In order to more accurately describe the change process of air flow and gas in the ventilation network system during the period of mine reverse ventilation, the asynchronous reverse wind problem of multiple wind Wells in the mine was studied by using TF1M3D simulation. The FDt(k) of all the air Wells is combined into a numerical code DF according to a certain algorithm to judge the working state of the fan of multiple air Wells,three possible working states of the ventilator, including shutdown, non-operation, single-machine operation and dual-machine operation, are defined, and corresponding codes representing these three states are given as the basis for judging whether the working state of the ventilator in TF1M3D is updated.Three possible working states of the ventilator, including shutdown, non-operation, single-machine operation and dual-machine operation, are defined, and corresponding codes representing these three states are given as the basis for judging whether the working state of the ventilator in TF1M3D is updated. Combined with the field example of the eighth mine in Hebi mining area, the simulation analysis is carried out by using TF1M3D simulation software. The reverse ventilation starts, the fan stops, and in a short pause, the mine relies on natural wind pressure ventilation, the airflow velocity is low, the air volume is small, and the gas accumulation occurs for a short time. Then, with the change of ventilation direction of the fan, the air flow changes accordingly. TF1M3D gives the corresponding changes of ventilation and working conditions. At the same time, the gas migration state changes, and the gas peak overlimit phenomenon appears on the working face. With the advance of the reverse ventilation work, the concentration of migrating gas showed a gradual decrease. The simulation of asynchronous reverse ventilation in multiple air shafts of mine accurately describes the corresponding changes of gas components in the mine system when the wind is stopped in the reverse ventilation operation, and improves the calculation accuracy of ventilation simulation during the disaster period.

To investigate the movement law of overlying strata in the fully mechanized top-coal caving face of extra-thick coal seams and further achieve the effective prevention and control of rock burst. Taking the 12240 working face of Gengcun Coal Mine as the engineering background, initially, the calculation formula for the collaborative deformation load of hard rock strata and the formula for the instability scale of hard rock strata are employed to precisely ascertain the position and instability scale of the key strata. Subsequently, a UDEC calculation model is established to comprehensively investigate the instability characteristics of the overlying hard rock strata. Simultaneously, a meticulous analysis of the working resistance of the on-site supports is carried out to validate the theoretical computations. Eventually, in light of the research findings, a rational determination of the cutting scale is made so as to offer a scientific foundation and efficacious guidance for the execution of roof pre-splitting blasting operations. The research results show that within 50 m above the 12240 working face, there is one low-position hard strata and three medium-position hard strata. Under the influence of mining, the displacement field of the overlying hard strata is symmetrically distributed with respect to the central axis of the goaf. The stress rise areas on both sides of the goaf present a “hyperbolic” distribution, and the stress of the goaf roof and floor is released, showing an “elliptical parabolic belt” distribution. The periodic instability occurred about 8 m after the initial instability of the low-level hard rock layer, and the periodic instability occurred about 20 m after the initial instability of the middle hard rock layer 1. The goaf of the 12220 working face in the north of the working face is a crucial reason for the support resistance of the upper part of the 12240 working face being greater than that of the middle part and the middle part being greater than that of the lower part. When the working face advances 20–30 m, a distinct pressure accumulation area emerges in the middle of the working face and is regularly distributed as the working face advances. The monitoring results indicate that microseismic events are concentrated in the basic roof and medium-position strata 1, and energy events of 103 J are prone to occur. The basic roof and medium-position strata 1 are the key points for rock burst prevention of the working face. A roof cutting scale of no more than 10 m for the low-position hard strata can effectively reduce the rock burst risk. The research results provide theoretical support for roof pre-splitting and rock burst prevention of the 12240 working face in Gengcun Coal Mine and can offer guidance for the prediction and prevention and control of rock burst risk in mines and working faces under similar conditions.

