Study on the disaster caused by the linkage failure of the residual coal pillar and rock stratum during multiple coal seam mining:mechanism of progressive and dynamic failure

Multi-seam mining often leads to the retention of a significant number of coal pillars for purposes such as protection,safety,or water isolation.However,stress concentration beneath these residual coal pillars can significantly impact their strength and stability when mining below them,potentially leading to hydraulic support failure,surface subsidence,and rock bursting.To address this issue,the linkage between the failure and instability of residual coal pillars and rock strata during multi-seam mining is examined in this study.Key controls include residual pillar spalling,safety factor(f.),local mine stiffness(LMS),and the post-peak stiffness(k)of the residual coal pillar.Limits separating the two forms of failure,progressive versus dynamic,are defined.Progressive failure results at lower stresses when the coal pillar transitions from indefinitely stable(f,>1.5)to failing(f,<1.5)when the coal pillar can no longer remain stable for an extended duration,whereas sud-den(unstable)failure results when the strength of the pillar is further degraded and fails.The transition in mode of failure is defined by the LMS/k ratio.Failure transitions from quiescent to dynamic as LMS/k.<1,which can cause chain pillar instability propagating throughout the mine.This study provides theoretical guidance to define this limit to instability of residual coal pillars for multi-seam mining in similar mines.

Coupling effects of coal pillars of thick coal seams in large-space stopes and hard stratum on mine pressure

Concerning the issue of mine pressure behaviors occurred in fully mechanized caving mining of thick coal seams beneath hard stratum in Datong Mining Area, combined with thin and thick plate theory, the paper utilizes theoretical analysis, similar experiments, numerical simulations and field tests to study the influence of remaining coal pillars in Jurassic system goaf on hard stratum fractures, as well as mine pressure behaviors under their coupling effects. The paper concludes the solution formula of initial fault displacement in hard stratum caused by remaining coal pillars. Experiments prove that coupling effects can enhance mine pressure behaviors on working faces. When inter-layer inferior key strata fractures, mine pressure phenomenon such as significant roof weighting steps and increasing resistance in support.When inter-layer superior key strata fractures, the scope of overlying strata extends to Jurassic system goaf, dual-system stopes cut through, and remaining coal pillars lose stability. As a result, the bottom inferior key strata also lose stability. It causes huge impacts on working face, and the second mine pressure behaviors. These phenomena provide evidence for research on other similar mine strata pressure behaviors occurred in dual-system mines with remaining coal pillars.

Structure instability forecasting and analysis of giant rock pillars in steeply dipping thick coal seams

Structure stability analysis of rock masses is essential for forecasting catastrophic structure failure in coal seam mining. Steeply dipping thick coal seams （SDTCS） are common in the Urumqi coalfield, and some dynamical hazards such as roof collapse and mining-induced seismicity occur frequently in the coal mines. The cause of these events is mainly structure instability in giant rock pillars sand- wiched between SDTCS. Developing methods to predict these events is important for safe mining in such a complex environment. This study focuses on understanding the structural mechanics model of a giant rock pillar and presents a viewpoint of the stability of a trend sphenoid fractured beam （TSFB）. Some stability index parameters such as failure surface dips were measured, and most dips were observed to be between 46° and 51°. We used a digital panoramic borehole monitoring system to measure the TSFB＇s height （△H）, which varied from 56.37 to 60.50 m. Next, FLAC^3D was used to model the distribution and evolution of vertical displacement in the giant rock pillars; the results confirmed the existence of a TSFB structure. Finally, we investigated the acoustic emission （AE） energy accumulation rate and observed that the rate commonly ranged from 20 to 40 kJ/min. The AE energy accumulation rate could be used to anticipate impeding seismic events related to structure failure. The results presented provide a useful approach for forecasting catastrophic events related to structure instability and for developing hazard prevention technology for mining in SDTCS.

