Numerical and theoretical study of large-scale failure of strata overlying sublevel caving mines with steeply dipping discontinuities

The deformation and fracture evolution mechanisms of the strata overlying mines mined using sublevel caving were studied via numerical simulations.Moreover,an expression for the normal force acting on the side face of a steeply dipping superimposed cantilever beam in the surrounding rock was deduced based on limit equilibrium theory.The results show the following:(1)surface displacement above metal mines with steeply dipping discontinuities shows significant step characteristics,and(2)the behavior of the strata as they fail exhibits superimposition characteristics.Generally,failure first occurs in certain superimposed strata slightly far from the goaf.Subsequently,with the constant downward excavation of the orebody,the superimposed strata become damaged both upwards away from and downwards toward the goaf.This process continues until the deep part of the steeply dipping superimposed strata forms a large-scale deep fracture plane that connects with the goaf.The deep fracture plane generally makes an angle of 12°-20°with the normal to the steeply dipping discontinuities.The effect of the constant outward transfer of strata movement due to the constant outward failure of the superimposed strata in the metal mines with steeply dipping discontinuities causes the scope of the strata movement in these mines to be larger than expected.The strata in the metal mines with steeply dipping discontinuities mainly show flexural toppling failure.However,the steeply dipping structural strata near the goaf mainly exhibit shear slipping failure,in which case the mechanical model used to describe them can be simplified by treating them as steeply dipping superimposed cantilever beams.By taking the steeply dipping superimposed cantilever beam that first experiences failure as the key stratum,the failure scope of the strata(and criteria for the stability of metal mines with steeply dipping discontinuities mined using sublevel caving)can be obtained via iterative computations from the key stratum,moving downward toward and upwards away from the goaf.

Stability analysis of subgrade cave roofs in karst region

According to the engineering features of subgrade cave roof in karst region, the clamped beam model of subgrade cave roof in karst region was set up. Based on the catastrophe theory, the cusp catastrophe model for bearing capacity of subgrade cave roof and safe thickness of subgrade cave roof in karst region was established. The necessary instability conditions of subgrade cave roof were deduced, and then the methods to determine safe thickness of cave roofs under piles and bearing capacity of subgrade cave roof were proposed. At the same time, a practical engineering project was applied to verifying this method, which has been proved successfu1ly. At last, the major factors that affect the stability on cave roof under pile in karst region were deeply discussed and some results in quality were acquired.

Failure of hanging roofs in sublevel caving by shock collision and stress superposition

Hanging roofs or high hang-ups.a common problem in sublevel caving mining,usually result in a large ore loss and undermine mining safety.This paper analyzed the formation of a hanging roof and showed that increased confining pressure and reduced free surface were its main characteristics.In order to break down a hanging roof,a new method based on shock wave collision and stress superposition was developed.In this method,two blastholes containing multi-primer at different positions are simultaneously initiated at first.By doing this,a new free surface and a swell room can be created.After these holes are fired,a long delay time is given to the next blasthole so that the fragments from the first twohole blasting have enough time to fall down.This new method was applied to three hanging roofs in one production area,and all of them were successfully broken down.Field inspection indicated that almost no damage was caused in the nearby drifts/tunnels due to the new method.In addition,the far field vibrations were found to be smaller than the maximum vibrations induced by some other blasts.

Numerical analysis of application for induction caving roof

New method for handling roof of the base successive mining is proposed, which is induction caving in the roof. The key is that it is made certain to the station of the space-time in the induction caving roof, as the stress is released with the mining process. And applying the catastrophe theory, the influencing factors of induction caving roof are studied in the emptied areas, such as the mechanical property of the surrounding rock, the area of the gob,the scope and dimension of tensile stress. The results show that the key factor is the area of the gob to the method of the induction caving roof. Then according to the geology and the ore characteristic, the three dimension FEM mechanical model is built in Tongkeng Mine, the laws of the tensile stress are analyzed to the space and the time in the roof with the mining, then it is rational design to the mine step and time of the handing the roof.