With the increase in mining depth and intensity, dynamic disasters such as rockburst in mines are becoming more severe. Deep resource extraction is characterized by a high in-situ stress geological environment, closely associated with geological dynamic disasters. However, there is currently no quantitative analysis method for the correlation between the two. In this study, an elastic energy density calculation method is employed, considering the dissipative effect of the self-weight stress field on the tectonic stress field. The remaining energy, referred to as impact energy, is used to classify the risk of coal seam impact, providing a computational method for rapid assessment of impact risk before mining production. The proposed calculation method is compared with 22 mine impact engineering practices in the literature, showing accurate predictions for 21 mines. Since measuring in-situ stress and coal seam physical and mechanical properties is a preliminary work in coal seam extraction, the comprehensive analysis of these data holds significant research and practical value.

During the coal and rock mass fracture process, elastic properties are released and vibration waves are radiated outward. The energy attenuation characteristics of these waves can describe the cumulative damage and elastic energy accumulation of the mass. To investigate coal and rock mass failure characteristics and energy attenuation rules during rockburst, numerical simulation and laboratory testing were utilized to study the energy transfer laws under various parameters. Six variables, including elastic modulus, Poisson’s ratio, bulk density, cohesion, internal friction angle, and void ratio, were selected to simulate the rockburst energy release process under different parameter combinations by adding surface pressure to the model. The coal and rock mass energy attenuation coefficient was obtained by fitting the node energy straight line using the least squares method. The six variables’ influence on vibration wave energy transfer was obtained using analytic hierarchy process program written in MATLAB, and a comprehensive calculation formula was proposed. Using the energy attenuation coefficient, the rock layer energy diffusion distance was calculated and compared with the roof collapse rock layer step distance, resulting in the roof rock layer cutting distance determination. By roof rock strata precutting, rockburst occurrence can be prevented, ensuring safe and efficient coal mine production.

With the shift of coal seam mining to the deep, the in-situ stress of coal and rock mass increases gradually. High ground stress can limit the generation of rock cracks caused by blasting, and blasting usually shows different crushing states than low stress conditions. In order to study the blasting expansion rule of rock mass with cavity under high ground stress and the rock mass fracture state under different side stress coefficients. In this paper, the effective range of blasting and the stress distribution under blasting load are analyzed theoretically. The RHT (Riedel-Hiermaier-Thoma) model is used to numerically simulate the blasting process of rock mass with cavity under different ground stress, and the influence of ground stress and lateral pressure coefficient on the crack growth of rock mass is studied. The results show that when there is no ground stress, the damage cracks in rock mass are more concentrated in the horizontal direction and the fracture development tends to the direction where the holes are located, which confirms the guiding effect and stress concentration effect of the holes in rock mass, which helps to promote the crack penetration between the hole and the hole. The length difference of horizontal and vertical damage cracks in rock mass increases with the increase of horizontal and vertical stress difference. Under the same lateral stress coefficient, the larger the horizontal and vertical stress difference is, the stronger the inhibition effect on crack formation is. For blasting of rock mass with high ground stress, the crack formation length between gun holes decreases with the increase of stress level, and the crack extends preferentially in the direction of higher stress. Therefore, the placement of gun holes along the direction of greater stress and the shortening of hole spacing are conducive to the penetration of cracks between gun holes and empty holes. The research can provide reference for rock breaking behavior of deep rock mass blasting.

High primary rock stress can limit the generation of rock cracks caused by blasting, and blasting usually shows different rock breaking states under different primary rock stress conditions. There are a large number of naturally formed joints in rock mass, due to the limitations of laboratory tests, a numerical model of jointed rock mass was established using LS-DYNA software to investigate the evolution of blasting damage under various in-situ stresses and open joints. In this simulation, using the Lagrange-Euler (ALE) procedure and the equation of state (JWL) that defines explosive materials, the study considered different joint thicknesses (2cm, 4cm, and 6cm), joint angles (0°, 30°, 60°, and 90°), and in-situ stress conditions (lateral stress coefficients of 0.5, 1, and 2, with vertical in-situ stresses of 10MPa and 20MPa), through stress analysis and damage area comparison, the relationship between damage crack propagation and horizontal and vertical stress difference is explored. The research aimed to understand the mechanisms underlying crack initiation and propagation. The results show that: (1) The presence of joints exerts a barrier effect on the expansion and penetration of cracks. When explosion stress waves reach the joint surface, their propagation is impeded, leading to the diffusion of wing cracks at the joint ends. When the lateral stress coefficient and joint angle are the same, an increase in initial in-situ stress results in a reduction in the area of the blasting damage zone. (2) Under the same initial in-situ stress conditions, the area of the blasting damage zone initially increases and then decreases with an increasing joint angle. However, it remains larger than that without a joint, and there exists an optimal angle that maximizes the damage area. In the simulated conditions, the area of damage cracks is greatest when the joint angle is 60° dip angle. (3) The presence of initial in-situ stress has a certain impact on the initiation and expansion of blasting cracks. The degree and nature of this influence are not solely related to the lateral stress coefficient but also depend on the joint’s angle and thickness. When in-situ stress is present, the initial in-situ stress field’s pressure is not conducive to the initiation and propagation of blasting cracks. However, the existence of a joint has a noticeable guiding and promoting effect on crack propagation, and the pattern of crack propagation is influenced by both joint and in-situ stress conditions.