Coal pillar design when considered a reinforcement problem rather than a suspension problem

Current coal pillar design is the epitome of suspension design.A defined weight of unstable overburden material is estimated, and the dimensions of the pillars left behind are based on holding up that material to a prescribed factor of safety.In principle, this is no different to early roadway roof support design.However, for the most part, roadway roof stabilisation has progressed to reinforcement, whereby the roof strata is assisted in supporting itself.This is now the mainstay of efficient and effective underground coal production.Suspension and reinforcement are fundamentally different in roadway roof stabilisation and lead to substantially different requirements in terms of support hardware characteristics and their application.In suspension, the primary focus is the total load-bearing capacity of the installed support and ensuring that it is securely anchored outside of the unstable roof mass.In contrast, reinforcement recognises that roof de-stabilisation is a gradational process with ever-increasing roof displacement magnitude leading to ever-reducing stability.Key roof support characteristics relate to such issues as system stiffness, the location and pattern of support elements and mobilising a defined thickness of the immediate roof to create(or build) a stabilising strata beam.The objective is to ensure that horizontal stress is maintained at a level that prevents mass roof collapse.This paper presents a prototype coal pillar and overburden system representation where reinforcement, rather than suspension, of the overburden is the stabilising mechanism via the action of in situ horizontal stresses.Established roadway roof reinforcement principles can potentially be applied to coal pillar design under this representation.The merit of this is evaluated according to failed pillar cases as found in a series of published databases.Based on the findings, a series of coal pillar system design considerations for bord and pillar type mine workings are provided.This potentially allows a more flexible approach to coal pillar sizing within workable mining layouts, as compared to common industry practice of a single design factor of safety(Fo S) under defined overburden dead-loading to the exclusion of other relevant overburden stabilising influences.

A physical model study of surrounding rock failure near a fault under the influence of footwall coal mining

A study of the deformation of the surrounding rock and coal pillars near a fault under the influence of mining is conducted on a physical model for the design of coal pillars to support and maintain the roofs of adjacent fault roadways.This research is based on the 15101 mining face in the Baiyangling Coal Mine,Shanxi,China,and uses simulation tests similar to digital speckle test technology to analyse the displacement,strain and vertical stress fields of surrounding rocks near faults to determine the influence of the coal pillar width.The results are as follows.The surrounding rock of the roadway roof fails to form a balance hinge for the massive rock mass.The vertical displacement,vertical strain and other deformation of the surrounding rock near the fault increase steeply as the coal pillar width decreases.The steep increase in deformation corresponds to a coal pillar width of 10 m.When the coal pillar width is 7.5 m,the pressure on the surrounding rock near the footwall of the fault suddenly increases,while the pressure on the hanging wall near the fault increases by only 0.35 MPa.The stress of the rock mass of the hanging wall is not completely shielded by the fault,and part of the load disturbance is still transmitted to the hanging wall via friction.The width of the fault coal pillars at the 15101 working face is determined to be 7.5 m,and the monitoring data verify the rationality of the fault coal pillars.

遗留煤柱边界下方撤面巷道合理位置确定研究

为了确定遗留煤柱边界下方撤面巷道合理布置位置,以内蒙古某矿为研究对象,基于遗留煤柱覆岩空间结构,探究应力在底板的传递规律,通过构建“上煤层遗留煤柱传递应力-下煤层巷道掘进应力”应力叠加计算模型,分析外错距离与撤面巷道围岩安全稳定之间的关系。研究结果表明:在遗留煤柱的“支撑”作用下,煤柱煤体及其上覆岩层近似成“T”型空间结构;10 m外错距离满足现场安全生产和巷道围岩控制的要求。研究结果可为相似开采条件下的巷道稳定评估和合理位置确定提供参考。

远距离下保护层开采遗留煤柱对被保护层回采的影响研究

为探究远距离下保护层开采遗留煤柱对被保护层回采效果的影响,以平顶山六矿戊_(8)煤层的戊_(8)-22310机巷和戊_(8)-32010风巷之间遗留宽度为4 m的区段煤柱和丁_(5-6)-22190工作面为研究对象,采用理论分析、数值模拟和现场实测结合的方法,研究了煤柱影响范围、煤柱区域垂直应力分布及影响区相关参数变化规律。结果表明:宽4 m遗留煤柱的影响范围为27.3 m,极限强度为25.9 MPa,塑性区宽度为2.26 m,弹性区宽度为2.68 m;煤柱两侧工作面均回采后,其承载的最大垂直应力87.9 MPa,远超煤柱本身的承载极限;在煤柱影响范围内测得的残余瓦斯含量、压力与卸压保护区无明显差异,并均小于突出危险临界值,煤柱影响区内瓦斯得到较好释放;煤柱影响范围内未产生明显应力集中现象,遗留小煤柱对被保护效果影响较小,保护效果在倾向上有连续性,对被保护层工作面回采影响小。

