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As the shallow coal seams are gradually being exhausted, mining of the lower group of coal seams in the shallowly buried proximity coal seam group becomes inevitable. Due to geological conditions affecting the upper group coal seams, various coal mining methods have resulted in a significant number of non-sufficient and sufficient collapse mining areas. The presence of water and gas in these mining areas can potentially impact the safe production of the lower group coal seams through inter-layer rock fractures. Therefore, studying the evolution characteristics and formation mechanism of inter-layer rock fractures is of utmost importance. In this paper, we employ theoretical analysis, numerical simulation, and similar simulation methods to investigate the evolution characteristics and formation mechanism of inter-layer rock fractures during the mining of the lower 3–1 coal seam. We specifically focus on the formation of different mining areas in the upper 2–2 coal seam, using the mining of 2–2 and 3–1 coal seams in the Hanjiawan coal mine as our research context. The study results reveal that the inter-layer rock fractures exhibit an overall positive trapezoidal distribution in different mining areas. We quantify the variation in density and width of inter-layer rock fractures in the two mining areas by utilizing the concept of rock damage degree. We analyze the overall damage degree, tensile damage degree, and shear damage degree ratios. The formation of inter-layer rock fractures is primarily influenced by the combined effect of various coal mining methods in both the upper and lower coal seams. The stress changes resulting from mining activities in these seams contribute to the development of transverse separation fractures and longitudinal breaking fractures within the inter-layer rock. The stress changes induced by mining operations in the upper and lower coal seams have been analyzed, and distinct stress regions have been identified. To understand the mechanism of inter-layered rock fracture development in the upper and lower coal seams, a mechanical expansion model of inter-layer rock fractures has been constructed for different stress regions. Highlights: The evolution process and development characteristics of inter-layered rock fractures, as well as their differences when different coal mining methods are employed in the upper coal seam, have been revealed. To determine the influencing factors and development patterns of inter-layered rock fractures under different coal mining methods. Quantitatively comparing and analyzing the damage of inter-layered rock and the percentage of damage degree under different coal mining methods, we found that the damage degree was 61% and 52%, respectively. The study shows that when mining the upper coal seam, the inter-layer rock primarily extends through primary microscopic fractures. On the other hand, when opening the lower coal seam, the main development is observed in macroscopic tension fractures and shear misalignment fractures. Based on the study of the development process of macroscopic fractures and microscopic fractures in the inter-layered rock, the mechanism of fracture development and evolution of the inter-layered rock under different coal mining methods has been determined. [ABSTRACT FROM AUTHOR]

The continuous large deformation of the surrounding rock during gob-side entry driving in a 1000-m-deep mine requires frequent renovation of the mining roadway, seriously restricting the normal mining operations at the coal face. Taking 11,216 headgate of a coal mine as the engineering background, this paper expounds that the large deformation factors of the gob-side entry driving of a 1000-m-deep mine are stress-field factors dominated by high in-situ stress, side abutment pressure, front abutment pressure and tectonic stress, and stress-related factors dominated by coal body crushing, unreasonable support, large section size and broken narrow coal pillar. The two-stage pressurization of the surrounding rock of the section roadway of the mine is governed by the following fundamental law: After pressurization, the peak area of the original abutment pressure of the roadway transfers to the lower parts of the surrounding rock, resulting in an increased stress in the surrounding rock at the side of the roadway. A targeted cooperative control involving external anchor-internal unloading of the surrounding rock of the roadway during gob-side entry driving is proposed, and the pressure relief mechanism in the surrounding rock during the static and dynamic pressure stages of the hole-making process is elucidated: the hole-making space transfers the peak stress to the deep during the static pressure stage, and it transfers the high stress to the deep during the dynamic pressure stage. The pressure relief laws of the surrounding rock were analyzed for different hole-making parameters during the static and dynamic pressure stages of deep gob-side entry driving, the hole-making parameters of the 11,216 headgate were as follows: the hole-making depth, length, and spacing were 8, 4, and 3.2 m. A "four-in-one" double space monitoring method was developed to evaluate the control effect of the support structure on the surrounding rock. The results indicated that the surrounding rock deformation remained within a controllable range, the support structure did not experience any breakage or detachment throughout the service period, and the hole-making space played a pressure relief role. The successful application of this technology to control the large deformations in the surrounding rock of coal roadways in deep mines via gob-side entry driving is of great scientific significance. Highlights: The influencing factors of surrounding rock stability during gob-side entry driving in a 1000-m-deep mine are of two types: stress-field factors and stress-related factors. During gob-side entry driving, the two-stage pressurization of the surrounding rock in the 1000-m-deep mine is influenced by the adjacent gob and front mining. The collaborative control of the surrounding rock of the roadway of a 1000-m-deep mine during gob-side entry driving can be facilitated using external anchor-internal unloading. Clarified the principle of pressure relief during the static and dynamic pressure stages of gob-side entry driving in the 1000-m-deep mine. A capsule-like system can monitor the deformation of the nonvisible spaces in pressure relief holes. [ABSTRACT FROM AUTHOR]

