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专题部分简析煤矿巷道断面设计思路摘要：煤矿巷道设计是煤矿技术工作的重要组成部分，正确合理的巷道设计对于保障安全、降低成本、提高工效具有重要意义。本文简要概括了煤矿巷道断面的设计思路，关键词：煤矿 巷道断面 设计思路1 引言煤矿巷道设计是矿井生产的前提和基础，一定程度上说，巷道设计情况决定了该巷道所服务的采场或矿井所能达到的技术、安全水平。巷道设计贯穿矿井设计的始末，从开拓方案的确定到巷道断面及井底车场的设计，接着采区布置，然后通风阻力的计算和风机的选择，还有最后工程量预算方面等等，都在和巷道打交道。所以能准确快速的绘制出巷道断面，以及准确的确定出工程量的大小，是一件非常重要的工作，这对矿井设计及方案的确定有很大的现实意义。本文对巷道设计工作进行简要的归纳，首先考虑巷道断面的分类，将巷道按运输方式、断面形状、支护方式、水沟形状和水沟支护分为五类，然后总结出以下九步的煤矿巷道设计流程，一要确定巷道运输方式，二要确定巷道断面形状，三要确定巷道净面积，四要计算道渣面的铺设高度，五进行风量的验算，六确定支护方式，七确定水沟的形状和位置，八进行管线布置，最后进行标注。2 分类及设计流程巷道断面的分类从以下几个方面考虑巷道断面的种类：常用的运输方式：轨道巷、运输巷、机轨合一巷、无轨运输巷、双轨巷常用的断面形状：半圆拱、圆弧拱、三心圆拱、梯形、矩形、不规则形常用的支护方式：锚杆、锚喷、有无网、砌筑、金属棚、木棚、U型钢常用的水沟形状：对称倒梯形、半倒梯形、矩形常用的水沟支护：砌筑有盖板、砌筑无盖板、无砌筑无盖板、大、中巷常用的设计类型：初步设计、改进设计每一种巷道的绘制中都要在这五部分中取其一进行排列组合，有2400多种，最终形成完整的巷道断面图。设计流程图2-1 巷道断面流程图设计流程如上图所示，首先考虑巷道的运输方式和断面形状来确定出巷道的净断面积，然后通过风量进行校核，即设计的巷道断面一定要满足通风的要求，其次进行支护方式的选择及水沟的选取，最后得出巷道的掘进工程量和材料消耗量，最后对其进行标注，完成巷道断面设计3 巷道断面的设计思路按照设计流程所需要考虑的因素我们分成十块内容逐一分析，以及设计中所注意的事项。它们分别是：1、巷道运输方式的确定2、巷道断面形状的确定3、巷道净面积的确定4、道渣面的铺设高度5、风量的验算6、支护方式的确定7、水沟的形状和位置8、管线布置9、标注确定巷道的运输方式巷道的运输方式是确定巷道净宽的主要因素，常用的运输方式有（轨道巷、运输巷、机轨合一巷、无轨运输巷、单轨吊）考虑断面的形状图3-1 巷道断面形状的选择矿上常用巷道断面的形状有（半圆拱形、圆弧拱形、三心圆拱形、梯形、矩形、马蹄形、圆形、椭圆形、不规则形）选择巷道断面形状主要考虑的因素如图1-1尺寸的确定巷道的净宽和净高要符合煤矿安全规程的相关条文：第二十一条 巷道净断面必须满足行人、运输、通风和安全设施及设备安装、检修、施工的需要，并符合下列要求：1、主要运输巷和主要风巷的净高，自轨面起不得低于2m。2、采区(包括盘区)内的上山、下山和平巷的净高不得低于2m，薄煤层内的不得低于1.8m。3、巷道净断面的设计，必须按支护最大允许变形后的断面计算。第二十二条 运输巷两侧（包括管、线、电缆）与运输设备最突出部分之间的距离，应符合下列要求：1、新建矿井、生产矿井新掘运输巷的一侧，从巷道道碴面起1.6m的高度内，必须留有宽0.8m(综合机械化采煤矿井为1m)以上的人行道，管道吊挂高度不得低于1.8m；巷道另一侧的宽度不得小于0.3m(综合机械化采煤矿井为0.5m)。巷通内安设输送机时，输送机与巷帮支护的距离不得小于0.5m；输送机机头和机尾处与各帮支护的距离应满足设备检查和维修的需要，并不得小于0.7m。巷道内移动变电站或平板车上综采设备的最突出部分，与巷帮支护的距离不得小于0.3m。2、生产矿井已有巷道人行道的宽度不符合本条第一款第(一)项的要求时，必须在巷道的一侧设置躲避硐，2个躲避硐之间的距离不得超过40m。躲避硐宽度不得小于1.2m，深度不得小于0.7m，高度不得小于1.8m，躲避用内严禁堆积物料。3、在人车停车地点的巷道上下人侧，从巷道道碴面起1.6m的高度内，必须留有宽1m以上的人行道，管道吊挂高度不得低于1.8m。第二十三条 在双轨运输巷中，2列列车最突出部分之间的距离，对开时不得小于0.2m，采区装载点不得小于0.7m，矿车摘挂钩地点不得小于1m。车辆最突出部分与巷道两侧距离，必须符合本规程第二十二条的要求。第三百五十六条 自轨面界起，电机车架空线的悬挂高度应符合下列要求：1、在行人的巷道内、车场内以及人行道与运输巷道交叉的地方不小于2m，在不行人的巷道内不小于1.9m。2、在井底车场内，从井底到乘车场不小于2.2m。3、在地面或工业场地内，在与其他道路交叉的地方不小于2.2m。第三百五十七条 电机车架空线与巷道顶或棚梁之间的距离不得小于0.2m。悬吊绝缘子距电机车架空线的距离，每侧不得超过0.25m。电机车架空线悬挂点的间距，在直线段内不得超过5m，在曲线段内不得超过表1-3规定值。根据设计手册中的要求列出巷道安全间隙表1-1和双轨中心距表1-2。表3-1 巷道安全间隙表项目规定数值人行侧从道碴面起1.6m高度范围内设备与拱、壁间综采矿井1000其他矿井800非人行侧设备与拱、壁间综采矿井500其他矿井300移动变电站或平板车上综采设备最突出的部分与拱、壁间300与输送机间700人车停车地点行人侧从道碴面起1.6m高度范围内设备与拱、壁间1000安设输送机巷道中输送机与拱、壁间500两列对开列车最突出的部分间200采区装卸点两列车最突出的部分间700电机车架空线与巷道顶或棚梁间200导电弓距拱、壁间300矿车债挂钩点两列车最突出的部分间1000导电弓距管子最突出部分间300运输设备距管子最突出部分间300设备上面最突出部分距巷道顶或棚梁间、壁间300用架空乘人装置运送人员时，蹬座中心至巷道一侧的距离700表3-2 双轨中心距表运输设备600mm轨距直线曲线1.0t矿车110013001.5t矿车 130015007t、10t、14t架线机车130016003.0t底卸式矿车150017005.0t底卸式矿车160018008t、12t蓄电池机车13001600表3-3 电机车架空线曲线段悬挂点间距最大值曲率半径/m25222119181615131211108悬挂点间/m4.543.532.52道床的铺设高度注：考虑到巷道在施工中出现超挖现象，因此将以设计掘进断面尺寸加上允许加宽值即超挖值(梯形棚子按25m，直墙取平均值75m)，作为计算掘进断面尺寸，并以此计算出巷道的掘进工程量和支架材料消耗量。常用道床铺设高度如表1-4。表3-4 常用道床铺设高度钢轨型号（kg/m）运输巷道（铺道碴）（不铺道碴）总高度（H）道碴高度h总高度H304102202802238022025015350220220进行风量的验算生产矿井的巷道通常兼作通风用，因此还应按照煤矿安全规程等规定的允许最大风速进行验算，即式中通过巷道风流的速度，m/s；通过巷道的风量，m/s；替道的净断面积，；巷道允许通过的最高风速，ms。煤炭工业设计规范规定：矿井主要进风巷的风速一般不大于6m/s，输送机巷或采区风巷一般不大于4m/s。设计时，应在不违反煤矿安全规程的原则下，按规范要求确定巷道断面，以留有余地。设计出的巷道净断面积，不符合上述风速要求的，需加大断面尺寸。第一百零一条 井巷中风流速度应符合表1-5要求。1、设有梯子间的井筒或修理中的井筒，风速不得超过8m/s；梯子间四周经封闭后，井筒中的最高允许风速可按表2-5规定执行。2、无瓦斯涌出的架线电机车巷道中的最低风速可低于表1-5的规定值，但不得低于0.5m/s。3、综合机械化采煤工作面，在采取煤层注水和采煤机喷雾降尘等措施后，其最大风速可高于表2-5的规定值，但不得超过5m/s。4、专用排瓦斯巷内风速不得低于0.5m/s。断面的支护形式当巷道断面形成后，紧接着的工序就是巷道支护，合理的支护是在规定的服务年限内满足通风、运输、行人的需要的前提下。支护工序简单、成本最低。在确定支护方式时，要考虑很多因素。结合设计手册，现简单归纳主要有以下几种。1、巷道的服务年限。2、围岩的物理力学性质、地质构造。3、巷道的断面形状和施工方法。4、支架的材料费，制造费，运输、架设和维护费用总和最低。目前，常用的断面支护设计方法一般分为3种：工程类比法，理论分析法和现场监控法。我们主要按照设计手册中的围岩分类及工程类比法来确定巷道的支护参数，主要的支护方式有锚杆、锚喷、砌筑、金属棚、木棚、U型钢。水沟的位置及形状根据流量、实际的需要选出适合的水沟。考虑的因素有（水沟的位置、水沟的断面和流量计算、水沟盖板的大小）。水沟布置的要求1、水平巷道及小于16倾斜巷道的水沟，一般布置在人行侧，当非人行侧有适当空间时，亦可布置。应尽量避免穿越轨道或输送机。2、在倾角大于16的巷道中，当涌水量小或巷道较窄时，水沟与人行台阶可在巷道同侧平行布置；当当涌水量较大或巷道较宽时，水沟与人行台阶可分设在巷道两侧。3、金属或木棚巷道的水沟为使柱腿牢固和流水通畅，水沟中线与柱腿之间的距离应大于0.5m，或者水沟与柱腿之间的距离应大于0.3m，而且水沟中线至轨道中心线不小于1100mm。4、专用排水巷道、中间设人行道的巷道、有底鼓的巷道和铺设整体道床的巷道、水沟也可布置在巷道中间。5、当涌水量大时，可并排布置两条水沟6、在倾角小于或等于10的行人及车辆来往频繁的主要巷道，水沟上面要加设盖板，盖板面应与道碴面齐平。水沟的砌筑根据水沟的服务年限可分为永久性水沟和临时水沟，永久性水沟应砌筑，临时性水沟可不砌筑。水沟一般可用混凝土现浇，也可采用钢筋混凝土预制。对称倒梯形水沟一般用混凝土砌筑。水沟断面确定1、常用的水沟断面形状有对称倒梯形、半倒梯形和矩形三种。根据支护的方式来确定断面的形状。对称倒梯形水沟适用于金属或木棚支护的巷道；半倒梯形水沟适用于砌碹支护的巷道；矩形水沟适用于锚喷支护的巷道。2、水沟断面尺寸主要根据流量、坡度来选择（如表2-11至2-13）水沟盖板1、为行人方便，大巷及小于15上（下）山的水沟，一般设置盖板；2、无运输设备的巷道，大于15上（下）山，和采区巷道的水沟一般可不设盖板；3、对称倒梯形水沟无盖板.；4、盖板的宽度一般比水沟净宽加宽150mm；5、盖板一般为钢筋混凝土预制板，厚度一般小于50mm。管线的布置方式管线布置的原则是保证安全和便于安装，检修。其布置方式要符合煤矿安全规程的相关规定第四百六十九条 电缆不应悬挂在风管或水管上，不得遭受淋水。