The stress state of primary rocks is redistributed during coal mining, which leads to deformation and failure of the floor strata. Coal seam floor experiences repeated additional stress from the mining of two overlying seams in close-distance mining. An assessment of floor stability is required in the presence of a confined aquifer within the floor strata. The failure characteristics of floor rock strata are crucial for analyzing water inrush from floor. This mechanical response of floor deformation and failure becomes particularly pronounced when the roof is treated using the full caving method. Floor strata were analyzed for the 220,105 working face at Xinji 2# Coal Mine, with an elastic half‑plane theory established. The stress distribution characteristics of the floor were investigated using software MATLAB. The width of the plastic zone and the maximum failure depth of floor were determined based on the slip line field theory of floor rock mass. A reference was provided for water damage control during repeated mining of the 1# coal seam. Continuous field observation of floor failure was performed through drilling combined with the network parallel electrical method. Apparent resistivity at the depth of 19.6 m below floor was approximately twice the background resistivity of the corresponding stratum. The apparent resistivity of rock strata below 19.6 m remained constant, indicating a protective zone. These results confirm the reliability of the theoretical analysis and provide a technical foundation for safe mining in the working face.

After coal extraction in underground mining, progressive fracturing occurs in both roof and floor strata, resulting in stress redistribution within these rock layers. Considering the fractures in the strata at roof and floor, an elliptical stress arch was introduced. This concept analyzed how the shape and structure of this arch evolved when mining progressed, as well as the evolution law of the strata at roof and floor fractured. Research results are as follows. (1) The elliptical stress arch bore most of the load on the surrounding rocks of the mining site, and the front and rear arch springing moved forward with the mining face. The surrounding rocks of the mining site redistributed stress, which formed tensile stress in the areas near the arch top and base. The stratum underwent active fracture and was compressed in its movement direction. Compressive stress was formed to clamp unbroken strata in the area far away from the arch top and base. Strata were passively broken along the arch trajectory due to their self-weight load, the compressive load of active-fracture strata, and the clamping effect of the compressive stress around the stress arch. (2) Unbroken hard strata inside the elliptical stress arch formed a cantilever beam. A hinged rock beam was formed after hard strata broke. Soft strata acted on lower strata in the form of loads. The support structures at the working face and the coal wall carried the weight of strata, which was transmitted through cantilever beams or hinged rock beams at roof and floor. The passive fracture of the cantilever beam and the instability of the hinged rock beam directly affected the stability of the working face support and coal wall. (3) The analysis focused on how the lengths of the long and short axes and the coordinates of the center of the elliptical stress arch changed as the working face advanced. The theory of composite beams was used to determine strata, rock loads, fracture distances, and fracture sequences of hard strata at the roof and floor of the elliptical stress arch. The stability at the roof and floor of the hinged rock beam was obtained using the criteria for compression deformation instability and sliding instability of the hinged rock beam. The research results can provide support for calculating the resistance of the working face support, the load borne by the coal wall, and the width of the plastic zone of the coal wall. Besides, they offer a theoretical basis for analyzing the stress of the working face support and the stability of the coal wall in the mining area, as well as production practice.

Focusing on the engineering characteristics of high stress, strong disturbance, and large deformation in deep coal mining. Borehole pressure relief technology, however, can control the stability of the surrounding rocks in mining faces. The work employed laboratory true triaxial tests on borehole pressure relief rocks, numerical simulations, and a secondary development approach incorporating representative volume elements (REV). The approach was to investigate the stability of surrounding rock structures under mining-induced conditions and the multiscale fracture evolution law of overlying strata in mining faces. The key findings are as follows. (1) True triaxial tests on rocks with holes were conducted to analyze the mechanical response and failure characteristics under different hole parameters. Stress concentration around holes significantly altered rock fracture modes, forming directional fracture zones aligned with the hole arrangement. (2) A multiscale fracture evolution model for rocks at macro- and micro-levels was established, employing REV and the Weibull function to characterize rock heterogeneity. The Mogi-Coulomb criterion was used to describe the damage evolution process of micro-level elements. A macro-micro numerical computation program for coal-rock masses was developed, with the progressive fracture evolution law of holed rocks under true triaxial conditions analyzed. (3) Stress concentration near holes and the X-shaped fracture propagation pattern were revealed to investigate the influences of hole quantity, size, and distribution on regulating rock fracture paths. The stability mechanism of surrounding rocks controlled by borehole pressure relief technology was elucidated.

