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The roof water inrush disaster induced by coal mining is becoming a vital bottleneck restricting mine safety. To accurately predict the water inrush position of the coal seam roof sandstone aquifer and accurately prevent and control it, this paper takes Lilou Coal Mine in Juye Coalfield as an example. Based on the comprehensive analysis of hydrogeological data, drilling data, and geophysiological data, this paper examines the water richness of the roof aquifer, the water insulation and geological structure, and the fracture development characteristics of the roof aquifer. Starting from the general side, nine factors, including fault strength index, tectonic intersection point, tip extinguishing point, development height of water conduction crack zone, roof sandstone aquifer thickness, aquifer drilling unit inrush, geophysical prospecting water-rich anomaly area, roof key layer thickness, roof aquifer thickness, brittle plastic rock thickness ratio, etc., form the roof plate roof. The main control factors for water inrush are deeply discussed. Through the spatial analysis function of GIS, unique drawings of different evaluation indicators are drawn, and the data are normalized. Combined with the AHP hierarchical analysis method, the corresponding weight is determined. Finally, a comprehensive water inrush risk assessment map of the roof of coal seam 3 in Lilou Coal Mine is obtained. Through the verification of the coal mine water inrush survey ledger over the years, it has been found that the evaluation results of the coal seam roof water inrush model are consistent with the actual situation. The evaluation results are reasonable and accurate, which can provide a reference basis for coal seam mining and water damage prevention and control in the future. [ABSTRACT FROM AUTHOR]

Copyright of Mine Water & the Environment is the property of Springer Nature and its content may not be copied or emailed to multiple sites or posted to a listserv without the copyright holder's express written permission. However, users may print, download, or email articles for individual use. This abstract may be abridged. No warranty is given about the accuracy of the copy. Users should refer to the original published version of the material for the full abstract. (Copyright applies to all Abstracts.)

Intelligent computing is transforming safety inspection methods and response strategies in coal mines. Due to the significant safety hazards associated with mining excavation, this study proposes a multi-source data based predictive model for assessing gas risk and implementing countermeasures. By examining the patterns of gas dispersion at the longwall face, utilizing both temporal and spatial correlation, a predictive model is crafted that incorporates safety thresholds for gas concentrations, four-level early warning method and response strategy are devised by integrating weighted predictive confidence with these correlations. Initially tested using a public dataset from Poland, this method was later verified in coal mine in China. This paper discusses the validity and correlation of multi-source monitoring data in temporal and spatial correlation and proposes a risk warning mechanism based on it, which can be applied not only for safety warning but also for regulatory management. [ABSTRACT FROM AUTHOR]

Longwall top coal caving technology is one of the main methods of thick coal seam mining in China, and the classification evaluation of top coal cavability in longwall top coal caving working face is of great significance for improving coal recovery. However, the empirical or numerical simulation method currently used to evaluate the top coal cavability has high cost and low-efficiency problems. Therefore, in order to improve the evaluation efficiency and reduce evaluation the cost of top coal cavability, according to the characteristics of classification evaluation of top coal cavability, this paper improved and optimized the fuzzy neural network developed by Nauck and Kruse and establishes the fuzzy neural network prediction model for classification evaluation of top coal cavability. At the same time, in order to ensure that the optimized and improved fuzzy neural network has the ability of global approximation that a neural network should have, its global approximation is verified. Then use the data in the database of published papers from CNKI as sample data to train, verify and test the established fuzzy neural network model. After that, the tested model is applied to the classification evaluation of the top coal cavability in 61,107 longwall top coal caving working face in Liuwan Coal Mine. The final evaluation result is that the top coal cavability grade of the 61,107 longwall top coal caving working face in Liuwan Coal Mine is grade II, consistent with the engineering practice. [ABSTRACT FROM AUTHOR]

This study aimed to investigate the internal dislocation law of coal seam working face under coal pillars with different widths and evaluate the safety of using the combined mining of close-distance coal seams under different layer spacing. Based on elastic–plastic mechanics and mine pressure theory, theoretical calculation and numerical simulation were implemented to establish a numerical model consistent with the actual production. Based on the theoretical and simulation results, the abutment stress distribution of surrounding rock in coal seam stope was analyzed and the stress transfer law of coal pillar floor under the influence of different coal pillar width and floor depth was studied. Finally, this paper analyzed the limit span of goaf under two different beam structures. The results indicate that with the same coal seam distance, the interior dislocation distance increases linearly with the increase of coal pillar width. With the increase of floor depth, the vertical stress under the coal pillar decreases, but the influence area increases slightly. Finally, the ultimate spans of two different beam structures were calculated. In summary, this paper explored the strata behavior, i.e., the deformation and failure mechanism of surrounding rock in coal seam mining. The research results demonstrated the feasibility of using the combined mining method in close-distance coal seams. The methodology proposed in this paper could be implied in similar working environments, and this study provides new insights into the mining technology in lower coal seams. [ABSTRACT FROM AUTHOR]

Based on the 7618 working face in Yaoqiao coal mine of Datun mining area, the activation mechanism of water-rich faults and the development characteristics of water-conducting fractures in overlying strata under the influence of faults are studied by theoretical analysis, numerical simulation and field measurement in this paper. The research results show that Anderson model and Mohr–Coulomb strength criterion are combined to establish the fault failure mechanical model, and the fault activation criterion under the influence of mining is obtained. FLAC3D numerical simulation results show that with the advance of the working face, the fault begins to be affected by the mining effect of the working face at the distance of 20 ~ 30 m from the fault. Meanwhile, with the advance of the working face, the overburden shear failure range also expands, and the fault fracture gradually expands from top to bottom. The failure zone of the working face roof is connected with the fault fracture zone. Then the fault is "activated" and causes the fault to become a water gushing channel, and finally the water gushing disaster occurs. Through numerical simulation and comparative analysis, the development height of water-conducting fracture is 73.2 m in the absence of fault, and 73.7 m in the presence of fault, indicating that the fault has little influence on the maximum development height of water-conducting fracture. The actual development height of the water-conducting fracture zone in the 7618 working face is 73.97 m and the fracture production ratio is 13.7. The research results can provide theoretical reference for the safe mining of similar working faces across faults. [ABSTRACT FROM AUTHOR]