In order to establish a quantitative relationship between fault activity and rock burst,realize the effective prevention and control of rockburst.We analyzed the geological structure environment of rockburst in Yima mining area theoretically,and analyzed the relative position relationship between panel 13200 and fault influence zone in Gengcun coal mine by taking the evaluation method of geo-dynamic conditions of rockburst.We further discussed the macro-control effect of the fault structure on rock burst.On this basis,we calculated the tectonic stress in the minefield and divided the tectonic stress into different zones,and analyzed the control effect of tectonic stress on rockburst.We designed the actual quantitative monitoring method of underground fault,monitored the activity of F16 fault in Gengcun mine field directly,analyzed the changes of fault displacement and stress increase during the preparation and occurrence of high-energy microseismic events quantitatively.We constructed a model of"the scale of coal and rock mass in the focal area is equal to that of the dy-namic core area,and the energy of the focal area gradually attenuates with the transmission distance,"which provides an important tool for establishing the relationship between large-energy microseismic events and fault activities,establishes the relationship between high-en-ergy microseismic events and fault activities,and determines the influence of fault activity on rock burst.The research results show that in Yima coalfield,the complex thrust nappe structure system constitutes the geological tectonic background of rockburst in Yima mining area.The width of F16 fault influence zone is 7 000 to 7 600 m,and the whole panel 13200 is within the influence zone of F16 fault,which further increases the risk of rockburst under the influence of mining activities.The areas controlled byⅠ-2 fault,Ⅲ-4 fault andⅣ-7 fault are the main areas for the occurrence of rockburst and large energy microseismic events in Gengcun Coal Mine.Most of the rockburst and large energy microseismic events are located in stress gradient areas.During the preparation and occurrence of large energy microseismic events,the displacement of F16 fault increased by 50 mm and 45 mm respectively.Before the two large energy microseismic events occur,the increase of fault activity tension are relatively the highest,which are 2.58 kN and 2.93 kN.The rapid increase of fault displacement and higher stress increase constitute the main energy source of large energy microseismic events.The occurrence of large energy microseismic events and rockburst are closely related to the activity of faults.The actual quantitative monitoring method of underground faults can be widely used in the guidance of mine rockburst prediction and prevention and control.

In recent years, with the increasing of mining depth and mining intensity, coal and gas outburst has become one of the most destructive and harmful dynamic disasters in coal mines. Natural earthquakes are the most destructive natural disasters in the earth’s crust. Both of coal and gas outbursts and natural earthquakes are caused by geo-dynamic processes. In order to research the correlation between coal and gas outburst and natural earthquake, in this manuscript, we took Pingdingshan eastern mining area as the research object. The principles of disaster prevention, seismology, statistics and geophysics were taken as the research basis, both the characteristics and laws of natural earthquakes, and the occurrence laws and characteristics of coal and gas outburst were systematically analyzed by geo-dynamic division method. At the same time, the relationship between the two disasters were established. The research results show that natural earthquake is the accumulation and release process of elastic properties in rock mass, which can induce the abnormal emission of gas in coal mines under specific conditions. Especially the “weak plane structure” in geological structure, which is more likely to lead to the occurrence of coal and gas outburst accidents. Tectonic activities and stress field changes have a unified mechanism for the occurrence of natural earthquake and coal and gas outburst, there are correlations between space and intensity. The research results can also provide new ideas for the prediction work of these two kinds of disasters.

