since it's symmetric. So I take to Google, and I come across this page that goes into some detail about how to do this with math. It is fairly easy(albeit long) to find this using repeated derivatives or expansion, but it would be much harder to program that for me. I didn't quite understand the second and third answers, since I have never encountered that notation or those concepts in my math studies before. Could someone please explain how I could write a program to do this, or explain the solutions given on that page, for my own understanding of combinatorics? EDIT: I'm looking for a mathematical way to solve this, that gives an exact theoretical number, not by simulating dice

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  • Remotely Schedule and Stream Recorded TV in Windows 7 Media Center

    - by DigitalGeekery
    Have you ever been away from home and suddenly realized you forgot to record your favorite program? Now Windows 7 Media Center, users can schedule recordings remotely from their phones or mobile devices with Remote Potato. How it Works Remote Potato installs server software on the host computer running Windows 7 Media Center. Once the software is installed, we’ll need to do some port forwarding on the router and setup an optional dynamic DNS address. When setup is completed, we will access the application through a web based interface. Silverlight is required for Streaming recorded TV, but scheduling recordings can be done through an HTML interface. Installing Remote Potato Download and install Remote Potato on the Media Center PC. (See download link below) If you plan to stream any Recorded TV, you’ll also want to install the streaming pack located on the same page. It isn’t required to stream all shows, only shows that require the AC3 audio codec. Click Yes to allow Remote Potato to add rules to the Windows Firewall for remote access. You’ll likely need to accept a few UAC prompts. When notified that the rules were added, click OK. Remote Potato will then prompt you to allow administrator privileges to reserve a URL for it’s web server. Click Yes. Remote Potato server will start. Click on the configuration button at the right to to reveal the settings tabs.   One the General tab, you’ll have the option to run Remote Potato on startup and minimized in the System Tray. If you’re running Media Center on a dedicated HTPC, you’ll probably want to enable both startup options. Forwarding Ports on Your Router You’ll need to forward a couple ports on your router. By default, these will be ports 9080 and 9081. In this example we’re using a Linksys WRT54GL router, however, the steps for port forwarding will vary from router to router. On the Linksys configuration page, click on the Applications & Gaming Tab, and then the Port Range Forward tab. Under Application, type in a name of your choosing. In both the Start and End boxes, type the port number 9080. Enter the local IP address of your Media Center computer in the IP address column. Click the check box under Enable. Repeat the process on the next line, but this time use port 9081. When finished, click the Save Settings button. Note: It’s highly recommended that you configure the home computer running Media Center & Remote Potato with a static IP address.   Find your IP Address You’ll need to find the IP address assigned to your router from your ISP. There are many ways to do this but a quick and easy way is to visit a site like checkip.dyndns.org (link available below) The current external IP address of your router will be displayed in the browser.   Dynamic DNS This is an optional step, but  it’s highly recommended. Many routers, such as the Linksys WRT54GL we are using, support Dynamic DNS (DDNS). What Dynamic DNS allows you to do is affiliate your home router’s external IP address to a domain name. Every time your home router is assigned a a new IP address by your ISP, the domain name is updated to point to your new IP address. Remote Potato’s user interface is accessed over the Internet is by connecting to your router’s IP address followed by a colon and the port number. (Ex: XXX.XXX.XXX.XXX:9080) Instead of constantly having to look up and remember an IP address, you can use DDNS along with a 3rd party provider like DynDNS.com, to sign up for a free domain name and configure it to be updated each time your router is assigned a new IP address. Go to the DynDNS.com website (See link at the end of the article) and sign up for a free Domain name. You’ll need to register and confirm by email.   Once you’ve signed in and selected your domain name click Activate Services. You’ll get a confirmation message that your domain name has been activated.    On the Linksys WRT54GL click on the Setup tab an then DDNS. Select DynDNS.org, or TZO.com if you prefer to use their service, from the drop down list.   With DynDNS, you’ll need to fill in your username and password you signed up with at the DynDNS website and the hostname you chose. Note: You can connect over your local network with the IP Address of the computer running Remote Potato followed by a colon and the port number. Ex: 192.168.1.2:9080 Logging in Remote Potato and Recording a Show Once you connect, you’ll see the start page. To view the TV listings, click on TV Guide. You’ll then see your guide listings. There are a few ways to navigate the listings. At the top left, you can click on any of the preset time buttons to jump to  the listings at that time of the day.  Click on the arrows to the right and left of the day and date at the top center to proceed to the previous or next day. Or, jump to a specific day with the date and date buttons at the top right.   