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.

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.

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.

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.

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

Given the difficulties encountered in roadway support under coal pillars,we studied the characteristics of stress distribution and their effect on roadway stability,using theoretical analysis and numerical simulation.The results show that,under a coal pillar,vertical stress in a floor stratum increases while horizontal stress decreases.We conclude that the increased difference between vertical and horizontal stress is an important reason for deformation of the surrounding rock and failures of roadways under coal pillars.Based on this,we propose control technologies of the surrounding rock of a roadway under a coal pillar,such as high strength and high pre-stressed bolt support,cable reinforcement support single hydraulic prop with beam support and reinforcement by grouting of the surrounding rock,which have been successfully applied in a stability control project of a roadway under a coal pillar.

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.

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.

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.

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。

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。

大巷煤柱工作面过空巷矿压规律及控制技术

【目的】由于矿井工作面布置方式的调整,煤柱工作面经常会面临通过废弃巷道时覆岩顶板难以控制的情况。【方法】为解决这一问题,以河南赵固二矿二盘区外侧煤柱工作面过空巷为工程背景,采用理论分析、数值模拟和现场试验等方法,研究工作面与空巷覆岩破断组合结构,模拟不同支护强度下顶板应力-位移全周期演化规律,分析工作面矿压显现特征,提出相应的控制技术。【结果和结论】结果表明,基本顶不同破断形式对矿压显现特征影响显著,关键块断裂位置可分为煤柱上方、空巷上方和实体煤上方3种类型。通过建立工作面过空巷力学模型,研究基本顶超前破断力学机理,基本顶受到空巷-煤柱-工作面支护系统支撑作用,形成“砌体梁”稳定承载结构,判定基本顶滑落失稳时空巷支护强度的临界值为4.6MPa。数值模拟显示,工作面超前支承压力与空巷应力集中产生的叠加效应对煤柱影响显著,当工作面推进至距空巷5m时,煤柱失稳破坏,基本顶易发生超前破断。在工作面过空巷过程中,煤柱超前支承压力分布特征由“双峰型”转变为“孤峰型”。不同支护强度下的顶板应力分布特征存在明显差异,确定空巷支护强度为4.5MPa能够防止基本顶超前破断。最后,在研究区二盘区外侧煤柱工作面采用“锚网索”支护方式对空巷顶板进行补强支护,过空巷期间液压支架工作阻力在研究区域处于安全范围内,未发生顶板垮落和压架等事故,解决了二盘区外侧煤柱工作面过空巷技术难题,可为类似工作面提供参考依据。

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。

《“三下”开采规范》中安全煤(岩)柱留设问题探讨

针对《建筑物、水体、铁路及主要井巷煤柱留设与压煤开采规范》中安全煤(岩)柱设计以及水压对顶板突水控制作用等问题进行探讨。首先分析了规范的附表4-3表头文字,指出保护层厚度的有关规定仅适用于松散含水层(含地表水体),至于基岩含水层下采煤应如何确定保护层厚度则未明确;进而以泥岩与黏性土层均具有阻水功能为桥梁,推导出基岩含水层下采煤可参照“松散层下黏性土层累计厚度大于采厚”条件执行,遵照“就高不就低原则”,基岩含水层下保护层厚度统一取采厚的4倍为宜;既然规范适用条件为单层采厚不大于3.0 m,则附表4-3中“松散层厚度小于采厚”的规定难以理解,建议删除。其次,基于“保护”一词的科学内涵对“保护层”重新定义,即导水裂缝带顶界面到含水层底界面之间的隔水岩层均具有“保护”功能,应统称为保护层H_(b);进而提出了保护系数B_(s)概念,即保护层厚度与单层采厚的比值;松散含水层下保护系数分区阈值(B_(i)=2、3、4、5、6、7),基岩含层下保护系数分区阈值B_(i)=4,据此对顶板水害风险进行等级划分:突水区(B_(s)≤0)、危险区(0<B_(s)<B_(i))、安全区(B_(s)≥B_(i))。此外,借鉴底板突水系数概念,将单位厚度保护层承受的水头压力称为保护层承压系数(T=P/H_(b)),通过对3种煤水结构条件下承压系数的分析,得出第4系松散含水层、非煤系基岩含水层下采煤可以不考虑水压,煤系基岩含水层下采煤随着开采深度增加、水压随之增大而带来突(涌)水风险。最后,分析了底板含水层顶部存在被泥质物充填隔水带时,《“三下”开采规范》给出的底板防水安全煤(岩)柱表达式(h_(a)≥h_(1)+h_(2)+h_(4))与附图相矛盾,正确的表达式应为h_(a)≥h_(1)+h_(2)-h_(4)。

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。

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。