Excessive strain energy accumulation in the coal seam is one of the essential conditions for rockbursts to occur. It is imperative to understand the parameters that affect the strain energy accumulation in retreating longwall mining to optimise the design and minimise rockbursts occurrences. In this paper, new analytical models were developed to calculate the strain energy accumulation considering the current state-of-the-art longwall machinery being used in the industry. Seven parameters, i.e., mining depth, length of the cantilever roof, coal seam thickness, thickness of the immediate roof, Young's modulus of coal and the roof, and Poisson's ratio of the coal seam, were identified as the parameters affecting the strain energy accumulation. A detailed statistical analysis of the parameters was conducted using the orthogonal testing method, which revealed that mining depth, Young's modulus of coal and the coal seam thickness significantly influence the strain energy accumulation within a 99% confidence interval. Highlights: An analytical model to calculate strain energy accumulation in longwall mining panels. Excellent agreement between analytical and numerical models. Parametric dominance was determined using the orthogonal testing method. Easy to use equations to calculate rockburst risk in longwall mining panels. [ABSTRACT FROM AUTHOR]

The Double-Yield (D-Y) model is a widely used method for predicting the mechanical properties of caved rock. However, determining the numerical simulation parameters for this model typically requires extensive trial and error. In this paper, a new approach for parameter inversion of the D-Y model is proposed and applied to a real case study involving repeated mining in multiple coal seams. The proposed framework combines FLAC3D, a finite element numerical simulation software, with a new computational tool developed using MATLAB® App Designer. By applying this framework to the case study, the accuracy and efficiency of parameter determination in the mining area are improved. In addition, a 3D finite element model of multiple coal seams is generated based on the geological information of the case study, and the stress evolution is analyzed. The simulation results are then validated through field measurements, providing valuable reference information for safe and efficient production. Overall, this improved method for parameter inversion of the D-Y model, along with the combined numerical simulation and field measurement approach, enhances the understanding of stress evolution in multiple coal seams and contributes to safer and more efficient mining operations. Highlights: An in-depth analysis was conducted to examine how variations in Double-Yield parameters impact the numerical simulation curve. A breakthrough approach for parameter inversion of the Double-Yield (D-Y) model is introduced, and the computational tool is developed. A 3D finite element model is crafted to analyse the dynamic stress evolution during repeated mining of multiple coal seams. The accuracy of simulation outcomes is bolstered by validation against field measurements of mining-induced stress. [ABSTRACT FROM AUTHOR]

Microseismic monitoring technology has developed rapidly in recent years, and effectively evaluating the risk of water inrush from coal seam floor using microseismic monitoring is a research method with development potential, which is of great significance for ensuring the safe and efficient mining of coal resources. Based on the research background of the threat of Ordovician limestone water inrush during the mining process of the 9212 working face in Dongpang Coal Mine, this paper proposes a new method for evaluating floor spatial failure characteristics and water inrush risk based on microseismic monitoring. First, three research stages are established according to theoretical analysis and comprehensive geophysical exploration, and the response characteristics and temporal and spatial distribution of the microseismic events in the different stages are analysed. Second, the weighted coefficient of variation ( CV W ) and three floor factors (S, e , and θ ) are introduced as new risk evaluation indices for the risk of floor water inrush in the vertical direction. The comprehensive weight is determined based on the analytic hierarchy process (AHP) and the entropy weight method, and then the water inrush evaluation model is established by the vulnerability index method. For the water inrush risk of the floor in the horizontal direction, the stratified cumulative kernel density estimation method is employed to determine the degree of water inrush risk and distribution location of different layers from the perspective of microseismic frequency and microseismic energy. Finally, the potential location of the water-conducting channel in the three-dimensional space of the floor is successfully predicted in the analysed working face, and targeted grouting reinforcement and other treatment measures are effectively implemented to ensure the safe mining of the working face. Highlights: Based on microseismic monitoring, a new universally applicable method for analysing the characteristics of floor failure and evaluating the risk of water inrush is proposed. The Weighted Coefficient of Variation ( CV W ) and Three Floor Factors (S, e , and θ) are introduced to analyse the migration characteristics of floor strata and used as the new indices to evaluate the risk of water inrush in vertical direction. The method of stratified cumulative energy kernel density and stratified cumulative frequency kernel density evaluation (SC-KDE) is introduced to comprehensively determine the connectivity degree and range of transverse fractures in the horizontal direction. The potential location of the water-conducting channel in the floor space has been successfully predicted in field engineering. [ABSTRACT FROM AUTHOR]