The coal-seam occurrence conditions of Yuandian No. 1 Coal Mine in Huaibei were taken as the engineering research background in the work to explore the distribution laws of abutment pressures ahead of the roadway under two mining processes. The theoretical analysis, numerical calculation, and engineering reality measurement method were used to study the peak position and suffering area of the advanced abutment pressure of the return airway under two mining processes at working face 824. The results showed the significant influence of the mining height on plastic and elastic zones. In the fully mechanized mining stage, the bearing pressure reached a peak at 2 m ahead of the coal face, and the maximum value was 20 MPa. At the stage of com-prehensive top-coal caving, the bearing pressure at 8 m ahead of coal mining reached a peak, and the maximum value was 17.38 MPa. The suffering areas were 42 and 53 m, respectively. The bolt (cable) spacing of the roadway was optimized to improve the stress environment and bearing capacity of surrounding rocks in the roadway. The research results can provide an advanced support basis for sublevel mining at working face 824.

为探究近距离煤层内错式开切眼水平错距的合理选取问题,以神东矿区某矿为工程研究背景,通过构建力学模型得到上覆煤层遗留停采煤层产生的附加应力对下方开切眼的应力解析解,采用数学分析得到下煤层的水平停采距离的合理范围,并运用FLAC3D数值模拟软件分析不同水平错距下切眼位置所处的应力环境,得到下煤层开切眼停采的水平距离为50 m时能够满足巷道布置要求.通过地质雷达进行开切眼巷道的围岩松动圈的测试,得到开切眼巷道松动圈的范围为3.1～3.8 m,切眼巷道左帮的破碎范围大于顶板,开切眼巷道围岩完整性较好.研究结果表明,下煤层内错式开切眼停采水平距离为50 m时,开切眼巷道能够满足工作面的正常回采.

The application of roof-cutting and pressure-relief gob-side entry retention plays a critical role in controlling the stability of the surrounding rock at the entry, easing continuity tension and improving resource recovery ratio. The excavation of the 360,803 airway in Xinji No. 1 Mine is affected by intense mining of the 360,805 working face. Hence, to address the stability problem of surrounding rock in the 360,803 airway, rock mass blast weakening theory was used in this study to analyze the blasting stress of columnar charged rock mass and obtain the radiuses of crushed, fractured, and vibration zones under uncoupled charging conditions. The reasonable array pitch, length, and dip angle of boreholes were determined according to the pressure-relief range of the blasting fracture. The migration laws of roof strata were explored based on a mechanical model of overlying roof strata structure on the working face. Subsequently, the horizon, breaking span, and caving sequence of hard roof strata were obtained to determine the roof-cutting height of this entry. On the basis of the theory of key stratum, the number of sequences at the roof caving limit stratum and hanging roof length in the goaf were calculated, the analytical solution to critical coal pillar width was acquired, the evaluation indexes for the stability of entry-protecting coal pillars were determined, and the engineering requirements for the 25 m entry-protecting coal pillars in the 360,803 airway were met. Moreover, various indexes such as roof separation fracture, displacement of surrounding rock, and loose circle of surrounding rock in the gob-side entry were analyzed. The stability and cementation status of surrounding rock in the 360,803 airway were evaluated, and tunneling safety was ensured.