电缆上严禁悬挂任何物件。电缆与压风管、供水管在巷道同一侧敷设时，必须敷设在管子上方，并保持0.3m以上的距离。在有瓦斯抽放管路的巷道内，电缆(包括通信、信号电缆)必须与瓦斯抽放管路分挂在巷道两例。盘圈或盘“8”字形的电缆不得带电，但给采、掘机组供电的电缆不受此限。井筒和巷道内的通信和信号电缆应与电力电缆分挂在井巷的两侧，如果受条件所限：在井筒内，应敷设在距电力电缆0.3m以外的地方；在巷道内，应敷设在电力电缆上方0.1m以上的地方。高、低压电力电缆敷设在巷道同一侧时，高、低压电缆之间的距离应大干0.1m。高压电缆之间、低压电缆之间的距离不得小于50mm。井下巷道内的电缆，沿线每隔一定距离、拐弯或分支点以及连接不同直径电缆的接线盒两端、穿墙电缆的墙的两边都应设置注有编号、用途、电压和截面的标志牌。第四百七十三条 井下下列地点必须有足够的照明：1、井底车场及其附近。2、机电设备用室、调度宣、机车库、爆炸材料库、候车室、信号站、瓦斯抽放泵站等。3、使用机车的主要运输巷道、兼做人行道的集中带式输送机巷道、升降人员的绞车道以及升降物料和人行交替使用的绞车道其照明灯的间距不得大于30m。4.主要进风巷的交岔点和采区车场。5.从地面到井下的专用人行道。6.综合机械化采煤工作面，照明灯间距不得大干15m。地面的通风机房、绞车房、压风机房、变电所、矿调度室等必须设有应急照明设施。布置位置1、在主要运输巷道中，大都将管路敷设在人行道侧顶部；当非行人侧宽度较大时，也可将管线敷设在非行人侧。2、动力电缆一般应敷设在非行人道一侧，并保证车辆掉道时不受撞击，而且电缆坠落时不至于掉到轨道或输送机上。3、当电缆和管路布置在巷道同一侧时，电缆应敷设在管路的上方，其间距应大于300mm，而且电缆上不得悬挂任何物件。4、通讯、信号电缆应与动力电缆分别挂在巷道的两侧。当受条件限制时，通讯、信号电缆敷设在动力电缆的上方，其间距应大于100mm。5、高、低压电缆相互间距应大于100mm，高压电缆之间或低压电缆之间的距离均不小于50mm。6、在有煤和瓦斯突出危险的煤层回风巷中，禁止设置动力电缆。7、而且巷道高度在1.6m至1.8m之间，不得架设管、线和电缆。电力电缆与管道应布置在巷道的不同侧。在梯形巷道内，电力电缆布置在人行道一侧的棚腿上部；管道则布置在另一侧下部，细管在上，粗管在下，与道渣面保持150mm距离，以利安装和检修，而且任何管子与运行车辆的距离都不得小于200mm。在拱形巷道内，管道布置在人行道一侧，而其下部与道渣面或水沟盖板面保持1.8m和1.8m以上的距离；电力电缆布置在另一例，距底板不得小于1m，与运行车辆的间距不得小于250mm，力求布置在车辆高度之上。电话和信号电缆应布置在电力电缆的另一侧，若不得已必须布置在同一测时则应在电力电缆上方100mm以外。当电缆与管道同侧布置时也应将电缆布置在管道之上不小于300mm的地方。固定方式1、砌碹支护的主要运输巷道，一般用槽钢或角钢将管子支托在人行侧的顶部；锚喷支护的主要运输巷道，也可将管路锚吊在人行侧的顶部。也可采用毛料石或混凝土墩柱支托管子。当设置多趟管路时，也可将管子架设在钢支架上。2、在倾角小于30的巷道中，电缆应用吊钩悬挂；在倾角大于30的巷道中，电缆应用夹持装置进行敷设。3、井下电缆的敷设应符合煤矿安全规程的有关规定。标注首先要在CAD中修改标注样式，方法如下：点击格式标注样式主单位比例因子（50），然后进行标注，主要标注三组尺寸。设备间隙尺寸、巷道净掘进尺寸、施工尺寸。出图时1:1打印。4 巷道断面的相关计算在巷道断面设计中我们要进行一下几种计算：1、巷道及水沟的净面积和掘进面积。2、支护参数的确定，如锚杆、锚索的间排距，喷浆厚度等。3、材料消耗，如混凝土消耗，锚杆锚索的消耗。净面积的确定巷道断面净宽度的确定（1）矩形巷道(直墙拱形巷道)的净宽度：系指巷道两侧内壁或锚杆露出长度终端之间的水平间距当其内设置运输机械或不设置、也不通行运输设备时，净宽度系指底板起1.6m高水平的巷道宽度。运输巷道净宽度，由运输设备本身外轮廓最大宽度和煤矿安全规程所规定的人行道宽度及有关安全间隙相加而得；无运输设备的巷道，可根据行人及通风的需要来选取。1）双轨（包括机轨合一巷）巷道净宽度：B=2）单轨（包括单输送机）巷道净宽度：B=式中B巷道净宽度，mm；、分别为非行人侧和行人侧轨道（或输送机）中线到巷道墙之间的距离。mm；轨道（或轨道与输送机）中线之间的距离，mm。按以上公式所计算的巷道净宽B值，应根据只进不舍的原则以100mm进级。3）无轨运输巷道净宽度：巷道一侧应留有1.2m以上的人行道；另一侧宽度也应不小于0.5m；两辆车对开最突出部分之间的距离不小于0.5m。在巷道转弯或交叉处，无轨运输车的间距必须满足安全运输的要求，此时巷道的宽度B应根据无轨运输车的转弯半径和运输间距来确定。B式中运输车转弯外半径，m；运输车转弯内半径，m。（2）对梯形巷道：当其内通行矿车、电机车时，净宽度系指车辆顶面水平的巷道宽度；车辆顶面距道渣面的高度可以取1150mm，上述取出的B值，即梯形巷道车辆顶面水平的巷道宽度，我们还要求出梯形巷道的底宽B2，顶宽B1,顶宽B1由梯形的高度确定。巷道断面净高度的确定矩形、梯形巷道的净高系指自渣面或底板至顶梁或顶部喷层面、锚杆露出长度终端的高度。煤矿安全规程规定；主要运输巷迫和主要风谊的净高，自轨面起不得低于1.9m。架线电机车的运输巷道的净高，必须符合有关规定，即；电机车架空线的悬挂高度，自轨面算起在行人的巷道内、车场内以及人行道同运输巷道交叉的地方不得小于2m；在不行人的巷道内不得小于1.8m：在井底车场内，从井底到乘车场不得小于2.2m。电机车架空线和巷道顶或棚梁之间的距离不得小于0.2m。对于采区(盘区)内的上山、下山和平巷的净高不得低于1.8m。根据上述规定及有关公式，便可求得巷道的净高度H和其他高度(要注意预留下沉量)。（1）拱形巷道净高度：式中净高度，mm；墙高，mm；从巷道底板到道碴面的高度，由铺轨参数确定，mm；拱高，mm。1）拱高：半圆拱形拱高为巷道净宽度B之半，既=B/2；圆弧拱形及三心圆拱拱高，煤矿一般采用=B/32）墙高：要满足行人安全、运输通畅、设备运送、安装和检修的需要，对于架线式电机车运输巷道，一般情况下可按架线高度和管子架设要求计算；其他如矿车运输、只铺设输送机和无运输设备的巷道，只按行人要求计算。但在行人道范围内不得敷设管、线、电缆及其他固定设施。其计算过程按手册25602561中的公式计算。计算结果必须按只进不舍得原则，以100mm进级。泛用两种以上方法计算的，取其最大值。（2）梯形和矩形巷道净高度：H2000mm；对于薄煤层H1800mm对于梯形巷道顶宽B1的计算公式如下：支护方式的确定通过可以确定出巷道的净面积，然后通过风量校核，如果不满足，对其宽度或高度适当的放大，最终满足其通风要求，第二步我们来确定支护方式锚杆或锚喷支护已知参数有：锚杆的直径、长度；锚索的直径、长度，我们需要设计的参数其参数有喷浆的厚度T，是否挂金属网，锚杆的根数和间排距。（1）锚杆间距和根数的确定考虑到具体的施工环境，不可能完全按照图纸的要求精确地确定锚杆位置，可以相差100mm，在施工中普遍使用一定长度的线绳来测量锚杆间距，从第一根锚杆等间距标出锚杆的位置。以往的施工方法是间距都是整值，如果设计是n根，量到最后时，发现距底板的距离过大就再补打一根。这种施工方法虽然考虑了均匀支护，但是不太科学，比如说在某一拱形巷道中，设计锚杆间距是800mm，以800mm的间距量过后还剩400mm。不得不补打一根。造成设计的图纸和实际实施时的差距很大。为了解决上述问题，根据采矿工程设计手册锚杆支护标准间距和挤压拱理论，按照等间距的打锚杆原则。结合实际对锚杆间距的设计做出了一些改变。其设计思路如图3-1。图4-1锚杆间距确定思路考虑到底板巷道角部应力集中，在巷道拐角处200300mm的地方要斜向上或下打一根锚杆，来加固边角或底板。过低则锚杆机无法施工，过高则起不到锚固地板的功效，所以在设计中默认250mm，具体情况现场具体考虑。现列出锚杆间距的标准值表1-6，在设计中以供参考。表4-1锚杆间距的标准值受采动影响的巷道围岩类 别巷道净宽B支护方式2000350035005000500650065008000锚喷锚杆锚深180018001800间距900900800喷射混凝土厚度50锚喷锚杆锚深1600180018001800间距800900800800喷射混凝土厚度505080锚喷锚杆锚深1600200020002200间距800800700700喷射混凝土厚度50508080锚喷锚杆锚深1600200022002200间距700700700600钢筋网加加加喷射混凝土厚度80100100120锚喷锚杆锚深1800200022002200间距700600600600钢筋网加加加加喷射混凝土厚度120120150150不受采动影响的巷道围岩类 别巷道净宽B支护方式20003500350050005000650065008000喷射混凝土厚度80100120锚喷锚杆锚深1800间距900喷射混凝土厚度50锚喷锚杆锚深1600200020002000间距800900800800喷射混凝土厚度50808080锚喷锚杆锚深1600200020002000间距800800700700钢筋网加加加喷射混凝土厚度80100100120锚喷锚杆锚深1800200020002200间距700700600600钢筋网加加加加喷射混凝土厚度120150150150回采巷道围岩类 别巷道B支护方式矩形断面回采200030003000400040005000锚杆顶锚杆锚深18001800间距900800帮锚杆锚深1800间距900锚杆顶锚杆锚深160018001800间距800800800帮锚杆锚深18001800间距900800塑料网加锚杆顶锚杆锚深180018002000间距800700700帮锚杆锚深18001800间距800800塑料网加加根据上述规范要求，在围岩类型确定的情况下采用区间的形式来确定锚杆的个数，实现程序化，具体参数如下表拱形巷道锚杆间距和根数的确定：对于拱形巷道的支护理论在实践中常采用均匀压缩带(拱)作用:该理论认为，在锚杆锚固力的作用下，每根锚杆周围形成一个两头带圆锥的筒状压缩区。各锚杆所形成的压缩区彼此联成一个有一定厚度的均匀压缩带，该带具有较大的承载能力如右图4-2。下列公式中遇到的字母代表的含意如下：图4-2C巷道净周长；B巷道净宽；L1锚杆标准间距；L锚杆实际间距；N锚杆数； n 等分弧的份数即分成了n份； N = n + 1 ；L = ( C B ) / n。