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专题浅析煤矿岩石巷道支护 含CAD图纸+文档 永城 矿区 陈四楼 1.5 Mta 设计 专题 浅析 煤矿 岩石 巷道 支护 CAD 图纸 文档

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专题部分 第19页浅析煤矿岩石巷道支护摘要：煤矿岩巷支护问题尤其是软岩巷道支护问题一直是困扰煤矿生产和建设的难题之一，由于巷道分类和变形破坏原因的复杂性和多样性，导致岩巷的支护问题很难解决。目前，我国大多数矿井已进入深井开采，所以岩巷支护日趋重要。近年来，随着科技的迅速发展和煤矿科研人员的不断努力,煤矿岩巷支护的技术与手段已趋于成熟。本设计主要从围岩破碎带范围和破碎范围等情况来论述岩巷支护形式。关键词：岩巷；支护形式；巷道变形原理；锚喷支护；锚注支护1问题的提出岩石巷道支护等理论，国内外尚无定论，给工程应用带来了诸多不便。大多数生产单位处理岩巷支护问题时，仍处在经验支护状态，这样导致盲目性大、成功率低，很多巷道出现了“前掘后修，前修后坏”的局面，每年的修复费用巨大，这样不仅造成巨大的经济浪费，使整个矿井生产陷于困境，甚至导致矿井关闭，并给煤矿安全生产带来极大的危害。巷道变形破坏、片帮冒顶等事故在地下工程中是最常见的。据不完全统计，煤矿事故中59%以上是巷道事故。究其原因，还是对巷道变形破坏规律认识不清、支护理论不完善，从而造成支护设计工程类比居多，缺乏科学的指导，巷道支护方式选择不合理，因而也就无法保证巷道在不同地质条件下稳定和安全使用。在软弱岩层中施工巷道,掘进较容易,维护却极其困难,采用常规的施工方法和支护形式、支护结构往往不能奏效。因此，软岩支护问题是井巷施工的关键问题。目前我国软岩矿井多是既软弱又有膨胀性，相当部分矿井为高应力状态复合型软岩矿井。在这种条件下，基建矿井每米巷道的掘进工程费已高达12万元以上；一个120万吨的生产矿井软岩巷道年维修费用高达1亿元以上，在目前市场经济条件下已经到了煤矿无法承受的地步。另一方面，随着我国新生代煤层的大力开发，软岩矿井数量与日俱增。此外，统计表明：我国立井的深度在20世纪50年代平均不到200m，而90年代平均已达600m。相当于平均每年以10m的速度向深部发展；生产矿井1980年平均开采深度为288m，而1995年平均深度为428m，平均年降深9.39m。事实上，大部分软岩生产矿均在向深部转移，一部分老矿区开采深度多为500700m。新建矿井深度多为600900m，在高应力作用下这些软岩矿井的巷道开掘和维护都将十分困难。所以，随着我国煤炭生产建设发展，软岩支护问题必将日益严重。2支护形式选择岩石支护是用来提高井下巷道围岩稳定性和维护围岩自撑能力，它分为岩石加固的主动支护和岩石支撑的被动支护。非预应力灌注式树脂锚杆是一中主动支护，它随围岩变形就张紧。型钢支架是一种被动支护，它只有当围岩失稳逐渐向巷道内移动时才起作用。井下巷道支护形式的选择，取决于围岩破碎带范围和破碎程度。I围岩属硬岩、低应力区，可采用光面爆破，不支护或树脂锚杆（图1a、b)或机械锚固式锚杆（图2a），挂金属网，防止小岩块冒落。树脂锚杆多为端头锚固型，即用树脂为粘结剂，在固化剂和加速剂作用下将锚杆的头部粘结在锚杆孔内。它具有凝结硬化快、粘结强度高、安全可靠、施工操作简便、适用范围广等优点，且控制围岩位移和抗震性能好，可以对围岩施加预压应力，在很短时间内便能达到很大的锚固力。树脂锚杆由树脂药包和杆体组成。树脂药包的外袋用双层玻璃纸或聚酯薄膜做成，内装树脂、加速剂和填料；内管为玻璃小管，玻璃小管内装有固化剂和填料。普通树脂锚杆我国使用的树脂锚杆以金属杆体为主。早期的锚杆杆体为圆钢，直径1418mm，其头部加工成反麻花形，以利增加粘结和搅破药包；为防止安装搅拌时树脂外流和保证锚杆固长度，在麻花结构尾部焊有挡圈。目前，锚杆杆体主要以螺纹钢为主，头部削成斜坡状以利搅破药包。钻孔直径为2842mm，杆体直径为1622mm，长度从1800mm到2400mm不等。树脂锚固剂为高分子化学材料，其粘结能力强、固化速度快、耐久性好、抵御环境和人为影响因素能力强。锚固剂是由两种不同组分的数值胶泥和过氧化物固化剂严格按科学配方分别包装而成，凝结时间可按设计要求在十几秒到几小时内准确调控，其结构见图2。树脂锚固剂的型号、规格见表1和表2。图2 树脂药包示意图 表1 树脂锚固剂规格型号规格/mm质量/g适用钻孔/mm使用范围35373537070010422井筒装备安装35303530055010422巷道锚喷支护端锚28352835040010322巷道锚喷支护及其它28502850064010322巷道支护和全长锚固23352335030010282巷道小直径支护和全长锚固23502350043010282巷道小直径支护和全长锚固 表2 树脂锚固剂主要型号和特征型号特性凝胶时间/min固化时间/min备注CK超快0.515在2010C环境温度下测定K快速1.527Z中速3412M慢速152040CM超慢120240等强度螺纹钢锚杆由于目前常用的锚杆杆体直径均大于锚杆杆尾螺纹部分直径，这就造成杆尾部分与杆体部分强度产生较大差异，因此在现场使用中，经常在未达到锚杆规定的极限强度前就在杆体与杆尾螺纹交界处产生拉断破坏。因此，要提高同等规格锚杆极限破断强度，就必须改变锚杆杆尾螺纹的加工工艺。为使锚杆整体强度真正达到等强。可采用滚压无热处理技术新工艺。该工艺简单易行，采用特殊机械滚压加工技术，使丝扣段螺纹配合较大型号螺母，如18mm、20mm的锚杆杆体配合M22、M24的螺母，从而使丝扣段强度与杆体强度保持一致（表3）。因此，等强度螺纹钢锚杆具有较好的支护性能。单向左旋无纵筋螺纹钢锚杆目前所使用的螺纹钢大部分分为双向纹两筋螺纹钢，虽然螺纹钢能与锚固剂有较好的结合，但在注入锚杆时，杆体纵筋螺旋转半径大于杆体螺纹钢旋转半径，从而造成杆体螺纹钢不能与树脂胶体紧密结合而产生较强的握裹力；同时，双向螺纹不利于锚固剂充填密实，因此可降低锚固强度。针对双螺纹钢存在的缺陷，可对锚杆杆体表面结构进行优化。将锚杆杆体专门轧制成单向无纵筋螺纹钢，取消纵筋，螺纹为单向左旋，与锚杆注入时旋转方向一致。在旋转注入锚杆时，在单向左旋螺纹旋转作用下产生强有力的压力向深部推进，在此压力作用下呈液体状态的树脂锚固剂可以充填孔中裂隙和排出孔中污水，增加锚固剂和锚杆杆体之间的握裹力以及锚固剂与岩体之间的粘结力，可以有效地提高锚杆的锚固力。单向左旋无纵筋螺纹钢可用于各类矿山和地下工程，特别是在不稳定地层巷道支护中取得较好的技术经济效果。 表3 普通锚杆与等强锚杆破坏强度对比表杆体直径/mm锚杆类型丝扣段直径/mm配合螺母/mm破坏部位破坏强度/kN加工工艺18普通锚杆16.2M18丝扣段117车削、损失加工等强锚杆18.2M20杆体段145车削、无损失加工20普通锚杆18.2M20丝扣段145车削、损失加工等强锚杆20.2M22杆体段178车削、无损失加工22普通锚杆20.2M22丝扣段178车削、损失加工等强锚杆20.2M24杆体段221车削、无损失加工玻璃钢锚杆玻璃钢锚杆是采用玻璃纤维作为增加材料、以聚酯树脂为基材，经专用拉挤机牵引，通过预成型模在高温高压下固化为全螺纹玻璃纤维增强塑料杆体，加上树脂锚固剂、托盘和螺母组成玻璃钢锚杆。玻璃钢杆体具有可割性，很适合于综采工作面临时支护使用，而且具有良好的防腐性能，可以部分取代钢锚杆以节约钢材。这种玻璃钢锚杆杆体可切割、轻质高强、便于安装；成本低，与金属锚杆相比成本可降低40%左右，可替代现有煤帮金属锚杆、木锚杆、竹锚杆等进行煤帮支护。II硬岩、低应力区、节理、层里面局部破碎，可采用机械锚固式锚杆加拉条钢带，对硐室和破碎机房，可用注浆锚杆支护。机械锚固式锚杆是井下最早的岩石加固方式，加拿大仍然常用它。若岩石对锚杆有足够的握裹力，安装良好的张壳式锚杆（图3a），可达到允许载荷力。若锚杆超载，托盘处螺纹或锚杆锚固端螺纹损坏，也不会由于锚杆滑动造成锚固失效。预应力锚杆对维护开挖巷道围岩松动岩块的稳定最有效。岩块松动可能是岩石节理、层理发育造成的，也可能爆破震坏导致的。松动岩块冒落都会危机工作环境安全，因此需进行某种形式支护的。由于围岩的松动波及不深，支护只需要承受松动岩石的自重。用机械式锚固和挂金属网可有效的支护。锚杆预应力要求极限断裂载荷的70%左右。机械锚固式锚杆存在的问题，一是锚固端随时间推移就滑移，可能是由于爆破震动引起的；二是锚体受含硫化物地下水的腐蚀，有时不到一年就失效，这时应采用全长注浆锚杆。图3 机械锚固式和机械摩擦锚杆由于爆破和防腐技术的改进，减少了围岩松动的范围，机械锚固式锚杆和金属网用量降低。注浆式或摩擦式锚杆能克服机械锚固式锚杆的缺点，采用全长注浆和全长摩擦锚杆，如树脂锚杆（图1）和管逢式锚杆（图3b）和胀管式锚杆（图3c）。即使锚杆发生滑移或托盘破裂，留在岩层中那段锚杆仍然锚固着，继续提供支护抗力。注浆式和摩擦式锚杆的缺点是不能张紧，所以必须在岩石有明显移动之前安装。爆破的精心操作，锚杆尽可能靠近掘进工作面安装，可为不同条件的围岩提供非常有效的支护。比机械锚固式锚杆应用范围大得多，安装锚杆紧跟工作面，相应减少围岩变形量，加强岩块间的结合。随着注浆式或摩擦锚杆的发展，非预应力锚杆的广泛应用于采矿业，并将成为岩石加固主要手段。井下用的大部分岩石锚杆和树脂锚杆都可以用钢丝绳砂浆锚杆代替，它不论是否张紧，都可以安装和注浆。近年来研究的钢丝绳简单的张紧技术，以取代以前复杂的张紧方法。用钢丝绳砂浆锚杆加固放矿溜道、研石溜道很有效。露出围岩的钢丝绳逐渐破损也不会减少留在岩层中那段注浆钢丝绳的锚固力。钢丝绳的柔性允许岩层在其强度没有明显削弱时的自撑。采场在开采和回填时，长钢丝绳砂浆锚杆可埋在要回采的矿物体内，以便支护回采工作面的上璧。随着钢丝绳的逐渐截短，钢丝绳剩余段由于工作面上部岩层下移而张紧，从而产生锚固力。当钢丝绳减少至2m时，新钢丝绳在第一批钢丝绳处搭接安装，用于承担第一批钢丝绳被截后的支护。当支护大范围潜在的失稳岩层时，若邻近的是断层带，张紧钢丝绳砂浆锚杆可在工作面回采前从准备巷道内安装。钢丝绳砂浆锚杆也成功地用土木工程，如地下大动力室吊车起重梁的锚固和大跨度隧道的支护。拉条带钢。当巷道围岩为层状结构时，即大部弱面为沿一个方向移动，岩石沿弱面方向的强度高于其垂直方向。采用拉条钢带比金属网更有效。拉条安装在锚杆之间，且垂直于弱面走向(图4)。平行于弱面走向安装拉条通常效果差些。图4 拉条III低应力区、爆破造成弱面和少量围岩破坏，采用预应力锚杆挂金属网支护。在使用的金属网中，选定锚杆间距为岩层弱面平均层间距的3倍。假设节理和弱面产生的岩块平均边长约0.5m，则理想锚杆间距应1.5m，锚杆长度为间距的2倍，则为3m。但如果节理平均间距为0.1m，锚杆安装中心距0.3m是不可能的，这时用金属网支护托盘间的小块岩石。金属网可以是链接，也可以是焊接，链接金属网(图5)易弯曲，有很高的承载能力，但安装困难，且不适宜喷射混凝土，这是喷射后，链接后面的混凝土空穴难以消除。焊接金属网在金属丝方格交叉点上施焊，它有很好的刚性比链节金属网好安装。网节小，不妨碍混凝土喷进它的后面，所以焊接金属网适宜于喷射混凝土。图5 链接金属网IV低应力区、节理发育，采用喷射混凝土支护，喷层厚度50mm，加硅粉和钢纤维。若喷层失效，打管缝式(图3b)或胀管式锚杆(图3c)，挂金属网。喷射混凝土支护，是以压缩空气为动力、用喷射机将细骨料混凝土以喷射方法覆盖到需要维护的岩面上从而凝结硬化后形成混凝土结构的支护方式。喷射混凝土可以单独使用，在岩石、土层面或结构面上形成护壁结构，成为喷射混凝土。 喷射混凝土也可以和锚杆、预应力锚杆（锚索）联合使用，形成以锚杆等为主的支护结构，简称锚喷支护。联合作用时，主要是用于避免锚头部位锚杆间岩土体的松脱和风化，可以起到加强锚杆等锚固构件的作用。由于喷射混凝土的施工工艺和普通混凝土有很大差别，因而其物理力学性能和对围岩的支护特性有以下特点：混凝土在高速喷射过程中,水泥颗粒受碰撞冲击,混凝土喷层受到连续冲实压密,而且喷射工艺又允许采用较小的水灰比,因此喷射混凝土层具有致密的组织结构和良好的物理力学性能。特别是其粘结力大，能同岩石紧密粘结，是形成喷射混凝土独特支护作用的重要因素。喷射混凝土能随巷道掘进及时施工，由于加入速凝剂后其早期强度成倍增长，因而能够控制围岩的过度变形和松弛。喷射较薄具有一定柔性，可以和围岩共同变形产生一定量径向位移。喷射混凝土支护的作用原理加固和防止风化喷射混凝土以较高速度射入张开的节理裂隙，产生如同石墙灰缝一样的粘结作用，从而提高岩体的粘结力和内摩擦角，即提高围岩强度。