Cropland is the foundation of food security. Coal is the guarantee of energy security. As China's demand for coal and grain continues to increase, so does the overlap area of their production bases. Unrestrained underground mining can cause serious damage to cropland, leading to increasing conflicts between coal mining and food production. Thus, this paper used a partial backfilling mining technology to control surface subsidence and thus protect cropland. The key to successfully implementing the technology is how to design the panel size. However, the design efficiency of the conventional enumeration method is low. Therefore, this paper proposed a design approach based on improved particle swarm optimization. The results indicated that the approach could quickly find the optimal size of the panel compared with the enumeration method and particle swarm optimization. Moreover, if the longwall panel is mined according to the size designed by the approach, the cropland will be protected, and the cost will be reduced. This study can provide technical support for the cooperative development of cropland protection and coal mining in a coal-cropland overlapping area. [ABSTRACT FROM AUTHOR]

Coal mine water outburst under fluid-solid coupling is a hot topic in coal geology. In this paper, the FLAC 3D simulation method is used to systematically study the water outburst problem of fluid coupling with coal mine floor. The results show that the vertical stress of the floor in the goaf has an obvious downward trend after coal seam excavation. Under the joint influence of mine pressure and bearing water pressure, the working face before the old roof pressure can be affected by two stages: the first stage is the advance pressure compression section on both sides of the working face, and the other stage is the pressure relief expansion section under the goaf. Before the old roof is pressed, the shear stress near the roof and floor is also affected to some extent. The front and rear displacement value of the goaf is negative, which is the result of the advance supporting pressure. The effective thickness of the floor waterproof layer decreases gradually with the advance of the working face. When the working face continues to advance, there will be a risk of water outburst. With the increasing width of the working face, the height of the guide belt increases. Therefore, in the process of mining activities, a variety of factors should be considered comprehensively, and a reasonable width of the working face should be selected to ensure the safety of mining. [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]

Automatic roadway formation by roof cutting is a sustainable nonpillar mining method that has the potential to increase coal recovery, reduce roadway excavation and improve mining safety. In this method, roof cutting is the key process for stress relief, which significantly affects the stability of the formed roadway. This paper presents a directionally single cracking (DSC) technique for roof cutting with considerations of rock properties. The mechanism of the DSC technique was investigated by explicit finite element analyses. The DSC technique and roof cutting parameters were evaluated by discrete element simulation and field experiment. On this basis, the optimized DSC technique was tested in the field. The results indicate that the DSC technique could effectively control the blast-induced stress distribution and crack propagation in the roof rock, thus, achieve directionally single cracking on the roadway roof. The DSC technique for roof cutting with optimized parameters could effectively reduce the deformation and improve the stability of the formed roadway. Field engineering application verified the feasibility and effectiveness of the evaluated DSC technique for roof cutting. Highlights: A directionally single cracking (DSC) technique for roof cutting in automatic roadway formation of the non-pillar mining method is proposed. The mechanism of the DSC technique to achieve single cracking relies on effective control of the blast-induced stress. DEM simulation and field experiments are effective methods to evaluate the DSC and roof cutting parameters to improve the stability of the roadway surroundings. [ABSTRACT FROM AUTHOR]

Emulsion pump station is widely used to provide power for full-mechanized coal mining face equipment. With the development of the mining technique of longwall face and the extensive use of large mining height and high working resistance of the hydraulic support, the fluid supply pressure is more unstable and the pressure impact is increasingly intensified. The intelligent control of the emulsion pump station is the key to the development of the stable fluid supply technology at the working face. This paper develops a new on-line adaptive learning technique to improve the response speed of pumping station based on the optimal control algorithm of adaptive dynamic programming (ADP). The state space model of the working face fluid supply system in the process of no-load column lifting of hydraulic support is first established using the flow relationship, and the effectiveness of the proposed control algorithm is then verified by simulation validation. Moreover, combined with a certain fluid supply system of a fully mechanized caving face in Shandong Energy Xinwen Group XinJulong Co., Ltd, a technical scheme of adaptive stable fluid supply is verified in engineering practice. The results indicate that the fluid supply system using the proposed control algorithm has a favorable tracking effect on the ideal trajectory, and the tracking error can converge to near zero. The output pressure of the fluid supply system using the stable fluid supply technology is more stable, and the practical effect is good, which meets the demand of stable fluid supply in the mining face. [ABSTRACT FROM AUTHOR]

Underground mining of minerals is accompanied by a change in the rock mass geomechanical situation. This leads to the redistribution of stresses in it and the occurrence of unexpected displacements and deformations of the earth's surface. A significant part of the civil and industrial infrastructure facilities are located within the mine sites, where mining and tunneling operations are constantly conducted. Irrational planning of mining operations can lead to loss of stability and destruction of undermined facilities. Therefore, it is important to study the earth's surface deformation processes during mining operations, which ensures safe and sustainable operating conditions. The research objective of this paper is to analyse the behaviour of a natural gas pipeline under the influence of underground mining activities, with a particular focus on understanding the effects of horizontal surface deformations and their potential impact on pipeline safety and structural integrity. Its performance and safety are determined on the basis of the found parameters of the earth's surface horizontal deformations and their comparison with permissible parameters characterizing the conditions for laying pipelines, depending on the mining-geological conditions and the degree of their undermining. Based on determined conditions for the safe undermining of the natural gas pipeline, it has been revealed that in its section between the PK212+40 and PK213+80 (140 m) pickets, the estimated parameters of the earth's surface horizontal deformations exceed their permissible values. This can cause deformation and damage to the pipeline. For the safe operation of the pipeline during the period of its undermining, in order to eliminate the hazardous impact of mining the longwall face, additional protection measures must be applied. It is therefore recommended that the gas pipeline between the PK212 and PK214+20 pickets be opened prior to the displacement process (200 m from the stoping face), thus reducing the density of the gas pipeline-soil system. Recommendations for controlling the earth's surface deformations within the natural gas pipeline route are also proposed, which will ensure premature detection of the negative impact of mining operations. [ABSTRACT FROM AUTHOR]