第一步：确定巷道的围岩特性，根据巷道的净宽度，在锚杆支护标准间距表中选出标准间距L1，确定第一根和最后一根锚杆的位置。如右图4-3注：AB = (200300)mm之间，为方便设计，一般取250mm，实际应用中可适当调整。图4-3 第二步：等分BCD及确定锚杆标注点。BCD = C B 2 AB , n = BCD / L1，则需锚杆数N = n 1, 把打锚杆点在图上标出，如图4-4 注： n只进不舍 图4-4第三步：确定锚杆实际间距L，L = BCD / n,打上锚杆，如图4-5结合CAD, 完善施工图表图4-5矩形巷道锚杆间距和根数的确定：H巷道的净高；Lb帮锚杆标准间距；Lb1帮锚杆标准间距；Ld顶锚杆标准间距；Ld1顶锚杆标准间距；n1巷道帮AB 等分别为n1份；N1帮锚杆数；n2巷道顶CD 等分别为n2份； N2顶锚杆数。 注：以矩形为例，梯形和不规则四边形原理同矩形。第一步：确定巷道的围岩特性，根据巷道的净宽度B，净高H，在锚杆支护标准间距表中选出标准间距L1，当回采巷道时，顶锚杆、帮锚杆间距要区别对待。由于在矩形巷道的四个角处都有应力集中，要进行加强支护，先把四角处的锚杆位置确定下来。注：边角处的锚杆距巷道顶和帮的距离均在(200300)mm之间，为方便设计，一般取250mm，实际应用中可适当调整。如右图4-6。图4-6图4-7第二步：等分AB、CD及确定锚杆标注点。n1 = AB / Lb， n2 = CD / Ld， N1 = ( n1 1) 2， N2 = n2 1则需锚杆数N = N1 N2， 把打锚杆点在图上标出。注： n1、n2只进不舍，如图1-7。图4-8第三步：确定锚杆实际间距Lb1、Ld1，Lb1 = AB / n1, Lb2 = CD / n2打上锚杆，如图1-8结合CAD,完善施工图表。图4-9砌筑、架棚支护其主要确定支护厚度T即可。U型钢支护U型钢支护中的参数很多，但我们只要知道其所需的面积就可以通过手册查出巷道断面的各个参数。通过净面积和支护方式的确定，然后再加上道渣的铺设厚度就可以求出巷道的掘进面积。式中 C：巷道的净周长减去底宽 T：支护厚度 d：巷道超挖值，取75mmB1：巷道净宽：道渣厚度水沟的相关计算 通过前面的设计，我们可以选取好水沟的形状，考虑水沟是否有砌筑，是否有盖板等，然后再根据流量行大巷的坡度通过下表1-71-9来确定水沟的净断面积、掘进面积、钢筋消耗、混凝土消耗的参数。表4-2对称倒梯形水沟编号支护方式流量（m3/h）净尺寸（mm）断面（）每米混凝土消耗量（m3）水沟超高（mm）坡度上宽B2下宽B1深H净掘进3451不砌筑0730840943602002200.060.06050混凝土砌筑07809001002301802600.050.1460.0932不砌筑7314684169941894502502800.100.100混凝土砌筑78118901361001522502203000.070.1740.1043不砌筑1462081692401892685202803000.120.120混凝土砌筑1181571361811522022802503200.080.1950.1104不砌筑2082832403262683655503003500.150.150混凝土砌筑1572431812802023133503003500.110.2360.1225不砌筑混凝土砌筑2433142803633134054203703500.140.2680.130表4-3 半倒梯形水沟大巷水沟编号巷道类别流量（m3/h）净尺寸（mm）断面（）每米材料消耗量水沟充满系数坡度宽B深H净掘进盖板水沟345上宽B2下宽B1钢筋（）混凝土（m3）混凝土（m3）1砌碹096011001233503003500.1140.1391.3360.02260.0990.752961971102271232544003504500.1690.2071.6330.02760.12031973402274032544505004505000.2380.2782.0360.03230.13743403974034584505125004505500.2610.3092.0360.03230.14553976294587265128126005506000.3450.3952.4360.03710.16266207277268408129396005506500.3740.4322.4360.03710.17077271018840117593913147006507000.4730.5333.0860.04270.18881018115011751320131414857006507500.5060.5743.0860.04270.195中巷及小于15上下山水沟编号巷道类别流量（m3/h）净尺寸（mm）断面（）每米材料消耗量水沟充满系数坡度宽B深H净掘进盖板水沟34551015上宽B2下宽B1钢筋（）混凝土（m3）混凝土（m3）1砌碹0630730810339048205952502002000.050.1311.0960.01740.0690.7526310673122811363395694828105959993002502500.070.1701.3360.02260.07931061591221831362053002503500.100.2131.3360.02260.095大于15上下山水沟编号备注巷道类 别流量（m3/h）净尺寸（mm）断面（）每米材料消耗量水沟超高（mm）坡度宽B深H净掘进盖板水沟1520上宽B2下宽B1钢筋（）混凝土（m3）混凝土（m3）1无盖板砌碹037904222001502000.040.0830.0485023795784426732502002000.050.0980.053表4-4 矩形水沟1.中巷及小于15上下山水沟编号巷道类别流量（m3/h）净尺寸（mm）断面（）每米材料消耗量水沟充满系数坡度宽B深H净掘进盖板水沟34551015钢筋（）混凝土（m3）混凝土（m3）1锚喷0470580630266038204582002000.040.141.0960.01740.0800.75247865897631122664713826694688203002000.060.1751.3360.02260.095386144971731121913003000.090.2251.3360.02260.1152.大巷水沟编号巷 道类 别流量（m3/h）净尺寸（mm）断面（）每米材料消耗量水沟充满系数坡度宽B深H净掘进盖板水沟345钢筋（）混凝土（m3）混凝土（m3）1锚喷08609701123003500.1050.1441.3360.02260.1140.75286172972051122274004000.1600.2031.6330.02760.13331723022053492273825004500.2250.2722.0360.03230.15243023743494323824725005000.2500.3062.0360.03230.16153745544326624727166005500.3300.3902.4360.03710.18065546626627487168466006000.3600.4292.4360.03710.1897662921748108384612067006500.4550.5283.0860.04270.2088921106910831249120613827007000.4900.5723.0860.04270.2143.大于15上下山水沟编号备注巷 道类 别流量（m3/h）净尺寸（mm）断面（）每米材料消耗量水沟超高（mm）坡度宽B深H净掘进盖板水沟1520钢筋（）混凝土（m3）混凝土（m3）1无盖板锚喷031203631502000.030.1050.0755023124683635512002000.040.1200.080图表的填写由于支护方式的不同，巷道断面特征和材料消耗也不近相同，在列出的表格中要包含以下几个主要方面。巷道断面特征表：主要包括净断面积、掘进尺寸、水沟断面积、支护方式、锚杆参数、粉刷面积等参数。每米巷道工程量及材料消耗表：主要包括锚杆数量、水沟材料消耗、混凝土量、金属网面积等。现列出四种比较典型的断面特征表。（1）锚喷巷道的巷道断面特征表和材料消耗表如图4-10图4-10 锚喷巷道的特征表2.架棚巷道的巷道断面特征表和材料消耗表如图4-11图4-11 木棚巷道的特征表由于金属棚和可缩性金属支架的材料消耗表及工程量表一致，下面就不在列出其工程量表。3.金属棚巷道的巷道断面特征表如图4-12图4-12 金属棚巷道的特征表4.可缩性巷道的巷道断面特征表和材料消耗表4-13。图4-13 金属棚巷道的特征表5 结论本专题主要分析和总结了巷道断面的设计思路，在认真研究巷道断面设计思路的基础上，对其进行完善和补充，形成一套比较严密的设计理论，设计思路具体总结如下。首先确定巷道运输方式，二要确定巷道断面形状，三要确定巷道净面积，四要计算道渣面的铺设高度，五进行风量的验算，六确定支护方式，七确定水沟的形状和位置，八进行管线布置，九进行标注，最后填写图表，通过以上十个步骤即可设计所需巷道。本专题是在查阅大量专业资料和书籍的基础上得出，但考虑作者并没有现场工作的实际经验，以上巷道断面设计思路依然存在很多细节问题，这些问题值得今后再去认真归纳总结。参考文献1中国煤炭建设协会.煤矿巷道断面和交岔点设计规范M.北京:中国计划出版社.2007:27-30.2 国家煤矿安全监察局编. 煤矿安全规程M. 煤炭工业出版社, 2010.3 国家安全生产监督管理局. 煤矿安全规程M. 北京：煤炭工业出版社.2010: 10-26.4德D.盖尔施因.金属矿地下开采的计算机辅助设计.国外金属矿山.1990.5 薛顺勋 .煤巷锚杆文护施工指南. 煤炭上业出版社 ,1999:136徐永圻. 煤矿开采学.中国矿业大学出版社.1999.7林在康,左秀峰,涂兴子.矿业信息技术基础M.中国矿业大学出版社. 209:12-14.8 侯朝炯、郭励生、勾攀峰.煤巷锚杆支护M.徐州飞中国矿业大学出版社，1999.。9 何水源,余索. 计算机进行巷道断面及支护参数的绘制与实践J . 河北煤炭,1994 (4) :31-34.10 单仁亮. 巷道断面的计算机辅助设计工矿自动化2009(2):85-87.11 毕映楷. 巷道断面设计软件的开发与应用. 