同时，喷射混凝土层封闭了围岩，能够防止因水和风化作用造成围岩的破坏和剥落。改善围岩应力状态巷道掘进以后即使喷射一层具有早期强度的混凝土，一方面可将围岩表面的凸凹不平处填平，消除因岩面不平而引起的应力集中现象，避免过大的集中应力造成围岩破坏；另一方面，可使巷道周边围岩由单向或双向受力状态转化为三向受力状态，可提高围岩强度。与围岩共同作用开巷后如能及时对暴露围岩喷射一层混凝土，使喷层与岩石的粘结力和抗剪强度足以抵抗围岩的局部破坏，防止个别围岩活石的滑移或坠落，那么岩块间的连锁咬合作用就能得以保持，这样不仅能保持围岩自身稳定，而且还可以与喷层构成共同承载的整体结构。喷射混凝土材料喷射混凝土要求凝结硬化快、早期强度高，故应优先选用硅酸盐水泥和普通硅酸盐水泥，水泥的强度等级不得低于32.5。为了保证混凝土强度和凝结速度，笨得使用受潮或过期结块的水泥。为了保证混凝土强度，防止混凝土硬化后的收缩和减少粉尘，喷射混凝土中的细骨料应采用坚硬干净的中砂或粗砂，细度模数宜大于2.5。为了减少回弹和防止管路堵塞，喷射混凝土的粗骨料粒径一般不大于15mm。喷射混凝土的强度一般要求不得低于15MPa；水灰比以0.40.5最佳。水灰比在此范围内喷射的混凝土强度高而回弹少。通过在普通喷射混凝土加入钢纤维的方法可有效的改善混凝土的整体力学性能和物理性质，特别是提高混凝土喷层的抗拉强度和变形能力、变脆性材料为塑性材料。国内外已经在隧道和地下工程中使用钢纤维混凝土，取代传统的普通混凝土加设钢筋网作为永久支护，这样可部分省掉钢筋网、节省大量的人力物力财力。喷射混凝土主要性能指标喷射混凝土抗压强度一般设计喷射混凝土强度要求达到1520MPa。喷射混凝土也和一般混凝土一样，强度随时间增长而增加，最终强度可以达到120%130%。粘结强度喷射混凝土的拌和料以高速冲击面层，不仅可以提高浆料密实度，而且还可以形成510mm的浆液层充满面层而接受后续的骨料。因此，无论喷层面是砖、混凝土或石料，均有较高的粘结强度。喷射混凝土厚度单独使用时，喷射混凝土厚度一般为50150mm；多次喷射时，喷层厚度也可以到250mm；与锚杆联合使用时，根据工程性质不同可以采用50120mm。一般考虑到防止围岩风化和工程特点，要求喷层厚度不小于80100mm。当岩体变形较大时，混凝土喷层将不能有效地进行支护。试验证明，当喷层厚度超过150mm时，不但支护能力不能提高，而且支护成本明显提高，因此应选用锚喷联合支护。这时支护以锚杆为主，喷混凝土只对锚杆间表面岩石进行局部支护和防止围岩风化。喷射混凝土施工工艺流程从混合料和施工工艺上可将喷射工艺分为干式喷射法和湿式喷射法两种。知道现在，干喷法在我国和其他国家仍是喷射混凝土的一种主要方法。用干式喷射工艺时，先将砂、石过筛，按配合比和水泥一同送入搅拌机，然后用矿车将拌合料运送到工作面，经上料机装入以压缩空气为动力的喷射机，同时加入规定量的速凝剂，再经输料管吹送至喷头处于水混合后喷敷到岩面上。用干式喷射机喷射混凝土时，装入喷射机的是干混合料，在喷头处加水后喷向岩石。喷射作业时粉尘大，水灰比不易控制，混合料与水的拌合时间短，因而混凝土的均质性和强度受到影响且回弹量大、喷层质量低。喷射混凝土质量检测主要检测内容包括喷射混凝土的强度和喷层厚度。喷射混凝土的强度检测采用点荷载试验法，也可采用喷大板切割法或凿方切割法，不得采用试块法。点荷载试验法是用混凝土钻取机从喷层中钻取圆柱体心样，然后用点载荷仪测试其点载荷强度，再根据点载荷强度确定喷层强度。拔出试验法是指在混凝土喷层中钻孔、切槽并安装扩拔器进行试验，根据极限拔出力确定喷射混凝土的抗压强度。最后用统计法或非统计法进行数据处理，以确定喷层中的抗压强度是否满足要求。喷层厚度，可在喷射混凝土凝固前采用侦探法检测，也可用打孔尺量法或取心法检测。图6 喷射混凝土V高应力区、节理围岩，采用树脂锚杆和胀管式锚杆支护，还可用钢丝绳砂浆锚杆(图1c)，不宜用机械锚固式锚杆。喷射混凝土用于永久性巷道支护，短期性巷道要挂金属网和钢带拉条。VI巷道处于高应力区，受采动影响.采用加密注浆钢筋锚杆。VII巷道处于承压区、围岩破碎、随时有冒顶的可能，采用管缝式锚杆，锚杆端间设拉条钢带。VIII断层破碎带的巷道，无法采用锚杆锚固，采用钢纤维喷射混凝土，在喷层中钻排泄孔，防止喷层内形成压力。在明显的高应力区和围岩持续移动区，则要求采用钢支架。在硬岩开挖中,钢支架使用有限,大部分支护可以使用机械锚固式锚杆和灌注式树脂锚杆、喷射混凝土和它们组合。掘进断层带、严重破碎岩带与断层结合带时，不能用机械锚固式锚杆和灌注式锚杆，需用型钢支架（图7），支撑巷道围岩冒落岩石的自重。图7 型钢支架此外，矿山地层条件的变化，使得单存锚杆支护或喷射混凝土支护技术不能满足巷道稳定和安全需要，从而在锚杆支护和喷射混凝土支护的基础上形成了一系列新的支护形式，如岩石支护中以锚杆、金属网、喷射混凝土、钢带等组成的联合支护形式，可简称为锚喷支护。随着技术发展，还逐步形成了预应力锚索支护技术和锚杆桁架支护技术等。目前这些技术已经在不同巷道中的支护取得了较好的技术经济效果。锚杆和喷射混凝土虽各有优点，但也都有不足之处。锚喷联合支护，恰能做到使二者相互取长补短、互为补充，是一种性能更好的支护形式。锚杆与其穿过的岩体形成承载加固拱，喷射混凝土层的作用则是在于封闭围岩、防止风化剥落，和围岩结合在一起能对锚杆间的表面岩石起支护作用。试验表明，用锚杆进行支护时，在两锚杆间的围岩表面附近会产生拉应力。如果岩石松软，在拉应力的作用下可能产生局部破坏和掉块，而局部小岩块的坠落又可能导致深部岩石的松动和破坏，这样将削弱岩石加固拱的稳定性和承载能力。因此，锚杆与喷射混凝土联合使用，就可以防止局部岩块松动和坠落，从而加固和提高岩石拱的承载能力。虽然喷射混凝土能有效控制锚杆间的石块掉落，但其本身是脆弱的，当岩石变形较大时易出现开裂剥落。解决方法之一就是在喷射混凝土之前敷设金属网，喷后形成钢筋混凝土层，提高了喷层的整体性，改善了喷层的抗拉性能，这就形成了锚喷网联合支护，能有效的支护松散破碎的软弱岩层。另外，为了克服喷网层的整体性差和刚度低的缺陷，可采用钢带或钢筋梯将全断面内的锚杆连接起来后在喷浆封闭，形成符合结构。为扩大锚杆支护使用范围，充分发挥锚杆支护的经济、快速、安全可靠等优点，在大断面、地质构造破坏段、顶板软弱且较厚、高地应力、综放巷道等困难、复杂的巷道中，为增加锚杆支护的安全可靠性，可使用小孔径预应力锚索进行加强支护。预应力锚索加强支护在岩巷支护中占有重要地位。由于它的锚固深度大，可将下部不稳定岩层锚固在上部稳定岩层中，可靠性较好且可施加预应力而实现主动支护，因而是支护技术中一种可靠有效的手段。对于节理和裂隙发育、断层破碎带等围岩松散、破碎的情况，近年来又开发了注浆锚杆技术，即利用锚杆注浆技术改变围岩松散破碎结构，提高其粘结力、内摩擦角和围岩的整体性，使围岩为锚杆提高可靠的着力基础，充分发挥锚杆对松散破碎 岩层的锚固作用。注浆锚杆既是锚杆又能用其进行注浆。围岩注浆后，一方面将松散破碎岩块胶结成整体，从而提高岩体的内聚力和内摩擦角，使岩体本事成为支护结构；另一方面，使普通端锚式锚杆变成全长锚固锚杆，使锚杆与围岩形成整体，充分发挥锚杆锚固作用，组成可靠的组合拱。利用浆液充填围岩裂缝，与锚喷网支护相结合，形成多层组合拱，可扩大支护结构的有效承载范围，提高支护结构的整体性和承载能力。另外，锚喷或锚喷支护加钢拱架或砌碹支护结构主要使用于破碎围岩松动压力和变形位移的情况以及其他复杂条件下。在这些条件下，锚喷或锚喷网支护作为一次支护，二次支护采用钢拱架或砌碹支护。3软岩巷道问题分析软岩巷道的矿压控制与巷道支护一直是长期困扰着地下采矿工业发展的难题之一。地压大、难支护的巷道已被列入软岩巷道研究之列,软岩已成为软弱、破碎、松散、膨胀、流变、强风化蚀变及高应力岩体的总称。我国许多矿区,目前都存在软岩巷道支护困难的问题,并成为影响矿区发展和矿井经济效益的主要因素。对软岩巷道的变形机理及支护技术进行深入研究有助于软岩支护理论的发展,改善主要凭经验支护的现状，使软岩支护技术更加科学。I软岩巷道变形机理及支护原则根据主要影响因素及工程岩体分级国家标准所依据的岩石坚硬程度和岩体完整程度两个岩体稳定性的基本共性因素，将软岩划分为5类,即软弱型软岩、破碎型软岩、高应力型软岩、软弱破碎型软岩、膨胀性软岩。由于组成岩体的岩块强度和结构面特性不同，不同类型软岩巷道围岩变形规律各不相同，实践中应根据不同类型的软岩确定巷道变形、失稳机理及其支护原则和方案。a.软弱型软岩巷道软弱型软岩的岩块强度低,岩石完整性较好，故变形以岩块变形为主,结构面的影响较小。这类软岩巷道破坏机理以塑性变形和流变变形为主，巷道变形的特点是变形持续时间长、变形速度快、变形量大,表现为明显的流变变形特征，破坏型式主要有持续性的挤压流动性底鼓、大变形量的顶板及两帮收敛变形。支护时，应采用高强度的全封闭高阻支护结构，并适当让压。常用方法有锚喷网加可缩性金属支架,锚注支护加喷砼和金属网法等。b.破碎型软岩巷道破碎型软岩的岩体完整性差,岩块强度较高，变形是由于岩块沿结构面滑移破坏。这类巷道变形以松动塌落变形和流变变形为主，破坏型式主要有顶板冒落、两帮片落鼓折、大变形量的顶板及两帮收敛变形。支护应能够加固岩体结构面，提高岩体整体强度,宜采用高强支护，并适当让压方式，如锚注支护加喷砼和金属网法等。c.高应力型软岩巷道高应力型软岩指岩体完整性差，处于高地应力和采动应力环境中的岩石。这类软岩巷道在岩块强度较高时，变形破坏以松动塌落为主，具体形式有冒顶、片帮；在岩块强度较低时，变形破坏以流变变形为主。高地应力地区软岩巷道的主要破坏形式有大变形和岩爆两种。当变形量很大且延续时间很长时，就产生了持续不断的破坏以致深入到围岩内部，使围岩塑性区逐渐增大造成硐室大规模坍塌,巷道围岩破坏具有明显的时效性。为了防止岩体破坏，必须适当卸压，且支护作用必须控制持续不断的变形和破坏，并维持巷道的稳定，可采用组合锚杆，金属网加喷砼方法等。d.软弱破碎型软岩巷道软弱破碎型软岩兼有软弱型与破碎型软岩变形的特点，岩块强度低，岩体完整性差。该类巷道变形机理十分复杂，表现为强烈的流变变形特性，且来压迅猛。巷道支护须采用高强度的全封闭支护结构来加固岩体，如锚注支护加喷砼和金属网，高强度喷网加高强度可缩性金属支架方法等。e.膨胀性软岩巷道膨胀性软岩多含有膨胀性颗粒和介质，如蒙脱石、伊利石等。膨胀性粘土矿物成分吸水后体积膨胀，产生很大的膨胀力或遇水发生物理化学反应，使围岩强度降低或丧失，从而致使巷道变形破坏。巷道变形以膨胀和流变变形为主，围岩变形量大，持续时间长，极易坍塌。巷道支护应封闭围岩表面，采用全封闭的高阻让压结构，如锚注支护加喷砼和金属网及高强锚喷网加高强可缩性金属支架法等。II软岩巷道支护技术a.锚杆支护技术锚杆支护把围岩视为主动支护体，在软岩巷道围岩发生较大位移变形前施行锚固，使围岩形成具有较大刚度的整体，充分利用围岩本身的强度和自承能力，变荷载为承载体，阻止和减少离层进一步发展。在巷道围岩上按一定网度布设锚杆形成锚杆群，其作用原理可分为以下三点：悬吊作用(如图8a)：在块裂结构岩体中,围岩往往会出现危险关键块体，可用锚杆群将其“悬吊”在稳固岩层上；组合作用(如图8b)：在层状岩体中，锚杆及连接螺栓，将薄岩层组成厚梁，增大其抗弯刚度和承压能力；挤压加固作用(如图8c)：在碎裂结构岩体中，以一定方式布置预应力锚杆群，各锚杆的预应力在锚杆两端产生一个锥形压缩区，并互相连接起来，两端锥形压缩区之间岩体受到挤压，形成一个挤压加固带，这个带内的岩体强度得以提高，整体性加强，好似一个承压拱，充分发挥围岩的自承载作用。锚杆使用过程中，必须选择合理的锚杆长度，使锚固端锚固在围岩松动圈外的稳固围岩上；确定合理的锚杆网度,达到最佳锚固效果；测定设计锚杆的锚固力，确保锚杆能够达到要求强度。同时，锚杆眼钻凿与锚杆安装也至关重要，直接影响到锚固质量。对于角砾状破碎不稳固的矿体，锚杆支护可以通过合理的网度或与金属网联合作用达到提高软岩巷道整体稳定性的目的。对于围岩松动圈半径不大的采场巷道，锚杆支护是一项整体效果良好的控顶措施；但如果围岩松动圈半径超过锚杆有效支护深度，锚杆支护也不能保证围岩的稳定性。图8 锚杆群的作用原理b.锚杆的联合支护技术为了进一步提高软岩巷道稳定性，锚杆被广泛应用于锚网喷技术及锚注技术等。锚网喷技术中，锚杆锚固在未破坏的岩体上，阻止围岩松动变形和破坏；喷射混凝土喷层封闭围岩表面,支护锚杆间围岩，防止表面岩层冒落；喷层中铺设钢筋网，可增加喷层的强度和柔性,提高支护的整体性。锚杆、混凝土喷层和钢筋网三者组成的支护体与围岩紧密结合，共同承载，既充分利用和发挥了围岩的自承能力，又在与围岩共同变形过程中及时提供支护抗力，限制围岩产生有害变形，从而保持巷道稳定。