工业广场保护煤柱开采井筒破损致因及防治技术

针对工业广场保护煤柱开采导致井筒破损的问题,以大社煤矿为工程背景,分析工广煤柱开采井筒围岩移动变形特征,揭示井筒破损致因;提出工广内后续工作面开采防治井筒破坏方案,形成相应的施工技术,并成功应用工程实践。研究表明:该矿工广内92606工作面邻近副井,开采引发的井筒竖向附加力是井筒破坏的致因;如不提前采取防治措施,后续92606外工作面开采,将导致副井井筒发生二次破坏;研究提出的不停产单卸压槽防治方案,具有不影响矿井生产、卸压率高、施工简单、成本低等优点。采用壁后注浆、开切卸压槽、架设密集井圈等技术,顺利完成了该矿不停产防治副井井筒破损工程。监测表明,该井筒竖向和环向应力,均小于井壁极限.承载力,处于安全状态,运行良好,达到了预期效果。

两次采动影响下小煤柱巷道切顶卸压围岩控制技术

为解决多次采动影响下留小煤柱巷道矿压显现剧烈、围岩变形量大的问题,提出了“切顶卸压+顶、帮锚索补强”联合控制技术。以贾家沟煤矿10115运输巷为工程背景,分析了切顶卸压位置、高度与巷道围岩垂直应力分布的关系,给出切顶卸压关键参数,并研究了留小煤柱巷道在切顶卸压后受两次采动影响下的矿压显现规律。结果表明:切顶卸压后,煤柱上的应力峰值随切顶深度的增加呈指数降低;10115运输巷在邻近工作面一次采动过程中巷道围岩变形依次呈现出无变形、缓慢变形、快速变形和围岩稳定的变化规律,本工作面二次采动过程中巷道围岩变形依次呈现出明显变形和剧烈变形的规律;巷道围岩在受二次采动影响时变形更加剧烈,巷道变形量为一次采动时的3.0~7.5倍。

坚硬顶板切顶卸压技术对巷道围岩变形规律影响

针对特厚煤层坚硬顶板、宽煤柱条件下临空巷道面临的高围岩应力、大变形等问题,以老石旦煤矿16403综放工作面为工程研究背景,从宽煤柱顶板侧向破断结构角度对临空巷道大变形的影响因素进行了理论分析,采用数值模拟方法研究了对16402运输巷实施不同切顶卸压方案时,临近采空区的16403回风巷侧向顶板采动应力传递规律,并在现场施工水力压裂钻孔进行切顶卸压,实现临空巷道围岩变形控制。研究结果表明:“低位坚硬岩层悬臂梁+高位坚硬岩层砌体梁”破断结构是特厚煤层宽煤柱临空巷道大变形的主要原因,可采用切顶卸压技术破坏宽煤柱顶板侧向破断结构来控制临空巷道围岩大变形;切顶角变化可使关键块B长度发生改变,切顶角越大,则关键块B长度越小,临空侧顶板载荷向煤柱传递的程度越弱,临空巷道围岩承受的采动应力越小,切顶角为100°时临空巷道围岩垂直应力与变形量最小;在16402运输巷以切顶角100°施工水力压裂钻孔后,16403回风巷顶底板变形量较未实施切顶卸压的16402回风巷减小86.5%,两帮变形量减小87.1%,临空巷道围岩稳定性得到极大提高。

高强度重复开采覆岩与地表移动机理

为了给神东等大型煤炭生产基地生态修复和“三下”开采提供理论和技术支持,以实测方式为基础,深入研究了高强度重复开采地表移动特征和机理。分别对老采空区下开采、工作面隔离煤柱下开采、跨老采空区和隔离煤柱下开采这3种工况布设地表移动观测线,开展快速密集监测,采集了丰富的地表移动数据。结合井下推进过程,通过数据分析,获得了不同时期下沉和水平移动值,绘制了地表移动曲线时序图。研究表明:隔离煤柱对地表下沉的发展过程具有阻碍性、带动性和活化性等特征,水平移动方向对井下推进方向具有追随性等现象。据此建立了高强度重复采动条件下形成的块体岩层结构模型,分析了块状岩体在井下推进过程中平衡—旋转平移—再平衡的运动过程,揭示了地表移动实测曲线形态背后的岩层移动机理。研究成果可为该矿区或类似矿区开采沉陷预计模型建立和生态环境修复治理提供参考。

