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沿空掘巷小煤柱合理宽度留设研究1 前言窄煤柱沿空掘巷是提高煤炭采出率的有效方法之一，本文通过数值计算方法，研究了沿空掘巷不同煤柱宽度和巷道支护强度时煤柱的应力场和位移场，提出沿空掘巷小煤柱的合理宽度的留设方法。研究结果表明: 巷道掘进期间，煤柱较窄时，煤柱内中心位置承受的最大垂直应力随煤柱宽度增加变化较大，当煤柱宽度达到5 m 后，增大煤柱宽度，最大垂直应力变化已不明显，平均应力有所降低; 采动影响阶段，煤柱内中心位置承受的最大垂直应力随煤柱宽度增大而提高; 当煤柱宽度在37 m 时，高支护强度对控制围岩变形效果更好。建议沿空掘巷煤柱的合理宽度留设标准: 软煤57 m，硬煤35 m。将本成果应用于现场实践，取得了良好的技术经济效果，具有很好的应用推广价值。随着煤矿开采机械化程度的提高，矿井产量和开采深度的不断加大，对煤炭采出率和回采巷道支护技术要求越来越高，传统的留设较宽的区段煤柱护巷的支护方式已不能满足要求。留窄煤柱沿空掘巷是提高煤炭采出率的有效方法之一。沿空掘巷技术由于巷道具有煤炭采出率高、容易维护等诸多优点，近些年来受到了广大学者和工程师的极大关注，在我国多个矿区逐渐推广应用。窄煤柱是沿空掘巷围岩的重要组成部分，其稳定性直接影响巷道整体的稳定性，所以窄煤柱宽度的确定成为沿空掘巷围岩稳定性的核心内容之一。巷道围岩的稳定性主要取决于围岩强度、应力状况及支护与围岩的相互作用关系。沿空掘巷围岩松软破碎、强度低，受采动影响时，上覆基本顶岩层三角块结构旋转下沉，塑性区、破碎区迅速扩展，导致巷道变形剧烈。国投新集刘庄煤矿主采11 煤层121102 工作面煤巷埋深超过630 m，属深井半煤岩巷，围岩为煤泥岩互层，为典型的三软巷道。由于煤层倾角的影响，掘进时需破底，整个巷道围岩为非均质层状赋存，在高地应力作用下表现为强烈的两帮移近和底臌。121102 工作面风巷按沿空掘巷设计，因此煤柱的宽度和巷道支护强度将直接影响巷道掘巷和回采期间的安全使用。本文通过数值计算，分析了不同煤柱宽度和支护强度条件下，煤柱应力场和位移场分布特征，为合理选取小煤柱宽度提供理论依据。2 沿空掘巷煤柱稳定性的数值模拟本文利用ANSYS 软件中死活单元技术、材料非线性问题的分析方法，模拟了沿空掘巷窄煤柱变形破坏机理。巷道埋深按600 m 计算，采用平面应变模型，模拟巷道的形状为矩形，宽4.5 m，高3.5 m。模型计算的范围为: 顶板28.5 m，底板22.0 m; 整个模型的宽度为72 m。边界条件: 地下巷道简化为平面应变问题，边界上作用有铅直原岩应力和水平原岩应力; 模型下边界位移全约束，岩体取二维平面应变元; 屈服准则选用Drucker-Prager准则。该准则考虑了围压对屈服特性的影响，并且能反映剪切引起膨胀扩容的性质，如图1所示。图1 计算模型及边界条件考虑到煤矿开采的实际，回采巷道围岩的强度一般比较低，所以模型中的弱结构体取强度较低的岩层，其他较坚硬岩层则取强度更大的岩层; 采空区、巷帮薄层弱结构采取改变帮部材质的方法实现，薄层弱结构体强度较两帮煤层还要低，模型中巷道围岩的部分力学参数见表1。2 1 煤柱应力分布考虑到煤层力学性质对煤柱应力分布产生的影响，在计算煤柱应力分布时把煤层力学性质按软硬分别分析。煤层不同性质参数取值如表2 所示。2.1.1 掘进期间煤柱应力分布特征巷道开掘后，底板附近原先处于三轴压缩状态的岩体中径向压力消失后碎胀导致体积增大，其应力状态的变化引起围岩裂隙不断扩展和其力学性质不断劣化，围岩的整体刚度和弹性模量会逐渐降低，岩体的破坏随着时间向深部扩展大; 当支护阻力不足时致使巷道围岩以较高的速度长时间地向巷内运移。由于掘进期间硬煤和软煤不同煤柱宽度时围岩垂直应力分布情况较为一致，这里只给出硬煤时的垂直应力分布情况如图2 所示，现分别提取硬煤和软煤时不同煤柱宽度下的垂直应力云图( 如图3、图4 所示) ，从图中可以看出掘巷期间煤柱的垂直应力基本沿中间轴线呈对称分布，中间位置垂直应力值较大，两侧临空面应力值较小。图2 掘进期间硬煤不同煤柱宽度时围岩垂直应力分布情况图3 掘进期间硬煤不同煤柱宽度时垂直应力分布情况图4 掘进期间软煤不同煤柱宽度时垂直应力分布情况为了进一步分析煤柱中应力分布情况，现取煤柱高度一半的中部层位研究煤柱内应力场分布情况。掘巷期间沿煤柱宽度方向的垂直应力分布如图5所示。图5 掘进期间煤柱垂直应力分布由图5 可见，掘进阶段沿空掘巷窄煤柱应力分布具有以下特征:( 1) 煤柱宽度对应力分布影响较大。软煤: 煤柱较小时其应力较小并且应力均匀，2m 的煤柱应力4.6 MPa; 随煤柱宽度增大，煤柱内最大应力增大，最大应力9.5 MPa; 煤柱宽度达到5 m 后，最大应力增加不明显，位置也相差不大，应力分布近似呈三角形。硬煤: 煤柱应力分布与软煤差别较大;煤柱宽度对最大应力影响不大，2m 煤柱最大应力为8 MPa，而5 m、7 m 煤柱的最大应力分别为13.9 MPa、15.9 MPa; 煤柱较小时( 25 m) 应力分布近似呈三角形。煤柱较大时( 715 m) 近似呈梯形分布。硬煤的煤柱应力均大于相应软煤煤柱的应力。( 2) 煤柱宽度对煤柱浅部应力的影响: 软煤煤柱2 m 时浅部应力较小，煤柱315 m 时浅部应力相差不大; 硬煤煤柱25 m 时浅部应力较小，5m以上煤柱浅部应力较大。硬煤煤柱浅部应力明显大于软煤煤柱浅部应力。2.1.2 回采期间煤柱应力分布限于文章篇幅，下面不再给出不同煤柱宽度时的垂直应力云图，仅取煤柱高度一半位置，煤柱内的垂直应力场分布情况做分析。回采期间沿煤柱宽度方向的垂直应力分布见图6。图6 回采期间煤柱垂直分布由图6 可见，回采阶段大采高工作面沿空掘巷窄煤柱的应力分布具有以下特征:(1)在采动影响阶段，煤柱最大垂直应力随煤柱宽度增大而提高。(2) 软煤和硬煤3 m 以下煤柱的应力均趋于均化，与掘进阶段相比，煤柱中部应力降低; 5 m 以上硬煤柱应力增加均较大，15 m 时最大应力达21.4 MPa; 而软煤应力增加不大，最大应力为10.2 MPa。(3) 与掘巷阶段相比较，煤柱较小时，最大垂直应力降低，其位置向巷道侧移动; 煤柱较宽时，最大垂直应力升高，其位置向采空区移动。煤柱应力分布基本都呈三角形。2. 2 煤柱变形机理2.2.1 掘进时期煤柱内位移场分布特征掘进期间煤柱水平位移分布如图7 所示。由图7 可见，掘巷期间沿空掘巷煤柱深部位移具有以下特征:图7 掘进期间煤柱水平位移分布(1) 煤柱向巷道内的位移随煤柱宽度增大而减小，达到一定宽度后再由小变大，然后趋于稳定。(2) 煤柱表面向巷道内的位移特征，软煤: 煤柱较小时，煤柱整体向巷道内移动，2 m 煤柱向巷道内位移量较大，715 m 煤柱向巷道内的位移量差别不大，但显著小于2 m 煤柱向巷道内的位移量。硬煤: 随着煤柱宽度增加而减小，但相差不大。(3) 软煤煤柱达到5 m，中部位移稳定并较小;硬煤3 m 煤柱中部位移稳定，随煤柱宽度增大位移减小。2.2.2 回采时期煤柱内位移场分布特征由图8 可见，回采期间煤柱内位移分布具有以下特征。图8 回采期间煤柱水平位移分布(1) 软煤: 煤柱较小时，煤柱基本呈整体状态向巷道内移动，煤柱宽度在23 m 时向巷道内位移量较大。煤柱宽度在515 m 时，从煤柱中间分开，一部分向巷道内的移动，位移量较大; 另一部分向采空区移动，位移量相对较小; 中间呈稳定状态。(2) 煤柱57 m 向采空区的位移较煤柱1015 m 大，但向巷道内的位移比其小。(3) 煤柱515 m 时，硬煤和软煤中部位移均较稳定。(4) 软煤煤柱宽度大于5 m 后，煤柱表面向巷道内的位移量显著增大，7 m 时达到最大; 硬煤煤柱表面向巷道内的位移量随煤柱宽度由大到小再趋于稳定，但减小量不显著，2m 时达到最大。(5) 煤柱23 m 时，由于硬煤已破碎，所以硬煤和软煤的位移量相差很小。2. 3 沿空掘巷煤柱宽度的合理确定煤柱是沿空掘巷围岩结构的一个重要组成部分，其稳定性决定沿空掘巷的稳定性，采用锚杆支护时窄煤柱宽度应满足以下几个要求:(1) 巷道处于应力降低区。当巷道位于应力降低区时，煤柱及巷道的稳定性均较好，所以应将巷道布置在应力降低区。(2) 煤柱内部有稳定的区域。(3) 有利于巷道围岩稳定。数值模拟结果表明，煤柱较小时，随煤柱宽度增大，巷道围岩变形量减小，煤柱宽度达到一定值后，随煤柱宽度增大，巷道变形量增加; 因此，煤柱宽度有一个合理的值。综上所述，从煤柱的稳定性和控制巷道变形及采出率方面考虑，以及煤柱的应力、位移分布特征及煤柱宽度对巷道变形影响综合分析，沿空掘巷煤柱的合理宽度为: 软煤57 m，硬煤35 m。3 工程应用刘庄煤矿东二采区11-2 槽煤层划分为五个区段开采，121101 工作面已回采结束， 121102 工作面上邻121101 工作面采空区，下方为实体煤，风巷沿采空区留7 m 小煤柱掘进; 断面为半圆拱形; 断面尺寸: 掘进断面宽度为5 200 mm，高度为4 100 mm。风巷支护参数如下:（1）顶板: 顺槽顶板支护采用22 2 500 mm的20MnSi 左旋无纵筋螺纹钢等强预拉力锚杆，配和走向钢带和三角形金属网联合支护，锚杆间排距为800 800 mm。顶板锚杆全部垂直于巷道表面，顶部锚杆共10 根。巷道掘出520 m 后，及时进行顶部锚索支护，锚索间排距为1 600 mm 1 600 mm，每断面布置3 根，中间锚索布置为纵向槽钢组合桁架锚索，纵向槽钢安装在巷道中部，槽钢梁长3.6 m，每根槽钢梁上安装3 根锚索。锚索倾角从下帮到上帮分别为56、72和90。（2）帮部: 两帮锚杆采用22 2 500 mm 的20MnSi 左旋无纵筋螺纹钢等强预拉力锚杆，高帮间排距为600 mm 800 mm; 帮锚杆共4 根，全部为水平布置。（3） 巷道两帮距巷道底板为300 mm 的底角锚杆俯斜角为30，底角锚杆共布置2 根。（4）锚杆均采用全长锚固+ 走向W 型钢带，钢带长2.5 m 风巷掘进期间巷道支护效果如图9 所示。4 结论（1）煤柱宽度对应力分布影响较大。巷道掘进期间，软煤煤柱较小时其应力较小并且应力均匀; 随煤柱宽度增大，煤柱内最大应力增大，煤柱宽度达到5 m 后，最大应力增加不明显，最大应力在煤柱内的位置也相差不大，应力分布近似呈三角形。硬煤煤柱应力分布与软煤差别较大; 煤柱宽度对最大应力影响不大，煤柱较小时应力分布近似呈三角形; 煤柱较大时近似呈梯形分布。硬煤的煤柱应力均大于相应软煤煤柱的应力。采动影响阶段，煤柱最大垂直应力随煤柱宽度增大而提高。软煤和硬煤3 m 以下煤柱的应力均趋于均化; 与掘进阶段相比，煤柱中部应力降低，5 m以上硬煤煤柱应力增加均较大，而软煤煤柱应力增加不大; 煤柱较小时，最大垂直应力降低，其位置向巷道侧移动; 煤柱较宽时，最大垂直应力升高，其位置向采空区侧移动。图9 掘进期间风巷支护效果图（2）不同锚杆支护强度对软煤变形的影响于硬煤，当煤柱宽度在37 m 时，高支护强度对控制围岩变形效果更好。（3）沿空掘巷煤柱的合理宽度为: 软煤57 m，硬煤35 m。