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.

Surrounding rock control of gob-side entry driving with narrow coal pillar and roadway side sealing technology in Yangliu Coal Mine

Gob-side entry driving can increase coal recovery ratio, and it is implied in many coal mines. Based on geological condition of 10416 working face tailentry in Yangliu Coal Mine, the surrounding rock deformation characteristics of gob-side entry driving with narrow coal pillar is analysed, reasonable size of coal pillar and reasonable roadway excavation time after mining are achieved. Surrounding rock control technology and effective roadway side sealing technology are proposed and are taken into field practice. The results showed that a safer and more efficient mining of working face can be achieved. In addition, results of this paper also have important theoretical significance and valuable reference for surrounding rock control technology of gob-side entry driving with narrow coal pillar under special geological condition.

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.

Stress distribution and surrounding rock control of mining near to the overlying coal pillar in the working face

The occurrence of overlying coal pillar(OCP)exerts a strong effect on the stress and strain distribution of the surrounding rock in the stope.In this paper,the stress distribution characteristics are analyzed via the numerical calculation with the account of OCP presence or absence.In addition,this study revealed the joint effect of side pressure relief area of the goaf and stress concentration in OCP on the final stress distribution.Furthermore,the rules of abutment stress distribution affected by three influencing factors,namely horizontal-vertical distances between OCP and working face and buried depth of OCP,are analyzed.The functional model linking the peak stress of surrounding rock with the above influencing factors is developed.The field application of the above results proved that the rib spalling and deformation of a 2.95 m-high and 5.66 m-wide roadway could be efficiently controlled by rationally adjusting working states of the support,and adopting the hydraulic prop coordinated with the p type metal beam and anchor cable to strengthen the surrounding rock of working face and roadway,respectively.The proposed measures are considered appropriate to satisfy the safe operation requirements.

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.

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.

Stability of coal pillar in gob-side entry driving under unstable overlying strata and its coupling support control technique

Considering the situation that it is difficult to control the stability of narrow coal pillar in gob-side entry driving under unstable overlying strata, the finite difference numerical simulation method was adopted to analyze the inner stress distribution and its evolution regularity, as well as the deformation characteristics of narrow coal pillar in gob-side entry driving, in the whole process from entry driving of last working face to the present working face mining. A new method of narrow coal pillar control based on the triune coupling support technique (TCST), which includes that high-strength prestressed thread steel bolt is used to strain the coal on the goaf side, and that short bolt to control the integrity of global displacement zone in coal pillar on the entry side, and that long grouting cable to fix anchor point to constrain the bed separation between global displacement zone and fixed zone, is thereby generated and applied to the field production. The result indicates that after entry excavating along the gob under unstable overlying strata, the supporting structure left on the gob side of narrow coal pillar is basically invalid to maintain the coal-pillar stability, and the large deformation of the pillar on the gob side is evident. Except for the significant dynamic pressure appearing in the coal mining of last working face and overlying strata stabilizing process, the stress variation inside the coal pillar in other stages are rather steady, however, the stress expansion is obvious and the coal pillar continues to deform. Once the gob-side entry driving is completed, a global displacement zone on the entry side appears in the shallow part of the pillar, whereas, a relatively steady fixed zone staying almost still in gob-side entry driving and present working face mining is found in the deep part of the pillar. The application of TCST can not only avoid the failure of pillar supporting structure, but exert the supporting capacity of the bolting structure left in the pillar of last sublevel entry, thus to jointly maintain the stability of coal pillar.

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.

Characteristics of stress distribution in floor strata and control of roadway stability under coal pillars

给在煤支柱下面在车道支持遇到的困难，我们在车道稳定性上学习了压力分发和他们的效果的特征，用理论分析和数字模拟。当水平应力减少时，结果证明在一个煤支柱下面，在一个地板阶层的垂直应力增加。我们断定垂直、水平的应力之间的增加的差别是为包围岩石的变丑和在煤支柱下面的车道的失败的一个重要原因。基于这，我们在一个煤支柱下面建议一条车道的包围岩石的控制技术例如高力量和高对的门栓支持，打电报加强支持，单个水力由包围岩石的 grouting 与横梁支持和加强支撑，它成功地在一个煤支柱下面在一条车道的一个稳定性控制工程被使用了。