特厚煤层垂直分层区段窄煤柱围岩稳定性及控制技术

针对特厚煤层垂直分层区段窄煤柱大变形失稳控制难题,采用理论分析、数值模拟及现场试验相结合的方法,对内蒙古老公营子煤矿特厚煤层窄煤柱围岩稳定性及控制技术展开研究。笔者在对窄煤柱内三向应力计算的基础上,基于摩尔−库伦破坏强度得到窄煤柱平面应变的屈服准则,定性分析了窄煤柱破坏特征及损伤程度演化规律,进一步分析了区段窄煤柱失稳机理及其尺寸效应影响规律,最后给出区段窄煤柱围岩控制关键技术,并通过数值模拟及现场试验进行验证。研究成果表明:①窄煤柱中部破坏程度大于两帮,中部自顶端向下煤体破坏程度由严重向轻微发展。随着宽高比增大窄煤柱内破坏程度及严重破坏区域占比逐渐减小,当宽高比大于1∶1,窄煤柱中下部开始出现大范围的轻微破坏区,当宽高比大于5∶3,轻微破坏区占比超过50%。当黏聚力C≥3 MPa,或内摩擦角φ≥20°时,窄煤柱两帮破坏程度转变为轻微。②工作面回采巷道一侧为窄煤柱低强度承载区,在上部高应力作用下,大范围的低应力承载区煤体向巷道位移,造成两帮持续性大变形,进而影响顶板的稳定性。煤柱高度是中、底分层区段窄煤柱稳定的主控因素,窄煤柱宽高比增大,对高强度承载区承载强度影响程度较小,而低强度承载区承载强度增加较明显,合理的窄煤柱宽高比可平衡煤柱内高、低强度承载区比例,并提高承载强度。③分层窄煤柱留设需考虑工作面两侧采空时围岩稳定性,保证煤柱内部高强度承载区范围大于煤柱宽度及高度的一半,并通过联合加强支护措施使煤柱内形成由浅入深的多重联合控制区,共同维护煤柱的自稳能力。

深井高应力大巷煤柱释能改性防冲机理及应用

针对当前我国深井高应力大巷煤柱频繁发生冲击地压的现状,采用现场调查、理论分析、数值模拟和工程试验等方法,研究了我国大巷煤柱冲击地压发生类型与机理,提出了一种高应力大巷煤柱释能改性防治冲击地压的新方法,基于该方法开展了防冲原理研究、技术体系设计、井下工业试验等一系列研究工作,得到:基于大巷煤柱冲击地压的主导力源类型,结合多起冲击地压典型案例,将我国大巷煤柱冲击地压划分为基础应力叠加型、附加应力诱发型和自身蠕变失稳型3类,分别建立并推导了3类大巷煤柱冲击地压发生机理的力学模型与工程判据,揭示了3类大巷煤柱冲击地压的发生机制。提出一种深井高应力大巷煤柱压裂−注浆协同防治冲击地压的方法,其防冲原理体现在“释能”和“改性”2个方面,井下工业试验发现弹核煤体起裂压力约40 MPa,整体最大压裂宽度达19.5 m,现场监测验证该方法能够降低大巷煤柱能量积聚水平,提高大巷煤柱整体冲击阈值。设计了深井高应力大巷煤柱释能改性防冲成套技术体系,包括孤立煤体冲击危险“评估−检验−校核”方法、弹性核水力压裂增缝释能技术和整体注浆加固改性技术。研发了大巷煤柱弹性核压裂“震动−应力−位移”三场监测设备和PS-Z-A/B型无机双组份速凝整体注浆改性材料。