To setup a recording, click on a program.   You can choose to record the individual show or the entire series by clicking on Record Show or Record Series.   Remote Potato on Mobile Devices Perhaps the coolest feature of Remote Potato is the ability to schedule recording from your phone or mobile device. Note: For any devices or computers without Silverlight, you will be prompted to view the HTML page. Select Browse Listings. Select your program to record. In the Program Details, select Record Show to record the single episode or Record Series to record all instances of the series. You will then see a red dot on the program listing to indicate that the show is scheduled for recording.   Streaming Recorded TV Click on Recorded TV from the home screen to access your previously recorded TV programs. Click on the selection you wish to stream. Click on Play. If you receive this error message, you’ll need to install the streaming pack for Remote Potato. This is found on the same download page as installation files. (See link below) The Begin from slider allows you to start playback from the start (by default) or a different time of the program by moving the slider. The Quality (bitrate) setting  allows you to choose the quality of the playback. We found the video quality on the Normal setting to be pretty lousy, and Low was just pointless. High was the best overall viewing experience as it provided smooth quality video playback. We experienced significant stuttering during playback using the Ultra High setting.   Click Start when you are ready to begin. When playback begins you’ll see a slider at the top right.   Move the slider left or right to increase or decrease the size of the video. There’s also a button to switch to full screen.   Media Center users who travel frequently or are always on the go will likely find Remote Potato to be a blessing. Since being released earlier this year, updates for Remote Potato have come fast and furious. The latest beta release includes support for streaming music and photos. If you like those nice network TV logos, check out our article on adding TV channel logos to Windows Media Center. Downloads and Links Download Remote Potato and Streaming Pack Find your IP address Sign Up for a Domain Name at DynDNS.com Similar Articles Productive Geek Tips Schedule Updates for Windows Media CenterUsing Netflix Watchnow in Windows Vista Media Center (Gmedia)Add a Sleep Timer to Windows 7 Media CenterStartup Customizations for Media Center in Windows 7Enable Media Streaming in Windows Home Server to Windows Media Player TouchFreeze Alternative in AutoHotkey The Icy Undertow Desktop Windows Home Server – Backup to LAN The Clear & Clean Desktop Use This Bookmarklet to Easily Get Albums Use AutoHotkey to Assign a Hotkey to a Specific Window Latest Software Reviews Tinyhacker Random Tips DVDFab 6 Revo Uninstaller Pro Registry Mechanic 9 for Windows PC Tools Internet Security Suite 2010 FoxClocks adds World Times in your Statusbar (Firefox) Have Fun Editing Photo Editing with Citrify Outlook Connector Upgrade Error Gadfly is a cool Twitter/Silverlight app Enable DreamScene in Windows 7 Microsoft’s “How Do I ?” Videos

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  • ASP.NET MVC 3 - New Features

    - by imran_ku07
    Introduction:          ASP.NET MVC 3 just released by ASP.NET MVC team which includes some new features, some changes, some improvements and bug fixes. In this article, I will show you the new features of ASP.NET MVC 3. This will help you to get started using the new features of ASP.NET MVC 3. Full details of this announcement is available at Announcing release of ASP.NET MVC 3, IIS Express, SQL CE 4, Web Farm Framework, Orchard, WebMatrix.   Description:       New Razor View Engine:              Razor view engine is one of the most coolest new feature in ASP.NET MVC 3. Razor is speeding things up just a little bit more. It is much smaller and lighter in size. Also it is very easy to learn. You can say ' write less, do more '. You can get start and learn more about Razor at Introducing “Razor” – a new view engine for ASP.NET.         Granular Request Validation:             Another biggest new feature in ASP.NET MVC 3 is Granular Request Validation. Default request validator will throw an exception when he see < followed by an exclamation(like <!) or < followed by the letters a through z(like <s) or & followed by a pound sign(like &#123) as a part of querystring, posted form, headers and cookie collection. In previous versions of ASP.NET MVC, you can control request validation using ValidateInputAttriubte. In ASP.NET MVC 3 you can control request validation at Model level by annotating your model properties with a new attribute called AllowHtmlAttribute. For details see Granular Request Validation in ASP.NET MVC 3.       Sessionless Controller Support:             Sessionless Controller is another great new feature in ASP.NET MVC 3. With Sessionless Controller you can easily control your session behavior for controllers. For example, you can make your HomeController's Session as Disabled or ReadOnly, allowing concurrent request execution for single user. For details see Concurrent Requests In ASP.NET MVC and HowTo: Sessionless Controller in MVC3 – what & and why?.       Unobtrusive Ajax and  Unobtrusive Client Side Validation is Supported:             Another cool new feature in ASP.NET MVC 3 is support for Unobtrusive Ajax and Unobtrusive Client Side Validation.  This feature allows separation of responsibilities within your web application by separating your html with your script. For details see Unobtrusive Ajax in ASP.NET MVC 3 and Unobtrusive Client Validation in ASP.NET MVC 3.       