The filling mining method is important in realizing the green mining of mineral resources. Aiming at the problems of land resource occupation, environmental pollution, and rational utilization of coal-based solid wastes such as coal gangue, fly ash, and desulfurization gypsum, a new paste filling material was developed with coal gangue, fly ash, and desulfurization gypsum as raw materials. The microstructure of the raw materials was analyzed by XRD and SEM. Combined with the Box-Behnken experimental design, the effect of each component on the fluidity of the filling slurry was analyzed through the response surface analysis. The significance of each component on its bleeding and fluidity was determined, and the optimal ratio of the filling slurry was obtained. Experimental results show that the microcosmic morphology of coal gangue, desulfurization gypsum, and gasification slag presents an irregular block and rough particle surface; the microcosmic morphology of fly ash and bottom slag presents first out spherical or quasi spherical particles. Moreover, obvious sintering traces exist on the surface of the bottom slag. The main crystal mineral of coal gangue and fly ash is SiO2, the desulfurization gypsum is composed of Ca(SO4) (H2O) and Ca(CO3) crystal minerals, the gasification slag is composed of carbon and nitrogen compounds, and the main crystal mineral components in the bottom slag sample are SiO2 and AlxSiyOz compounds. The order of significance of each key factor on slurry fluidity is as follows: C (desulfurization gypsum) > D (gasification slag and bottom slag 1:1) > A (coal gangue) > B (fly ash). The order of the significance of each key factor on slurry bleeding is as follows: B (fly ash) > C (desulfurization gypsum) > D (gasification slag and bottom slag 1:1) > A (coal gangue). Considering the material preparation, field application, and other conditions, the mass percentage of each factor content of the new paste filling material is as follows: 49.5% coal gangue, 8.3% fly ash, 4.1% desulfurization gypsum, 6.2% gasification slag, and 6.2% bottom slag.

为探究多源煤基固废充填体力学特性,制备29组尺寸为70.7 mm的立方体试件,采用单轴压缩实验系统得到不同龄期充填体试件抗压强度,利用XRD,SEM微观观测技术分析原材料的微观形貌及物质成分组成,并采集不同龄期试件断面微观形态结构,揭示充填体强度演化规律;基于响应面法研究各组分对不同龄期(3,14,28 d)充填体强度的影响,应用声发射测试系统监测试件破坏过程中声发射特征.研究结果表明:各因素对试件强度影响的显著性大小为:粉煤灰＞矸石＞脱硫石膏＞气化渣与炉底渣的1∶1混合物;响应面法得到粉煤灰及矸石对充填体试件的抗压强度影响较大;新生成的棒状结晶物及蜂窝棉絮状胶凝物质随龄期的增长逐渐增多,从而使充填体强度增高;试件3 d强度最小,3～14 d强度增加幅度最大,14～28 d强度增加幅度次之;塑性变形阶段,有少量能量释放,试件表面出现细微裂纹,并逐渐演变成贯通裂隙;屈服破坏阶段,声发射振铃计数陡然增大,试件内部集聚的能量突然释放,试件表面出现明显贯通式裂隙,并逐渐裂开、崩落;峰后阶段,试件依然具有一定的强度,声发射事件依然存在,并伴随能量释放,研究成果可为工程现场充填材料配比优化提供依据.

Rock excavation has experienced complex stress paths. The development of the original crack under the path of principal stress magnitude and principal stress direction is a key scientific problem that needs to be solved in rock underground engineering. The principal stress magnitude dominates the initiation and propagation of the crack and increases rock damage under the action of principal stress rotation. In this study, the theoretical calculation and numerical analysis method have been combined with the crack propagation conditions to study the stress-driven mechanism of brittle rock crack propagation under principal stress rotation. The results show that the “relative initial angle” of crack angle is being updated in time during the principal stress rotation process; once the stress is rotated, it will become the next initial crack angle; the crack propagation direction is deviated under the applied shear load, and it is always in the direction of minimum shear load, leading to a certain degree of inhibition of crack propagation depth in the initial direction. According to the results of numerical simulation, the effect of principal stress rotation caused by mining excavation is obvious and has a certain range of influence depth, the stress of surrounding rock of roadway is the highest within the depth range of 1∼2 m, and the maximum principal stress is as high as 26.89 MPa. The rotation of principal stress direction on the roadway surrounding rock surface is the strongest, which makes the surrounding rock more fragmented, and the middle principal stress and the maximum principal stress rotate about 90° counterclockwise along the Ox axis. Studying the action mechanism of principal stress rotation on fractured rock masses can provide scientific basis for geotechnical engineering design and rock mass surrounding support.