Coal and gas outburst is a type of coal mine dynamic disaster with complex causes and serious consequences. In China, the number and scale of outburst disasters are among the highest in the world. It is an inevitable stage for outburst that the mechanical instability of the combination of intact coal and tectonic coal (CICTC) in strata. The dip angle of the interface has an important influence on the instability of CICTC-containing gas. To further understand the instability process of CICTC, the uncoordinated deformation of CICTC was defined. According to this, three strain conditions were delimited. The variations of stress under three strain conditions were analyzed, the expressions of additional interfacial stress were established. Then, based on the Mohr–Coulomb criterion, the strength characteristics and failure forms were analyzed. It was found that the dip angle deflects the principal stress in the interface, and the influences of additional stresses on the value of principal stress are opposite. These two reasons make the principal stress state and instability process of the combination complicated. Finally, combined with actual parameters, the instability processes of CICTC under different stress paths and buried depths were studied. It can be found that the failures of CICTC under different stress paths were concentrated appearing in the tectonic coal body, intact coal interface, and tectonic coal interface. With the increase of buried depth, the initial maximum principal stress increases and the reduction from additional stresses to the strength of intact coal interface enhances, which makes the combination damage easily and the failure form change. The research results in this paper can provide a theoretical basis to prevent and control the outburst of the combination containing gas considering the dip angle of the interface. Highlights: The uncoordinated deformation of combination considering the dip angle was first defined. The expressions of additional interfacial stress of combination considering the dip angle were established The influences of different stress paths and buried depths for the instability of combination were discussed quantificationally. [ABSTRACT FROM AUTHOR]

The instability control of deep mining roadway is currently dominated by ejection and grouting, as well as floor support, which contradicts the economic and excavation efficiency of production. Taking the failure site of Xin'an Coal Mine in China as an example, this paper systematically analyzes the failure characteristics of supporting components, crack propagation, and surrounding rock deformation in situ, and the impact of failure on the rights and interests of staff, enterprise benefits, and low-carbon environmental protection. Then, a three-way continuous anchorage theory and the derived control technology are proposed based on the exploration of the instability mechanism. Besides, the UDEC Trigon model is adopted to reproduce the instability process and demonstrate the effect of the proposed control technology. According to the subsequent industrial test, the plastic deformation developing from shallow to deep parts leads to the bending, subsiding, swelling, and deformation of rock mass within 5.5 m of the roof, and the surrounding rock failure mainly features the shear failure in the middle of the tail of the supporting components and anchorage relaxation. Meanwhile, it wrecks huge havoc on the rights and interests of staff, enterprise benefits, and CO 2 emission. Following the smooth application of the new technology to roadway excavation, the roof subsidence and two side displacements are reduced by 91.18% and 58.5%, respectively, compared with the original scheme, the rock mass fractures within 1.71 ~ 3.92 m of the roof is effectively sealed, the evolution depth of bed separation is down by 78.6%, and the excavation efficiency rises by 50% while ensuring the supporting cost. In addition, the defects of the new technology in engineering application are also discussed. Accordingly, the R&D and application of this new technology provide a positive reference for the efficiency control of deep mining roadway. Highlights: A three-way continuous anchorage theory and relevant technique are proposed. Multiscale original position analyses are conducted on the failure features of deep coal roadway. The impact of surrounding rock instability on sustainability was considered for the first time. The new scheme reduces roof subsidence by 91.2% and sides' displacement by 58.5%. [ABSTRACT FROM AUTHOR]

In the N00 coal mining method, one of the most critical issues is ensuring the safety of the entry behind the working face. To address this challenge, various roof control techniques, such as precise directional roof-cutting technology (PDRCT), constant-resistance cable supporting technology (CRCST), and dynamic pressure-bearing supporting technology (DPBST), have been developed. However, the design method for these technologies has not been clearly defined, resulting in a lack of theoretical basis for parameter selection. This paper seeks to address this gap by further studying the mechanisms of PDRCT, DPBST, and CRCST in the N00 coal mining method, and proposing corresponding mechanical calculation models, design ideas, and calculation methods for these technologies. The results reveal that the primary function of PDRCT is to expand the height range of the caved rock mass in the mined area, thus reducing the subsidence space of the high-level roof and mitigating the dynamic pressure effect of the entry surrounding rock. The height of the roof cutting is crucial to the design, and its theoretical value must satisfy the condition that the broken expansion volume of the caved rock mass is equal to the volume that is mined. DPBST is essential for protecting the entry immediate roof and preventing internal damage during the movement of the upper roof. The dynamic pressure-bearing support strength should be appropriately designed to prevent the short cantilever roof structure from breaking within the entry's width during the movement of the high-level roof. Similarly, CRCST's most critical function is to maintain the long-term stability of the entry immediate roof by relying on the cable support forces and the stability properties of the high-level roof strata while it is not moving. The anchor cable support strength should be reasonably designed to prevent the short cantilever roof structure from collapsing under its own gravity. The proposed design method was applied to the Ningtiaota coal mine, leading to a 15.6% reduction in maximum roof deformation and a 12.0% decrease in average deformation, demonstrating the favorable impact of the approach. Highlights: The mechanisms of precise directional roof cutting technology (PDRCT), constant-resistance cable supporting technology (CRCST), and dynamic pressure-bearing supporting technology (DPBST) in the N00 coal mining method were revealed. Corresponding mechanical calculation models, design ideas, and calculation methods for the key parameters of PDRCT, DPBST, and CRCST were obtained. [ABSTRACT FROM AUTHOR]