Numerical simulation of roof cavings in several Kuzbass mines using finite-difference continuum damage mechanics approach

An essential stage of mine design is an estimation of the steps of the first and periodic roof caving in longwall mines.Generally,this is carried out using the field experience and can be much enhanced by numerical simulation.In this work,the finite-difference method was applied coupled with the continuum damage mechanics(CDM)approach to simulate the stress-strain evolution of the rock mass with the underground opening during coal extraction.The steps and stages of roof caving were estimated relying on the numerical simulation data,and they were compared with the field data from several operating mines in the south of the Kuznetsk Basin,Russia.The dependence of the first roof caving step in simulation linearly correlates with field data.The results correspond to the actual roofs of longwall panels of the flat-dipping coal seams and the average rate of face advancement is approximately 5 m/day.

A new method determining safe thickness of karst cave roof under pile tip

Cave roofs are used to support pile foundation in many engineering projects. Accurate stability analysis method of cave roof under pile tip is important in order to ensure the safety of the pile foundation structure. Firstly the mechanical model to analysis the stability of cave roof under pile tip is founded aiming to solve the problems that the simplified mechanical model has. Secondly, the boundary of cave roof is simply supposed to be supported according to the integrity of the rock mass in the boundary of cave roof. Thirdly, based on the theory of plates and shells, the simplified model is calculated and the theoretical calculation formula to determine the safe thickness of cave roof under pile tip can be obtained when the edges of the cave roof are simply supported. In the end, the analysis of the practical engineering project proves the feasibility and the rationality of the method which can be a new method to calculate the safe thickness of cave roof under pile tip.

Similar material simulation of time series system for induced caving of roof in continuous mining under backfill

With the help of similar material simulation test,time series system for induced caving of roof in continuous mining under complex backfill in ore body No.92 of Tongkeng Tin Mine was studied. According to the similarity theory,a two-dimensional similar simulation test-bed was constructed. The stress and displacement that change along with the advance of mining were acquired and analyzed automatically by data system. The processes of continuous mining of ore-block in 5 intervals and artificial induced caving of roof were simulated. The results of the test show that ore body remained as safety roof in thickness of 15 m guarantees the safe advance of stoping work face. Caving of safety roof puts in practice at the first two mining intervals when the third interval of continuous mining is finished,and one interval as the safety distance should be kept all the time between stopping and caving. While mining in the last interval,pre-slotting should be implemented first of all,and the roof of the last two mining intervals is caved simultaneously. Only this kind of time series system can be an efficient and safe way for induced caving of roof in continuous mining.

Similar material simulation research on movement law of roof over-lying strata in stope of fully mechanized caving face with large mining height

Similar material simulation test was carried out in a geological model ofW_(9-15)101 fully mechanized caving face with large mining height in the Liuhuanggou Colliery,in Xinjiang Uigur Autonomous Region.The roof overlying strata movement law in thestope of a fully mechanized caving face with large mining height was studied and showthat the roof overlying strata in the stope of a fully mechanized caving face with large miningheight can be formed into a stable arch structure; the fracture rock beam is formed resemblinga 'bond beam', but it has essentially the structure of 'multi-span beams' underthe big structure of the stable arch.The roof overlying strata movement law in the stope ofa fully mechanized caving face with large mining height is similar to that of the common,fully mechanized caving stope, which is determined by the deformation and instability ofthe structure of 'multi-span beams'.But because of the differences between the miningheights, the peak pressure in the stope of a fully mechanized caving face with large miningheight is smaller while the affected area of abutment pressure is wider in the front of theworking face; this is the obvious difference in abutment pressure between the stope of afully mechanized caving face with large mining height and that of the common.

STUDIES ON THE LAW OF ROOF-COAL MOVEMENT BY USING THE ROOF-COAL CAVING METHOD

In this paper, the law of roof-coal movement has been investigated through the field measurement, theoretical analysis and numerical calculation. Several results, which are of important values for caving process, design of the supports, controlling end-face stability, raising recovery rate, realizing working face high output and other related aspects in practice, have been obtained. These results mainly include the following: roof-coal breaking curve of soft-coal seam, roof-coal movement curve of soft-coal and medium-hard coal seam, and roof-coal movement equation. The roof-coal caveability has been analyzed.