科技情报开发与经济.2006 (2):232-234.12 刘勇,何元东. 标准巷道断面自动查询绘制系统J . 煤炭工程,2001 (10) :59-60.13刘文清. 论巷道断面的确定方法. 煤炭工程2009（2）:12-13.14王如林. 巷道断面设计最优方案探讨煤2002(4 ):47-49.15张园园，杨胜强. 巷道断面优化新方法 煤矿现代化2010 (2):20-21.16卞恩林, 王利民. 巷道断面最佳设计煤炭技术1999 (2):14-16.17 胡建华. 巷道围岩稳定性分类及其支护对策智能研究. 武汉理工大学硕士论文. 2001.5.18董方庭，宋宏伟，郭志宏等.巷道围岩松动圈支护理论J.煤炭学报，1994. 任务书学院 专业年级 学生姓名 任务下达日期：20xx年1月8日毕业设计日期：20xx年3月12日 至 20xx年6月8日毕业设计题目：平煤四矿5.0 Mt/a新井设计毕业设计专题题目：简析煤矿巷道断面设计思路毕业设计主要内容和要求：以实习矿井平煤四矿条件为基础，完成平煤四矿5.0Mt/a新井设计。主要内容包括：矿井概况、矿井工作制度及设计生产能力、井田开拓、首采区设计、采煤方法、矿井通风系统、矿井运输提升等。专题内容是关于煤矿巷道断面设计思路的简要归纳总结，在查阅大量资料的基础上，简要总结出了煤矿巷道断面的设计思路和设计步骤。完成2009年国际岩石力学与采矿科学杂志上与采矿有关的科技论文翻译一篇，题目为“Determination of the most effective longwall equipment combination in longwall top coal caving (LTCC) method by simulation modeling”，论文3763字符。院长签字： 指导教师签字：翻译部分英文原文Determination of the most effective longwall equipment combination in longwall top coal caving (LTCC) method by simulation modelingFerhan Simsir*, Muharrem Kemal OzfiratMining Engineering Department, Dokuz Eylul University, 35160 Buca, Izmir, TurkeyReceived 12 January 2007; received in revised form 2 November 2007; accepted 21 November 2007Available online 4 March 2008 1. IntroductionIn order to recover limited resources in underground efficiently, the most suitable production equipment and method must be applied to a colliery. Trying all alternatives and equipment combinations would be a very expensive challenge. Computer simulation, on the other hand, is a cost-effective tool for evaluating what-if scenarios in mine development and ore production. It is a useful tool for analysing complex systems such as factories, health care networks, logistics, and service-type operations. There are many different computer simulation types, such as discrete event, continuous, hybrid, and so forth. Discrete event simulations process discrete events that occur at random times through a central processing unit. In discrete event systems many events can occur simultaneously 1.Discrete event simulation languages, available since the 1960s, provide general facilities to model the operations involved in processing and manufacturing as discrete events occurring in time. The case under study, the fully mechanized longwall operation in Omerler, a colliery of the state-run Western Lignites Corporation, is suitable for discrete event simulation since movements of shearerloader and roof supports are occurring simultaneously.In Turkey, caving methods are mostly employed in mining of thick coal seams as long as the roof strata are suitable for their use. Longwall with caving is always preferred to stowing faces because of its simplicity, favourable economics, and high productivity. It is assumed that the upper bound of applying single-pass longwall (SPL) method as a mechanized system in thick coal seams is about 6m 2. If the thick coal seam cannot be mined by SPL, then multi slice longwall (MSL) can be employed. However, for thick seams, MSL is less convenient, less economic, and requires more labour compared to longwall top coal caving (LTCC) method. When choosing which method to employ, the features of the seam also need to be considered.The LTCC method offers a viable means of extracting up to 7580% of seams in the 59m thickness range. Suitable geomining conditions in Europe and many other countries have led to a wide range of applications of caving coal mining faces during underground working including sublevel caving of thick seams. The LTCC method is increasingly used in thick seam mining, for example, there are over 70 LTCC faces operating in China 3. In addition, in Australia, the LTCC method has successfully been used by several Chinese companies. The initial thickness, typically 3m, is cut and loaded conventionally with a shearer and front AFC. The remaining top thickness of coal, typically an additional 39 m, is allowed to cave into the rear AFC. By this way, coal recovery is increased to 85% from a 9m seam 4.In 1988, Senkal et al. found the coal loss to be 24.3% in the same underground mine. However, at that time, longwall equipment consisted of hydraulic props+steel roof bars as roof support, an AFC, and loosening blasting+pneumatic picks as winning method. The most important disadvantage of the LTCC method is that significant coal losses may appear when drawing the coal through the support window. Therefore, in this study, firstly the coal loss is figured out. After that, coal production sequence in the colliery is modelled dynamically by simulation. Real data collected from the system (i.e. the colliery itself) as well as the coal loss computed are used in the simulation model. The whole longwall operation is simulated using computer software, and daily production figures from face and the top are achieved.Here, the longwall, approx. 