The grand-scale suspended roofs overlying the goaf are commonly developed under thick-hard rock strata (THRS) during the mining process of the deep coal seam. As a result, strong mine seismicities triggered constantly by deep mining have impeded severely the safety and efficient operation of deep mines. The working-face deep-hole blasting technology has been utilized to weaken the roof for releasing the mine pressure, thereby reducing the risk of violent dynamic hazards. In this paper, Dongtan deep coal mine is taken as an example, the radius of rock failure regions induced by blasting is calculated based on the expansion theory of cylindrical charges. LS-DYNA and UDEC7.0 softwares are employed to simulate the rock pre-splitting and overburden migration under the working-face deep-hole blasting, and the roof pre-splitting and pressure relief mechanism of the working-face deep-hole blasting are revealed. According to the micro-seismic, support working resistance and real-time stress monitoring results, the working-face deep-hole blasting technology could actively break the integrity of the roof. Using the blasting, high-energy mining-induced seisms are decreased dramatically or disappeared fully. The advanced abutment pressure of the coal seam decreases with the increase of the working-face deep-hole blasting time, which indirectly indicates that the weakening degree of the thick-hard roof gradually increases under the action of multiple blasting. The working-face deep-hole blasting technology has achieved the desired effect in the pressure relief in hard suspended roofs for preventing and controlling strong mining-induced seismicity. Highlights: The distribution of mining-induced seismicities before and after the working-face deep-hole blasting is revealed by the micro-seismic system. The rock pre-splitting and overburden migration influenced by the working-face deep-hole blasting are numerically simulated. The effectiveness of strong seismicity prevention and control is systematically evaluated by a series of monitoring technology. The working-face deep-hole blasting technology could achieve the desired effect in the pressure relief of roofs. [ABSTRACT FROM AUTHOR]

Rock and gas outbursts occur most frequently in coal mining. Gas-induced geodynamic events in dolomite are a new and unexplored phenomenon. In the last time, two outbursts of gas and dolomite have occurred in the Legnica-Głogów Copper Region in southwestern Poland, which is one of the world's largest copper ore mining regions. Gas-induced geodynamic events are dependent on the co-occurrence of two factors: rock porosity, and the presence of gas under significant pressure in the pores and crevices of the rock. This paper presents a condensed energy balance for the phenomenon of outbursts of dolomites and gases. The quantity of energy depends on the type of transformation taking place during decompression. The work produced enables first the disintegration of the rock, and then its transportation along the mine working space. This paper has focused on the estimation of the energy required to disintegrate the rock. It has been estimated that 0.38 MJ is needed for the size reduction of 1 m3 of ejected mass. Highlights: In the case of a dolomite outburst, the energy is accumulated primarily in the form of compressed gas present in the porous structure of the rock and causes the rock to disintegrate and its transport along the mine working space. Gas decompression during a gas-induced geodynamic phenomena can be between isothermal and adiabatic. Existing disintegration theories: Rittinger's, Kick's and Bond's are based consists chiefly of crushing, compression or abrasion, caused by the action of external factors. Author's theory of disintegration of research on tensile stress and is based on Brazilian tests. The disintegration of 1 m3 of rock requires the production of energy of approximately 0.38 MJ. [ABSTRACT FROM AUTHOR]