鹤壁十矿南翼大巷埋深大、地质条件复杂、服务年限长，对其应用此项技术，有效地控制了围岩变形，保证了巷道稳定，取得了明显的技术经济效益。锚注支护是兼有锚杆支护与注浆加固共同优点的一种支护方式。在巷道开挖以后，对巷道围岩进行喷浆封闭，防止围岩进一步风化；然后在围岩中打入注浆锚杆进行注浆加固。注浆锚杆既有锚杆支护的特点，又能通过此锚杆对围岩进行注浆。通过注浆、浆液充填、压密裂隙空间,使围岩由注浆前的无约束松散状态变为由锚杆、浆体和围岩共同作用的具有承受抗压、抗剪切、抗拉等适应复杂应力、应变状态的支护体。焦作矿区古汉山矿主要运输大巷布置在软岩层位上，巷道变形破坏十分严重。应用锚注支护技术提高了围岩本身的承载能力，达到治理围岩的目的。c.砌碹支护及喷砼技术砌碹支护是软岩支护的传统方法，利用支护体自身的支护强度来支撑来自围岩的初期矿山压力，待平衡后，支护体和围岩一起抵抗来自围岩层的压力。这种支护方法适合于巷道围岩非常破碎，矿压较大，采用锚喷支护优越性不显著；巷道围岩很不稳定，顶帮岩石极塌落，砼喷不上、粘不牢，锚杆的锚固力明显下降的含油泥岩、粘土岩及断层破碎带。砌碹支护属刚性支护，由于巷道本身成形不规则，当应力重新分布时，支护体是局部而不是全部接触岩层。先接触岩层的支护体必将先产生形变，在地压过大时从某一支护薄弱点开始遭受损坏，碹体很容易被压碎，崩落。因此，探索用“刚柔结合，先柔后刚”的方式克服各种压力的不良影响，在碹体与围岩间预留掘进断面一般取1.02.2 m2，两帮预留间隙30100 mm，拱顶预留间隙120300 mm，并用砼、砂或碎矸石等填实(图9)，形成碹体均匀受压的缓冲层，在保证巷道有一定的可缩性情况下，尽量提高巷道整体承载能力。图9 壁后喷砼的砌碹支护技术d.U型钢可缩性支架壁后充填层技术U型钢可缩性支架是广泛应用于煤矿岩巷的一种被动支护，其最大优点是当围岩作用于支架上的压力达到一定值时，支架便产生屈服缩动,缩动的结果使围岩作用于支架上的压力下降，从而避免了围岩的压力大于支架的承载力而导致支架的破坏，保证了巷道的正常使用。但是由于施工技术、岩性条件等的限制，任何刚开挖出来的巷道周边都是凹凸不平的，与光滑的U型钢支架出现点接触现象，引起支架的受力不均匀，造成支架在复杂力系作用下工作，出现应力集中导致支架局部屈服从而影响整个支架的性能。根据国内外的试验结果和使用经验表明：U型钢可缩性支架壁后充填技术可以使支架均匀受力，有效地发挥支架的性能;在壁后密实充填的情况下，U型钢支架的承载能力可比不进行壁后充填时提高2.53倍。因为，将一定厚度的胶结硬化材料进行壁后充填，可使支架与围岩紧密接触，保证支架能及时承载和均匀承载。当围岩来压后通过充填层的压缩变形产生让压作用，提高围岩的自承载能力，控制围岩的变形。实施壁后充填后，巷道围岩与支架相互作用体系从无壁后充填情况下的“支架围岩”作用体系变成了“支架充填层围岩”三位一体的作用体系(见图10)。支架与充填层组成了复合支护结构，将无壁后充填情况下支架单独对围岩作用变成了支架与充填层共同对围岩的作用。通过壁后充填层将支架与围岩联成一体，形成共同承载体，对于发挥围岩的自承载能力更为有利。U型钢支架壁后充填采用砼喷射机直接喷射充填，不但充填密实、效果好，而且施工工艺简单、技术要领容易掌握，大大降低了工人的劳动强度，加快掘进速度。该支护方法主要适用于强膨胀岩层及断层破碎带，淮南矿业集团望峰岗煤矿-817 m副井车场北绕道，围岩主要以泥岩和粉砂质泥岩为主，岩石破碎松散，推广以水泥、石膏和粉煤灰为原料的充填材料进行工业试验，效果良好。e.离壁支护技术所谓的离壁支护是指在料石碹或U型支架与巷道毛断面之间留有一定的间隙，且不进行壁后充填的一种打破常规的特殊支护形式。常规的料石碹或U型支架人工壁后充填，不仅工人的劳动强度大、工艺复杂、工作效率低；而且充填不匀称，造成充填物受力不均匀，在多数情况下是以“点”或“局部”形式传递压力，使碹体产生局部破坏；同时，围岩施加于支护体的压力具有一个峰值，峰值过后压力减小，然后稳定下来，稳定下来的载荷为围岩松动圈内岩石重量，而常规支架壁后充填支护在峰值到达之前完成，从而使得支架和充填物经受峰值压力的作用，容易破坏。巷道掘进以后，打破了岩体的原始平衡状态,产生次生应力场。为达到新的平衡，围岩就要产生变形、破坏和冒落，壁后间隙就起到了释放应力的让压作用。假如壁后间隙的尺寸预留得合理，最终能导致巷道自然冒落，形成一定的冒落拱体，在其形成的过程中，拱内的应力得到了卸载(冒落的岩石重量由支架来支承)，拱外的压力转移到拱壁，大大减弱了支架上的压力，最大限度地发挥了岩石自身支承能力。同时，自然冒落拱的形成过程完成了壁后自然填充，而且离壁支架基本上是在巷道压力达到高峰值后才承受压力，故可充分发挥围岩的自承能力。该技术构成一个“先柔后刚”的支护结构，具有支护效果好、施工安全、速度快、效率高、成本低和维护简单等优点。建议浅部开采矿山在无地质构造影响、无采动影响和无膨胀性的软岩中推广使用。半罗山煤矿永安矿区软岩巷道应用离壁支护技术，经过6年多时间的使用和观察，支护体未出现异常变化和损坏现象，支护效果良好。图10 “支架充填层围岩”三位一体作用体系f.二次耦合支护技术一般情况下，软岩巷道具有初期变形速度快、变形量大、蠕变延续时间长等特点，因而巷道开挖后表现为来压快、压力大、巷道一次支护成功率低。根据支架与围岩共同作用原理，软岩巷道的围岩控制技术可采用具有一定变形量的柔性支护，并进行二次支护。一次柔性支护让压，围岩体受力达到较低变形速率下的力学平衡，充分发挥围岩承载力；大刚度二次支护，减少巷道岩体偏应力，使巷道围岩切向应力相对降低，径向应力相对升高，促进围岩应力向稳定应力状态转化。柔性支护可采用锚杆支护、锚注支护及锚杆与金属网等联合支护方式；大刚度支护可采用料石碹或U型支架等支护方式。该方法适用于深部松软破碎，具有高应力、低强度、膨胀性、流变性的软岩巷道的支护。程村矿深部巷道围岩被软弱面(断层、裂隙、节理、层理)切割，导致围岩的各向异性与不连续,致使围岩松动破裂引起危岩滑落。支护方式采用锚网喷一次支护让压，料石碹大刚度高强度的二次支护是成功的。4结语岩巷支护是一项系统工程，包括设计、施工、监测三个大的方面，设计是支护的重要前提，施工是巷道支护的关键环节，监测是支护技术的重要组成部分。岩巷支护是一项特殊工程，只有在合适的时候采用合理的支护方法才是最好的支护方法。介绍了现有的岩巷支护设计方法，研究表明在现场中，往往是几种设计方法配合使用，互相检验，互相支撑。探讨了岩巷支护施工工艺与全面质量管理，以及支护质量监测等问题。此外，软岩巷道的支护技术是一项复杂的工作，尤其是矿井向深部转移后，在高应力作用下，严重影响矿井的经济效益和安全状况。因此，必须对软岩巷道变形机理进行研究，确定合理的支护技术方式。在具体的支护过程中，应具体问题具体分析，对于用单一的支护方法难以得到理想支护效果的情况，要采用多种技术方法联合使用，以实现安全、经济、合理、高效的支护。参考文献1 杜计平.煤矿深井开采的矿压显现及控制.徐州：中国矿业大学出版社，20002 左秀峰.高产高效矿井与深井的开拓部署及决策支持系统.徐州：中国矿业大学出版社，20003 徐永圻.煤矿开采学.徐州：中国矿业大学出版社，19914 尹传理，李化敏等.我国煤矿深部开采问题探讨.煤矿设计，1998（8）5 董小石.煤矿深部开采可能出现的问题及对策.煤炭技术，2003（7）6 涂敏，等.深部开采巷道矿压显现及控制.矿山压力与顶板管理，1994（2）7 付国彬，姜志方等.深井巷道矿山压力控制.徐州：中国矿业大学出版社，19968 钱鸣高，刘听成等.矿山压力及其控制.北京：煤炭工业出版社，19919 李化敏，胡劲松等.深井巷道矿压与支护问题的探讨.焦作矿业学院学报，1994（6）10 段绪华，等.浅谈深部开采工作面支护与矿压控制.煤矿开采，1999（6）11 王元仁.深井困难条件下的巷道支护技术.煤炭科学技术，2003（2）12 薛顺勋，聂光国等.软岩巷道支护技术指南.北京：煤炭工业出版社，200113 王继良，等.中国矿山支护技术大全.南京：江苏科学技术出版社.199414 郑永学.矿山岩体力学.北京：冶金工业出版社，199515 刘会文，高树堂等.井巷工程.徐州：中国矿业大学出版社，198916 黄福昌，等.兖州矿区煤巷锚网支护技术.北京：煤炭工业出版社，200017 王建军.锚索支护在井巷支护中的应用.矿山安全与环保，2003（6）18 李家鳌，王圣公等.煤巷锚索支护技术.煤炭科学技术，1997（12）19 李洪占,王洪代等.锚网支护技术在回采巷道中的应用.煤炭技术，2003（3）20 胡成忠,宋申华等.锚网支护在软岩巷道中的应用.矿山压力与顶板管理，2003（2）21 .希罗科夫等著,王秀容等译.锚杆支护手册.北京：煤炭工业出版社，199122 G.不霍依诺，矿山压力和冲击地压，煤炭工业出版社，1985。23 唐春安，王述红，傅宇方 岩石破裂过程数值试验J 科学出版社,2003。24 钱鸣高,刘听成.矿山压力及其控制M.北京:煤炭工业出版社,1996。任务书学院 矿业工程学院专业年级 采矿工程 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： 陈四楼矿 1.5 Mt/a 新井设计毕业设计专题题目：浅析煤矿岩石巷道支护毕业设计主要内容和要求：以实习陈四楼煤矿条件为基础，完成陈四楼煤矿 1.5 Mt/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。结合煤矿生产前沿及工作单位情况，撰写一篇关于煤矿岩石巷道支护专 题论文。完成与采矿有关的科技论文翻译一篇，题目为“A method for the design oflongwall gateroad roof support”。院长签字：指导教师签字：翻译部分英文原文A method for the design of longwall gateroad roof supportW.Lawrence Geowork Engineering,Emerald,QLD,AustraliaAbstract：A longwall gateroad roof support design method for roadway development and panel extraction is demonstrated. It is a hybrid numerical and empirical method called gateroad roof support model（GRSM）, where specification of roof support comes from charts or equations. GRSM defines suggested roof support densities by linking a rock-mass classification with an index of mining-induced stress, using a large empirical database of Bowen Basin mining experience. Inherent in the development of GRSM is a rock-mass classification scheme applicable to coal measure strata. Coal mine roof rating（CMRR）is an established and robust coal industry standard, while the geological strength index（GSI）may also be used to determine rock-mass geomechanical properties.An elastic three-dimensional numerical model was established to calculate an index of mining induced stress, for both roadway development and longwall retreat. Equations to calculate stress index derived from the numerical modelling have been developed. An industry standard method of quantifying roof support is adopted