为研究坚硬覆岩的结构失稳运动规律,以耿村煤矿13200工作面为工程背景,构建坚硬岩层失稳结构模型,对坚硬岩层结构特征、结构形成条件、结构失稳尺度、结构失稳释放能量等进行了分析,并分别给出了相应的计算公式,确定坚硬岩层失稳对冲击地压的控制作用.研究结果表明:13200工作面低位和中位坚硬岩层下方存在自由空间,具备结构形成的条件;高位坚硬岩层下方无自由空间,不具备结构形成的条件.根据13200工作面钻孔柱状及其物理力学参数确定了其上覆岩层关键层个数和所在层位,并对该工作面坚硬岩层结构失稳尺度和失稳能量释放进行了计算,计算结果表明:在13200工作面回采期间,低、中位坚硬岩层初次失稳和周期失稳的释放能量均超过冲击地压发生的临界能量,是工作面发生冲击地压的主要影响因素;高位坚硬岩层上部的巨厚砾岩不会出现垮落失稳而释放冲击能量,不直接控制冲击地压的发生,是工作面发生冲击地压的次要影响因素.确定了13200工作面坚硬岩层结构失稳的防治措施,关键是通过减小低、中位坚硬岩层结构失稳尺度和失稳后释放的能量,以降低冲击地压的发生几率.研究结果可为其他具备类似坚硬岩层条件矿井的冲击地压控制提供重要参考.

Abstract In order to solve the serious secondary disasters of the gas overrun or gas explosion, which is caused by coal and gas outburst in mine. In this paper, the equations of fluid transients, convection diffusion and source-containing ventilation network were established respectively to research the dynamic effect and gas migration process of gas outburst in ventilation system of coal mine. In order to research the unsteady movement characteristics of mine airflow under the joint action of gas outburst source flow power and fan ventilation, we completed the experiment of ventilation pipe network system for gas outburst. The fitting equation of gas dispersion parameters under different wind speeds was obtained, which was used as the simulation calculation parameter. In this paper, NC2.0, one simulation software of coal and gas outburst, was developed, and the 10.27 coal and gas outburst accident in Jiaozuo Jiulishan coal mine was reappeared by it. Throughout the simulation process, unsteady air flow and gas distribution in Jiulishan coal mine after coal and gas outburst were simulated respectively by the given fluid changing curve. The research results indicate that the outburst pressure after coal and gas outburst, which is the main factor leading to the changing airflow, and radiates in the form of wave. The high-intensity air flow from the gas outburst source changes the air flow movement of the mine system, causes over air flow phenomenon in return air roadway, suppresses the air volume in air inlet roadway at the same time. Even countercurrent occurs in severe cases. The research results provide the theoretical basis for disaster prevention and aid, reducing loss and preventing secondary disasters in outburst mine.

Abstract In order to solve the safety problem of rockburst in coal mining under fault conditions and realize safe and efficient production of coal mine, this paper analyzes the existing mechanism and geological environment of fault, constructs fault monitoring scheme and field construction method, and carries out underground monitoring of F16 fault. The stress and displacement of F16 fault in Gengcun Coal Mine are monitored, the deformation and activity of the fault are analyzed, and the influence on energy when the stress and displacement change is observed. The experimental results show that : Before and after the two large energy microseismic events, the stress and displacement changed significantly, and the energy increased first and then decreased sharply. The mining of working face had an important impact on the activity of F16 fault. The deep mining of coal mine would disturb the coal seam and roof and floor strata, destroy the stress state of the original rock mass, increase the in-situ stress, complicate the geological structure, and develop various faults one after another, It will destroy the integrity and integrity of the coal seam, cause the destruction of surrounding rock near the fault, release energy, cause large energy microseismic events, and in serious cases, it will lead to rockburst accidents. The research results have guiding significance for monitoring the impact of F16 fault activity on Rockburst in Gengcun Coal Mine, and provide experimental methods and theoretical basis for other coal mines to study the change process of energy during fault activity.