Fractures have significant impact on the stability of underground projects such as coal mining. The identification of geological features based on drilling parameters is a promising option for intelligent detection of rock formations. A series of studies conducted to identify fractures under unstable drilling conditions are presented in this paper. First, the forces acting around the drill bit are analyzed to further improve the force model of the two-wing PDC drill bit. Then, based on the self-developed borehole drilling device, the drilling parameters are collected in real time while drilling in concrete of different strengths and different fracture widths. Finally, the response characteristics of drilling parameters are analyzed, and the multi-parameter voting method is proposed for fracture identification. The experimental results show that when fracture width increases, the crushing area, which forms around the borehole when the drill bit encounters the fracture, tends to increase. The rate of penetration (ROP) increases suddenly, the revolution per minute (RPM) decreases, and the torque increases at fracture encounter. The sudden changes are recovered after passing the fracture. The multi-parameter voting method has a high recognition rate for fractures of width greater than 2 mm, but a relatively low recognition rate for fractures of 1 mm width. In addition to detecting fractures, the method performs well in predicting fracture location and fracture width. Highlights: The response characteristics of multiple drilling parameters at the fracture were analyzed. As the thrust (torque) increases, the amount of change in penetration rate caused by the unit change in thrust (torque) decreases. Results show that the proposed multi-parameter voting method has a high recognition rate for fractures with width greater than 2 mm. [ABSTRACT FROM AUTHOR]

A thorough excavation of the deep ultra-thick mines is highly challenging in rock engineering, which mainly depends on the rheological properties of coal-rock mass, high-stress concentrations, and complex geological conditions. Under heterogeneous settings, a detailed understanding of microseismic (MS) spatio-temporal response to rock mechanical behavior is essential for the efficient and safe-yield of deep ultra-thick mines. This paper utilizes real-time passive seismological data of 36 months in an ultra-thick coal mine through LLTCC mining to quantify the coal-rock mass mechanical behavior. In the post-processing of geophysical data, Power Spectral Density (PSD) was performed and the MS source parameters were computed containing over 18,000 events. The events were classified into different energy-levels to evaluate the increased-pressure rock mass. The Gaussian results reveal that the distribution of MS events coincides with the increased rock pressure. The high energy events (≥ 50,000 J) accounted for 57.93% of the microseismic-induced earthquakes. Moreover, the cloud energy density maps were obtained to identify the zones of high-stress concentration and enhanced seismicity. It is inferred that the overburden pressure and intense fracturing developed the high-stress concentration zones during excavation. Furthermore, the temporal parameters computed from MS data show that sharp-rise and sharp-drop variation can be regarded as early warning indicators for increased rock pressure. High deviation in energy and frequency also evidenced the increased rock pressure and micro-fractures to macro-fracture development. Based on these precursory parameters, a comprehensive early warning method was proposed. Besides, 3D visualization of the fracturing process and identification of new faults assisted in the accurate evaluation of the ultra-thick mine. Several risk zones were identified based on the seismic energy and fractures correlation. Finally, with the help of a deep learning LSTM approach, a prominent peak in time-series data is predicted that defines the increased rock pressure and further characterizes the coal-rock mechanical response. Compared with the existing studies, our integrated geophysical and mine engineering workflow can provide accurate insights to depict the rock mass mechanical behavior of the deep mines as an aid for disaster prediction and early warning. Highlights: A detailed understanding of microseismic Spatio-temporal response to rock mechanical behavior is presented. Field recorded microseismicity is quantitatively correlated with regions of increased-rock pressure and stress concentration. Temporal trends in the microseismic source parameters evidenced increased-rock pressure and a comprehensive early warning model is proposed. A novel microseismological approach is presented to model the mining-induced fracturing process and correlated with seismic energy to identify risk zones. Machine learning based LSTM approach pioneered and resulted in the prediction of rock mass response on time-series data. [ABSTRACT FROM AUTHOR]

High-intensity longwall top coal caving (LTCC) mining of super-thick coal seam in high stress conditions often leads to significant movement and failure of overburden. The competent roof above the super-thick coal seam is easy to form large unstable structures, and the breaking and movement of the roof strata often induce high weighting strength and large dynamic disasters which severely affect the mine safety. The key to control the behavior of competent roof is to understand the fracturing law of the roof strata. In this paper, a case study based on microseismic (MS) monitoring was performed at LW8309 working face in Tongxin coal mine, where super-thick coal seam under competent sandstone roof in high stress conditions was mined by LTCC method. The spatial distribution characteristics of MS events were studied and four impact zones in the horizontal direction and 'four-zones' in the vertical direction were identified. Multifractal analysis of MS events was carried out to study the multifractal characteristics of the fracturing of competent overburden. In addition, a large and small periodic law of fracturing of competent roof was revealed from the temporal distribution of MS events and on-site monitoring of hydraulic support pressure. A key stratum (KS) structural model was then proposed to explain the strong behaviors of competent strata during high-intensity mining. The research results are of great significance for safe and efficient mining of super-thick coal seam under competent roof in high stress conditions. Highlights: The impact zones in front of the mining face and four vertical zones were identified and the fracture characteristics of the competent overburden in the impact zones and the vertical zones were revealed. Multifractal characteristics of the fracture of the competent overburden were obtained. A large and small periodic law of fracture of competent overburden was revealed from the temporal distribution of MS events and on-site monitoring of hydraulic support pressure. A key stratum (KS) structural model was proposed to explain the behavior of the fracture of competent overburden. [ABSTRACT FROM AUTHOR]