as a base template（GRSUP）.The statistical analyses indicated that an improved quantification of installed support can be gained by simple modifications to the standard formulation of GRSUP. The position of the mathematically determined stable/failed boundary in the design charts can be changed depending on design criteria and specified risk.Keywords: Coal mine；Roof control；Support Design1. IntroductionLongwall gateroad strata stability is essential to ensure uninterrupted production. In Central Queenslands Bowen Basin, immediate gateroad roof lithology varies from coal to weak interlaminated material, to strong almost massive sandstone, with localised areas of weak fault affected strata. It is usual for roof conditions within any one mine to vary significantly. Typically, longwall mines in the Bowen Basin have specified gateroad roof support based on past practice. Modifications to gateroad support are generally reactive, due to encountered difficult strata conditions, and less proactive. Current gateroad support design approaches have limitations, which have restricted their applicability and adoption as mine site design tools.A prototype for an improved gateroad support design methodology has been developed that is integrated and systematic, based on rock engineering principals, but requires engineering judgement and experience 1. There were several broad objectives for the design methodology. A consistent and unambiguous definition of strata conditions and behaviour was required. Gateroad roof support needed to be assessed and specified. The method had to provide design calculations and justification for compliance and statutory purposes, and could serve as a frame work for a mine strata management system. Mine site support designers must be able to readily use the method to manage uncertainty and risk. The method must be able to be reviewed, modified and expanded.2. Current roof support design methods for longwall gateroadsNumerous roof support design methods have been proposed over the years, but none have gained widespread acceptance by the coal mining industry 2. There are empirical databases, some proprietary, based on industry practice, which specify gateroad primary and secondary support densities, using a statistical approach 3,4. Analytical methods are not appropriate when rock-mass yield due to high mining induced stresses occurs, but may be applicable and adapted to low stress environments 5. The application of complex post-yield numerical modelling in the design process for excavation support is valid although contentious, and requires a more comprehensive justification and better industry understanding of its strength and limitations 6. The complete mathematical representation of rock-mass properties and behaviour is a complex issue, which is still outside the capability of current numerical modelling code 6.Engineers and mathematicians do not have the current capability to fully define rock-mass geomechanical properties and their mathematical representation. Elasticplastic numerical modelling is a useful tool if used appropriately. It is not exclusively correct or unique, or always superior to other available and accepted design techniques. These aspects have been recognised during recent collaborative Australian Coal Association Research Program research on longwall microseismics 7, where it was considered that current 3D numerical models lack sufficient validated constitutive relationships, and are forced to make compromises when dealing with complex rock-mass behaviour.Simplified elastic numerical methods 8,9 have merit and are certainly applicable for more massive sedimentary rock-masses 5. An assessment of their applicability to weaker, laminated clastic rock-masses is required. Hybrid numerical and empirical methods have been developed for the geotechnical design of undercut and production level drifts of block caving mines 10.3. Geotechnical roof classification of longwall gateroadsTwo classification schemes were considered appropriate. Firstly, the coal mine roof rating （CMRR） 11, which is an established coal industry standard. Secondly, the Geological Strength Index, GSI with strength parameters included 12. A recent publication 13 has contended that GSI estimates of rock-mass strength should not be used for coal mine roof problems, where the geometrical scale of the problem is similar to discontinuity spacing. A distinction needs to be made between the GSI classification and the related HoekBrown failure criterion. This scale effect and situations where the failure criterion should not be used have been discussed 14 However, this does not mean that a classification of the rock-mass cannot be made. Indeed, this scale issue is a problem inherent in any rock- mass classification scheme, not just GSI, and for any failure criterion. For example, some mines appropriately use unconfined compressive strength （UCS） as an index or failure criterion, but UCS is also scale dependent and has the same limitations.Within the support design methodology, the rock-mass classification schemes will link mining-induced stresses （or stress index） and required installed roof support. Therefore, the classifications should be independent of environmental and geometrical factors, such as mining induced stresses and excavation orientation and size. A rock-mass