Instability characteristics of the cracked roof rock beam under shallow mining conditions

To establish the movement relationship for the roof breaking under shallow mining conditions, the mechanical model of the roof rock beam was built, then the structure instability process of the roof rock beam was analyzed. The changing criterion of the vertical displacement was established and the relationship between the deflection and the rotary motion of roof block was determined. Regarding a mining face in Shangwan Mine, the responsing laws of the deflection and horizontal thrust of the roof rock beam were obtained through FLAC3D numerical analysis. The results show that the structure instability of the cracked roof rock beam depends on the interaction between the vertical load and the horizontal thrust.For the roof rock beam, when the vertical load keeps constant, the horizontal thrust fluctuating rises with increasing deflection. The horizontal thrust increases constantly with the deeper buried depth and the smaller span.

Experimental and numerical simulation on the sliding process of roof rock blocks triggered by dynamic disturbance

With the excavation of underground opening,the roof rock block may fall under external dynamic disturbance.In this paper,the roof rock mass was simplified as a rock block system composed of trapezoidal rock blocks.A series of experiments and numerical simulations were performed to study the sliding process of the key block under the horizontal static clamping load and vertical impact disturbance.The propagation of stress wave in the block system were captured and analyzed by using high-speed camera and digital image correlation technique.The results reveal that a pendulum-type wave was generated due to the propagation and superposition of stress waves in the block system.Therefore,the governing mechanism of the sliding displacement of the key block was clarified based on the propagation of incident stress waves or pendulum-type waves.Meanwhile,it is found that the sliding distance of the key block decreases in a power function with the increasing friction coefficient,or decreases in a parabolic function with the increasing trapezoid internal angle.Finally,a case study on the roof block sliding of the roadway at a gold mine was conducted,and it is concluded that the sliding of the key block resulted from the coupled effects of"shear driving"and"low friction"driven by stress wave propagation,regardless of a single or multi-layer rock block system.These results may provide technical guideline for preventing rock-falling accidents caused by blasting distur-bances in underground mining.

Numerical simulation study on hard-thick roof inducing rock burst in coal mine

In order to reveal the dynamic process of hard-thick roof inducing rock burst, one of the most common and strongest dynamic disasters in coal mine, the numerical simulation is conducted to study the dynamic loading effect of roof vibration on roadway surrounding rocks as well as the impact on stability. The results show that, on one hand, hard-thick roof will result in high stress concentration on mining surrounding rocks; on the other hand, the breaking of hard-thick roof will lead to mining seismicity, causing dynamic loading effect on coal and rock mass. High stress concentration and dynamic loading combination reaches to the mechanical conditions for the occurrence of rock burst, which will induce rock burst. The mining induced seismic events occurring in the roof breaking act on the mining surrounding rocks in the form of stress wave. The stress wave then has a reflection on the free surface of roadway and the tensile stress will be generated around the free surface. Horizontal vibration of roadway surrounding particles will cause instant changes of horizontal stress of roadway surrounding rocks; the horizontal displacement is directly related to the horizontal stress but is not significantly correlated with the vertical stress; the increase of horizontal stress of roadway near surface surrounding rocks and the release of elastic deformation energy of deep surrounding coal and rock mass are immanent causes that lead to the impact instability of roadway surrounding rocks. The most significant measures for rock burst prevention are controlling of horizontal stress and vibration strength.

Rock characterization while drilling and application of roof bolter drilling data for evaluation of ground conditions

Despite recent advances in mine health and safety, roof collapse and instabilities are still the leading causes of injury and fatality in underground mining operations. Improving safety and optimum design of ground support requires good and reliable ground characterization. While many geophysical methods have been developed for ground characterizations, their accuracy is insufficient for customized ground support design of underground workings. The actual measurements on the samples of the roof and wall strata from the exploration boring are reliable but the related holes are far apart, thus unsuitable for design purposes. The best source of information could be the geological back mapping of the roof and walls, but this is disruptive to mining operations, and provided information is only from rock surface.Interpretation of the data obtained from roof bolt drilling can offer a good and reliable source of information that can be used for ground characterization and ground support design and evaluations. This paper offers a brief review of the mine roof characterization methods, followed by introduction and discussion of the roof characterization methods by instrumented roof bolters. A brief overview of the results of the preliminary study and initial testing on an instrumented drill and summary of the suggested improvements are also discussed.