Landslides may be triggered by a variety of external factors and can lead to dramatic human and economic consequences. Systematically investigating all possible parameters that could potentially influence the extent of a landslide is costly and generally not practical. In this paper, a methodology to numerically assess the impact of potential landslides is proposed. In this approach the landslide simulations were carried out using finite element model estimates of potential failure volume as an input into a discrete element model which is used to assess run-out and consequences. A statistical design of experiments was implemented to determine the most important factors influencing the landslide extent thereby significantly reducing the number of simulations required. The method is illustrated on a case study of potential failure of an overburden dump at a coal mine. Over the parameter ranges considered, the particle–particle (rock block to rock block) friction coefficient and the volume of failing material were determined to be the most important factors influencing the extent of the landslide. A systematic analysis of varying the particle–particle friction coefficient was then undertaken to better understand the dynamics of the collapse and different types of collapses were identified between low and high inter-particle friction coefficients. The methodology proposed in this paper should be of interest to practitioners as a way to thoroughly yet efficiently identify and assess the main factors influencing a potential landslide on sites at risk. [ABSTRACT FROM AUTHOR]

Data integrity and reliability during seismic event detection are the main factors limiting the effectiveness of seismic monitoring systems in underground mines. As it is impossible for current seismic monitoring systems to record all mining-induced seismic events, the incomplete seismic dataset may cause significant bias during data analyses and interpretation processes. Evaluation of seismic data integrity in mine seismology and eliminate its impact on seismic analyses is a critical issue that needs to be addressed. Therefore, this paper presents the results of a study that investigated the characteristics of seismic data and its integrity by assessing the event detection probability of the seismic monitoring system in an underground coal mine. The wave picking capacities of individual geophones are first evaluated, and the detection probabilities for seismic events within a specific monitoring area are then calculated. The results showed that geophones presented different wave picking capacities for seismic events at various locations and energy magnitudes: a larger event detection probability and a broader detection range is observed for higher energy seismic events. A method to correct seismic data is also proposed to improve the completeness of the recorded seismic data, which provides more information on the frequencies of seismic events than the raw data. This paper enhances the method of assessing seismic data reliability and improving the accuracy of seismic analyses in underground mines. [ABSTRACT FROM AUTHOR]

Mine tremors and rockbursts triggered by thick-hard roof fracturing are major problems in safe production of coal mines, and the corresponding numerical simulation is also a great challenge due to complexity of rock failure and wave field characteristics radiated from sources. In this paper, based on the engineering background that strong mine tremors frequently occurred in the 7301 working face of the Baodian coal mine due to thick-hard stratum fracture, a novel method is used to successfully simulate the whole process of tremor generation caused by thick-hard roof fracturing and triggered rockburst in a roadway. Much interesting information is studied in detail, including the characteristics of the source radiation wave field, the variation of the stress field caused by seismic wave diffusion, and the changes in stress, crack and displacement throughout the process of the roadway rockburst from inoculation to explosion under the disturbance of seismic waves. The effects of tremor disturbance intensity are studied based on the relationship between the relative spatial orientation of the roadway and fault plane. The disturbance to the roadway is largest in the direction of 45° of the fault plane. The research results provide a useful reference for understanding the mechanisms of rockburst induced by mine tremors. The simulation method gives a valuable insight into the disturbance effects and focal mechanism of mine tremors. Highlights: Characteristics of source radiation wave field of mine tremors induced by shear fracture are numerically simulated. Influences of fault plane geometric properties, radiation wave field, and relative position between roadway and source on rockburst are discussed. Rupture types of thick-hard stratum in different areas are investigated by moment tensor inversion. [ABSTRACT FROM AUTHOR]

Fully grouted rock bolts are one of the main forms of rock reinforcement used in underground mines, especially in coal mines. Analytical models describing the deformation behaviour of rock bolts under various loading conditions are used in numerical models for designing appropriate rock reinforcement. A reliable analytical model capable of accurately predicting rock bolts' responses under different loading conditions is, therefore, important for more realistic numerical models and effective reinforcement design. In this paper, a new analytical model for shear behaviour of rock bolts is proposed which can better predict rock bolt behaviour under large shear deformations. A method to consider the effect of post-elastic strain hardening on bending stiffness of rock bolt is presented. The analytical model is validated with results from laboratory shear tests and found to be consistent with the laboratory results. To analyse the effect of post-elastic strain hardening on shear behaviour of a rock bolt, results from the proposed analytical model is compared with analytical model without strain-hardening effect. Without strain hardening, the analytical models overestimate the ultimate shear displacement by up to 35% in the examples shown in the paper. [ABSTRACT FROM AUTHOR]