The primary key stratum (PKS) is an important research object in ground control, and all strata up to the surface are controlled by it. However, the geological conditions of coal mining are complex, and coal mining under different geological conditions will cause different degrees of impact on PKS. This paper takes the Meihuajing coal mine as an example, theoretical analysis, numerical simulation and geological condition analysis are comprehensively applied to study the subsidence mechanisms of PKS under the influence of different factors. The results show that subsidence of the PKS is directly proportional to the coal thickness (M) and buried depth (D), and inversely proportional to the distance between the coal seam and the primary key stratum (H). The order of the disturbance degree with each factor on the maximum subsidence of the PKS is M > D > H. It is demonstrated that the disturbance impact of coal seam mining on the PKS in an area with thick coal seams needs to be paid more attention. The research results guided adjusting and reducing the disturbance impact of mining on the PKS. [ABSTRACT FROM AUTHOR]

Shearer is an essential equipment in mechanized coal mining process. Roof cutting often turns up for shearer during working process, leading to extra vibration on shearer and reducing its service life. Coal and rock interface recognition is a fundamental technical way for shearer to avoid continuous roof cutting. In this paper, an effective method for coal and rock interface is proposed and it is mainly based on the cutting loads on shearer drum in three different kinds of working states. The constitutions of cutting force on shearer drum are analyzed under conditions of different proportions of rock thickness in the cutting range and then model for calculation of cutting force in the coal–rock mixed seam are proposed which is a function of rock proportion in the cutting range. To verify the model, hard roof with different material strengths and different interfaces are established with the numerical method and rock proportions in the cutting range are derived based on the mentioned method. In terms of the situations where rock material properties are close to coal and the results are obtained with lower accuracy, enlarging the difference of rock proportions in the second and third states is proposed and the relations between the recognition accuracy and the differences of rock proportions in different states are explored as well. The results show that the estimation method works well with qualititative analysis of coal and roof interface even if coal and rock material properties are various and not able to be known in advance. In addition, the adjustment height of shearer drum is the important factors highly related with the estimation accuracy and it would be helpful for accuracy improvement in conditions where the material properties of coal and roof are closer to each other or larger roof cutting thickness happens in coal mining. The method provides the possibility for qualititative analysis of coal and roof interface with easy accessibility and wide adaptability. [ABSTRACT FROM AUTHOR]

Copyright of Mine Water & the Environment is the property of Springer Nature and its content may not be copied or emailed to multiple sites or posted to a listserv without the copyright holder's express written permission. However, users may print, download, or email articles for individual use. This abstract may be abridged. No warranty is given about the accuracy of the copy. Users should refer to the original published version of the material for the full abstract. (Copyright applies to all Abstracts.)

The method of filling mining can solve the problem of surface subsidence caused by coal mining. Among them, it is crucial to study the mechanism of filler strength improvement timeliness and filler mining to control rock movement for filler mining. In this paper, by combining theoretical analysis and similar simulation experiments, compressive strength is used as the research parameter to conduct proportioning test research on paste filling similar materials such as coal gangue, fly ash, and cement. The results prove that the strengths of the test ratios can meet the strength design criteria and lay the foundation for the requirements of similar simulation experiments. In order to study the characteristics of overburden failure, stress and displacement in the process of filling mining, the key technical parameters of overburden movement are determined. Similar simulation experiments were conducted to study the movement and deformation of overburden rock and the displacement and stress distribution law of overburden rock in the coal mine under different filling rates and filling steps conditions. The results show that the filling rate and filling step are the keys to preventing the overlying rock from rupture and collapse, and the larger the filling rate is, the smaller the stress and displacement of the overburden; the larger the filling step is, the larger the displacement and stress change of the overburden, and vice versa. In addition, the displacement curve along the strike is basically an "arch" type distribution, and the stress variation trend is "large-small-large" with a "Z" type distribution. The research results are of great significance to guide the practice of filling mining and can provide the theoretical basis for its further promotion. [ABSTRACT FROM AUTHOR]

To build a digital twin model of intelligent longwall coal mining, this paper presents a new full-parameter position and attitude digital twin model of hydraulic supports. A full-parameter matrix describing the position and attitude of a support is established first. Then, the position and attitude digital twin model of a support based on this matrix is built. Afterward, a position and attitude digital twin model of the hydraulic support system in the workface is established using the Unity3D software and the Runge–Kutta algorithm as a motion calculation method. Finally, a simulation experiment is conducted based on the digital twin model of hydraulic supports, and results show its effectiveness. This digital twin model has the potential to control the hydraulic support system in future intelligent coal mining. [ABSTRACT FROM AUTHOR]

To explore the stability control mechanism of surrounding rock during coal seam mining, the rock mechanical properties under the mining stress path were studied in this paper. The study begins by summarizing the mining roof and floor triaxial stress (MRTS and MFTS) paths through numerical simulations. The MRTS path is characterized by axial loading and confining unloading, while the MFTS path involves axial loading and confining loading. Subsequently, triaxial tests were conducted on rocks subjected to the mining stress path. The results indicated that compared to the standard triaxial stress (STS) path, the peak strength, peak axial strain, elastic modulus, and Poisson's ratio under the MRTS path were lower, whereas those under the MFTS path were higher. The MRTS path exhibited a more forward dilation deformation, rendering it more susceptible to damage, whereas the MFTS path displayed a more stable and backward dilation deformation. Tensile failure was primarily observed under the MRTS path, whereas the MFTS path resulted in shear failure. Furthermore, the study analyzed the variation of rock permeability with axial and volumetric strain. The relationship between permeability and axial strain follows a √ shape, initially decreasing and subsequently increasing. The relationship between permeability and volumetric strain is found to be complex. Before the dilation onset point, the permeability decreases under both the STS and MFTS paths, while it decreases and increases slowly under the MRTS path. After the dilation onset point, the permeability increases under the STS and MRTS paths but initially decreases and then increases under the MFTS path. The research findings provide valuable theoretical guidance for mining strata control and water conservation mining practices. [ABSTRACT FROM AUTHOR]