Rock burst induced by roof breakage and its prevention

Based on the research on rock burst phenomenon induced by the breakage of thick and hard roof around roadways and working faces in coal mines,a criterion of rock burst induced by roof breakage(RBRB) was proposed and the model was built.Through the model.a method calculating the varied stresses induced by roof breakage in support objects and coal body was proposed and a unified formula was derived for the calculation of stress increment on support objects and coal body under different breaking forms of roof.Whilst the formula for calculating dynamic load was derived by introducing dynamic index Kd.The formula was verified in Huating Mine by stress measurement.According to the formula for stress increment calculating,the sensitivities of dynamic load parameters were further studied.The results show that the thickness and breaking depth of roof,width of support object are the sensitive factors.Based on the discussion of the model,six associated effective methods for rock burst prevention are obtained.

Numerical study of rock bolting parameter variation effect on stability of stratified, composite roof structure of drift in a coal mine

Described the outcomes of a comprehensive numerical modeling on the rock bolting performance for preventing the deformation of stratified, composite roof structure of drift in a coal mine. The investigation was undertaken in the adverse geological condi- tions, with variation of bolt parameters, including length, density, distribution, pretension, as well as the geometry of opening,so as to determine the effect of bolting parameter variation on roof deformation and stability. The outcomes clearly demonstrated that a sig- nificant improvement of roof stability can be achieved associating with bolting parameters optimization, and indicated the importance of flexible geotechnical designation of rock bolting reinforcement in mining practice.

Comprehensive assessment on dynamic roof instability under fractured rock mass conditions in the excavation disturbed zone

The damage process of fractured rock mass showed that the fracture in rocks induced roof collapse in Yangchangwan Coal Mine, China. The rock mass was particularly weak and fractured. There occurred 6 large-scale dynamical roof falls in the excavation disturbed zone （EDZ） with the collapsing volume of 216 m^3. First, the field detailed geological environment, regional seismic dynamics, and dynamic instability of roadways were generally investigated. Second, the field multiple-index monitoring measurements for detecting the deep delamination of the roof, convergence deformation, bolt-cable load, acoustic emission （AE） characteristic parameters, total AE events, AE energy-releasing rate, rock mass fracture, and damage were arranged. Finally, according to the time-space-strength relations, a quantitative assessment of the influence of rock-mass damage on the dynamic roof instability was accomplished.

Physical Modeling of Influence of Rock Mass Structure on Roof Stability

Based on the developing degree of structure planes in coal roof, whole, blocky and heavily fractured structure models are built up. Through simulation test of similar materials, the distribution of deformation, failure and underground pressure induced by coal mining in coal roof with different rock mass structures are analyzed. The test results indicate that the distances of first and periodic weighting of main roof and the height of caving and fracture zone decrease with the increment of fractures in roof rock mass. From whole to blocky and heavily fractured structures, abutment pressure ahead of working face reduces and the peak value of abutment pressure migrates to inside of roof rock mass.

深部沿空巷道锚固围岩破坏失稳能量驱动机理

深部沿空巷道受侧向顶板断裂所产生的动载影响,锚固围岩易产生大变形,甚至破坏失稳。以山东省孙村煤矿31120工作面上平巷为工程背景,首先采用YTJ20型岩层探测记录仪获得了锚固顶板裂隙发育规律及以脆性张裂破坏为主的破坏方式。然后,通过相似材料模拟试验方法获得了深部沿空巷道侧向顶板前期、过渡期和后期3个运动阶段典型特征,并分析了不同阶段锚固围岩破裂演化及能量释放规律;其中过渡期运动阶段锚固围岩内部应力、变形量急剧增大,裂隙发育明显,能量释放显著,对沿空巷道锚固围岩稳定性影响最大。最后,构建了侧向顶板断裂运动下沿空巷道结构力学模型,给出了锚固围岩输入能量与可抵抗能量定量计算方法,揭示了侧向顶板断裂诱发锚固围岩破坏失稳能量驱动机理,并定义了失稳能量判据,即当作用在沿空巷道锚固围岩上的能量大于锚固围岩可抵抗能量时,将发生破坏失稳。进一步地,提出了锚固围岩失稳风险等级划分方法和相应地控制技术,计算结果表明,31120工作面沿空巷道锚固围岩失稳风险等级为中风险。采取加强支护措施后,沿空巷道顶底板及两帮移近量分别减小35.47%和35.71%,锚索受力减小23.43%,变形速度明显降低,锚固围岩能量积聚程度减小。