参考文献1 张俊云，柴敬 沿空留巷研究中若干问题分析J 矿山压力与顶板管理，2000，( 1 ) : 38-392 徐乃忠，涂敏 厚煤层沿空掘巷底臌机理及控制 安徽理工大学学报( 自然科学版) ，2004，24 ( 2) : 1-43 徐乃忠，涂敏，徐仁海 深井大断面沿空掘巷底臌变形控制技术研究J 安徽理工大学学报( 自然科学版) ， 2007，27 ( 3) : 17-216 FLAC3D( Fast lagrangian Analysis of Continua in Dimensions)version 2 0 Users ManR USA: ItascaConsulting Group Inc， 19977 廖秋林，曾钱帮，刘彤，等 基于ANSYS 平台复杂地质体FLAC3D 模型的自动生成J 岩石力学与工程学报，2005， 24 ( 6) : 1 010-1 0138 侯恩科，吴立新，李建民，等 三维地学模拟与数值模拟的耦合方法研究J 煤炭学报， 2002， 27( 4) : 388- 3929 许能熊，何满潮 层状岩体三维构模方法与空间数据模型J 中国矿业大学学报， 2004， 33( 1) : 103-10610 王明华，白云 层状岩体三维可视化构模与数值模拟的集成研究J 岩土力学，2005，26 ( 7 ) : 1 123-1 126Prevention and forecasting of rock burst hazards in coal minesAbstract：Rock bursts signify extreme behavior in coal mine strata and severely threaten the safety of the lives of miners, as well as the effectiveness and productivity of miners. In our study, an elastic-plastic-brittle model for the deformation and failure of coalrock was established through theoretical analyses, laboratory experiments and field testing, simulation and other means, which perfectly predict sudden and delayed rock bursts. Based on electromagnetic emission(EME), acoustic emission (AE) and microseism effects (MS) in the process from deformation until impact rupture of coal-rock combination samples, a multi-parameter identification of premonitory technology was formed, largely depending on these three forms of emission. Thus a system of classification for forecasting rock bursts in space and time was established. We have presented the intensity weakening theory for rock bursts and a strong-soft-strong (3S) structural model for controlling the impact on rock surrounding roadways, with the objective of laying a theoretical foundation and establishing references for parameters for the weakening control of rock bursts. For the purpose of prevention, key technical parameters of directional hydraulic fracturing are revealed. Based on these results, as well as those from deep-hole controlled blasting in coal seams and rock, integrated control techniques were established and anti-impact hydraulic props, suitable for roadways subject to hazards from rock bursts have also been developed. These technologies have been widely used in most coal mines in China, subject to these hazards and have achieved remarkable economic and social benefits.Keywords：rock bursts; elastic-plastic-brittle model; multi-parameter premonitory; intensity weakening; strong-soft-strong structure; directional hydraulic fracturing; anti-impact hydraulic prop1 IntroductionCoal resources are the main source of energy in China and 95 of the coal produced comes from underground mines. As the mining depth increases (about 20m per year)and geologica1 conditions deteriorate, the mechanical environment and basic behavior in deep-leve1 mining is significantly different from that in shallow mining and shows obvious characteristics of nonlinear dynamic instability, which may easily lead to an increase in dynamic disasters, such as rock burst, roofs collapsing over large areas and other problems which pose serious threats to the safety of coal production in mines. The rock burst is one of the typical dynamic hazards in coal mining, which is caused by elastic energy emitted in a sudden, rapid and violent way in a coal-rock mass and even can increase the possibility of other dynamic accidents such as coal and gas outburst, explosions.etc. Rock burst hazards exist in over 100 coal mines in China, especially in Fushun, Fuxin, Xinwen, Yanzhou, Kailuan, Datong, Xuzhou and Huating. For example, a serious rock burst and gas explosion accident occurred in the Sunjiawan coal mine in Fuxin Liaoning province on Feb, 14, 2005. After the ML=2.5 rock burst occurred, a large amount of gas was emitted, which then induced a serious gas explosion and resulted in many injuries and loss of life. Thus, the safety and highly efficient production of the underground mine had been severely impacted by the rock burst.The rock burst mechanism is a quite complicated problem. Although much significant research has been carried out around the world, from rock burst mechanism studies to rock burst forecasting and hazard contro1, there are still many key issues requiring further research. Our study mainly presents recent progress in research on the prevention and control of rock bursts conducted at the China University of Mining Technology.2 Tendencies in rock burst of compound coal-rock samplesFrom the analysis of roof and floor structures in previous rock bursts, it appears that a considerable number of rock bursts occur under conditions of hard roof and floor structures. Especially hard thick sandstone roofs overlying coal seams is the one of major factors affecting rock bursts. Under conditions of “two-hard” (hard