classification scheme must also provide rock-mass geomechanical properties to enable the calculation of mining induced stresses.It is anticipated that CMRR will be the principal classification scheme used. However, the single rock-mass classification scheme that is best suited is the GSI derived global rock-mass strength. For numerical or analytical models, HoekBrown failure criterion parameters, modulus of deformation and rock-mass strength can be estimated from GSI 15,16.Direct utilisation of either CMRR or GSI is included within the design methodology.4. An index of mining induced stressAn index of mining induced stress in the gateroad roof at a location of interest is required. The three-dimensional （3D） stress distribution about a longwall panel including goaf reconsolidation, and the continuous stress redistribution that occurs during panel retreat, is a complex and difficult phenomena to quantify. One approach would be to construct a full elasticplastic, 3D numerical model. This approach would have limitations to a verified, unique and readily achieved calculation of stress, for several reasons. Generalised model roadway and goaf geometry may not always match the actual geometry. Generalised model roof lithology may not always match the actual lithology and variations. The roof/seam/floor interaction is a complex system and is difficult to model accurately. Rock-mass geomechanical properties, in particular post-yield cannot be fully defined. The geomechanical properties of the goaf, extent and behaviour of strata fracturing and caving, and goaf stress reconsolidation are largely unknown. The model may take many days to complete just a single scenario.While calculated mining induced stress from a detailed elasticplastic, 3D numerical model may be an appropriate parameter, there is little justification to improved accuracy compared to other methods. An alternative approach is to calculate mining induced stress from elastic 3D numerical models. Calculated mining induced stress in the immediate gateroad roof just outbye of the face-line may not be accurate if rock-mass yield occurs, but as an index of stress, it may be appropriate. An important criterion of its suitability would be how reasonable its relative variation is with changes in input parameters. A significant advantage is that it could be readily calculated for variable scenarios and would be within the range of capability of more geotechnical engineers.Maximum elastic tangential stress in the roof of a modelled gateroad could be considered a better indicator of rock-mass failure than the residual post-yield stress. Undoubtedly, significant rock-mass failure and subsequent stress redistribution do occur, which are not reflected in an elastic model. In the immediate roof of the gateroad, these failures are initiated at a critical mining induced stress. The stress index is a reasonable and appropriate measure of this critical stress, even if it may not agree in absolute magnitude after stress redistribution occurs. For mining induced stresses from an elastic 3D numerical model to be a reasonable representation, several issues influencing the stress distribution must be considered, which include strata fracturing and caving and goaf reconsolidation.For bulking-controlled caving, empirical relationships are used to predict the height of caving （goaf） and fracturing 17: （m） （1） （m） （2）where Hc is the caving（goaf） height above top fextracted horizon, Hf is the thickness of the fractured zone above top of caving zone, h is extraction thickness,and c1, c2, c3, c4, c5 and c6 are coefficients depending on lithology （Table1）.Table 1 Coefficients for average height of caving zone 17LithologyCompressive strength/MPaCoefficients /mC1C2C3C4C5C6Strong and hard402.1162.51.22.08.9Medium strong20-404.7192.21.63.65.6Soft and weak206.2321.53.15.04.0Weathered-7.0631.25.08.03.0Goaf stressstrain behaviour can be been defined 18 （Eq. （3）, based on earlier work 19, as follows: （MPa） （3） where, and are the vertical goaf strain and stress,respectively, E0 is the initial tangent modulus,and m is the maximum possible strain of the bulked goaf material.The initial bulking factor, BF, defines m as follows: （4） The initial tangent modulus, E0, can be defined as a function of the compressive strength of rock pieces, c, and the bulking factor, BF18,20: （MPa） （5）The FLAC3D double-yield constitutive model is used to simulate a strain-stiffening material with irreversible compaction,i.e.volumetric yield,in addition to shear and tensile failure.Upper-bound tangential bulk and shear moduli are specified21, with the incremental tangent and shear moduli evolving as plastic volumetric strain takes place.In addition to the shear and tensile strength criteria,a volumetric yield surface or cap has to be defined. The cap surface,defined by the cap pressure, pc, is related to the plastic volume strain, pv. The cap pressure, pc, is not the goaf vertical stress, v. The relationship between cap pressure and plastic volume strain is derived from an iterative FLAC2D compression test model,using a one element,1m1m,grid. Loading was simulated by applying a velocity to the top of the element, which has confined sides and base.The constitutive equation was derived from the iterative results by a Microsoft Excel Solver regression analysis,assuming a linear