底板动压巷道压裂弱结构体应力转移控制技术

为了满足煤矿安全生产的需要,许多巷道都会布置在煤层底板中,如部分运输大巷、排水巷道、瓦斯抽采巷道等。采动应力容易造成底板巷道围岩应力升高,加剧底板巷道围岩变形,造成支护永久失效、顶板下沉、巷道底鼓、两帮收敛等破坏。针对此,提出了在应力传递路径上实施水力压裂,在指定的区域制造出一定空间形态的水压裂缝网,形成压裂弱结构体,实现区域范围内的应力转移,从而降低巷道区域范围内的应力,控制巷道的围岩稳定性的控制技术,并通过理论分析及现场工程验证等方式,揭示了底板动压巷道压裂弱结构体应力转移的力学机制,建立了相应的力学模型,对压裂弱结构体的合理位置、范围等影响因素进行了求解。得出:(1)压裂弱结构体使局部应力场发生明显变化,出现应力升高区和应力降低区,应力降低区主要分布在弱结构体与采动应力连线的方向上,主要集中在一个拱形的范围内;由于膨胀效应,在与应力来源垂直的方向上产生应力集中,出现应力升高区。(2)最大主应力变化幅度与压裂弱结构体的长轴长L、短轴长H、到巷道的距离P、与巷道连线的水平夹角β、压裂层的强度C及内摩擦角α、压裂的损伤变量D等有关。其中到巷道的距离P对卸压效果影响最大,损伤变量D对卸压效果影响最小。(3)采用提出的计算方法设计了淮北矿业集团袁店一矿的103运输集中巷的卸压方案,工程应用结果表明,底板动压巷道变形速率明显减缓,验证了底板强动压巷道压裂弱结构体应力转移模型的合理性。

矿震与冲击地压防治研究进展

矿震与冲击地压是采矿领域亟待解决的热点、难点和瓶颈问题,为了有效控制矿震与冲击地压,综述矿震定义与分类、矿压假说及矿震与冲击地压防治技术,回顾切顶卸压理论及其在中厚金矿采空区处理与卸压开采、中厚及以下煤矿沿空留巷中的应用,并比较其与110工法的差异。据此,给出矿震定义,指出矿震与冲击地压发生的条件类似,针对厚矿体开采提出深埋坚硬顶板控制爆破切槽放顶技术。研究表明:控制爆破切槽放顶技术仍然是未来矿震与冲击地压防治中释放并转移高地压的主要方法,它还可将深埋厚矿体的“砌体梁”简化成弹簧岩梁承载体系。

巷帮煤体整体滑脱型冲击地压锚杆防冲支护原理及工程实践

冲击地压是目前严重影响煤炭安全有效开采的灾害之一,研究锚杆防冲支护原理和技术对防治巷道冲击地压灾害具有重要意义和价值。通过对巷帮煤体整体冲入型冲击地压发生的地质条件以及破坏特征进行总结分析,认为坚硬顶板与坚硬煤层是此类型冲击地压的重要地质特征,而巷帮煤体整体滑脱是其主要冲击破坏特征。在此基础上,以巷帮滑脱煤体为研究对象,建立了顶板-巷道-底板复合结构体力学模型,建立了巷帮煤体发生水平滑移的极限平衡方程,并对各个参数进行分析。结果表明:由于顶板反弹使巷帮煤体竖直方向压力降低,巷帮煤体被构造应力推入巷道发生冲击地压。基于该发生机制模型,认为目前的锚杆支护设计体系在防治巷帮煤体冲入型冲击地压存在不足,并基于其发生和破坏特征,建立了针对巷帮煤体整体滑脱型冲击地压的锚杆防冲支护设计原则,即将顶和底帮锚杆锚固端分别穿层打入稳定的顶底板内,并使用长锚索取代中部帮锚杆,提供锚杆支护的防冲作用。基于新建立的锚杆防冲支护设计方法,以大屯矿区孔庄煤矿7305工作面防冲支护为工程背景,在宽煤柱段巷帮锚杆支护采取了防冲设计,帮顶、底部锚杆及补强锚索均锚固于顶底板内部,能够有效吸收煤体滑动动能,提高安全性。