Dependency Resolver:             Dependency Resolver is another great feature of ASP.NET MVC 3. It allows you to register a dependency resolver that will be used by the framework. With this approach your application will not become tightly coupled and the dependency will be injected at run time. For details see ASP.NET MVC 3 Service Location.       New Helper Methods:             ASP.NET MVC 3 includes some helper methods of ASP.NET Web Pages technology that are used for common functionality. These helper methods includes: Chart, Crypto, WebGrid, WebImage and WebMail. For details of these helper methods, please see ASP.NET MVC 3 Release Notes. For using other helper methods of ASP.NET Web Pages see Using ASP.NET Web Pages Helpers in ASP.NET MVC.       Child Action Output Caching:             ASP.NET MVC 3 also includes another feature called Child Action Output Caching. This allows you to cache only a portion of the response when you are using Html.RenderAction or Html.Action. This cache can be varied by action name, action method signature and action method parameter values. For details see this.       RemoteAttribute:             ASP.NET MVC 3 allows you to validate a form field by making a remote server call through Ajax. This makes it very easy to perform remote validation at client side and quickly give the feedback to the user. For details see How to: Implement Remote Validation in ASP.NET MVC.       CompareAttribute:             ASP.NET MVC 3 includes a new validation attribute called CompareAttribute. CompareAttribute allows you to compare the values of two different properties of a model. For details see CompareAttribute in ASP.NET MVC 3.       Miscellaneous New Features:                    ASP.NET MVC 2 includes FormValueProvider, QueryStringValueProvider, RouteDataValueProvider and HttpFileCollectionValueProvider. ASP.NET MVC 3 adds two additional value providers, ChildActionValueProvider and JsonValueProvider(JsonValueProvider is not physically exist).  ChildActionValueProvider is used when you issue a child request using Html.Action and/or Html.RenderAction methods, so that your explicit parameter values in Html.Action and/or Html.RenderAction will always take precedence over other value providers. JsonValueProvider is used to model bind JSON data. For details see Sending JSON to an ASP.NET MVC Action Method Argument.           In ASP.NET MVC 3, a new property named FileExtensions added to the VirtualPathProviderViewEngine class. This property is used when looking up a view by path (and not by name), so that only views with a file extension contained in the list specified by this new property is considered. For details see VirtualPathProviderViewEngine.FileExtensions Property .           ASP.NET MVC 3 installation package also includes the NuGet Package Manager which will be automatically installed when you install ASP.NET MVC 3. NuGet makes it easy to install and update open source libraries and tools in Visual Studio. See this for details.           In ASP.NET MVC 2, client side validation will not trigger for overridden model properties. For example, if have you a Model that contains some overridden properties then client side validation will not trigger for overridden properties in ASP.NET MVC 2 but client side validation will work for overridden properties in ASP.NET MVC 3.           Client side validation is not supported for StringLengthAttribute.MinimumLength property in ASP.NET MVC 2. In ASP.NET MVC 3 client side validation will work for StringLengthAttribute.MinimumLength property.           ASP.NET MVC 3 includes new action results like HttpUnauthorizedResult, HttpNotFoundResult and HttpStatusCodeResult.           ASP.NET MVC 3 includes some new overloads of LabelFor and LabelForModel methods. For details see LabelExtensions.LabelForModel and LabelExtensions.LabelFor.           In ASP.NET MVC 3, IControllerFactory includes a new method GetControllerSessionBehavior. This method is used to get controller's session behavior. For details see IControllerFactory.GetControllerSessionBehavior Method.           In ASP.NET MVC 3, Controller class includes a new property ViewBag which is of type dynamic. This property allows you to access ViewData Dictionary using C # 4.0 dynamic features. For details see ControllerBase.ViewBag Property.           ModelMetadata includes a property AdditionalValues which is of type Dictionary. In ASP.NET MVC 3 you can populate this property using AdditionalMetadataAttribute. For details see AdditionalMetadataAttribute Class.           In ASP.NET MVC 3 you can also use MvcScaffolding to scaffold your Views and Controller. For details see Scaffold your ASP.NET MVC 3 project with the MvcScaffolding package.           If you want to convert your application from ASP.NET MVC 2 to ASP.NET MVC 3 then there is an excellent tool that automatically converts ASP.NET MVC 2 application to ASP.NET MVC 3 application. For details see MVC 3 Project Upgrade Tool.           In ASP.NET MVC 2 DisplayAttribute is not supported but in ASP.NET MVC 3 DisplayAttribute will work properly.           ASP.NET MVC 3 also support model level validation via the new IValidatableObject interface.           ASP.NET MVC 3 includes a new helper method Html.Raw. This helper method allows you to display unencoded HTML.     Summary:          In this article I showed you the new features of ASP.NET MVC 3. This will help you a lot when you start using ASP MVC 3. I also provide you the links where you can find further details. Hopefully you will enjoy this article too.  