function.Goaf deformation and material strength parameters are defined as follows（Table2）. Table 2 FLAC3D goaf reconsolidation parameters1Upper bound tangent modulus230 MPa2Poissons ratio0.303Density1.7 gm/cc4Cohesion0.001 MPa5Friction angle256Dilation27Tensile strength0 MPa8Table 3 FLAC3D numerical model geometrical, geomechanical and geotechnical para meters1ParameterRangeUnitage2Roadway height2-3.4m3Roadway width4.8-6.5m4Longwall panel width200-300m5Pillar width15-45m6Depth60-330m7Immediate roof USC8-62MPa8Ratio of in situ horizontal to vertical stressRange from 1.2 to 2-9Rock-mass stiffness-10Rock-mass poissonratio0.25 for stone ,0.3 for coal-There are many theories on goaf reconsolidation, based on sound principles. Results from the various formulations do vary significantly. Which, if any, are correct is unknown, as goaf stresses have not been measured 18,22. For no other reasons than it is well described, and includes more of the parameters perceived to be important, the goaf stressstrain behaviour as defined is utilised in the calculation of a stress index 18. The elastic FLAC3D numerical model simulates a single two-heading longwall. Roof and floor strata are composite, uniform continuum. Strong contact is assumed between the coal seam and roof and floor. No discontinuities were modelled. Pillars will always be stable, which means that the actual pillar design must be appropriate and pillars adequately sized for the strata conditions .A range of geometrical,geological and geotechnical parameters must be specified,with the database distribution of some parameters listed in Table3. Some rock-mass geomechanical properties may be derived from thegeological strength index 15. Model 3D geometry may be visualised in Figs.1 and Figs.2 In these figures, scale may be judged from the seam thickness（3m）and thickness of immediate roof and floor. Axes of geometric symmetry are used, e.g. only half of the total goaf width is shown. Fig. 1. Typical 3D model geometryentire modelFig. 2. Typical 3D model geometryhorizontal section taken from the top of seamStress measurements are typically taken in discrete, more competent, and stiffer strata. Defining in situ horizontal stresses in all strata units of different stiffness is a difficult issue. There are problems associated with stress measurements in less competent strata and coal. The approach taken was simply to define in situ horizontal stresses in terms of rock-mass competency or classification. This does assume a correlation between the rock-mass parameter and elastic modulus. Quite rightly there are limitations with this approach. Model output （stress index） is simply the elastic mining-induced major principal stress in the immediate maingate roof just outbye of the longwall face-line . Example of FLAC 3D model output shows a longwall retreat situation, where the plane shown is horizontal, in the immediate roof. Note that the orientation of the major principal stress may be near-horizontal （immediate roof of gateroad） or near-vertical （pillar or face）. In this particular case, the stress index used in GRSM is 31 MPa. In the context of this design methodology and the development of a stress index, it is not critical that mining induced stress magnitudes agree. It is important is that relative changes in magnitude are reasonable and occur appropriately as parameters change. The effect of the intermediate （or minor） in situ horizontal principal stress must also be considered when assessing this model. This stress component will superimpose its own mining induced stress. 5. Characterisation of installed roof support A standard measure of the intensity of installed support, widely used within the Industry is GRSUP （ground support rating）, given by 4 （kN/m） （6） where Lb is the thickness of the bolted horizon defined by roof- bolts （m）, Nb is the average number of roof-bolts in each bolt row, Cb is the ultimate tensile strength of roof-bolts （kN）, Sb is the spacing between roof-bolt rows （m）,Nt is the average number of cables in each cable row, Ct is the ultimate tensile strength of cables （kN）, St is the spacing between cable rows（m）, w is the roadway width （m）, and 14.6 is a constant that is needed to convert from the original NIOSH equation, which was in Imperial units, to SI units; this will allow for compatibility with all USA data using the standard NIOSH equation.6. Database Data points have been collated from underground mines throughout the Bowen Basin coalfields. Nonstable, or failed data points are not restricted to situations where roof falls have occurred. Any situation where installed support was not sufficient was classified as a nonstable data point. Such situations include roof falls, where supplementary support was required for strata stabilisation, and where the area was mapped or observed to have experienced excessive deformation and deterioration. There are currently 280 defined data points, of which 106 are non-stable, and cover a range of strata conditions. Depth of cover ranges from 60 to 330 m. Roof conditions vary from weak interbedded to strong sandstone, also thick coal. There is high magnitude in situ horizontal stress in stone roof and relatively low magnitude stress in coal roof. Bolting densities range from a minimal four-bolt primary support in strong roof conditions, to intensive bolt plus cable