86m long, is set up on the floor part of the seam, and coal left on top is drawn through the roof support chute on the gob shield onto the front AFC. The thickness of the seam is variable due to layer formation, but it is 8.5m on average (Fig. 1). Of this 8.5, 3m is mined from the longwall, the rest 5.5m caves in. The longwall equipment used consists of a double-ranging shearer-loader (Eickhoff EDW-150-2L), an AFC (SGZ-730/264, Chinese manufacture) and 56 roof supports(CMEC ZYD 4000/18/32, Chinese manufacture) with chutes on the gob shield to draw the top coal. Fig. 2 gives the plan (a) and the cross-sectional views (b) of the longwall.In this study, equipment used in the longwall and effective on the whole panels operation have been changed in the simulation model one by one creating 320 distinct points to be executed. Sixty additional design points have been created by adding the cut numbers of the shearerloader to these two equipment, too. After executing these points in the model, best results for equipment combination and number of cuts of shearer-loader have been found out.2. SynopsisSimulation modelling turned out to be very suitable for this study since there is no chance to test alternative production methods during the ongoing work of the mine.It would be very costly to set up a pilot face to test alternatives, also. By the use of simulation, it is possible to test the effects of new methods and factors economically and quickly. Imitating the operations of real-life systems or processes is the main purpose of computer simulation.Operational scenarios can be tested and evaluated without the need or expense of physical experimentation. Applications have been developed to simulate the space and time relationships between mining equipment, mainly in connection with transport systems 68. Tsiflakos and Owen 9 discussed the philosophy, methodology, objectives,mining process logic, program structure and its stateof-the-art for a recently developed mine simulation model.By Connor et al. 10, a two-dimensional rigid block computer model was used to simulate discontinuities within the strata overlying a longwall coal mine. Hunt11 simulated the ore haulage in Henderson molybdenum mine. The purpose of the simulation was to demonstrate to mine management how simulation could be used to assist in optimizing ore transport using existing trains, trucks,and ore passes.Today, there exist specially designed high-level simulation programs. The most commonly used simulation languages are GPSS (General Purpose Simulation System),GPSS/H (event-driven version of GPSS) and ARENA.Also, Vagenas 12 and Sturgul 13,14 have applied discrete event simulation to both underground and openpit mining operations using GPSS and GPSS/H. SIMAN was used by Tan and Ramani to study belt networks. Kolonja 15 used SIMAN for studying the various dispatch criteria for open pit mines. Many of the mine simulations done using ARENA were proprietary in nature and did not appear in the literature.In addition to these studies, web-based simulation programs have also been made presenting a newly developed,user-friendly visual simulation computer tool which helps mine operators to plan the optimum mining sequence for different mine geometries and equipment layouts.Another study contributes to the body of knowledge by developing a robotization and stability control (RASC) model for operating dump trucks. Konyukh and Ramazanov 16 aimed in their study the optimal use of LHD vehicles working in underground by an operator at surface using the GPSS/H simulation language.When searching the studies carried out up to now, it can be seen that especially haulage equipment have been simulated at the underground mining methods room-and pillar and sublevel caving so far, and, longwall mining has not widely been subject to simulation.3. Calculation of coal loss In order to find out the coal loss, ash content of samples taken from face coal, top coal, and from the belt conveyor in the main gate to represent the whole panel, are analysed.To determine the real seam thickness and its properties,samples are taken from the face at 1015m intervals