roof and floor), the strength and thickness of the coa1 seam also has certain effect on the distribution of secondary stress after the excavation of a coal-rock mass. Therefore, the research on the tendency of rock bursts in compound coal-rock samples in the system of “roof-coal seam-floor”, as well as the effects of the strength and thickness of coal seams on rock burst occurrences, will greatly benefit the prevention and control of rock burst hazards. From the laboratory research on compound coal-rock samples, our results indicate that the higher the proportion of the roof component in the compound coal-rock, the higher the modulus of elasticity and degree of breakage and the greater the tendency of rock bursts and unconfined compressive strength (UCS) in the samples. as seen in Fig.1. Additionally, the index of bursting energy decreases gradually, fo1lowed by an increase in the proportion of coal components in compound coal-rock. In contrast, the index of elastic energy increases gradually as the proportion of coal components increases, i.e. the higher the coal height in the compound coa1-rock, the larger the index of elastic energy.Fig.1 Relationships between indices of rock burstand ratio of coal and roof heightThe 1arger the hardness of the roof, the higher the degree of stress concentration and vertical stress around the working face. Fig.2 shows the simulated distribution of vertical stress around a workface when the mining depth is 700m, the thickness of the roof 6, 20 and 10m. the bulk modulus l6, 27 and 16GPa and 4, 3 and 4GPa, respectively. As shown in the figure, the vertical stress in the coal-rock mass is much higher under conditions of hard roof than under soft roof conditions. The difference in maximum vertical stress under the two conditions is nearly 44 MPa.The maximum vertical stress in soft roofs decreases approximately 45. The concentration of local stress appears in the goaf roof near the coal mass side under hard roof conditions. In addition, the variations of hardness and thickness of coal seams also have a large effect on vertical stress distribution of a coal-rock mass under the “two-hard” condition, as seen in Fig.3. The vertical stress in the coal body decreases, with a decrease in the hardness of coal seams, with the difference between hard and soft coal conditions approaching 44.8 MPa. The maximum vertical stress decreases as the thickness of the coal seam increases. From the data of our numerical simulation, a regression analysis shows that the relationship between the maximum vertical stress F and the thickness of coal seam h can be expressed as a quadratic function, i.e. F =ah2 bh+c. Therefore, in the system of “roof-coal seam-floor”, the smaller the thickness of the coal seam, the larger the maximum vertical stress of the coal body. In other words, the smaller the proportion of the coal seam thickness，the easier a rock burst can be induced under the “two-hard” condition.Fig.2 Vertical stress distribution of coal mass ahead of the working faceunder different hardness conditions of the roofFig.3 Relationship between vertical stress distributionon the coal mass side and stiffness of the coal seam3 Technique of multi-parameter classification forecasting of rock bursts3.1 Effects of AE, EME & MS in the deformation and failure of a coal-rock massThe electromagnetic emission (EME), acoustic emission (AE) and microseism (MS) effects in the process of rock burst failures of a compound coal-rock mass were tested in the State Key Laboratory of Coal Resource and Mine Safety, China University of Mining& Technology.For the compound coal-rock samples, the count rates(or pulse numbers)of the AE EME include both the deformation and failure of roof and coal. When the applied load approaches the ultimate strength of the coal samples, the roof begins to yield and unload. For the roof part, this will not cause a burst failure in the roof because of its high compressive strength and the deformation and failure of the coal mass will result in an elastic resilience of the roof. In the period of deformation resilience, the count rates of the AE EME signals decrease.Meanwhile, before the burst failure of the coal mass, much energy is released because the roof starts to unload and springs back，which will accelerate the deformation and failure of the coal mass and the count rates of the AE & EME signals will reach their maximum valuesTherefore, the rules for the relations of AE EME in the premonitory information of rock burst can be written as：nt=nu1.u1+nu2.u2 (1)nt=nu2.u2 (2)where nt is the time series of the accumulated count rates of the AE & EME signals, u1, u2 are the variations in velocity of the deformation of roof and coal mass, respectively, nu1, nu2 are the parameters which are closely correlated with the brittleness and UCS of roof and coal body, respectively.Fig.4 shows the distribution of the count rates (or pulse numbers) of the AE & EME signals by the deformation and failure of the compound sampleFig. 