大倾角采场围岩应力分布及矸石充填特征的倾角效应研究

煤层倾角是造成大倾角采场采动力学行为呈现复杂性、特殊性,诱发众多灾害事故的重要因素之一,为揭示煤层倾角对大倾角采场围岩控制及矿压显现特征的影响规律,采用物理相似模拟和数值计算相结合的研究方法,在综合分析了大倾角工作面顶板破断运移及矸石滑滚充填特征的基础上,利用有限元−离散元(FLAC2DPFC2D)耦合算法建立了不同倾角的大倾角采场耦合数值模型,研究了大倾角采场围岩应力分布及矸石充填特征的倾角效应。结果表明:①采动作用下,大倾角采场顶底板内围岩应力均呈非对称拱形分布,随着煤层倾角增大,拱形垂直应力释放区范围和向上部偏移程度逐渐增大,但水平应力释放区范围和应力值逐渐减小,无论是垂直应力还是水平应力都易在工作面上下端头顶板处出现应力集中,但最大集中应力会随煤层倾角的增大而减小;大倾角工作面顶底板内应力大小和传递方向均存在非对称特征,随着煤层倾角增大,工作面顶底板应力拱高逐渐降低,围岩应力的传递方向以围岩涌向采出空间为主,由初始近似竖直方向逐渐偏转趋于工作面垂向。②工作面顶板破断及矸石的滑滚充填具有时序性和分区演化特性,并随煤层倾角的改变而呈现一定的倾角效应。随着煤层倾角增大,直接顶的初次破断位置将逐渐向工作面倾向上部区域转移,同时由于重力沿工作面倾向的分力变大,矸石沿倾向的充填程度更为密实,但充填长度减小,工作面中上部区域高位岩层的破碎程度和空洞范围增大。③采空区内矸石对围岩的作用机制主要体现在提供侧向应力和竖向支撑2个方面,且受矸石重力作用影响较大,会随煤层倾角改变呈现较强的倾角效应。

软破矿岩条件下胶结充填法转分段崩落法研究及应用

金川集团龙首矿西二采区上部中段充填采场发生失稳事故后,采矿方法由下向分层进路式胶结充填法转为无底柱分段崩落法时,面临着软破矿岩条件下覆盖层形成和采场稳定性问题。采用数值计算和现场试验等方法,对这2个关键技术问题进行了研究。数值模拟结果表明,随着崩落法首采分段回采工作的进行,采场复合顶板经历了破坏裂纹萌生、零星散块冒落、拱形批量冒落和柱塞状整体冒落等阶段,当首采分段回采结束时,顶板复合岩层冒落能够为崩落法采场提供足够厚度的覆盖层,且在崩落法采动地压下软破矿岩采场能够保持稳定。现场工业试验表明,在崩落法采场首采分段采用“阶梯式退采均匀扩展采空区+总量控制出矿”的技术方案,能够安全、高效地形成覆盖层,并通过增补“钢拱架+钢筋网+锚杆”的技术方案,解决在崩落法采动地压下维持破碎矿岩采场稳定性的问题,为无底柱分段崩落法在西二采区的扩大应用奠定了技术基础。

浆液套管增强沿空掘进煤柱稳定性数值模拟研究

为解决采动作用下沿空掘进巷道一侧窄煤柱变形等问题,现场采取浆液充填套管加固煤柱底板以改善煤柱的稳定性。采用数值模拟分析浆液充填套管对底板岩体变形的控制作用机制。数值计算结果显示,在横断面方向煤柱两侧测点变形高、中间测点变形低,近采空侧的测点变形大于近巷道侧测点变形。煤柱基底变形不一致会导致煤柱受力不均,顶板荷载作用下煤柱变形形态更加复杂。施加浆液充填套管后,煤柱基底测点的变形量减小。巷道侧煤柱底板测点变形减少20.5%,大于采空区一侧的改善效果19.2%。采动作用下随着深度增加底板岩层变形逐渐减小,施加浆液充填套管后底板岩层变形减小。对类似矿井的沿空掘进巷道一侧煤柱变形研究有指导意义。