and every web. Also thickness of top coal is found by the help of samples taken from the beginning of top coal up to overlying roof strata (Fig. 3). To find out the ash content of top coal, and so to determine top coal recovery, samples are taken from belt conveyor in the main gate every 2 h. Samples taken from the face, top coal, and belt conveyor are separately brought together and mixed.These joined samples are quartered and 3040 kg of them is used to determine ash content and density. Totally six experiments are performed on samples and the outcomes are averaged to give the final result. In addition, samples taken from overlying roof strata are put into drying oven to find the ignition loss. Except tail and main gates, the length of longwall is 86 m. Since there is coal on top of both main and tail gates, the length of top coal including these ways is 93.2 m. It is assumed that coal from longwall face is mined with a recovery of 100%. Therefore, coal loss is caused during drawing of top coal only. To determine the coal loss, the weighted average of ash content of different materials is used. Five field experiments are carried out and the values are given in Table 1.4. Simulation study Simulation is setting up a model to represent a system.This model provides a chance to test operations that are infeasible to test on the real system. Simply, it can be defined as computer trials. Simulation has many applications in mining and mineral processing. In underground mining, its benefits include increased production,capital and operational cost savings, and improved prediction 17.Firstly, the flow chart of operations in the mine is figured out. Once the shearer-loader performs two cuts,AFC is pushed towards the face, and then, roof supportsstart moving forward. As the first roof support moves two webs forward, the top coal caves in and is drawn through the roof support chute. Then the second roof support starts moving, and this operation repeats until all roof supports are moved ahead, then the shearer-loader starts cutting a new web. The assumptions, parameters, variables, and attributes of the model are given in the following. In the real system, the shearer-loader moves continuously and hence feeds the AFC continuously. In this study, it is assumed that coal cut by shearer-loader is loaded onto the AFC once during one cut which takes place in the middle of the cut. The coal amount loaded onto the conveyor is equivalent to the coal produced by one cut of shearer-loader.The flow diagram is turned into a simulation model in ARENA 2.2. In the model, first of all, the motion is created(with the CREATE block) and the shearer-loader starts its first cut. Region I expresses the operations of shearerloader and the roof supports, in other words, coal production operations. During this cut, amount of coal produced is computed and assigned to a variable (AMOUNTCOAL1).When the shearer-loader finishes its first cut, it is released (RELEASE block) and it spends a certain time to set up its drum for a new cut. Then the program controls whether this was the first (CUTDIRECTION ? 1) or the second cut (CUTDIRECTION ? 2) of shearer-loader (BRANCH block). If it is the first cut, the motion goes back to the shearer-loader and makes the second cut. If it is already the second cut, the coal produced is sent out to belt conveyor. Region II expresses the movement through the chain conveyor, stage loader, crusher and belt conveyor,respectively.After the shearer-loader, the motion moves to roof supports and the first roof support moves forward. The chute is opened and top coal is drawn. At this point, the amount of coal drawn is assigned to a variable(TOPCOAL). Then, similar to face coal, top coal produced is sent out through conveyors. At this point, the program controls the number of roof supports (T) moved forward(BRANCH block). If 56 roof supports have finished their operation, then face-end and T-junction supports are moved. If not, the motion moves to the next roof support. After all roof supports are moved, the motion goes to the very beginning and the shearer-loader starts cutting again. The motion never stops since the mine works 24 h continuously. After creating the model,it is tested and compared with the current operating values. Average daily production value obtained by the model is 1046 tons and the real daily production of the mine is 922.93 tons. Since these results are close to each other, the model is verified. In addition, the model is tested with extreme values for parameters and its validity is proved.5. Creating the simulation modelWhile creating the frame of the experimental design, all equipment in the panel have been accepted as effective factors. As seen in Table 2, the shearer-loader has five different types including the existing one currently used, and the roof support has four different types including the existing one. For the remaining equipment (AFC, stage loader, crusher, belt conveyor), two different types have been investigated. All combinations of these values of longwall equipment give a total number of 5_4_2_2_2_2 ? 