To investigate the temporal and spatial evolution process of overlying karst mountain instability caused by mining, the case of "8.28" mega landslide in Pusa, Nayong County, Guizhou Province, China, is taken as the example. Based on the results of the field investigation after the disaster, the complete process of rock strata fracture and slope instability under the descending and ascending mining conditions of gently inclined coal seam is reproduced by similitude model experiments. According to the key motion features of overburden deformation and fracture evolution law, the failure process of slope collapse controlled by deep and large karst fissures under mining action is summarized. And the change law of overburden pressure and displacement in the mining process is analyzed as well as the failure mode of slope collapse is proposed. The results show that the slope experiences four failure stages in the process of coal seam descending mining: coal seam roof caving, overburden cantilever tension fracture, overburden integral dumping subsidence and slope instability collapse, and the slope undergoes deformation toward the free surface and downward. During the ascending mining process, the slope experiences four failure stages: periodic caving of coal seam roof, upward transmission of separated fractures, overburden integral bending subsidence and slope instability collapse, and the slope undergoes deformation toward the inside of the slope and downward. The rock mass ahead of the advancing direction of working face has the phenomenon of tension–compression stress transformation. Along the mining direction, the overburden sequentially experiences the change process of pressure increase, decrease, or increase again, and stabilization. Deep and large karst fissures at the slope top are the structural basis for the occurrence of collapse. Coal mining activities have a certain promoting effect on mountain cracking and rock fragmentation. The failure mode of slope can be summarized as tension–shear–slip–toppling collapse failure. Improved understanding of the deformation movement characteristics and damage mode of slope can aid in early identification and timely warning of geo-hazards in karst area. Highlights: This paper focuses on the deformation movement features and failure modes of mountains controlled by deep and large karst fissures under mining action. The full process of rock strata fracture and slope instability under descending and ascending mining conditions is reproduced in a simulated model. Extensive karst fissures at the slope top are the main structural basis for the occurrence of collapse. Coal mining activities have a certain promoting effect on mountain cracking and rock fragmentation. The failure mode of slopes can be summarized as tension-shear-slip-toppling collapse failure. [ABSTRACT FROM AUTHOR]

While large-scale rock strata affected by underground coal mining have been widely studied through numerical modeling, there are still some aspects that can be better understood. Specifically, researchers do not fully utilize borehole logs and test results of rock specimens to reveal rock mass property variation along horizontal directions and strata lateral thickness variation at bed level. In this paper, we address this knowledge gap by proposing a data-intensive numerical modeling method (DINMM) that can make full use of these data with consideration of the modeling limitations of BlockRanger, a grid generation tool. Both the proposed method and the conventional numerical modeling method (CNMM) are applied to the Ying-Pan-Hao coal mine via the FLAC3D (Fast Lagrangian Analysis of a Continua in 3 Dimensions) package, and their predictions are then calibrated and compared to discuss the validity. Results show that, compared to the CNMM-based predictions, the root mean square error of 70 monitoring points is decreased at least by 27.4% in the DINMM-based prediction, and the relative error of maximum subsidence is reduced by 5.1% with a reduction rate of 66.5% on average, even though the CNMM-based model was originally better calibrated. We also find that a DINMM-based model is more in line with field observations and theoretical understanding in terms of displacement, stress, and failure propagation. The notion of data-intensive modeling seems to be quite promising and the DINMM should be useful for a better understanding of strata movement and subsidence prediction. Highlights: • We propose a data-intensive modeling method (DINMM) to build numerical models based on multiple borehole logs, rather than on a single or generalized borehole log. • We realized rock mass property variation along horizontal directions and strata lateral thickness variation at bed level when modeling large-scale rock strata. • We find that the root mean square error of 70 monitoring points is decreased at least by 27.4% in the DINMM-based prediction, even though the control model was originally better calibrated. [ABSTRACT FROM AUTHOR]