support in weak and fault affected roof conditions. Longwall gateroad development contributes 166 data points （60 failed, 106 stable）, mains development contributes 24 data points （nine failed, 15 stable）. Bord and pillar first workings contribute seven data points （three failed, four stable）. Longwall retreat in the maingate belt-road contributes 83 data points （34 failed, 49 stable）.7.Design methodology7.1 IntroductionThe design methodology is tailored for roadway development and longwall gateroads. Situations considered are mains and gateroad development and longwall retreat in the maingate belt-road. Roof support for bord and pillar first workings can also be assessed. Evaluating roof support using GRSM incorporates several design steps. An initial roof characterisation or classification is required, followed by a calculation of a stress index. Suggested minimum GRSUP is then determined. Finally, primary and secondary roof support patterns are proposed, also considering the influence of factors not assessed by GRSM.7.2. Rock-mass characterisationA classification is required for the immediate 2m of roof, and if a longwall retreat assessment is required, the 4m section above that. Typically, it would be expected that most practitioners would calculate CMRR. Alternatively, the GSI global rock-mass strength may be calculated. It is important when calculating CMRR or GSI global rock-mass strength not to overestimate, particularly when assessing bore-core. Consider the in situ, asmined condition of the rock-mass. When dealing with bore-core it is easy to overlook the detrimental effect of joints that may not appear in the core, and weak bedding laminations that may remain intact in the core. Similarly, the intact rock strength should not be overestimated, particularly when using a geophysical correlation.7.3 Stress indexTo effectively use GRSM it is important to be able to quickly and accurately calculate a stress index, without having to resort to a FLAC 3D numerical model. Equations to calculate stress index have been developed for two situations; roadway development and longwall retreat. A series of Microsoft Excel Solver analyses were conducted to define equations that could replicate this elastic numerical modelling calculation of stress index. It is recognised that there may be situations where the calculated stress index could be varied. At this stage in the development of GRSM no Table 4 Stress index equation for roadway developmentinput parameters and constantsParametersConstants-CMRRGSIa-7.66-7.43X1Immediate 2m roof （e（CMRR/40） or GSI global rock-mass strength）b10.0330.0086X2Roadway or excavation height （m）b20.2270.227X3Roadway or excavation width （m）b30.00130.0041X4Depth of cover （m）b40.006770.00681X5Solid or rib-to-rib pillar width （m）b5-0.0013-0.0013X6Ratio of in situ horizontal stress to vertical stress for immediate 2m roofb60.7670.772X71+sin（）:where is the angle between the roadway orientation and the in situ major principal horizontal stress.only use a positive number between 0and 180b70.2800.282Table 5 Stress index equation for longwall retreatinput parameters and constantsParametersConstantsParametersConstants-CMRRGSIc-22.40-22.10y1mmediate 2m roof （eCMRR/40 or GSI global rock-mass strength）d10.7970.168y2Upper 4m roof （eCMRR/40 or GSI global rock-mass strength）d2-1.046-0.150y3Roadway or excavation height （m）d30.8170.749y4Roadway or excavation width （m）d4-0.406-0.215y5Depth of cover （m）d50.01080.0101y6Solid or rib-to-rib pillar width （m）d6-0.0129-0.0098y7Longwall panel width , rib-to-rib （m）d7-0.000210.00057y8Ratio of in situ horizontal stress to vertical stress for immediate 2m roofd80.6860.690y9Ratio of in situ horizontal stress to vertical stress for immediate 4m roofd91.5530.953y101+sin（-）:where is the gateroad oritentation looking inbye and the in situ major principal horizontal stress.Clockwise is positive. The is taken as 20,for the angle - ,only use a positive number between 0and 180.d100.6740.637guidance can be offered about any adjustments. Intersections, both for roadway development and longwall retreat, have different mining-induced stress compared to roadways. Longwall start-up, before regular caving occurs, and major weighting events along the longwall face may have higher abutment stress. As a longwall approaches intersections, there may be an increase in mining-induced stress.There are 197 roadway development data points and 78 longwall retreat datapoints. The proposed roadway development stress index calculation is given by Eq.（7）, with the input parameters and constants defined in Table 4 , The proposed longwall retreat stress index calculation is given by Eq. （8）, with the input parameters and constants defined in Table 5. Both Eqs. （7） and （8） have a correlation coefficient （R2）of 0.99. （MPa） （7） （MPa） （8）where, SIDEV is the stress index for roadway development, SILR is the stress index for longwall retreat, x1, x2, y1, y2, are independent input parameters, and a,b1,b2,c,d1,d2, are con- stants.Note that both Eqs.