In order to improve the simulation accuracy of the process in coal and gas outburst disaster period,time step method was put forward to describing the movement and accumulation of coal flow in roadway network when coal and gas outburst happened.Downward dip angles of tunnel were taken as the independent variables,model for distribution weight of coal output was built based on "Downward Priority" principle,and the model could adapt to downward roadway,nearly horizontal roadway and upward roadway.Strength change of outburst coal was established based on the collapse model.Combining with the coal and gas outburst accident and outburst monitoring facts in Jiulishan Coal Mine,process of coal flow movement was simulated,and the process was that coal flow swarmed into haulageway 16301,passing through reverse No.2 ventilation door and much gas was desorbed and becoming part of the outburst gas source at the same time.Gas of high concentration was monitored by methane sensor when passing through No.2 ventilation door under the influence of counter current wind pressure.The results show that the rapid rise of gas in roadway could be reasonably described,considering the simulation calculation of coal flow.

The paper has managed to simulate and analyze the changing process with the TF1 M3D method by taking the Jiulishan coal mine as a case study sample in hoping to investigate and clarify the changing process of the gas migration and the concentration distribution in the ventilation network system during the period of the mining reverse ventilation.As a result of our simulated study,we have found that the gas emission in the mining system mainly comes from the working face of the goaf due to the reverse ventilation,and,in turn,the gas distribution and the characteristic features of the mine are mainly dependent on the gas emission amount atthe minimum level of the said process.Therefore,our simulation and analysis results are helpful for disclosing the regularity of the gas concentration change in the panel in the reverse ventilation period,that is,the gas tends to return to the airway of the panel through the reverse ventilation and pile up with the would-be-emitted gas from the panel,repeatedly,hence leading to the gas concentration pile-up and affecting the gas return airflow in the initial inlet airway.Moreover,due to the abrupt increase of the gas concentration at the beginning end of the reverse ventilation,the gas concentration would decrease gradually bit by bit after reaching the peak ventilation point.The simulation results have also proven that there exists a superlative position between the reflux gas and the emission gas in the panel during the reverse ventilation period.In addition,the peak point of the gas concentration piling-up must have been related with the reverse ventilation system structure of the mine.More exactly speaking,the safety of the reverse ventilation of the mine tends to be affected by the peak gas concentration airflow,which is of the actual continuous need of supervision and the monitoring activities for the gas concentration changing rate in the panel on the one hand,and the countermeasures and technological regularity facilities for the coal production should be well made to the point in advance on the other.

为更准确描述矿井反风时期通风系统网域中风流与瓦斯的变动过程,定义了通风机可能出现的停机、单机运行和双机运行等3种工作状态,给出对应状态编码.结合鹤壁矿区八妒实例,运用TF1M3D进行仿真分析,仿真结果与现场实测各特征点的瓦斯浓度相拟合.反风开始,通风机停机,矿井依靠自然风压通风,风量小,出现短时间的瓦斯积聚.通风机反风启动后,工况点发生变劾风流随之变化,工作面瓦斯出现高峰超限现象;随着反风的进行,运移瓦斯的浓度呈阶段性的降低.TF1M3D描述异步反风时矿井系统瓦斯气体组分变化的仿真.

为仿真描述有源风网上外源气体涌入运移过程,创新引入事件步长法,把风网中外源气体组分运动到达某一节点作为一个事件,同时取同期其他分支事件所经历的最短时间作为下一个时间步长,考察记录网络分支中外源气体到达的位置.为减小截断误差,对过大的事件时间步长作出限定,形成所谓“时间-事件混合步长”法.使每个迭代时间步长都能至少有外源气体到达某一个节点,采用节点分支汇流法计算节点浓度,使网络节点上气体浓度得到及时更新.结合矿井瓦斯突出流在网络中的扩散给出了算例,展示了不同时刻瓦斯达到网络各个节点的过程.研究结果表明:新方法能够自动截取网络中组分气体流动的微小事件变化,防止因过大的时间步长导致的截断误差;与单纯的时间步长法相比,计算精度高,模拟仿真更准确.