4 Time series of AE & EME signalsAs shown in Fig.4, the count rates of AE and pulse numbers of EME have distinct, two-stage failure features in the process of repeated loading and unloading until impact rock burst failure occurs. The count rates or pulse numbers of AE & EME reach their extreme points before the burst failure of the samples, after which the intensity of the signals decreases suddenly. The signals in the second stage are mainly the results of the deformation and failure in the post-peak phase, with poor and steady signals.As a consequence of these results, the risk of rock bursts in mining and extraction processes of working faces can be monitored and predicted in real time according to the multi-parameter premonitory features of AE & EME in the process of burst failures in the compound coal-rock.Technique of classification forecasting of rock bursts Based on a theoretical analysis, laboratory testing and extensive field trials, the zero, weak, middle and strong risk of rock bursts can be classified quantitatively by a risk index of rock bursts. According to the various danger levels of rock bursts, correspondingly controlling measures can be taken, as shown in Table 1.Table 1 Danger levels of rock burstsDanger levelRisk stateRisk indexControlling measureANon-risk0.75Stop mining work and withdraw workers from dangerous locations. Take rock burst intensity weakening measures. After the checking of intensity, weakening effort by monitoring again, further mining work can continue until the risk of rock bursts has been eliminated.For mines and mining districts with a danger of rock bursts, the geology and mining conditions are in first instance, analyzed by a comprehensive index method and then the danger zone of rock bursts and key monitoring regions are marked of. Thus early rock burst forecasting can be achieved. Based on early forecasting, micro seismic monitoring is used for regional monitoring in real-time. For regions with abnormal micro seismicity，an electromagnetic emission (EME) method is adopted for further local monitoring. Again, a drilling method can be used for forecasting and effort verification in dangerous local regions. Therefore, the danger level of rock bursts can be determined comprehensively by this classification forecasting technique, where dangerous areas and spots can be controlled using intensity weakening technology. Fig. 5 shows the method of classification forecasts of rock bursts and steps of implementation during field practice.Fig. 5 Method of classification forecasting rock burst and steps of implementationFor dangerous area with rock bursts, the work flow of classification forecasting and controlling technique is as follows: Early comprehensive forecasting (Using the comprehensive index method to determine key monitoring regions) Real-time forecasting Regional forecasting(Using continuous micro seismic monitoring for regional forecasts of rock bursts over time) Local forecasting (Using continuous EME monitoring for local forecasts of rock bursts overtime) Local point forecasting (Validation of the accuracy of regional and local forecasts using a drilling method and, as well, carrying out point forecasting) Gradually eliminate and confirm the danger rank Relieve risk process (Intensity weakening control of coal-rock mass, eliminate risk of rock bursts) Test controlling effort (Weakening effort verification using MS, EME and drilling)4 Intensity weakening theory of rock burstsWhen taking the dynamic load impact of coal-rock mass by mining-induced tremors into account, the variation of accumulated elastic strain energy ahead of the working face, or around the roadway, can be expressed with the following function. where U is the accumulated elastic strain energy of the coal-rock mass at any time, Uf the disturbance energy by mining-induced tremors, U0 the difference between the initial accumulated and dissipated elastic strain energy of the coal-rock mass, Uj the limited elastic stored energy of the coal-rock mass, Ut the increment of accumulated elastic strain energy at any time of coal-rock mass, while Ue denotes the released elastic energy by the relieve shot. Fig. 6 presents the rock burst intensity weakening mode. As shown in the figure, the decrease in intensity and thickness of the entire roof and floor and the decrease in intensity in the coal mass using a relieve shot, can reduce the rate of energy