Heat from gas explosions affects the physical and mechanical properties of rock and reduces the stability of tunnels, such as in coal mines. This paper mainly studies the physical and mechanical properties of sandstone after heating and investigates the influences and mechanisms involved. Acoustic emission (AE) monitoring, X-ray diffraction (XRD), scanning electron microscopy (SEM) and polarized light microscopy are used to show that the uniaxial compressive strength (UCS) and peak strain of sandstone change significantly with temperature. In particular, the UCS decreases rapidly at 500 ~ 600 °C due to a transition from α -quartz to β -quartz. The XRD analysis shows that the physical and mechanical properties of sandstone are affected by its mineral composition, as physical and chemical reactions of quartz, muscovite and kaolinite at high temperatures weaken the sandstone. The AE signals increase with temperature and correspond well with observed loading stages, as they reflect the spatiotemporal evolution in internal micro-cracks. The degree of crack opening is low at low temperatures, then cracks become connected with increasing temperature, further explaining the changes in mechanical properties. Polarized light microscopy shows that grain size and pore characteristics are related to mechanical strength; generally, higher strength is associated with smaller grains with a less even size distribution. In addition, thermal and mechanical damage were quantitatively characterized by deriving thermal, comprehensive and dynamic damage factors. The results provide a theoretical reference for mining engineering in high-temperature environments. [ABSTRACT FROM AUTHOR]

The deformation and permeability of coal are largely affected by the presence and distribution of natural fractures such as cleats and bedding planes with orthogonal and abutting characteristics, resulting in distinct hydromechanical responses to stress loading during coal mining processes. In this research, a two-dimensional (2D) fracture network is constructed based on a real coal cleat trace data collected from the Fukang mine area, China. Realistic multi-stage stress loading is designed to sequentially mimic an initial equilibrium phase and a mining-induced perturbation phase involving an increase of axial stress and a decrease of confining stress. The geomechanical and hydrological behaviour of the fractured coal under various stress loading conditions is modelled using a finite element model, which can simulate the deformation of coal matrix, the shearing and dilatancy of coal cleats, the variation of cleat aperture induced by combined effects of closure/opening, and shear and tensile-induced damage. The influence of different excavation stress paths and directions of mining is further investigated. The simulation results illustrate correlated variations among the shear-induced cleat dilation, damage in coal matrix, and equivalent permeability of the fractured coal. Model results are compared with results of previous work based on conventional approaches in which natural fracture networks are not explicitly represented. In particular, the numerical model reproduces the evolution of equivalent permeability under the competing influence of the effective stress perpendicular to cleats and shear-induced cleat dilation and associated damage. Model results also indicate that coal mining at low stress rates is conducive to the stability of surrounding coal seams, and that coal mining in parallel to cleat directions is desirable. The research findings of this paper have important implications for efficient and safe exploitation of coal and coalbed methane resources. [ABSTRACT FROM AUTHOR]

Mining-induced stresses in underground coal mines play a significant role in pillar and ground support design, hence in the safety of mining operations. In the US, Analysis of Longwall Pillar Stability (ALPS) and in Australia, Analysis of Longwall Tailgate Serviceability (ALTS) software are used for designing Longwall coal mine layouts; and in the US, Analysis of Retreat Mining Pillar Stability (ARMPS) software is used to design retreat room-and-pillar mine layouts. All these software determine the adequacy of the design by comparing the estimated loads to the load-bearing capacity of the pillars and they use the "abutment angle" concept and a square decay stress distribution function to calculate the magnitude and distribution of the mining-induced loads. The abutment angle concept has been successfully applied to US longwall coal mines with the use of ALPS and ALTS in Australia. ARMPS uses the same concept for retreat room and pillar coal mine design in the US. The suggested abutment angle for coal mines in the US was derived as 21° by the back analysis of underground stress measurements from the 1990s and implemented in ALPS and ARMPS. The ALPS methodology was re-examined and calibrated for Australian conditions with additional Australian stress measurements and resulted in the original ALTS methodology which has been continually improved and expanded with additional cases. In this paper, some recent stress measurements are back-analyzed, and the abutment angles are investigated to verify the applicability of using 21° in retreat room and pillar mines with different depths and mining dimensions. For shallow mines, the derivation of the 21° abutment angle is supported by the new case histories. However, at depths greater than 200 m, the abutment angle was found to be decreasing with increasing depth. In this study, a new equation for the calculation of abutment angle for moderate and deep cover cases was constructed and tested for its applicability in retreat room and pillar mines. The differences in the mechanism of complete side abutment loads in shallow and deep cover mines are further analyzed by applying the finite volume modeling (FVM) approach to two case study mines, one shallow, and one deep cover. A 2D model of each mine is created and one-side and two-side abutment loads of consecutive panels are analyzed. Analysis of the deep cover mine indicated that the prior panel gobs provide a considerable amount of support to the overburden strata. These higher gob loads prevent a higher percentage of overburden loads from being transferred to the active panel workings, and this is in agreement with the lower abutment angles observed for deep cover mines. The findings of this study should only be used for retreat room-and-pillar mines' production pillar loads since these are calculated geometrically using the abutment angle concept. [ABSTRACT FROM AUTHOR]