The severe deformation of surrounding rock and the easy failure of the anchoring structure of roadway under large dynamic pressure influence in the short distance are the urgent issues to be paid more attention. Based on this problem, this paper is a case study on the instability mechanism and control technology of surrounding rock in the deep roadway under dynamic mining loading in the Gubei Coal Mine of China. The vertical distance from the − 575 m roadway of 11–2 mining area and mining working face is only 68 m. Firstly, the main factors influencing the surrounding rock's severe deformation are determined by the analysis of site investigation, surrounding rock composition and structure, overlying strata failure measurement, and borehole camera detection. The stress distribution, displacement distribution, plastic zone, and lining stress distribution of roadway of three reinforcement schemes, including nine working conditions, were calculated by numerical simulation for ordinary bolt support, bolt-support + grouting of roof and two sides, and bolt-support + full section grouting. Then the combined support technology system of "36 U-shaped steel + bolt + anchor cable + grouting" was proposed. The field monitoring results showed that the deformation of two sides and roof were only 20 mm and 16 mm, respectively, during the 65 day monitoring period. Furthermore, the cracks in the surrounding rock are almost filled with grouting materials, and the overall stability of roadway surrounding rock was completely achieved. This successful case study demonstrated that the optimized supporting structure is reasonable and the deformation of surrounding rock has been effectively controlled. [ABSTRACT FROM AUTHOR]

Longwall mining operation is mainly accompanied by the roof fall phenomenon and out of seam dilution (OSD). These occurrences are strongly dependent on the quality of the face's roof and floor. Therefore, the existence of discontinuities in the floor and roof reduces their quality, increasing the intensity of roof fall and OSD. This paper suggests a synthetic rock mass (SRM) model which is able to simulate roof behavior by initiating and tracking of fractures in cave mining, in order to predict the roof fall and OSD and to evaluate their event risk. For this purpose, longwall mining, including excavation of coal, installation of powered support, and caving the roof, has been successfully established in this numerical model using the PFC2D approach. The risk analysis in a panel of Tabas mine using the matrix risk model indicated that the risk level of roof fall is moderate and the risk level of OSD is high. The results of SRM modeling are consistent with previous studies and field observations at the Tabas coal mine, indicating that the SRM model is a powerful technique for studying the behavior of discontinuities in mining. [ABSTRACT FROM AUTHOR]

This paper constructs a geophysical model of hidden disaster-causing anomalies in coal rock for simulation experiments based on the structural design of coal rock seams at the mine working face to study the geophysical response characteristics of geological anomalies in coal rock seams. The material ratio and petrophysical parameters of the coal seam, roof, floor, collapse column, and fault in a mining area in Shanxi were tested using homogeneous and orthogonal test methods, and their petrophysical parameters, such as Vs,Vp, resistivity, permeability, and dielectric constant, were obtained. The coal dust and sand–cement contents are the main factors affecting the wave velocity, whereas the dielectric constant and resistivity are highly influenced by the water content of the material. The response characteristics of the petrophysical parameters of various types of geological structures measured via experiments are consistent with the corresponding geophysical responses of actual geophysical exploration methods. The experimental data were comparatively analyzed with the original coal rock samples and field in-situ probing results, and the best construction ratios of each geological structural unit were as follows: coal seam ratios of 61.11%, 11.11%, and 27.78% were selected for coal dust, sand, and cement content, respectively; sand–cement and gypsum–cement ratios of 2:1 and 2:8, respectively, were selected for the roof; and sand–cement and gypsum–cement ratios of 3:1 and 4:6, respectively, were selected for the floor. The optimal ratios of 75% sand, 10% gypsum, and 15% cement were chosen for the fault; those for the collapse column were 66.7% sand, 6.67% gypsum, and 26.6% cement. This study focuses on the ratios of various geotectonic physical models and their physical properties, it provides data support for subsequent model construction and a good experimental platform for the demonstration of various geophysical exploration methods and techniques. [ABSTRACT FROM AUTHOR]

Since the western region of China, which is typical of extraordinary resource endowments, has gradually emerged as the major mining zone in China, the mining of thick coal seams and roadways with coal-rock composite roof have become more and more common in this region. However, it is extremely difficult to realize safe and effective maintenance and control of such roadways due to the differences in natural endowments of coal-rock masses. With the roadway with coal-rock composite roof of Hulusu Coal Mine in western China as the engineering background, experiment research on large-scale similarity model was conducted through comprehensive measures such as the pneumatic loading system, the surrounding rock stress monitoring system, the roadway deformation monitoring system, the bolt load monitoring system, and the displacement field monitoring system in this paper. According to the results of the experiment, the control effects of the three support systems on the roadway with coal-rock composite roof were significantly different. When the single support of short anchor bolts was applied, the comparatively low initial anchor-hold failed to constrain the initial micro deformation of the roof. Consequently, wide-range fractures of the roof were triggered at a loading pressure of 0.8 MPa. In the meanwhile, the deep surrounding rocks witnessed a downward inflection point in stress, accompanied by the possibility of the collapse of the thin-layer anchorage zone at any time. As for the support combining both short anchor bolts and long anchor cables, though a reinforced effect on the bolt anchorage zone could be achieved with the help of the cables, the active reinforcement capacity of the bolt was limited. The bolt anchorage zone was the first to be damaged at a loading pressure of 0.9 MPa, which would subsequently affect the effective bearing capacity of the deep surrounding rocks. In the application of the single support of high-strength long anchor bolts, the long bolts with high pre-tightening force were able to lock multiple groups of coal-rock strata to form a thick-layer anchorage bearing structure capable of withstanding a load as high as 1.0 MPa. The crash and collapse of the coal wall eventually caused the subsidence of the roof. Based on the intense dynamic load experiment and the feedbacks of engineering application outcomes in the field, it was concluded that the high-pretension thick-layer (HPTL) anchoring technology can effectively constrain the deformation of roadways with coal-rock composite roof with favorable application outcomes. [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]