大倾角大采高软弱顶底板工作面分区矿压规律研究

为研究内蒙古长城二矿大倾角大采高软弱顶底板复杂开采地质条件下,1305S工作面正常区、过渡区、卸压区的矿压显现规律及特征,建立了倾斜煤层采空区冒落矸石滑移充填数值模拟模型,采用数值模拟和现场实测相结合的研究方法进行研究。研究结果表明:沿倾斜方向,1305S工作面中上部所受垂直应力呈现出过渡区>正常区>卸压区的特点,工作面下部呈现出卸压区>过渡区>正常区的特点;1305S工作面推进过程中,平均周期来压步距为13.3 m,周期来压时支架工作阻力在26.75~33.05 MPa之间,沿工作面倾斜方向顶板所受垂直应力总体呈现出工作面上部>中部>下部的特点;巷道工作面一侧的边帮变形量较大。现场实测结果表明,回风巷顶板监测最大移进量为860.18 mm,运输巷顶板监测最大移进量为367.84 mm,两巷围岩移进量呈现出回风巷>运输巷的特点。

营盘壕煤矿深部多煤层开采地表移动变形规律及最优错距研究

鄂尔多斯东胜煤田为侏罗系含煤地层,诸多深部矿区2-2煤层和3-1煤层全区可采,并且煤层间距较小。以营盘壕煤矿为地质原型,通过FLAC3D和UDEC数值模拟软件构建研究区域数值模型,围绕2-2煤层和3-1煤层开采引起的地表移动变形及应力场演化规律进行研究。结果表明:多煤层重复采动将引起沉陷盆地中心投影向下煤层采空区方向移动,进而逼近煤层初采中心投影,并最终越过之;两煤层连续开采后,地表最大下沉值小于两煤层单独开采最大下沉值之和,煤层复采下沉系数小于初采下沉系数;煤柱错距与地表移动参数关联密切,错距系数与地表下沉系数、水平移动系数和主要影响角正切分别呈对数正态函数、凹抛物线和凸抛物线函数关系,并且与宽深比作为共同因子对煤层初采及复采表征参数的比值产生耦合影响;另外,同时随着错距的增大,地表沉陷盆地范围先减小后增大的趋势发展,水平移动近盆地等值线由椭圆形向梭子形演化;综合卸压错距经验公式、地表损伤程度及煤柱卸压效果确定的研究区域多煤层开采最优错距为30 m。

某磷矿房柱法开采空区治理技术研究与应用

某磷矿采用脉内开拓盘区房柱法开采,在开采过程中形成了一定规模的采空区,给未来生产造成了较大的安全隐患。为了消除或减小采空区对矿山安全开采的危害,针对采空区存在的现状,进行工程地质调查,采空区治理及安全监测措施等研究,提出了采用废石干式充填方法进行空区治理方案,并对充填准备和充填步骤进行详细部署;同时,对空区封堵工艺及封堵墙规格进行设计,根据采空区分布特点布置位移监控报警仪和地压监测报警仪,实现采空区地压活动实时监测和预警预报。研究结果为矿山的采空区治理和安全有序生产提供决策依据和技术保障,也为类似条件矿山提供一定的借鉴。

近直立煤层群水平分段综放开采充填卸压防冲研究

目前近直立煤层开采冲击地压防治措施主要有爆破、水压致裂、建立保护层等,或破坏层间岩柱和顶底板,或难以解决大采深情况下的层间岩柱应力集中问题,且会导致较大的地表沉降。以乌东煤矿为工程背景,针对近直立煤层群水平分段综放开采方法,提出了充填采空区技术,以支护层间岩柱及顶底板,降低开采分段周围煤岩体的应力集中现象。设计了3种充填方案:方案1为第一开采分段采空区使用高强度材料充填,其余分段采用普通材料充填;方案2为第一开采分段采空区使用高强度材料充填,其余分段交替采用高强度材料和普通材料充填;方案3为每一分段采空区均使用高强度材料充填。通过数值模拟研究了3种充填方案的卸压防冲效果,结果表明:与未充填相比,3种充填方案下层间岩柱最大垂直应力分别下降25.07%,26.57%,29.23%,下一分段煤体最大水平应力分别下降10.63%,10.79%,12.34%。综合考虑卸压效果和经济效益,优选间隔充填的方案3。指出可结合高应力区域实时智能监测技术,及时支撑层间岩柱,减少层间岩柱及下分段煤体的应力集中,防止冲击地压发生。