不连续结构面对爆炸应力波在岩体中的传播有很大影响,研究应力波作用下含大型结构面岩体动力垮塌失稳机制,为含结构面等地质构造条件下岩体开采及稳定性控制提供理论依据.以冬瓜山铜矿56-8#矿房采场回采诱发相邻56-7#矿柱采场垮塌为工程背景,建立结构面剪切滑移失稳力学模型,基于莫尔-库仑强度准则,采用理论分析方法计算了爆炸应力波通过结构面的透、反射系数,分析应力波通过结构面的应力响应特征;推导结构面发生剪切滑移的能量判据及应力判据,获得结构面的发生剪切失稳的条件.通过对结构面抗剪强度参数进行不断弱化,研究了微差爆破对结构面的破坏规律.研究表明:(1)入射角在一定范围内,反射波不发生相位延迟现象,并且在此入射角范围内,能量耗散系数大于0,结构面发生剪切滑移;(2)结构面是否发生剪切滑移失稳与结构面抗剪强度参数(内摩擦角、黏聚力)及应力波入射角、爆源距结构面距离有关;爆源距结构面的距离越远,结构面的抗剪强度参数越大,越不容易发生剪切滑移失稳.(3)微差爆破造成结构面强度参数不断弱化,使其发生剪切破坏变得容易,当结构面参数达到了发生剪切滑移的临界条件时,则56-7#矿柱采场发生垮塌.

为了解决刘庄煤矿150802胶带顺槽快速掘进过程中巷道帮角局部变形严重的问题,通过FLAC3D数值模拟软件确定了巷道异常区域的关键部位.结果表明:巷道在帮角处产生较大剪切应力集中区域,高帮角的剪切应力集中值大于低帮角,巷道顶板的位移变形量大于两帮,通过数值模拟确定巷道支护的第一关键部位是巷道顶板的两个肩窝,其次是顶板,最后是两帮.针对异常区域提出了优先支护的工艺顺序,并对巷道围岩采用"锚带网索"加强支护,锚杆(索)应力传感器现场实测应力是顶板两肩窝＞顶板＞两帮,与数值模拟结果相吻合,且巷道采用异常区域优先与"锚带网索"联合支护后,在巷道掘进110 m后,巷道围岩整体变形逐渐趋于稳定,巷道两帮的移近量和顶板的下沉量均低于45 mm,异常区域变形得到有效控制,实现了巷道平均日进尺16.8 m的快速掘进目标.

An overlying rock structure plays a key role in controlling the roof deformation of nonpillar gob-side entry retaining by roof cutting. On the bases of the actual geological conditions of II 632 Haulage Roadway at the Hengyuan coal mine, a similar three-dimensional simulation experiment of roof precutting is conducted. Thereafter, the caving characteristics and migration law of the roof strata in the strike and dip directions are obtained. Moreover, the roof of the retained roadway and key strata of the goaf can form a hinge structure of the key blocks. By monitoring the deformation of the surrounding rock and stress distribution of the roof, the skew deformation characteristics of roadway roof are obtained. By observing the borehole peeping technology, the roof subsidence near the goaf is determined to be greater than that of the solid coal side, and the roof subsidence of the gob-side entry retained by roof cutting is greater than that of the floor heave and two sides approaching. Results of the three-dimensional similar simulation experiment indicate that the mechanical structure model of the key block of the retained roadway roof is constructed, and the mechanical analytical solution of the required support resistance of the retained roadway roof is obtained. This study proposes the constant resistance and large deformation anchor cable reinforcement support method to control the roof deformation of the retaining roadway. Through engineering application, the maximum value of the roof and floor movement of the retained roadway is stable at approximately 650 mm. The retained roadway can meet the demand of the next mining face.