accumulation of the compound coal-rock mass. When the total strain energy can be released, the intensity weakening of a rock burst is achieved, which is the theoretical basis of the intensity weakening theory for rock bursts. Fig. 6 Energy accumulation and release before and after intensity weakening of coal-rock massThe concrete terms, we should mainly reflect on two aspects. (1) In a dangerous district of rock bursts, reducing the intensity and integrity of the thickness of the roof by loosening the coal-rock mass, causes a decrease in the tendency for rock bursts. (2) After intensity weakening of the coal-rock mass, the peak area of the stress moves deep into the heart of the coal mass, causing a reduction in the rate of energy accumulation of the coal-rock mass. (3) After taking intensity weakening measures, a burst of energy of the coal-rock mass is induced and the intensity of the rock burst reduced.5 Strong-soft-strong structure effort of rock surrounding roadwaysThe stress within a rock mass surrounding a roadway will be redistributed by the excavation of the roadway. When the superimposition of an in situstress field and the stress shockwave from the external hypocenter exceeds the limited bearing capacity of the rock surrounding the roadway, the state of balance of the rock mass is broken and the surrounding rock will suffer transient fracturing or cumulative damage from the iterative effort of tension and compression of the stress wave. On the other hand, although the shockwave from the external hypocenter is, at times, not strong enough, rock burst hazards continue to exist if the superimposed stress field exceeds the limited bearing capacity of the rock surrounding the roadway. Therefore, the transmission and disturbance of shockwaves induced by underground mining is a key factor which may be related to impending rock burst hazards in roadways, that is to say, as long as the total superimposed stress intensity is beyond the utmost carrying capacity of the roadway supporting system, rock bursts may happen around the roadway. Based on the disturbance of external shockwaves, a strong-soft-strong structure model for controlling rock surrounding roadways is established. In the strong-soft-strong structure model shown in Fig. 7, the relative, intact status of the external strong structure, causes the attenuation index of the seismic waves to be relatively small. When the seismic waves are transmitted in this structure, there is no significant attenuation of the seismic energy and only a small part of the energy is absorbed. Therefore, the external strong structure is presented as a weak characteristic in its ability to dissipate energy. Because of its poor integrity, continuity and high porosity, the seismic waves can be scattered and absorbed largely in the middle soft structure. Therefore, the middle soft structure is presented as a strong characteristic in its ability of energy dissipation. The stronger the characteristic, the more favorable the protective effect on the surrounding rock. Because of the compact and intact status of internally strong structures, the entire structure just moves following the distortion of the middle soft structure and its own deformation is relatively small. For the capacity of limited energy dissipation, the strong internal structure is also presented as a weak characteristic of energy consumption. Hence, from the point of energy dissipation of shockwaves, a strong-soft-strong structure can be presented as a soft-strong-soft characteristic.Fig. 7 Strong-soft-strong structure model for controlling surrounding roadwaysTherefore, the prevention of rock bursts in roadways can be achieved by reducing the external load of the seismic source, by properly setting a soft structure, by improving support reinforcements and by other measures. 6 Key technology for rock burst preventionDeep hole relieve shot of coal-rock massAs the main method for intensity weakening，the deep hole relieve shot of a coal-rock mass plays an important role in rock burst prevention in underground mines. So far, this technology has been widely applied in more than ten mining areas, such as the Sanhejian coal mine in Xuzhou, the Jisan coal mine in Jining, the Huating and Yanbei coal mines in Gansu and so on. Fig. 8 shows the layout of blasting holes of deep hole relieve shots in the roof above an auxiliary roadway of the 6303 working face in the Jisan coal mine. Fig. 8 Layout of blasting holes of deep-hole relieve shot in the roof(2) Directional hydraulic fracturing technology of hard roofsDirectional hydraulic fracturing of hard roofs mainly aims at intensity weakening of rock bursts induced by hard roofs. Especially a dangerous working