Development and selection of fuzzy variable weight vector and its application on fuzzy comprehensive decision-making (FCDM) are studied in detail in this paper. Firstly, a parameter contained variable weight vector is developed based on the definition of variable weight vector, and then further extended to obtain the fuzzy variable weight vector. Its capability on weights changing can be adjusted by changing the parameter. In order to make the fuzzy variable weight vector for different changes of states to be selected and applied easily, the concept of fuzzy dispersion degree, which can be used to measure the balance of decision factors' states, is proposed; And then the concept of fuzzy average relative adjustment degree, which can be used to measure the capability of fuzzy variable weight vector on weight changing, is also proposed. Based on the concepts of fuzzy dispersion degree and fuzzy average relative adjustment degree, relationship between the weight changing capability of the developed fuzzy variable weight vector and the parameters it contains is analyzed. With consideration of keeping balance among states, steps on selecting fuzzy variable weight vector with its corresponding capability on weight changing are presented. Based on the fuzzy combined weighting method, a FCDM model with fuzzy variable weight is developed. The fuzzy structured element is used to express fuzzy numbers and fuzzy operations in the model. An analytical algorithm of the FCDM model based on the fuzzy structured element is presented. The developed model was then applied on evaluating and predicting for coal and gas outburst of coal mining panels. The research finding provides an alternative way for fuzzy operations and determination of weights in other fuzzy analysis models. [ABSTRACT FROM AUTHOR]

In the process of coal mining, the residual goaf will destroy the stress balance state of the overlying strata, which will result in a large number of fractures. The development of overburden fractures not only affects the movement of strata and the subsidence of surface, but also determines the dominant channel of gas, groundwater and other fluid migration. Therefore, this paper uses 3DEC software to simulate the temporal and spatial evolution of fractures in overlying strata during the mining of coal seam #15 in Pingshu Coal Mine, Qinshui Coalfield, and studies the fractal characteristics of mining overburden fissure development, then we use the drilling hole backwater method to verify the development height of mining-induced fractures. The results show that the overlying strata bend and collapse obviously with the advance of the working face. The development of overburden fractures shows very regular fractal characteristics, and the fractal dimension can be divided into four stages in time and two stages in space. The research results have guiding significance for mastering the law of overburden movement, accurately identifying groundwater and gas enrichment areas, and strengthening the prevention and control of mine water hazards and the effect of gas drainage. [ABSTRACT FROM AUTHOR]

The characteristics of overlying strata movement and the evolution of surface subsidence-related parameters are essential in preventing geological hazards and environmental alterations. This paper comprehensively analyses surface subsidence, overlying strata movements, and dynamic surface characteristics above the No. 24213 working face of the 1930 Coal Mine in Xinjiang, China. A numerical simulation model was established to predict the surface subsidence under a larger dip angle. Various time functions were used to determine the suitable function to predict subsidence speed. The in situ surface subsidence results showed prominent asymmetric characteristics. The tilt value was larger toward the footwall than in the opposite direction. However, the curvature evolution was not overly asymmetric and fluctuates greatly. The maximum surface subsidence value in tendency and inclination were 1800 mm and 433 mm, respectively. Moreover, the movement and advance angles were 77.4° and 66°, respectively. The numerical results were in good agreement with the on-site results; the overlying strata movement shape from underground to the surface was similar to an inclined funnel, and the stable draw angles of both sides were 41° and 82°, respectively. The logistics time function can describe the dynamic subsidence characteristics well for a representative point under large dip angle conditions. The parameters in the logistics time function were related to the rapid deformation stage duration time and initial subsidence speed. This study can offer guidance in predicting asymmetric characteristics in inclined coal seams. [ABSTRACT FROM AUTHOR]

Combined mining is when two working faces separated by a certain distance simultaneously advance in the same direction, which commonly results in stress concentration and rockburst. This paper investigates the rockburst mechanism due to inclined coal seam combined mining under complex geological conditions in the Da'anshan Coal Mine, China. Based on the experimental analysis, there is an impact tendency of the No.10 coal seam. A theoretical model is established to evaluate rockburst risk due to combined mining. In addition, a 3D multifactor numerical model of inclined coal seam combined mining is established. The novel "PyFlac-Exc" calculation code and the "PyFlac-Eng" elastic strain energy calculation script are compiled. The evolution of elastic strain energy is obtained during combined mining. Theoretical and numerical results show that a rockburst occurs in the western fourth working face of the No. 10 coal seam, which is consistent with the on-site rockburst accident. In addition, the kernel density estimation method is proposed to reduce stress concentration by changing the coal pillar widths and misalignment distances. Numerical results show that the elastic strain energy levels are low for the same coal pillar width, and the misalignment distance is arranged from 50 to 80 m. With the same misalignment distance, a coal pillar width greater than 45 m can effectively reduce the rockburst risk. [ABSTRACT FROM AUTHOR]

To analyse the impact of underground pressure on the variation in floor porosity and pore water pressure distribution during the coal seam mining process, this paper uses closed and open water pressure sensors to analyse rock pressure, water pressure zoning and layer distribution characteristics of the coal seam floor under mining effects based on a simulation analysis method using similar rock materials; the results determine that the separation characteristics of the pore water pressure distribution are affected by the rock pressure 'barrier' function, which is caused by changes in rock permeability under the coupling effect. To quantitatively analyse the rock pressure barrier's impact on the floor mechanical structure, this paper proposes the 'virtual separation layer' concept and defines the virtual separation layer index to determine the scope of the floor tension fracture zone. Through monitored data of pore water pressure at different depths of the floor during field mining, the variation between the water pressure, porosity, mining distance and floor depth is obtained; similar material simulation results are verified, and a zonal distribution model of pore water pressure in the floor is established, providing a reference to investigate the distribution of pore water pressure in the floor under mining disturbances. The results show that the influence of changes in mining stress will also lead to a differential distribution of the floor mechanical properties and ultimately damage the floor and affect the inrush of the confined water. The research results play an important role in risk classification and prevention of water pressure distribution in coal seam floors. [ABSTRACT FROM AUTHOR]