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  • Complete Guide to Symbolic Links (symlinks) on Windows or Linux

    - by Matthew Guay
    Want to easily access folders and files from different folders without maintaining duplicate copies?  Here’s how you can use Symbolic Links to link anything in Windows 7, Vista, XP, and Ubuntu. So What Are Symbolic Links Anyway? Symbolic links, otherwise known as symlinks, are basically advanced shortcuts. You can create symbolic links to individual files or folders, and then these will appear like they are stored in the folder with the symbolic link even though the symbolic link only points to their real location. There are two types of symbolic links: hard and soft. Soft symbolic links work essentially the same as a standard shortcut.  When you open a soft link, you will be redirected to the folder where the files are stored.  However, a hard link makes it appear as though the file or folder actually exists at the location of the symbolic link, and your applications won’t know any different. Thus, hard links are of the most interest in this article. Why should I use Symbolic Links? There are many things we use symbolic links for, so here’s some of the top uses we can think of: Sync any folder with Dropbox – say, sync your Pidgin Profile Across Computers Move the settings folder for any program from its original location Store your Music/Pictures/Videos on a second hard drive, but make them show up in your standard Music/Pictures/Videos folders so they’ll be detected my your media programs (Windows 7 Libraries can also be good for this) Keep important files accessible from multiple locations And more! If you want to move files to a different drive or folder and then symbolically link them, follow these steps: Close any programs that may be accessing that file or folder Move the file or folder to the new desired location Follow the correct instructions below for your operating system to create the symbolic link. Caution: Make sure to never create a symbolic link inside of a symbolic link. For instance, don’t create a symbolic link to a file that’s contained in a symbolic linked folder. This can create a loop, which can cause millions of problems you don’t want to deal with. Seriously. Create Symlinks in Any Edition of Windows in Explorer Creating symlinks is usually difficult, but thanks to the free Link Shell Extension, you can create symbolic links in all modern version of Windows pain-free.  You need to download both Visual Studio 2005 redistributable, which contains the necessary prerequisites, and Link Shell Extension itself (links below).  Download the correct version (32 bit or 64 bit) for your computer. Run and install the Visual Studio 2005 Redistributable installer first. Then install the Link Shell Extension on your computer. Your taskbar will temporally disappear during the install, but will quickly come back. Now you’re ready to start creating symbolic links.  Browse to the folder or file you want to create a symbolic link from.  Right-click the folder or file and select Pick Link Source. To create your symlink, right-click in the folder you wish to save the symbolic link, select “Drop as…”, and then choose the type of link you want.  You can choose from several different options here; we chose the Hardlink Clone.  This will create a hard link to the file or folder we selected.  The Symbolic link option creates a soft link, while the smart copy will fully copy a folder containing symbolic links without breaking them.  These options can be useful as well.   Here’s our hard-linked folder on our desktop.  Notice that the folder looks like its contents are stored in Desktop\Downloads, when they are actually stored in C:\Users\Matthew\Desktop\Downloads.  Also, when links are created with the Link Shell Extension, they have a red arrow on them so you can still differentiate them. And, this works the same way in XP as well. Symlinks via Command Prompt Or, for geeks who prefer working via command line, here’s how you can create symlinks in Command Prompt in Windows 7/Vista and XP. In Windows 7/Vista In Windows Vista and 7, we’ll use the mklink command to create symbolic links.  To use it, we have to open an administrator Command Prompt.  Enter “command” in your start menu search, right-click on Command Prompt, and select “Run as administrator”. To create a symbolic link, we need to enter the following in command prompt: mklink /prefix link_path file/folder_path First, choose the correct prefix.  Mklink can create several types of links, including the following: /D – creates a soft symbolic link, which is similar to a standard folder or file shortcut in Windows.  This is the default option, and mklink will use it if you do not enter a prefix. /H – creates a hard link to a file /J – creates a hard link to a directory or folder So, once you’ve chosen the correct prefix, you need to enter the path you want for the symbolic link, and the path to the original file or folder.  