face with a thick roof is difficult to collapse. At present, this technology has been successfully practiced in the Xinzhouyao and Meiyukou coal mines in Datong. Fig. 9 shows the numerical simulation of the stress distribution around the initial crack tip of the controlled fracture drilling. As shown in the figure, the tensile stress concentration is generated in the hard rock after water injection, where the maximum tensile stress is about 52 MPa. The initial crack tip starts to fracture under the action of tensile stress and the crack extends along the direction of the rock level under the action of 30 MPa high-pressure water. According to the directional hydraulic fracturing practice in the roadway of the 8l008 working face in the 410 panel of the Meiyukou coal mine, the fracturing radius can reach 10 m when the pressure of bumping station drops from 35 to 30 MPa. Fig. 9 Numerical simulation of stress distribution aroundinitial crack tip of a controlled fracturing drilling(3) Anti-impact single hydraulic propIn the advance support section of both roadways of the dangerous working face, subject to rock bursts, a single hydraulic prop protection device has been de signed. This device is composed of a hinged top cover a moulding expansion column, a guard pin, a buffer tank, a buffer spring and protective pads. The moulding expansion column is connected to the buffer tank through the guard pin and the buffer tank is set to the hydraulic prop directly, while the buffer spring is placed inside the buffer tank. When a rock burst occurs suddenly, the protection device is activated simultaneously and the guard pin is cut of automatically, thus the hydraulic prop can be protected.At present, the device has obtained a national utility model patent and obtained effective results in the Gucheng Coal Mine of the Linyi Mining Group, Shandong Province.7 ConclusionsWe have investigated the tendency for rock bursts in compound coal-rock samples and a multi-parameter identification of premonitory technology that mainly depends on MS, EME and AE was formed. Thus a system of classification forecasting rock bursts in space and time was established. As well, we established the intensity weakening theory for rock bursts and a strong-soft-strong structural impact model of controlling rock surrounding roadways, which lays a solid theoretical foundation for the weakening control of rock bursts. In addition, an active control technique based largely on directional hydraulic fracturing and deep hole relieve shots in coal seams and rock has been formed. An anti-impact hydraulic prop, suitable for roadways, subject to rock burst hazards, has been also developed. So far, these research results have been widely used in more than ten burst-prone mining areas, specifically in mines such as the Sanhejian coal mine in Xuzhou, the Jisan coal mine in Yanzou, the Muchengjian coal mine in Beijing, the Huating and Yanbei coal mines in Gansu and others, where practical and effective results have been obtained.The behavior of stress wave propagation in rock walls and the process of rock bursts were simulated by application tests of material similar to rock. 8 AdditionWith the increase in mining depth, the frequency and intensity of rock bursts increase continually, seriously threatening the safety in coal mine production. Controlling the mechanism of rock bursts has attracted much attention of scientists who have proposed a series of theories on the mechanism of the origin of rock bursts and some countermeasures for preventing them from taking place. Classic theories includes, intensity theory, energy theory, impact tendency theory, three guideline theory and others. However, there is not one accepted theory about the mechanism of rock burst formation because it has a rather complex dynamic instability. It is therefore urgent to study the mechanism of the formation of dynamic hazards such as rock bursts.It is well known that the main reason for rock bursts lies in the deformation energy in hard rocks and the mechanical processes involved in rock bursts are generally regarded as static. However, static load theory cannot explain all the mechanisms about rock bursts. The accumulation of strain energy is a necessary condition for rock bursts but not a necessary and sufficient condition. Therefore, an external disturbance is necessary for a rock burst. Moreover, stress waves may be produced by driving, blasting, roof breaking, weighting of the working face, and seismic waves in mining processes, external disturbances are often introduced before the occurrence of a rock burst. It is necessary to study the effect of stress waves in rock burst. Meanwhile, simulation is an important