The reserved thickness of top coal has an important influence on the stability of a large section open-off cut under gob in the thick seams slicing mining. The destabilization extremum conditions of the open-off cut top coal were derived from by elastic–plastic theory, and the optical fibre sensing technology was utilized to monitor the top coal deformation law with different thicknesses (3, 3.5, and 4 m) in the physical similar simulation experiment in the paper. The results show that the top coal thickness is greater than 3.4 m without tension cracks. In the vertical direction, the top coal of the large open-off cut is divided into mining and excavation disturbance zones under the influence of the upper slice coal mining and the excavation disturbance. In the direction of the span of the top coal can be divided into the roof fall risk zone and the warning zone. The deformation changes from exponential to linear to logarithmic in the roof fall risk zone, and it changes from linear to logarithmic in the roof fall warning zone as the number of excavations increases. The sinking amount in the two zones is smaller as the thickness of the top coal becomes larger. It is comprehensively determined that the thickness of the top coal of open-off cut is set as 3.5 m, the stability is moderate, and the field application shows that the integrity of the top coal is good after support, and the maximum off-layer value is 6 mm, which can satisfy the production requirements. [ABSTRACT FROM AUTHOR]

Detecting the development height of the water-flowing fractured zone in the overburden of coal seam is of great significance for the overburden safety of working face with developed sandstone and thick loose layer. In this paper, the dual-mode parallel electrical method was used for dynamic monitoring of overlying rock in shallow and large mining height working face. Under the influence of mining, the apparent resistivity of initial mining fracture was high. The apparent resistivity began to decrease after the crack was understood by water entering the surrounding rock, and the apparent resistivity of fracture water-flowing to goaf was high again. According to the apparent resistivity monitoring results, the maximum height of the water-flowing fracture was 122 m. The ratio was 11.1 between the fractured zone and mining height. The numerical simulation results showed that the height of the water-flowing fractured zone above the working face was 122 m. The field monitoring results and numerical simulation results verify each other. With the influence of mining, the loose layer near the surface above the working face was cracked. There would exist a series of geological problems, such as soil erosion and ground subsidence. [ABSTRACT FROM AUTHOR]

Characterizing the permeability evolution and methane release is of great significance for the safe mining of the high gas outburst seams, as well as coal and gas simultaneous extraction. It contributes to reduce methane emissions from coal mining for greenhouse effect control. Theoretical analysis, laboratory testing, and numerical simulation are widely used methods to characterize the permeability and methane release with the treatment process of pressure-relief mining. However, these methods cannot fully reflect the complexity of filed practice. In this study, we report the effectiveness of protective coal seam (PCS) mining and the pressure-relief area in the protected coal seam (PDCS) based on detailed and integrated field measurements in a Chinese coal mine. To the best of our knowledge, it is the first time to measure the permeability coefficient and gas pressure evolution in the PDCS during the process of PCS longwall mining. The evolution of the permeability coefficient in the pressure-relief area during PCS mining can be divided into four stages: slowly decreasing, sharply increasing, gradually decreasing, and basically stable. The maximum permeability coefficient is 322 times of the initial value and stabilized at 100 times after the goaf compacted. The gas pressure evolution in the PDCS indicates that the strike pressure relief angle is 52.2° at the active longwall face zone, and 59.3° at the installation roadway side. The inclined pressure relief angles at the lower and upper sides of the longwall face are 75° and 78.9°, respectively. The residual gas content and gas pressure of the PDCS in the pressure-relief area are reduced to less than 6 m3/t and within 0.4 MPa, respectively. The field measurements further prove that pressure-relief mining can prevent coal and gas outbursts in PDCSs. The field observations in this paper can serve as benchmark evidence for theoretical analysis and numerical simulations, and also provide insights into realizing safety mining in similar conditions. [ABSTRACT FROM AUTHOR]

This paper presents a study on the stability of a mine entry of an innovative approach to non-pillar coal mining—gob-side entry retaining using cantilever beam theory (GERMC). This method is on the rise in China and it has shown to increase the coal seam rate of recycling and mining productivity. The method utilizes the caved-in material resulting from the mining activities to fill the gob area and ensure the stability of the roof of the already mined area. The caved-in material will form the gob-side wall of the retained entry. This characteristic is unique to this method. Thus, there are not many studies that analyze the stability of the filling material and of the gob-side entry wall under these circumstances, which is a vital element for the safety of the mining operations. This paper studies the stability of the gob-side retained entry of the Hongjingta underground coal mine in China. The authors perform the analysis using a discrete fracture network (DFN) model developed by the Massachusetts Institute of Technology (MIT), GEOFRAC, in combination with the discrete-element method (DEM) software UDEC. GEOFRAC estimates the rock blocks (fracture networks) in the filling body, from fracture traces measured along the gob-side wall of the entry. Statistical methods are used to estimate the fracture intensity and the mean fracture areas, which are inputs to the DFN. The generated fracture networks are then inputted into the DEM code, UDEC, to evaluate the stability of the mine-retained entry. For this study, we developed two types of models in UDEC, one considering the fractures generated by the DFN model and another considering the gob-filling material as a continuum. This methodology—combined DFN–DEM—is unique and provides a more realistic representation of the gob-filling material by considering the fractures in the filling body and by estimating those fractures (interfaces between rock blocks) using field data. In addition to this, we also considered the effects of uncertainties that are associated with estimating three-dimensional (3-D) fractures' networks/rock blocks from two-dimensional (2-D) field data. Finally, the results of the simulations were compared with the measurements from the mine and are in good agreement. [ABSTRACT FROM AUTHOR]

The year-after-year deepening of North China-type coal mines has resulted in frequent water inrush disasters at coal seam floors, causing significant casualties and property loss. The key factor for assessing water inrush hazards is the depth at which the coal seam floor will fail. It is therefore of great importance to predict this failure depth so as to evaluate the potential for floor water inrush incidents and establish effective controls in a timely manner. This paper takes a deep-mining mine in North China Coalfield, as an example, a comparative study is made using four methods: theoretical calculation method, back propagation neural network (BPNN) method, field measurement method, and empirical formula method. The results show that the new theoretical calculation method and the back propagation neural network method proposed in this paper are closer to the true floor failure depth under deep seam mining conditions. Therefore, this study has played an important reference role for prevention and control of water inrush from coal seam floor in deep mining. [ABSTRACT FROM AUTHOR]