层状复合夹煤岩石组合体试样强度试验与破坏特征研究

针对巷道复合夹煤组合结构顶板易冒落问题,以某矿取芯和钻孔窥视探测获得的巷道层状复合夹煤顶板结构组合类型为实验背景,对砂质泥岩、中砂岩以及煤厚度为20、30、40 mm不同结构组合类型的复合试件进行了单轴压缩试验,研究了复合夹煤岩石组合体的力学特性与破坏特征。试验结果表明:复合夹煤岩石组合体的抗压强度均小于单一岩石,中砂岩夹煤厚度为30mm时抗压强度达到单一中砂岩抗压强度的85.95%,抗压强度随着煤厚变薄而下降;砂质泥岩夹煤厚度为20 mm时仅为单一砂质泥岩抗压强度的32.5%,随着夹煤厚度增加抗压强度也增大,夹煤厚度超过30mm后抗压强度增大不明显。复合夹煤岩石组合体整体呈X状共轭斜面剪切破坏、单斜面剪切破坏、部分拉伸破坏,随着复合夹煤岩石组合类型不同呈现不同的破坏特征等。

冲击地压矿井充填工作面超前采动应力对充填体充实率的反馈机制

高应力是深部矿井冲击地压灾变频率走高的主要原因,充填开采则是控制岩层运动,缓解采动应力集中程度,降低围岩破坏和冲击风险的有效手段。为研究采空区充填体对超前采动应力的控制能力,以山东古城煤矿1123充填工作面为工程背景,采用理论分析、室内试验、现场实测手段探究充填开采工作面采动应力分布规律,揭示超前采动应力对充填体充实率的反馈机制,指导冲击地压矿井充实率确定。结果表明:工作面开采初期充填体充实率低于80%,坚硬顶板下沉量大,超前采动影响范围大于30 m,应力集中系数达到1.5,断层影响区采动应力影响范围和集中系数分别增至60 m和1.65,片帮冒顶等围岩失稳现象增多;实测了采空区充填体承载应力全程动态演化特征,承载应力分布曲线划分为“快速降低—短暂稳定—快速升高—缓慢降低—二次稳定”5个阶段,低充实率条件下采空区上、中、下3个区域承载应力稳定值分别为1.9、5.2、2.8 MPa;将采空区充填体划分为非充分压实区和充分压实区,构建了充填体支撑作用下坚硬顶板连续沉降模型,得到了坚硬顶板“ʅ”型沉降曲线,非充分压实区范围随充实率近似呈线性减小;试验得到充填体弹性模量和单轴抗压强度随凝固时间的演化曲线,结合顶板沉降曲线和推进速度得到工作面前后采动应力全区域分布曲线;建立了超前采动应力集中程度与充填体承载能力的负指数函数关系,揭示了超前采动应力对充填体充实率的负向反馈机制,实现充填开采降载防冲效果的定量评价;提出了充填体充实率“三位一体”协同提升措施,将1123工作面充实率升高至90%,增强了充填体承载能力,超前采动应力集中系数降至1.3,厚顶煤膨胀变形量减少至50 mm,坚硬顶板破断致冲风险显著降低。

工作面单双巷顶板爆破应力演化规律研究

坚硬顶板是影响工作面冲击危险的重要因素,实践中常采用爆破技术对工作面顶板进行预裂和卸压。以田陈煤矿3煤层上方存在坚硬顶板的7202工作面为对象,采用理论分析、数值模拟、工业试验等方法,研究了不同爆破高度、工作面单巷和双巷顶板爆破下围岩应力演化及微震规律。结果表明:针对低位亚关键层采用单巷和双巷爆破时,爆破高度越大,巷道围岩峰值应力越低,卸压效果与爆破高度呈正相关关系;但采用双巷爆破时,当爆破高度增加至6倍采高(亚关键层中部位置)以上时,工作面中部区域会出现应力升高现象,且工作面越短应力升高越明显;在7202工作面初采阶段进行高度为3倍采高的单巷顶板爆破时,微震多集中在巷道区域,卸压效果一般,改为双巷顶板爆破并提高爆破高度至6倍采高后,微震多集中到工作面中部区域,无大能量微震事件发生,应力监测预警次数也大幅降低,巷道区域卸压效果较好。