Focal mechanism caused by fracture or burst of a coal pillar

As a regional, real-time and dynamic method, microseismic monitoring technology is quite an appropriate technology for forecasting geological hazards, such as rock bursts, mine tremors, coal and gas outbursts and can even be used to prevent or at least reduce these disasters. The study of the focal mechanisms of different seismic sources is the prerequisite and basis for forecasting rock burst by microseismic monitoring technology. Based on the analysis on the mechanism and fracture course of coal pillars where rock bursts occur mostly, the equivalent point source model of the seismicity caused by a coal pillar was created. Given the model, the seismic displacement equation of a coal pillar was analyzed and the seismic mechanism was pointed out by seismic wave theory. The course of the fracture of the coal pillar was simulated closely in the laboratory and the equivalent microseismic signals of the fractures of the coal pillar were acquired using a TDS-6 experimental system. The results show that, by the pressure and friction of a medium near the seismic source, both a compression wave and a shear wave will be emitted and shear fracture will be induced at the moment of breakage. The results can be used to provide an academic basis to forecast and prevent rock bursts or tremors in a coal pillar.

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。

宽高比对煤柱型冲击地压影响规律的实验研究

我国深部煤炭开采日趋复杂,区段煤柱在采动、构造或坚硬顶底板影响下极易诱发煤柱型冲击地压,煤柱型冲击地压的防治已成为煤矿安全高效开采的难题,研究不同宽高比条件下区段煤柱的力学性能及冲击破坏特性对煤柱型冲击地压防治具有积极意义。采用不同宽高比煤样的“岩-煤-岩”组合体进行了单轴压缩实验,通过分析煤岩组合体的冲击倾向性、动态破坏特征、分形维数和声发射特征参数等,研究了宽高比对煤柱冲击破坏的影响规律。结果发现:①煤柱的宽高比对煤岩组合体的冲击倾向性具有显著影响,宽高比不小于2∶1时,其冲击能量指数K_(E)为1.82~2.65,冲击倾向性无明显变化;小于2∶1时冲击倾向性呈先升高后降低的趋势,1∶1时K_(E)最大,达到了15.43。②随宽高比减小,煤岩组合体破坏特征依次表现为:拉压破坏—压剪破坏—拉剪破坏。煤柱宽高比为5∶1~3∶1时,煤岩组合体破坏较为缓慢;宽高比为2∶1时开始出现片状煤屑弹出,冲击破坏剧烈程度较低;宽高比为1∶1和0.75∶1时,具有明显的冲击破坏特性;宽高比为0.5∶1时,煤岩组合体整体稳定性下降,相对0.75∶1煤岩组合体煤柱破坏的剧烈程度降低。③峰后声发射能量释放率与分形维数D变化规律相似,均随宽高比减小呈先升高后降低的趋势。宽高比不小于2∶1时,煤岩组合体破坏过程中能量持续释放时间较长,煤柱破坏平缓;宽高比为1∶1时,能量释放率和D值明显增大,相较宽高比不小于2∶1煤岩组合体的能量释放率增大了约4倍,D值增大了0.18~0.23,煤柱冲击破坏最为剧烈;宽高比0.75∶1煤岩组合体能量释放率与D值分别降低了约10%和0.01,破坏剧烈程度与1∶1煤岩组合体相近;宽高比减小至0.5∶1时,相关参数的降低幅度约为0.75∶1煤岩组合体的6倍,破坏剧烈程度相对较小。研究表明:宽高比对煤柱的冲击破坏具有显著影响,整体上,煤柱冲击破坏剧烈程度随宽高比减小(5∶1~0.5∶1)呈先升高后逐渐降低的趋势;煤柱宽高比大于3∶1时,煤柱冲击危险性相对较小,煤柱宽高比为1∶1和0.75∶1时冲击危险性较大,0.5∶1次之。