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。

高瓦斯综放面开切眼架前膨胀预裂坚硬顶板切顶技术研究

针对高瓦斯坚硬顶板综放工作面初采期来压步距大及顶煤损失的问题,提出代替火药爆破的开切眼架前及两顺槽煤柱侧静态膨胀预裂切顶技术;采用数值模拟方法确定了静态膨胀预裂切顶的关键参数,设计了切顶方案及工艺,并通过现场工程应用验证了静态膨胀预裂切顶技术的有效性。结果表明:综放工作面初采前在开切眼架前及两顺槽煤柱侧实施静态膨胀预裂切顶,既保证了开切眼设备安装期间的安全性,又可在后续采动应力的作用下顶板及时破断,减小坚硬顶板初次来压步距和强度;当综放工作面初采推进至开切眼宽度时发生初次来压,人工减跨效果良好,不仅提前回收了工作面的顶煤资源,而且还消除了坚硬顶板大面积悬顶带来的诸多重大安全隐患,研究成果可在类似条件下代替炸药爆破进行推广应用。

桥梁荷载作用下巨型岩溶洞穴顶板稳定性研究

某高速公路桥墩基础完全置于巨型溶腔上部,通过定性、半定量、定量的方法对岩溶洞穴顶板稳定性进行综合分析评价,确定了桥型方案,然后通过地表变形监测,验证稳定性综合评价方法的合理性。

采空区顶板垮落摩擦诱发瓦斯爆炸试验研究

大采高工作面采空区顶板垮落时岩体相互撞击摩擦产生火花已成为诱发采空区瓦斯爆炸事故的主要隐患。为研究砂岩摩擦速度、摩擦力、瓦斯体积分数、SiO_(2)含量等因素对砂岩摩擦诱发瓦斯爆炸的机制,采用岩石摩擦效应引燃瓦斯试验系统对各因素进行定量试验研究。结果表明:瓦斯体积分数为7.0%~11.5%时,砂岩摩擦诱发瓦斯爆炸时间随瓦斯体积分数增加先缩短后延长;砂岩摩擦高温热表面是诱发瓦斯爆炸的主要原因;摩擦诱发瓦斯爆炸分为砂岩摩擦面燃烧瓦斯与瓦斯爆炸扩散2个阶段;摩擦诱发瓦斯爆炸时间随摩擦力增大而缩短,随砂岩中SiO_(2)含量增大而缩短;摩擦低于最低临界速度不会诱发瓦斯爆炸,高于最低临界速度且低于最高临界速度时最大摩擦速度与摩擦诱发瓦斯爆炸时间呈负相关。

南阳煤业综放工作面坚硬顶板控制技术研究

针对南阳煤业3#煤层顶板岩层厚度大、硬度大不易垮落的问题,提出在开切眼和回采巷道同时布置爆破孔,对顶板岩层进行爆破弱化处理。采用理论分析、数值模拟的方法确定了爆破方案及爆破孔参数,并且在南阳煤业3203综放工作面进行了现场试验。结果发现,对顶板进行爆破弱化处理后,顶板岩层垮落效果良好,顶煤冒放性明显提升,顶煤回收率可达86%。

厚硬顶板刀把形采空区下综放工作面冲击地压机制研究

不规则的开采空间结构可增大煤矿冲击地压危险,研究多煤层复杂开采空间结构下工作面冲击机制对煤炭资源安全采出具有重要意义。综合采用理论分析、数值模拟、现场监测等方法,研究了某矿井W1123综放工作面的动静载分布特征,分析了冲击显现机制,提出了针对性的防冲措施并进行现场应用。结果表明,上覆刀把形采空区和厚硬顶板综合作用下W1123工作面动静载分布具有区域性特征,当工作面仅受到刀把形采空区的“刀柄”作用或同时受“刀柄”和“刀身”作用时,工作面中下部和区段煤柱静载集中、动载扰动剧烈,工作面回风巷围岩也存在动静载集中,且距离刀把形采空边界越近则围岩动静载越高;仅受“刀身”作用时,工作面回风巷围岩静载集中受动载扰动。工作面静载集中区和动载扰动区重合,动静载叠加造成冲击显现。依据刀把形采空区结构特征,W1123工作面回采期间冲击类型依次为强动载型、高静载强动载型、高静载型。数值模拟和现场监测揭示的冲击危险区与理论分析结果一致。针对W1123工作面静动载分布特征,优化设计了冲击地压防治措施,增大了厚硬顶板处理高度和破碎程度、增加了工作面运输巷煤体爆破和煤柱切顶以降低顶板和煤体应力和能量集中;优化措施实施后,工作面微震事件日均能量总体下降54%,大能量事件大幅减少,冲击地压危险显著降低。该研究结果可为相似条件矿井的冲击地压防治提供借鉴。