320 design points to be simulated.After first investigations it has been seen that different types of coal hauling equipment (AFC, stage loader, crusher, belt conveyor) are not effective on production figures. The reason for this is the low difference among the haulage durations and the capacities of these equipment, and that these values do not cause a bottleneck in coal production. In the second stage of the experimental design, the effects of shearer-loader and roof support types, including number of face cuts, have been investigated. In Table 3, it can be seen that the second stage comprises three distinct factors and 60 design points, which are simulated for a period of 30 days (Table 4). Tables 5 and 6 give the features of different roof support types used in the simulation, where shearer-loader features remain the same as in Table 2.6. Evaluation of computational resultsHaving determined the effective shearer-loaders and roof supports at design points, these are combined with number of cuts of shearer-loader. These design points have been computed for the three working rhythms (i.e. drawing the top coal after one, two or three shearer travels along the face) and results are obtained (Figs. 4 and 5). As can be seen in Fig.4, the design points delivering the biggest coal production figures are 6, 10, and 14 for the single-cut system, 26, 30 and 34 for the double-cut system, and, 46, 50 and 54 for the three-cut system, respectively. The equipment selected for these points and their features are given in Table 7 and in Table 3 in detail.The results indicate that the most efficient production system is the one in which top coal is drawn after two face cuts performed by shearer. In three cuts production system,top coal is left quite far from the roof support, and the overlying strata cave avoiding top coal to cave in through the roof support frame. On the other hand, in one cut production system, dilution of coal with dead rock increases and work organization within shifts becomes unstable.In all three types of production systems, RS1 turns out to be the roof support providing optimum production amounts. RS1 is similar to the current roof support but differs in terms of frame and support dimensions. Type of shearer-loader is not very decisive on the production amount, therefore, roof support is the primary factor effective on production efficiency of the LTCC method.7. Discussion and conclusionBy investigating the results it can be seen that equipment except the shearer-loader and the support units are not effective on the production amount, and the design points delivering the biggest coal production values are 26, 30 and 34. The shearer-loaders used at these points are SL1, SL2, SL3, whereas the best roof support unit has been RS1 in view of obtaining the highest value of top coal production.Drawing of top coal after two face cuts being currently applied in the colliery delivers bigger production values compared to other systems (drawing top coal after single or three face cuts). Therefore, also in case of alternative longwall equipment to be selected, this working rhythm should be kept in the mine. The coal production figures of this working style points to a high top coal recovery, which is a significant factor in applying this method efficiently.In the future research of this study, geomechanical properties of top coal and overlying strata should be included. So, whether the cavability properties and breaking distance of top coal and overlying strata are close to each other should be examined. In case the cavability values of top coal and overlying strata are close to each other, pressured-water injection can be applied to top coal in order to increase its cavability. By this way, loss of top coal can be decreased as well as decreasing dilution with dead rock. Also, the probability of roof support frame to be plugged by big coal blocks would be decreased.References1 Schriber T, Brunner DT. Inside discrete-event simulation software: how it works and why it matters. In: Proceedings of the winter simulation conference, Atlanta, 1997. p. 1422.2 Kose H, Tatar C. 