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。

松软厚煤层区段煤柱剪切滑块运动机理及协同控制技术

松软厚煤层区段煤柱高、煤壁暴露面积大,加之煤质松软、裂隙发育,强采动作用下极易造成煤柱失稳,巷道维护难度极大。以山西伏岩煤业3号煤层开采为工程背景,基于剪切滑块理论,探究采掘扰动下煤柱变形破坏机理,求解煤柱剪切滑块运动范围及应力分布规律,揭示煤柱侧帮剪切滑块运动机理,提出煤柱稳定性协同控制对策并在现场进行工程实践验证。结果表明:①采用极限平衡理论与叠加理论,确定了煤柱剪切滑块运动范围及煤柱垂直应力分布规律,阐明煤柱剪切滑块安全系数分布规律:0~1.26 m深度,煤柱上部安全系数较小;在1.26~3.95 m处,煤柱中线部分大面积安全系数较小,易受顶板来压破坏。②提出了1种以“注浆加固—锚索强化—切顶卸压”为主体的区段煤柱协同控制技术,煤柱侧裂隙较无支护条件及原支护条件分别减少62.89%和46.26%,巷道围岩完整性大幅提高,形成了强承载结构,有效控制了煤柱变形及底臌。③根据松软厚煤层区段煤柱条件,合理确定了协同控制设计参数,并对煤柱防控效果进行试验监测评估。现场试验结果表明,煤柱裂隙得到充分填充,注浆后煤体强度提高63%以上;巷道位移、锚杆索受力、离层等均在可控范围,表明协同控制技术明显提高了煤柱承载力,回采巷道围岩变形得到有效控制,为工作面安全高效开采提供了空间保障。

基于岩体质量的井筒保安矿柱优化与应用

为兼顾井筒稳定和减少资源的不必要浪费,在常规的岩层移动角井筒保安矿柱圈定法基础上,采用工程类比法,将Ⅱ级岩体质量围岩部分的井筒保安矿柱尺寸优化为80m,形成上部台体下部圆柱体的保安矿柱。同时对优化后的保安矿柱周边压覆资源采取控制采场顶板暴露面积和分层充填的保护性开采措施。数值模拟结果表明:按优化后的保安矿柱范围开采周边矿体后,地表井架、卷扬机房及井筒内壁变形量均未超过最大允许变形值,说明优化方案具有可行性。

变区段煤柱工作面诱冲机理及防治技术研究

针对某矿变区段煤柱工作面开采冲击危险性较高的问题,运用数值模拟方法对工作面采掘过程中煤柱区域垂直应力分布规律进行研究,分析了不同宽度区段煤柱应力积聚特征,揭示了变区段煤柱工作面诱冲机理,并制定了“近场-远场”协同卸压方案。研究结果表明,较宽的区段煤柱(30 m和55 m)受采空区侧向支承压力和超前支承压力影响,煤柱应力集中程度较高;受“近场高静载+远场动载”叠加影响,变区段煤柱工作面回采期间易诱发冲击地压;采取近场高承压煤体强卸压+远场高位坚硬顶板超前预裂断顶卸压协同控制方案后,微震能量事件始终维持在104J以下,表明控制方案能够有效降低工作面采掘期间的冲击危险性。

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。

特厚煤层沿空掘巷围岩支卸协同控制技术

针对特厚煤层小煤柱沿空掘巷大范围塑性全煤巷道锚网索的支护难题,以塔山煤矿8204-2小煤柱工作面为工程背景,采用现场实测、理论分析等方法,揭示了锚杆、锚索对沿空大范围塑性全煤巷道的支护机理。锚杆支护主要作用于浅部二次破碎区煤体,形成浅部连续承载结构;锚索主要作用于深部处于三向受力状态的稳定煤体,形成更大范围的连续稳定承载结构;进一步明确了锚杆、锚索有效长度的计算方法,形成了以高预紧力、高强“锚-网-索”支护为基础,以坚硬顶板井下磨料水射流切顶卸压、巷帮大直径钻孔卸压、底板卸压槽卸压为辅的“支卸协同”小煤柱沿空巷道围岩控制技术体系。工程应用结果表明,试验巷道围岩最大变形量小于700 mm,能够满足使用要求。

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。