综采工作面坚硬顶板深孔预裂爆破技术应用研究

井工煤矿进行煤炭开采时,采煤工作面的矿压显现主要受工作面上方顶板的影响,特别是坚硬顶板,影响更大。坚硬顶板造成的综采工作面采空区顶板悬顶面积过大对综采工作面的矿压显现及安全生产有很大影响。回采过程中,坚硬顶板不能随着工作面向前推进及时垮落,会形成大面积悬顶,如果大面积顶板突然垮落,不仅会砸毁设备,破坏工作面,还会产生强烈的冲击波,造成人员伤亡,形成矿难。为了防治井下采煤时坚硬顶板突然垮落造成的安全隐患,塬林煤矿根据W1-108工作面实际情况,决定采用深孔预裂爆破强制放顶技术。

矿井综放工作面安全高效回采技术研究

针对山西某矿13102综放工作面在进行放顶煤作业时,在初采阶段存在悬顶面积大,在回采阶段存在顶煤放煤效果差及在推采作业中易发生推采上窜下移等不足,为有效解决上述问题,基于13102综放工作面工程概况,提出了可通过进行水力压裂切眼顶板,对推采放煤工艺进行改进优化等措施来提升放煤效果,经综合应用这些措施后,13102综放工作面很少再出现大面积悬顶现象,顶煤垮落也更及时、更均匀,放煤头尾安全出口也更畅通,实现了安全高效放煤,取得了较好的放煤效果,可为类似工况下的综放工作面安全高效开采提供参考。

龙岩市非煤矿山顶板事故分析与管理措施

顶板事故在非煤地下矿山事故中占比最多,顶板管理是非煤地下矿山安全管理的重点和难点。统计分析龙岩市2008—2023年共15a的顶板事故情况,总结非煤矿山顶板事故的特点。研究得出,矿井爆破工艺技术水平低及在大暴露空间下进行作业是顶板事故原因。提出改善措施,将中深孔爆破分段崩落采矿法代替空场法,但受到矿石价格影响,该方法成本高,当前可行性低;通过推广应用凿岩台车、撬毛台车等机械工具,严格控制井下采掘作业空间宽度和高度等措施,加强顶板管理,努力实现顶板零事故。

综放开采回采巷道切顶护巷技术研究

厚煤层综放开采时由于一次开采厚度大,导致开采引起的采动压力影响范围、显现程度等均较为明显,给邻近巷道围岩控制工作带来一定影响。2301综放工作面运输巷与邻近的2303回风巷间留设25 m保护煤柱,受2301综放工作面开采影响导致2303回风巷变形量较大,难以满足后续使用需要。结合现场条件及以往研究成果,提出在2301运输巷靠近煤柱侧施工切顶钻孔,通过切顶较少2301综放工作面对2303回风巷影响,并对切顶布置方案、乳化炸药布置以及封孔方式等进行详细设计。现场应用后,2303回风巷围岩变形量受2301综放工作面开采影响较小,回风巷围岩变形量在允许范围内,可满足后续使用需要。

工作面停采煤柱宽度的确定与实践

燕子山矿采用宽煤柱方式护巷,煤柱宽度约60 m,造成资源的严重浪费。通过对回风巷的受力变化规律进行监测分析,得出工作面回采对巷道压力影响的临界距离为35 m。以此为依据,在综合考虑8210工作面特殊性的基础上,分析得出该工作面最佳停采煤柱宽度为38~45 m。为检验8210工作面煤柱宽度,该矿在回风大巷开展了顶板钻孔窥视工作,检测顶板煤层回采期间和停采后的变化情况,窥视检测结果表明,工作面留设38~45 m煤柱宽度能够保证巷道围岩的稳定性,是安全可靠的。