means in the study of rock bursts because it can simulate the rock burst process and provide some important information about the mechanism of its incidence, failure spots and the methods of failure. The effect of rock failure caused by a superimposed sine wave, concluding that the disturbance of stress wave decreases the strength of rocks. The relationship between the disturbance and rock burst needs to be further investigated. Xi simulated the effect of static load on the propagation of known cracks and their unstable failure by using a precrack in the rock and thus obtaining the distribution of its speed of propagation. However, the propagation of cracks under a dynamic load has not been investigated. In our investigation, the incidence and development of rock bursts were simulated with material similar to rock and the mechanism of splitting type rock bursts was studied.In geotechnical engineering, selection and experiments with similar material are relatively mature for static problems within an elastic stage. However, few investigations have been carried out on the characteristics of this material beyond the elastic stage especially those of dynamic unstable failures.In our experiment, sand, gypsum, calcium carbonate cement, borax and water were used to prepare material similar to rocks according to the proportions listed in Table 1 and the mechanical properties of this material were determined. For the measurement of mechanical properties, standard specimens with 50mm in diameter and h=100mm in height were prepared, followed by drying and then tested by using the test system material used in the simulation were prepared according to the proportions of the material numbered 18. The dimensions of the mould were 520 mm 400 mm 70 mm, producing a physical model of the same size as the mould. A module with dimensions 50 mm40 mm70 mm was placed into the mould to simulate the rectangle roadway in making the physical model and a cylindrical module with a hole of 15 mm diameter for an explosion. After the initial set of our simulated material, the module of roadway and the cylinder of the explosion hole were drawn out to form the roadway and the explosion hole in the physical model. The explosion point in the model was arranged vertically above the roadway.The physical model was placed in the test rack and the load of the overlying strata was applied using a lever with a load proportional to the weight of the overlying strata. The loading device is shown in Fig. 2 and the disturbance load was applied at the explosion point. To analyze the effect of material properties and charge location on the roadway structure, three experiments with the proportions of #1#3 material were carried out. The surface of the physical model was painted white to make the results more visible (Fig. 3). The arrangement of measurement points is shown in Fig. 4. The vertical distance between the measurement points is 40mm and the signal of acceleration was picked up by the accelerometer attached to a piece of iron based in the physical model. The vertical acceleration of the piece of iron and that of the physical model can be approximately regarded as identical, since the acceleration transducer was attached to the piece of iron. The output signal of the acceleration transducer was amplified by using an integrated amplifier, recorded and analyzed by a dynamic virtual analyzer and relevant software developed by Chongqing University. The wave form is shown on the computer monitor. Table 2 shows the peak displacement, peak velocity and peak acceleration at each measurement point. In addition, from the three fitted curves, the maximum acceleration attenuates most rapidly, which is correlated with the material properties. There is a specially big difference between the acceleration for the #1 and #3 materials. The variation in vertical acceleration as a function of distance in the surrounding rock can be explained by stress-strain behavior of the material. When the wave propagates in the #1 material, because the material is harder and the ultimate stress is relatively high, the stress range before elastic ultimate stress becomes relatively small, corresponding to soft material.Therefore, the energy absorbed and dissipated is relatively small and the wave amplitude relatively large when the wave propagates in hard material.Fig. 8 shows the breakage of the surrounding rock under explosive stress waves and the same charge to time for the #1 material. It can be seen from Fig. 8 that the cracks appear in the upper boundary surrounding the rock of the road-rib where the stress was concentrated and they propagated deeply in the boundary area parallel to the road-rib. Due to the initial