为研究煤与瓦斯突出后整个矿井发生的灾变过程,基于有源通风网络理论和数值仿真技术,利用TF1M3D矿井灾变通风仿真系统软件,对2004年郑煤集团大平煤矿"10·20"煤与瓦斯突出事故进行仿真推演.模拟得到,突出高峰期,以突出点为中心,东翼区域发生大规模逆流,逆流瓦斯渗透通过11轨道石门中的风门进入西大巷;在突出衰退后,逆流消失,进入西大巷的瓦斯随进风流进入13,15采区;同样因风门的限流,11轨道石门中出现瓦斯滞留,当架线电机车经过交汇点时引发瓦斯爆炸.推演结果与实际情况相吻合.研究表明,瓦斯突出灾变时期,防止瓦斯滞留是预防二次爆炸事故的关键;积极推广使用远程控制的自动风门,用它进行调控,能最大限度降低灾害损失.

The mathematical model of disaster process in mine ventilation system during mine fire period was built based on source-containing ventilation network and previous fire studies,and the software TF1M( 3D) was developed based on MATLAB. Combined with the typical mine ascensional airflow fire example,the airflow movement during fire period,the distribution of fume concentration and temperature,the ventilation system variations were simulated in a whole mine scale. Simulation results show that the air quantity in main airway increases with fire when ascensional airflow fire occurred in mine,while the airflow in the side branch of fire source decreases,stagnate and reverse; driven by fire burning power( fire ventilation pressure),overflowing quantity and reverse air quantity have consistent symmetry; and in complex ventilation network,reverse airflow in side branch orderly occurs. A series of changes of ventilation system during fire period are all produced by the interaction of ventilation power and fire ven-tilation pressure,which also lead to the dynamic drift of mine system total wind drag. This paper provides the relationship between fire severity and wind drag drift value. The software TF1M( 3D) offers a good platform for analyzing mine fire because of its large amount of information,good animation and intuitive phenomenon.

采用平衡解吸法研究了西辽河沉积物不同天然有机组分对氨氮解吸特征的影响.结果表明,去除有机质后的沉积物氨氮解吸比例(Dr=0.70)大大增加,同时,解吸迟滞性指数显著降低(Tn =0.016),有机质是影响沉积物氨氮解吸特征的重要因素;重组有机组分中的紧结态腐殖质(胡敏素)对抑制氨氮解吸起关键作用,它不但解吸比例较低,而且解吸迟滞性指数较大;氨氮在轻组有机组分上的表面分配作用吸附是导致解吸比例增大,解吸迟滞性指数减小的原因;考查沉积物吸附态氨氮的解吸特征不但要考虑有机质的含量,更要考虑有机质的存在形态,它也是影响沉积物吸附态氨氮解吸特征的重要因素,轻组有机组分、重组有机组分以及重组有机组分中的稳结态腐殖质和紧结态腐殖质所吸附的氨氮对上覆水体的扩散通量可分别按其饱和吸附量的0.97倍、0.41倍、0.25倍和0.17倍进行估算.

针对巷道火灾绕火源火焰的局部阻力计算及阻力系数的确定问题,建立按火焰断面积来计算的模型,其中的火焰局部阻力系数是火焰截断面积的函数,指出火焰面积是在特定巷道火灾条件下火源燃烧强度与风流速度相互作用的结果,确定阻力系数值必须考虑到巷道风流对火焰的压制作用.分析火焰局部阻力与风流速度、火源燃烧强度的关系,即随着火源燃烧强度的增大,火焰面积和局部阻力增大,风速降低,出现所谓的节流现象.当风速等于富氧燃烧临界风速时,火源燃烧充分,火焰面积和火焰局部阻力最大;当风速小于该临界风速时,火源燃烧减弱,火焰面积和局部阻力变小;当风速过大时,风流压缩火焰,火焰截断面积和火焰局部阻力减小,火焰局部阻力在风速中低时为最大.上述变化规律与点火源和线火源两组经典的实验结果相吻合;依据实验数据拟合确定出模型的参数,依火源强度不同,最大火焰面积系数取值一般在0.285 ～0.410.