The productivity of coal mines is seriously threatened by the combined disasters of gas and coal spontaneous combustion, which have become a common disaster mode. It is unclear how the gas and coal spontaneously combusted in the roof cutting along gob working face. The goal of this study is to identify the distinctive features of combined disasters in gob from two different types of roof cutting along working faces. In these two different types of roof cutting along gob working faces, the paper constructs the permeability model of the gob. The findings demonstrate that the data from the field experiment and the simulation results agree, which validates the simulation's reliability. In contrast to single sided roof cutting along gob working faces, double sided roof cutting along gob working faces clearly has a thinner oxidation zone. Moreover, the oxidation zone of the double side roof cutting along gob working faces is closer to the working face, which is located in the shallow area of the gob 50 m behind the working face. The gas explosion area and the coal spontaneous combustion area are divided by the double side roof cutting along gob working face, which reduces the risk of compound disasters. Important theoretical direction for the prevention and control of gob disasters in the roof cutting along gob working face is provided by the simulation results. [ABSTRACT FROM AUTHOR]

Rock burst disaster is still one of the most serious dynamic disasters in coal mining, seriously restricting the safety of coal mining. The b value is the main parameter for monitoring rock burst, and by analyzing its changing characteristics, it can effectively predict the dangerous period of rock burst. This article proposes a method based on deep learning that can predict rock burst using data generated from microseismic monitoring in underground mining. The method first calculates the b value from microseismic monitoring data and constructs a time series dataset, and uses the dynamic time warping algorithm (DTW) to reconstruct the established b value time series. A bidirectional short-term and short-term memory network (BiLSTM) loaded with differential evolution algorithm and attention mechanism was used for training, and a prediction model for the dangerous period of rock burst based on differential algorithm optimization was constructed. The study used microseismic monitoring data from the B1+2 fully mechanized mining face and B3+6 working face in the southern mining area of Wudong Coal Mine for engineering case analysis. The commonly used residual sum of squares, mean square error, root mean square error, and correlation coefficient R2 for time series prediction were introduced, which have significant advantages compared to basic LSTM algorithms. This verifies that the prediction method proposed in this article has good prediction results and certain feasibility, and can provide technical support for the prediction and prevention of rock burst in steeply inclined thick coal seams in strong earthquake areas. [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 angle α between the fault strike and the axial direction of the roadway produces different damage characteristics. In this paper, the research methodology includes theoretical analyses, numerical simulations and field experiments in the context of the Daqiang coal mine located in Shenyang, China. The stability control countermeasure of "pre-splitting cutting roof + NPR anchor cable" (PSCR-NPR) is simultaneously proposed. According to the different deformation characteristics of the roadway, the faults are innovatively classified into three types, with α of type I being 0°–30°, α of type II being 30°–60°, and α of type III being 60°–90°. The full-cycle stress evolution paths during mining roadway traverses across different types of faults are investigated by numerical simulation. Different pinch angles α lead to high stress concentration areas at different locations in the surrounding rock. The non-uniform stress field formed in the shallow surrounding rock is an important reason for the instability of the roadway. The pre-cracked cut top shifted the high stress region to the deep rock mass and formed a low stress region in the shallow rock mass. The high prestressing NPR anchor cable transforms the non-uniform stress field of the shallow surrounding rock into a uniform stress field. PSCR-NPR is applied in the fault-through roadway of Daqiang mine. The low stress area of the surrounding rock was enlarged by 3–7 times, and the cumulative convergence was reduced by 45%–50%. It provides a reference for the stability control of the deep fault-through mining roadway. [ABSTRACT FROM AUTHOR]

The study of water-conducting fracture zone development height is key to the scientific prevention and control of water damage in mines. Based on the geological conditions of the Wenjiapo coal mine in Binchang, China, this paper investigates the development of water-conducting fracture zone in overlying bedrock during mining under large buried depth and huge thick aquifer by combining on-site well-location microseismic monitoring and laboratory similar material simulation. To overcome the limitation of the " limited outlook " of water-conducting fracture zone investigation, the spatial development characteristics of roof fissures in coal seam mining were determined by on-site " the underground - ground" combined microseismic monitoring and follow-up monitoring, and the development of overlying rock fracture under the large depth of burial was concluded. The fractures were mainly distributed in the upper part of the protective coal pillar on both sides of the working face, but less in the upper part of the working face, and primarily distributed in the protective coal pillar on the side of the working face and the adjacent mining area. To verify the accuracy of the conclusion, the overlying bedrock movement and deformation characteristics and the development process of the hydraulic fracture zone during coal seam mining were analyzed by simulating similar materials in the laboratory, using the monitored area as a prototype. The results show that the development height of the mining fracture zone obtained from microseismic monitoring is basically consistent with the simulation results of similar materials. The research finding have significant implications for the study of fracture distribution characteristics and the evolution law of mining overburden, and provide a foundation for scientific prevention and control of water damage on the roof. [ABSTRACT FROM AUTHOR]

Gas is one of the main factors causing coal mine accidents in China. Understanding the mechanism of the influence of mine pressure on coal seam gas emission has important theoretical guiding significance for coal mine enterprises to be able to control gas. This paper chose the 72,909 working face in the Du'erping coal mine as the engineering background, analyzed the characteristics of the mine pressure in the working face and the driving face through numerical simulation, and found that the power of the gas emission in the working face is the mine pressure, while that of gas emission in the driving face is the gas pressure gradient. Based on the field monitoring data of the hydraulic support in the working face and the single pillar in the mining roadway, the law of mine pressure was obtained. Combined with the monitoring data of the number of drill cuttings and gas in different positions of the mining roadway and the gas concentration change in the upper corner of the working face during mining, the influence of the advanced support pressure and periodic pressure on the gas emission of the coal seam was analyzed. The results show that the advanced support pressure affects the coal seam gas emission by affecting the coal seam permeability coefficient. With the increase in the support pressure, the coal seam permeability coefficient decreases, and the coal seam gas emission decreases. During the period of periodic pressure, the advanced support pressure moves forward, forming a pressure relief zone in front of the working face, which is conducive to the emission of coal seam gas, and periodic pressure provides power for the gas emission. [ABSTRACT FROM AUTHOR]