stress field, many cracks were formed, parallel to the road-rib. These cracks propagated, resulting in macroscopic cracks along the direction of the principal stress under the stress waves.Although a certain amount of difference exists between our similar material and actual rock, the process and phenomenon of instability of rock can be accurately simulated when suitable material and their proportions are selected. The attenuation characteristics of stress waves are related to material properties. The stress waves attenuated more quickly for the softer material. For the explosion load applied at the top of the roadway, the number and length of cracks increase with a decrease in the distance between the source of the explosion and the roadway. When distance h=280 mm from the roof of roadway there were no cracks near the roadway. However, when distance h=210mm along the roadway, some minute cracks appeared near the road-rib and when distance h=140mm, larger cracks appeared. Under a given amount of pressure, the appearance of cracks is related to the intensity of the stress wave. The intensity of the stress wave can be controlled in order to decrease the appearance of cracks in the surrounding rock and prevent the formation of a layered crack structure.中文译文：冲击地压的预防和预测摘要：冲击地压，煤层一种极端活动行为的显示，严重威胁着矿工的生命安全。冲击地压同样影响着矿工的工作效率和生产力。在我们的研究中，借助煤岩弹塑性脆性变形破坏模型，通过理论分析，实验室实验，现场测试、模拟和其他手段，很好地预测了瞬间和延迟冲击地压。基于电磁辐射、声发射、微震等先进技术，从煤岩样品的变形到挤压破坏，一个多参数的预测技术随之形成。在很大程度上取决于这三种形式的共同作用。这样，在时间和空间上预测岩石脉冲的分类系统就建立起来了。我们已经提出了岩石强度弱化理论和“强-弱-强”脉冲结构模型来控制冲击地压对岩石巷道的影响。奠定了坚实的理论基础，达到了弱化、控制冲击地压的作用。为达到预防的目的，定向水力压裂关键技术参数得到了显示。基于以上的结论，以及在煤层、岩层中打的深层爆破钻孔，综合防治技术得到了建立；在有冲击地压危险的巷道中，抗冲击液压支柱也得到了开发。这些技术在中国的大部分煤矿得到了应用，有效控制了灾害，并取得了很好的经济和社会效益。关键词：冲击地压；岩弹塑性脆性变形破坏模型；多参数预测；强度弱化；“强-弱-强”结构；定向水力压裂；抗冲击液压支柱1 介绍煤炭资源是我国主要的能源资源，95%的煤炭产量来源于井工煤矿。随着开采深度的增加（大约20m/年），和地质条件的恶化，深部开采和前部开采在开采环境和基本理论方面呈现出显著的不同，并且呈现出明显的非线性动态不稳定特征。从而容易导致动力灾害数量的增加，例如：冲击地压、顶板大面积离层垮落等一系列严重问题，严重威胁着煤矿的安全生产。冲击地压就是煤矿生产的典型危险源之一，源于围岩弹性能量的突然发出，煤岩质量的迅速、激烈增加甚至会增加其他动力事故的可能性，比如：煤与瓦斯突出，瓦斯爆炸等。在我国，100多座煤矿有冲击地压危险。特别是在抚顺、阜新、兖州、开滦、大同、徐州和华亭。例如，2005年2月14日，在辽宁省阜新市孙家湾煤矿发生了一次巨大的冲击地压和瓦斯爆炸事故，在强度为2.5ML的冲击地压发生之后，大量的瓦斯气体被激发了，诱导一次严重的瓦斯爆炸，造成大量的人员伤亡。因此，煤矿井下的安全、高效生产受到冲击地压的严重威胁。冲击地压发生机理是一个相当复杂的问题。尽管在世界范围内，从冲击地压作用机理的研究到灾害的预测和控制进行了大量的探讨，但仍存在许多关键性的问题需要进一步研究。本文介绍了中国矿业大学近年来在冲击地压的预防和控制上进行的探索和研究。2 煤岩化合物样品冲击地压倾向通过对以往发生冲击地压的巷道顶底板结构的分析，呈现出大量的冲击地压发生于坚硬顶底板结构下的趋势。特别的，煤层上部厚而坚硬的砂岩顶板是影响冲击地压的一个主要因素。在坚硬顶底板的条件下，煤层的强度和厚度同样对煤岩二次开挖后产生的压力有所影响。因此，在“煤层顶底板”系统中，通过对煤岩混合物样品冲击地压的倾向性的分析调查，同时，调查煤层的强度和厚度影响冲击地压的结果将有利于对冲击地压危害的预防和控制。通过在实验室里对煤岩混合物的研究，我们的结论表明，煤岩混合物中顶板的比重越高，样品的弹性模量，损坏度越大，含水层样品抗压强度和冲击地压的趋势也更大。如图1所示，另外，随着煤岩混合物中煤的组分含量的降低，爆破能量指数逐渐降低。相反地，随着煤岩混合物中煤的组分含量的升高，爆破能量指数逐渐升高。煤岩混合物中煤的含量越高，爆破能量的指数也越高。图1 冲击地压指标和煤岩高度比之间的关系顶板的坚硬程度越高，工作面所受的集中应力和垂直应力也越大。图片2表示了开采深度达到700m时，工作面所受到垂直应力的模拟分布。顶板的厚度为6m、20m和10m时，体积变化模量一般为16GPa、27GPa和16 GPa ； 4GPa、 3GPa和 4 GPa。如图所示，煤岩体中，坚硬顶板的条件下所受的垂直应力大于松软顶板条件。两种条件下，垂直应力的最大差值在44兆帕左右。松软顶板的条件下，最大水平应力减少近45%。在坚硬顶板条件下，局部集中应力出现在采空区顶板下的煤附近。另外，在坚硬顶底板的条件下，煤层的坚硬度和厚度的变化同样对作用在煤壁上的垂直应力分布起着重要的作用。如图3所示，随着煤体集中应力的减小，煤层的坚硬程度随之减小，两者最大情况下相差44.8兆帕。最大水平应力随着煤层厚度的增加而增加。我们从数值模拟中得到的数据进行分析，我们得知最大水平应力F与煤层厚度h的关系可用一个二次函数表示F =ah2 bh+c，因此，在煤层顶底板系统中，煤层的厚度越小，煤体所承受的最大水平应力越大。换句话说，在坚硬顶底板的条件下，煤层厚度的比例越小，冲击地压发生的几率越大。图2 不同顶板强度条件下工作面前方煤体垂直压力分布图3 煤体边缘垂直压力分布和煤层刚度之间的关系3 多参数分类预测冲击地压技术3.1 电磁发射、声发射、微震等技术在煤岩样品的破坏、变形中的应用在中国矿业大学的煤炭资源和安全开采重点实验室，电磁发射、声发射、微震等技术在煤岩混合物的实验样品在发生冲击压作用下发生破坏的过程中进行了实验。对于煤岩混合物样品来说，声发射或电磁发射技术在岩层顶板或煤层中会产生数字频率（或脉冲数字）。当加载的负荷达到煤样的最大承载能力的时候，顶板开始弯曲并逐渐脱落。对于顶板部分，这不会因为冲击地压而造成破坏，原因在于它的高的抗压和抗弯强度。煤壁的破坏会给顶板带来弹性应力。在破坏变形的期间，声发射或电磁发射的数字频率信号会有所减弱。期间，在煤壁发生变形破坏之前，由于顶板开始垮落并产生弹性恢复变形，很多能量得到了释放，这会加速煤壁的变形破坏，声发射或电磁发射的数字频率信号会达到最大强度，因此，冲击地压的预警信息之前，声发射或电磁发射的关系可以描述为： nt=nu1.u1+nu2.u2 (1)nt=nu2.u2 (2)其中，nt是电磁发射、声发射信号的计数率在时间上的积累。一般地，u1, u2是顶板和煤体变形速度的变化，而nu1, nu2是与顶板和煤体脆性与抗压强度密切相关的参数。图片4展示了电磁信号和声发射信号在化合物样品破坏变形中计数率的变化情况。图4 电磁发射、声发射信号随时间变化趋势如图片4所示，电磁信号的计数率和声发射信号的脉冲数在试件样品重复的加载和卸载直到冲击地压发生的变化过程中，两个阶段的破坏特征是有所区别的。在试件发生破坏特征之前，电磁信号和声发射信号的计数率和脉冲数会达到极限点，之后，信号的强度迅速地减弱。第二阶段的信号主要是破坏变形达到峰值之后的所反映出的微弱和平稳的信号。由于这些结果，发生在采煤过程中和工作面掘进过程中发生岩石破坏的风险，可以根据煤岩化合物在出现冲击特征的过程中所反映出的征兆和多参数电磁信号和声发射信号，在正确的时间被检测和预测到。3.2 冲击地压的预测和分类技术基于理论分析、实验室实验和广泛的现场实践，零、微弱、中等、强冲击地压可以用冲击地压风险指数定量分类出来。根据冲击地压危险等级的变化，如表1所示，可采用相应的控制措施。表1 冲击地压的危险等级危险等级风险阶段风险指数控制措施A无风险0.75停止作业，从危险地点撤离工作人员。采取冲击地压强度减弱措施。强度监测之后，强度再次弱化之后，直到冲击地压危险完全消除之后，采煤工作可以继续通过煤矿不同采煤方法产生冲击地压危险的区分，地质和煤层赋存条件是首要因素，通过综合指数方法的分析之后，冲击地压的危险区域和关键监测区域被标注出来，这样，冲击地压的早期预测工作就完成了。基于早期预测，区域性预测中，在合适的时间微震技术得到了采用。微震活动异常区域，电磁发射技术在进一步监测工作中得到了采用。另外，可采用钻孔方法来预测和确认危险区域。因此，冲击地压的危险等级可以被这些不同的预测技术所广泛地确定，对于危险的区域和危险点，可以通过强度弱化技术来控制。图5展示了现场实践过程中，冲击地压的分类预测方法以及在现场加以落实的措施。图5 测冲击地压的分类方法和应用步骤对于有冲击地压危险的区域，分类预测方法和控制技术如下所示：早期综合预测（使用综合指数方法来决定主要的监测区域）实时监测区域预测（之后使用连续的微震技术进行区域性预测）局部地点的预测（验证区域性预测的准确性，使用钻孔方法、推进点的预测等）逐步确认和减弱等级危险减轻风险过程（弱化控制煤岩体的强度，消除冲击地压的危险）测试控制效果（使用声发射、微震技术、钻孔方法来验证弱化结果）4 验证冲击地压的削