For example, if I wanted a folder in my Dropbox folder to appear like it was also stored in my desktop, I would enter the following: mklink /J C:\Users\Matthew\Desktop\Dropbox C:\Users\Matthew\Documents\Dropbox Note that the first path was to the symbolic folder I wanted to create, while the second path was to the real folder. Here, in this command prompt screenshot, you can see that I created a symbolic link of my Music folder to my desktop.   And here’s how it looks in Explorer.  Note that all of my music is “really” stored in C:\Users\Matthew\Music, but here it looks like it is stored in C:\Users\Matthew\Desktop\Music. If your path has any spaces in it, you need to place quotes around it.  Note also that the link can have a different name than the file it links to.  For example, here I’m going to create a symbolic link to a document on my desktop: mklink /H “C:\Users\Matthew\Desktop\ebook.pdf”  “C:\Users\Matthew\Downloads\Before You Call Tech Support.pdf” Don’t forget the syntax: mklink /prefix link_path Target_file/folder_path In Windows XP Windows XP doesn’t include built-in command prompt support for symbolic links, but we can use the free Junction tool instead.  Download Junction (link below), and unzip the folder.  Now open Command Prompt (click Start, select All Programs, then Accessories, and select Command Prompt), and enter cd followed by the path of the folder where you saved Junction. Junction only creates hard symbolic links, since you can use shortcuts for soft ones.  To create a hard symlink, we need to enter the following in command prompt: junction –s link_path file/folder_path As with mklink in Windows 7 or Vista, if your file/folder path has spaces in it make sure to put quotes around your paths.  Also, as usual, your symlink can have a different name that the file/folder it points to. Here, we’re going to create a symbolic link to our My Music folder on the desktop.  We entered: junction -s “C:\Documents and Settings\Administrator\Desktop\Music” “C:\Documents and Settings\Administrator\My Documents\My Music” And here’s the contents of our symlink.  Note that the path looks like these files are stored in a Music folder directly on the Desktop, when they are actually stored in My Documents\My Music.  Once again, this works with both folders and individual files. Please Note: Junction would work the same in Windows 7 or Vista, but since they include a built-in symbolic link tool we found it better to use it on those versions of Windows. Symlinks in Ubuntu Unix-based operating systems have supported symbolic links since their inception, so it is straightforward to create symbolic links in Linux distros such as Ubuntu.  There’s no graphical way to create them like the Link Shell Extension for Windows, so we’ll just do it in Terminal. Open terminal (open the Applications menu, select Accessories, and then click Terminal), and enter the following: ln –s file/folder_path link_path Note that this is opposite of the Windows commands; you put the source for the link first, and then the path second. For example, let’s create a symbolic link of our Pictures folder in our Desktop.  To do this, we entered: ln -s /home/maguay/Pictures /home/maguay/Desktop   Once again, here is the contents of our symlink folder.  The pictures look as if they’re stored directly in a Pictures folder on the Desktop, but they are actually stored in maguay\Pictures. Delete Symlinks Removing symbolic links is very simple – just delete the link!  Most of the command line utilities offer a way to delete a symbolic link via command prompt, but you don’t need to go to the trouble.   Conclusion Symbolic links can be very handy, and we use them constantly to help us stay organized and keep our hard drives from overflowing.  Let us know how you use symbolic links on your computers! Download Link Shell Extension for Windows 7, Vista, and XP Download Junction for XP Similar Articles Productive Geek Tips Using Symlinks in Windows VistaHow To Figure Out Your PC’s Host Name From the Command PromptInstall IceWM on Ubuntu LinuxAdd Color Coding to Windows 7 Media Center Program GuideSync Your Pidgin Profile Across Multiple PCs with Dropbox TouchFreeze Alternative in AutoHotkey The Icy Undertow Desktop Windows Home Server – Backup to LAN The Clear & Clean Desktop Use This Bookmarklet to Easily Get Albums Use AutoHotkey to Assign a Hotkey to a Specific Window Latest Software Reviews Tinyhacker Random Tips DVDFab 6 Revo Uninstaller Pro Registry Mechanic 9 for Windows PC Tools Internet Security Suite 2010 Gadfly is a cool Twitter/Silverlight app Enable DreamScene in Windows 7 Microsoft’s “How Do I ?” Videos Home Networks – How do they look like & the problems they cause Check Your IMAP Mail Offline In Thunderbird Follow Finder Finds You Twitter Users To Follow

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