回采参数是影响无底柱分段崩落法损失贫化的主要因素之一,其中崩矿步距与边孔角调整较灵活,对其进行优化,是降低矿石损失贫化的有效途径.在分析回采参数优化基本依据的基础上,针对毛公铁矿回采指标不佳的问题,采用物理模拟实验方法,开展大结构参数条件下崩矿步距与边孔角的优化研究.实验结果表明,在分段高度20 m、进路间距18 m的条件下,采用3.6m的崩矿步距和45°边孔角,不仅有利于回采矿石,也有利于控制岩石混入,可减少矿石资源的浪费,实验结果可为类似矿山优化回采参数提供参考.

崩落体、放出体和崩矿步距之间的相互制约关系影响着无底柱分段崩落法放矿过程中的损失贫化.以大结构参数单进路放矿试验为基础,进行了3.6 m崩矿步距条件下5种不同铲入深度的物理模拟放矿试验.通过放出标志颗粒的位置信息获取内部滑移面的位置,研究了铲入深度对放出体发育形态的影响,并分析了铲入深度与放出体之间的关系以及放出体与崩矿步距的匹配关系.结果表明:①3.6 m崩矿步距、5种不同铲入深度条件下,沿进路方向放出体发育的轮廓都大于崩落体的轮廓;②铲入深度对放出体形态的发育有一定影响,放出体发育高度随铲入深度增大呈先增加后减小的趋势,放出体沿进路方向的发育宽度呈增加趋势,从纯矿石回收量、矿石回收率方面分析,铲入深度为5.32 m较好,相应的回收率为57.74％.

Aiming at the technical problem of large loss and dilution in non-pillar sublevel caving mining method ,the effects of shoveling depth on the index of the amount of drawing out ore ,the ore recovery ratio and rate of rock mixing researched via using a single funnel ore drawing model with big structure parameters .The results show that the waste rock at the front face of the end of the ore drawing will invade the ore drawing process earlier than the top waste rock;when the shovel depth is small ,the rate of rock mixing rate increases slowly and the hanging-up arch often occurs at the ore drawing hole;with the increasing of shoveling depth ,the increasing rate of rock mixing rate enlarges and it is difficult to form the hanging-up arch in the ore drawing hole .Meanwhile ,it is discovered that the inhomogeneous flow of ore and rock at the drawing hole is an important factor induced the waste rock at the front face of the end of the ore drawing invade the ore drawing process earlier and massively .

针对无底柱分段崩落法放矿损失贫化大这一技术难题,以大结构参数单进路放矿模型为基础,通过物理模拟实验进行了3种崩矿步距、7种铲入深度下的21组放矿实验.从放矿口出现废石的早与晚、出现废石时顶部废石漏斗凹陷深度、放出纯矿石量、损失与贫化等方面分析了铲入深度与崩矿步距之间的关系.结果表明:崩矿步距越大,正面废石侵入时间越晚;铲入深度越大,正面废石侵入时间越早.4.6 m崩矿步距条件下,铲入深度为5.32 m时,矿石回收率达到最优值59%左右,此时混岩率为16%左右.

覆盖层是满足无底柱分段崩落法回采工艺条件和确保生产安全的重要保障.在归纳覆盖层形成方法和覆盖层的作用特征的基础上,分析了组成覆盖层岩石的粒度及其流动特、覆盖层的阻风特性、抗冲击特性和抗渗特性等因素对覆盖层厚度的影响,对研究覆盖层的合理厚度意义重大.

In view of the issues of the lower ore recovery rate and less ore drawing in Maogong Iron Mine,the physical simulation experiments based on the stope structure parameters of the iron ore were carried out for exploring ore recovery,the shape of the residual ore body and its forming process at the conditions of the ore caving slice with 1. 6 m and the 1. 8 m. The results showed that the ore recovery rate was ranging from about 40% to 50% at the first caving sublevel and is at about 70%~90% in the following sublevels with mixing rate of about 35%. The ridge residual ore was gradually recovered in the next sublevel;Waste rock in front of the caved ore body was mixed earlier into the process of ore drawing and caused ore dilution. Comprehensive analysis showed that the ore recovery rate of 1. 8 meter caving step space is better than that of 1. 6 meter.