托顶煤切顶留巷围岩控制技术研究

为了解决托顶煤巷道围岩变形难以控制的问题,运用理论分析、离散元数值模拟及现场工业试验相结合的方法对义棠煤业100602工作面切顶成巷围岩控制技术进行研究,分析了不同切顶高度、采高对留巷围岩变形影响规律。研究发现,适当增加切顶高度并降低采高有利于巷道围岩稳定,并确定开采技术参数为切顶高度24 m、采高6 m。现场监测结果显示,在合理的支护策略及开采工艺参数下,托顶煤切顶留巷围岩变形控制效果良好。

Evaluation of load transfer mechanism under axial loads in a novel coupler of dual height rock bolts

The effective reinforcement of two or more overlying layers of mine openings in a single installation is usually done by coupling of two standard rock bolts mainly during the extraction of medium-thick coal seams.However,field observations show that the couplers of multiple bolts often degrade or break mostly at their connections.These types of failures can be avoided by strengthening the couplers of such multi-bolts assemblies.To achieve this,a novel threaded coupler system with an expansion shell was suggested in this paper.The newly designed coupler consists of a threaded tapered-plug-cumconnector with an expansion shell for connecting and tightening two standard rock bolts.An analytical model for evaluating the load distribution along the coupler subject to axial load was derived.Numerical analysis was performed to analyse the load transfer,deformation,and strains across the coupler including the factor of safety for the bolt-coupler-resin and bolt-coupler-expansion shell.The results validated the analytical model of the proposed coupler design,which provides better anchorage near the interface of the host rock mass.Thus,the developed coupler design would reduce the failures of the proposed coupler and stabilize laminated roof strata above the medium-thick coal seams in underground mines.

深井岩巷空顶区失稳及安全空顶距研究

为解决深井岩巷掘进过程中空顶区冒顶、片帮的安全问题,以吕家坨矿-950四采区回风道为工程背景,综合采用理论分析与数值模拟相结合的方法开展研究。运用卡斯特纳公式计算了巷道开挖之后塑性区半径为3.19 m,以此为基础建立了一端简支一端固支的空顶区岩梁力学模型,推导出了最大空顶距表达式。研究结果表明,-950四采区回风道的安全空顶距为2.6 m;运用GDEM模拟分析了不同工况下围岩的应力和损伤情况,模拟结果表明前方岩壁的支撑作用要明显高于后路锚杆对顶板的支撑作用,Ⅲ型工况下围岩的应力和损伤情况与Ⅱ型工况相差不大,考虑炮孔深度受限,空顶距为2.4 m时较为合适。

甜水堡煤矿煤巷支护参数与设备工艺优化研究

目前巷道快速掘进技术研究主要针对巷道快速掘进的影响因素、设备优化等,对巷道空顶距、支护参数、施工工艺联合优化的研究较少。针对该问题,以甘肃省环县甜水堡煤矿2号井1309工作面回风巷为研究对象,对煤巷支护参数与设备工艺优化方法进行研究。分析了巷道掘进各工序的用时特征,得出掘进、永久支护、临时支护用时最多,占比分别为25.3%,49.9%,6.2%;以耗时最长的3个工序为重点优化方向,构建了掘进工作面空顶区顶板力学模型,得出掘进工作面理论最大空顶距为2.32 m,考虑现场受设备、地质、工艺等因素影响,确定空顶距为2.0 m;根据不同支护方案下巷道围岩应力、变形、塑性区的分布特征,结合巷道高效掘进需求,确定最佳锚杆间排距为800 mm×1000 mm。结合巷道实际的地质条件,配套优化了掘进设备、临时支护工艺与施工工艺。现场试验结果表明,优化后最大日进尺由8 m提高到10 m,巷道掘进速度提高了25%;巷道围岩变形基本处于稳定状态,最大变形量为226 mm。优化方案不仅保证了巷道的安全稳定,还显著提高了巷道的掘进速度。

综采工作面超前巷道变形机制及围岩控制技术研究

针对采煤工作面超前巷道变形量大等问题,研究了采煤工作面端部覆岩运动特性对超前巷道变形的影响,分析了超前巷道切顶卸压机理,确定了合理的切顶参数,进行了现场试验。结果表明:工作面超前巷道的变形的根本原因是工作面推进过程中弧形三角板在工作面推进方向和倾斜方向下沉运动导致其重量及其荷载作用在巷道顶板造成的,并在此过程引起超前支承压力和侧向支承压力的集中。有效切断回采巷道上方顶板与工作面顶板的联动效应后,超前巷道替代了单体架棚支护,超前巷道顶底板累计变形量和两帮累计移近量分别为未切顶的1/5和1/3。