（7）and（8）suggest that the some parameters have little influence on the magnitude of the stress indices and need not be considered, e.g. pillar and panel width. This is not surprising, as the stress index is calculated in the maingate roadway not the adjacent travel road, and does not consider any variation in upper strata caving and weighting for any one mining scenario. Roadway width had little influence for the roadway development equation, but this is not the case for longwall retreat. An interesting outcome of the equation optimisation is the parameter, 1+sin（-）. Essentially it says that the minimum mining induced stress does not occur when the gateroad orientation is parallel with the in situ major principal horizontal stress. The value of 20was derived by trial and error, minimising the sum-of-squares. The result is in general agreement with published data derived from both numerical modelling and stress measurements 23,24.7.4 Design chartsThe stablefailure boundary is determined mathematically using logistic regression, which is a statistical regression model for binary dependent variables. Logistic regression model para- meters are determined using a maximum likelihood function called logitfit 25 contained within the mathematical software package MATLAB. Orientation of the stablefailure boundary is determined by the logistic regression analysis. Fixing the position of the boundary is done by using cumulative distribution functions to minimise the number of misclassifications.General design charts and equations utilising all data points, and probability of a stable outcome may be constructed. The logistic regression analyses considered three independent vari- ables; roof classification, stress index, and GRSUP. The initial stablefailure boundary was determined by cumulative distribution functions where there was the same proportion of mis- matched data points either side of the boundary. However, the position of the stable/failed boundary can be changed depending on design criteria and specified risk.As the desired probability of stable outcome is increased the number of incorrectly assigned failed data points decreases.All failed data points report correctly to the failed region with a probability of stable outcome of 0.9. Therefore, given a design within the parametric bounds of the database there appears to be no benefit using a probability of stable outcome greater than 0.9. A reasonable conservative design could be obtained using a probability of stable outcome of 0.85 or 0.8, for a CMRR or GSI global strength classification, respectively. This level corresponds to all roof fall data points correctly assigned. Estimates of GRSUP can be calculated as follows: （kN/m） （9）（kN/m） （10）Where GGRS = GSI global rock-mass strength （MPa）, and SI = stress index. Alternatively, the design charts of Figs.4 and 5 may be used to evaluate required GRSUP.The logistic regression analyses indicate that an improved quantification of installed support can be gained by simple modifications to the standard formulation of GRSUP（Eq.（11）. Use effective installed primary support, not as-installed, but there is some subjectivity to specifying this. Use averaged bolt spacing. This is calculated as the mean of the bolt row spacing, and the average distance between bolts within a row.（kN/m） （11）Where NBe is the effective installed primary support, and SBav is the average bolt spacing, both within and between rows. Empirical charts may be unreliable outside the parameter ranges within the database. Extrapolation outside of the database parameter ranges should not be done. Investigation of the database composition has indicated that the size of the database appears to be sufficient. At low levels of stress index there are more data points, therefore the stablefailure boundaries can be expected to be more reliably positioned. However, there are fewer data points at higher values of stress index, which means that the upper portion of the boundaries may not be well positioned, and this is confirmed by the logistic regression analyses.There are several deficiencies in the database. More longwall retreat data points at higher stress index are required. While there are a number of data points for depth of cover greater than 250 m, more are required. The database is particularly deficient of data points for depth of cover greater than 300m. There is limited experience of mining in the Bowen Basin at depths greater than 300m. The database requires more longwall retreat data for relatively weak roof （CMRR402.1162.51.22.08.9中等坚硬20-404.7192.21.63.65.6软弱206.2321.53.15.04.0风化的-7.0631.25.08.03.0通过下式可以确定采空区的应力应变：（MPa） （3）式中 和 分别是采空区的垂直应力和垂直应变，E0是初始正切模量，m是采空区岩层可能的最大变形量，通过碎胀系数计算变形量，计算公式如下： （4）通过岩石抗压强度c和碎胀系数BF计算初始正切模量E0，计算公式如下： （MPa） （5） 应用双倍屈服三维本构模型通过单轴压缩模拟应变钢化材料的剪切和拉伸破坏。作为弹性应变，正切模量的最大值是通过不断加载来确定的。为了得到抗剪强度和抗拉强度必须确定变形体积。表面体积通过表面应力pc计算，它与采空区垂直应力v不同与弹性应变pv有关。把1m1m的试件作为测试模型，通过对其反复压缩得到两者之间的关系。加载过程中对有固定侧和底座的试件进行快速填充。制定的公式是以采空区变形和岩石强度为基础，利用EXCEL软件反复进行衰减实验分析得到的。具体参数见表2.表 2 采空区变形三维数值模拟参数1最大正切模量230 MPa2泊松比0.303密度1.7 gm/cc4内聚力0.001 MPa5内摩擦角256扩容27抗拉强度0 MPa8弹性应变表 3 三维数值模拟力学参数1参数RangeUnitage2巷道高度2-3.4m3巷道宽度4.8-6.5m4工作面200-300m5煤柱宽度15-45m6埋深60-330m7直接顶等级8-62MPa8水平应力与垂直应力之比1.22-9岩体刚度-10泊松比岩石0.25,煤0.3-基于声发射原理有多种采空区变形的理论，根据不同的理论公式得到多种结果。由于采空区应力没有确定，任何一种结果的正确性都不得而知。采空区应力是最重要的参数，利用采空区应力应变计算应力除此之外，对力学参数的理解也很重要。三维弹性数值模型简化模拟了长臂工作面，顶底板岩层是复杂的连续统一体，通过模型模拟了煤层与顶底板之间的联系。 图1 典型三维几何模型-整体模型 图2 典型三维几何模型-沿煤层顶部的截面对煤层和顶底板进行连续模拟，煤柱设计必须准确而且尺寸应适应地层条件，煤柱才能稳定。一系列的工程地质几何参数必须通过资料库精确确定，参数设置见表3。岩石力学参数应取自于地应力参数。三维几何模型如图1和图2，图中煤层厚度为3m顶底板厚度均为2m，沿采空区中心线划分。应力数值来自更加坚硬的岩层。不同岩性岩层的水平应力较难确定，软弱岩层与煤层中应力数值更难确定。可以根据岩性与岩石等级确定一个应力范围。假设岩石力学参数与弹性模量之间存在一种关系，这种方法存在局限性。应力参数就是采煤工作面直接顶