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Simulation of mining systems by objectoriented graphical modelling. Int J Rock Mech Min Sci 1993;30:A108.10 Connor O, Dowding KM, Engng CH. Hybrid discrete element code for simulation of mining induced strata movements. Int J Rock Mech Min Sci 1993;30:A67.11 Hunt C. Simulation model of ore transport at the Henderson mine. Comp Geosci J 1994;20:7584.12 Vagenas N. Applications of discrete-event simulation in Canadian mining operations in the nineties. Int J Surf Min 1999;13:778.13 Sturgul JR. Discrete mine system simulation in the US. Int J Surf Min 1999;13:3741.14 Sturgul JR. Mine design: examples using simulation. Colorado: SME,2000.15 Kolonja B. Simulation analysis of dispatching strategies for surface mining operations using SIMAN. MSc thesis, Penn State University,1992.16 Konyukh VL, Ramazanov RA. Control of underground loader-hauldumpers from the surface. J Min Sci 2004;40:3749.17 Hustrulid WA, Bullock RL. Underground mining methods. Colorado: SME; 2001.中文翻译在长壁放顶煤采煤法中通过仿真模型优化长壁工作面设备组合Ferhan Simsir*, Muharrem Kemal Ozfirat采矿工程系，Dokuz Eylul大学，35160 Buca,伊兹密尔,土耳其于2007年1月12日收到初稿；并于2007年11月2日收到修改文章；此后于2007年11月21日录用该文；2008年3月4日可上网查询1、简介为了有效地利用有限的地下资源，最合适的生产设备和采煤方法必须应用到煤矿。然而尝试所有可选择的设备组合将花费巨大。而另一方面，计算机模拟在矿山开发和控制生产成本方面非常有效地工具。这是一个在工厂、卫生系统、物流业以及服务型企业中分析复杂系统的非常有用的工具。现在有许多不同的计算机模拟类型，如离散事件，连续，混合类型等等。离散事件模型通过一个中央处理单元来处理离散事件。在离散事件系统的许多事件可以同时发生1离散仿真语言自1960年问世以来，可以及时提供一般的设施来模拟包括加工和制造作为离散事件发生时所涉及的操作。根据研究所得，在奥莫洛的一个国营的西方褐煤企业的煤矿中，综合机械化长壁采煤法中滚筒采煤机和顶板支护的同时发生适合作为离散事件来仿真。在土耳其，放顶采煤法大多应用于开采厚煤层并且顶板要适于放顶煤开采的条件。长壁放顶采煤法作为首选是因为的它的操作简单、经济效益高、且有较高的效率。据推测，单一走向长壁采煤法在应用机械化系统开采厚煤层时的上限是6m左右。如果厚煤层不能应用单一走向长壁采煤法，那么可以分层开采。然后厚煤层分层开采相比长壁放顶煤采煤法操作复杂、经济效益低且需要更多劳动力。当考虑选择哪种方法采煤时，对于煤层的特征也应加以考虑。长壁放顶煤采煤法提供了一种在开采5-9m的厚煤层时可以达到75%-80%采出率的方法。在欧洲和其他许多国家因条件适宜，放顶煤采煤法在厚煤层开采中已得到了广泛的应用。长壁放顶煤采煤法已经日益广泛用用于厚煤层开采中，例如在中国已经有超过70个长壁放顶煤采煤工作面。此外，在澳大利亚，长壁放顶煤采煤法也已经在几家中国公司中成功应用。一开始，通常是3米左右的煤被滚筒采煤机和刮板输送机割下并被运走。然后额外的3米到9米的顶煤被放在后面的刮板输送机上。通过这种方式，9米厚的煤层的回收率达到85%左右。而在1988年，Senkal发现在同一个煤矿煤炭损失率竟达到24.3%。然而在那个时候，长壁工作面装备由液压支柱、工字钢支护顶板、刮板输送机和松动爆破炸药加风钻组成。而长壁放顶煤采煤法的最大的缺点是当通过放煤窗口放煤时会出现煤炭损失。因此，此项研究首先计算煤炭损失。然后建立煤矿煤炭生产序列动态仿真系统。系统实时收集数据与此同时在仿真系统中计算煤炭损失。整个长壁工作面使用计算机软件模拟，并且从工作面获得日常生产数据。这里，顶煤从支护顶板支架的斜槽中漏在工作面长达86米的前刮板输送机上。煤层厚度的变化时由于层位的形成不同，但平均厚为8.5米。（图1）这8.5米中的3米厚的煤从工作面上采下，而其余5.5米厚的煤从顶上放落而下。长壁工作面装备由一个双滚筒采煤机（Eickhoff EDW-150-2L）、一个刮板输送机(SGZ-730/264, 中国制造)，以及56个带放煤口支架(CMEC ZYD 4000/18/32, 中国制造)组成并将顶煤从上方采下。图2中的a图为采煤方法图，b图是长壁工作面的剖面图。在这项研究中，在长壁工作面中使用并使整个盘区受益的装备已经一个又一个的改变了仿真模型的320个节点。60个额外的设计点也已经添加到采煤机的截割数据中。在模型中执行这些点后，对于这些装备组合和采煤机割煤的设计点的最好结果是被发现。2、概要仿真模型被证明是非常适合本研究，因为当工作正在采煤时没有机会实验其它可能的采煤方法。而设立一个实验工作面来实验可能的设备花费巨大。通过使用仿真模型，他可以最经济和快速的前提下实验这些新方法和特征的益处。模拟真实系统或过程的操作是电脑仿真系统的主要目的。操作场景可以测试和评估,而不需要或费用的物理实验。应用程序已经发展到可以模拟采矿设备与时间和空间的关系，其中主要模拟运输系统的连接810。Tsiflakos和Owen11讨论了哲学、方法论、目标、采矿程序的逻辑，程序结构以及新近开发的采矿仿真模型。Connor 12则通过一个二维的刚性块电脑模型来模拟长壁采煤工作面覆盖的不连续地层。Hunt则模拟亨德森钼矿的运输系统。模拟的目的是为了证明开采管理如何模拟可以最优化用矿车和胶轮车运输矿石。今天,有特别设计的高层次仿真程序。最常用的仿真语言是GPSS (通用目的的模拟系统), GPSS / H(事件驱动版本的GPSS)和舞台。Vagenas14和Sturgul15,16应用离散事件仿真对井工矿和露天矿作业都是用GPSS和GPSS/H.SIMAN被Tan和Ramani17用来研究带网络。Kolonja 18使用SIMAN研究各种不同露天矿山的调度标准。除了这些研究,基于网络的仿真程序也呈现出一个新开发的、用户友好的视觉模拟计算机的工具,这个工具帮助矿井经营者面对不同的煤矿采用最优的顺序开采并安排开采设备。另外一项研究通过开发一个稳定控制系统模型来操作自卸卡车20。Konyukh和Ramazanov21计划在他们的研究中由在地面上的操作者使用GPSS/H仿真语言来使井下的机械得到最优化的利用。当研究进行到了目前为止，尤其是运输设备模拟了在地下采矿方法的支柱和无底柱崩落法，然而到目前为止,长壁开采并没有广泛受到仿真。3、计算煤炭损失为了计算煤炭损失，灰分的标本来自面临煤炭、主要煤炭,从带式输送机在大门前代表整个面板,进行了分析。为了确定真实煤层的厚度极其属性，样品在工作面每间隔10-15米取样。此外，放煤厚度发现有助于样品从来自上覆岩层，以找出放煤的灰分，从而确定放煤循环，在主石门的胶带输送机上每隔2小时取样。样品取自工作面，顶煤和胶带输送机上，然后分别聚集并混合。这些混合的样品按30-40千克分为4份然后用来确定灰分和密度。总共六个实验样本，结果取自六个实验的平均值。此外，从上覆岩层取来的样品放入烤箱中干燥，以得到损失数据。除了主石门，长壁工作面的长度为86米。由于有煤在主巷和尾巷上方，有放顶煤的工作面的长度包括这些巷道的长为93.2米。假定长壁工作面的煤炭采出率达到100%。因此，煤炭的损失仅仅是由于放煤而造成的。为了确定煤炭损失，则用到了灰含量不同的材料。进行的五次试验数据列在表1中。表一研究领域全部煤层稀释百分比（%）煤炭损失百分比（%）顶煤1全部煤层25.5823.64顶煤38.5934.752全部煤层20.710.83顶煤35.4416.483全部煤层34.8715.73顶煤41.6723.494全部煤层23.7615.84顶煤38.2423.095全部煤层29.617.16顶煤52.4225.83平均值全部煤层24.916.64顶煤41.2724.734、仿真研究仿真研究是通过建立一个模型来表示系统。这个模型提供一个机会来测试那些不可在真实系统中测试的操作。简单的说，这可以被定义为计算机试验24。仿真系统在采矿和矿物加工中有很多应用。在地下采矿中，它带来的好处包括增加产量，节约运营费用，并改善预测。首先，矿井操作流程图已在图四中展示出来了。当滚筒采煤机割过两刀煤后，刮板输送机被推倒工作面的前方，然后液压支架开始向前移动。随着第一个液压支架向前移动两个步距，顶煤从液压支架后面的卸煤口中放下来。然后第二个液压支架开始移动，这些操作重复进行直到