Taking coal mine dynamic disaster as research background in this paper, five samples of coal–rock composite bodies (CRCBs) with different coal thicknesses were designed, and the uniaxial loading tests were carried out on them by using the MTS uniaxial loading instrument and the DS5 acoustic emission instrument, and the damage process of the samples were analyzed from the perspective of energy conversion. The results were as follows. With increase in coal thickness of CRCBs, the uniaxial compressive strength and elastic modulus of CRCBs decreased while the peak strain increased, and the overall bearing capacity of the samples decreased, resulting in a decreasing trend of AE peak ringing count and peak energy. According to the theory of conservation of energy, it was found that the dissipated energy of the samples in the compaction stage accounted for a large proportion, and the elastic stage was dominated by the accumulation of elastic energy. After the plastic stage, the energy conversion rate in the samples accelerated, and the dissipated energy increased rapidly, leading to the gradual failure of the samples. The energy storage limit of samples decreased logarithmically together with increase in coal thickness. Finally, it was found that coal was the main energy storage structure of the whole coal and rock composite system by analyzing the energy accumulation mechanism of coal and rock composite structure in practical engineering. Therefore, to prevent and control underground dynamic disaster in practical engineering, the internal energy storage of a coal seam should be released and the clamping effect of roof and floor on coal body should be weakened. The achievements of this study will be a theoretical guidance for preventing and controlling dynamics. [ABSTRACT FROM AUTHOR]

The slow excavation speed of coal roadways has always been a key factor restricting the safe and efficient production of large-scale coal mines in China, and the problem of unbalanced mining replacement caused by this is widespread. This paper takes the S1231 heading face of the Ningtiaota coal mine of Shaanxi Coal and Chemical Industry Group as the research object, analyzes the characteristics of the stress evolution of coal roadway driving, reveals the principle of low-density cross-border support, and proposes a low-density cross-border support plan. Using FLAC3D to study the roadway stress-displacement evolution law of the new support scheme during the driving and mining phases, the results show that the peak stress during the driving is 5.3 MPa, and the coal pillar side stress concentration is the most obvious during the mining period, with the peak value being 7.9 MPa. The moving distances of the two banks are both 10 mm, which verifies the feasibility of low-density cross-border support. Field application shows that during roadway excavation, the amount of roof subsidence and the displacement of the two sides are 9 mm and 11 mm, respectively, and the development depth of roof cracks is controlled within 0.5 m. The overall control effect is good, and the speed of coal roadway driving is increased by 77.19% compared with the original. The new support builds a thick roof anchor structure to ensure the safety and stability of the roadway. At the same time, by reducing the number of bolts, the bolt support time has been greatly reduced, effectively alleviating the tight situation of mining replacement, and providing solutions for mines under the same conditions. [ABSTRACT FROM AUTHOR]

Based on the S1201-2 large height mining in the 2–2 coal seam of Ningtiaota colliery with on-site microseismic measurement, physical simulation and theoretical analysis methods, this paper explores the rule of roof movement in thick coal seams with roof cutting and non-pillar (hereinafter referred to as RCN-P) mining, so as to obtain scientific and effective theoretical basis for entry support and to summarize the regional structural characteristics and dynamic periodic fracture characteristics. As can be seen from microseismic events, the entry roof is featured by "two zones and one line" along the horizontal direction, namely, the crack generation area, the roof movement area. Additionally, and the obvious lateral breaking of the entry roof on the coal wall is a typical feature of the thick coal seam with RCN-P mining. The roof is vertically divided into "three zones", the crack generation area, the roof movement area and the crack development area. The roof cutting activity mainly affects the overburden activity within the basic roof height range, which is also the roof movement area. In addition, the distribution frequency and the intensity of microseismic events indicate the roof periodic breaking characteristics. The "breaking pressure relief," "advanced crack development," and "the limit breaking state" of roof breaking corresponds to the initial, middle, and final stage of breaking in the periodic weighting process, respectively. Compared with the normal mining, the RCN-P mining reduces the periodic weighting length and increases the pressure strength. As is shown in the physical simulation experiment, the basic roof and the cutting control layer in the "regional structural characteristics" constitute the "large" and "small" structures with RCN-P mining. The basic roof key layer is the core to control the stability of the strata, and the breaking process from the cantilever beam to the short masonry beam of the roof-cutting control layer is the main cause of the entry stress. Correspondingly, the basic structure model of "short masonry-hinged" roof was proposed and the calculation method of support was established for the entry with RCN-P mining in thick coal seam, providing a research foundation for scientific and effective rock formation control. [ABSTRACT FROM AUTHOR]

The No. 3 coal seam is one of the primary mineable coal seams in Du'erping coal mine of Shanxi Coking Coal Group Co., Ltd. The direct roof of No. 3 coal seam is a compound roof with two layers of 0.3–0.5 m extremely soft coal seams in the unstable rock stratum. The mining roadway has a large amount of deformation and is difficult to support. This paper takes 63,409 ventilation roadway of Du'erping coal mine as the engineering background. The RFPA2D numerical simulation software was used to compare and analyze the distribution characteristics of surrounding rock fracture zone and surrounding rock deformation in the mining roadway of the general roof and compound roof under the conditions of coal body–coal body and coal body–coal pillar. The supporting form and main supporting parameters of the mining roadway with compound roof were determined. Through simulation and comparison, the deformation of the mining roadway with compound roof under unsupported and three different support conditions were analyzed, and the combined support method of "bolt + anchor cable + metal mesh" was field tested. By analyzing the monitoring data of the roadway roof subsidence, floor heave, ribs displacement and roof abscission layer, the results show that the combined support method of "bolt + anchor cable + metal mesh" effectively controls the large deformation of the mining roadway with compound roof, improves the stability of the surrounding rock, suppresses the occurrence of roof fall and rib spalling, and ensure the safety of the roadway. [ABSTRACT FROM AUTHOR]

To ensure the safety of coal mining, lots of coal pillars were reserved in the coal mine. The stability of coal pillar and roof composite structures determines the safety of stope. Under the action of cyclic loading and unloading such as mining face, chamber blasting, and roadway excavation, the coal pillar, and roof composite structure will gradually instability, which will affect the safety production of the coal mine. Therefore, this paper takes the dynamic disaster of coal pillar and roof composite structure as the research background, carries on the cyclic loading and unloading test of the coal-rock composite structure with different coal-rock height ratios, and studies the influence of coal-rock height ratio on the progressive instability characteristics, mechanical properties and damage variables of the coal-rock composite structure. The crack initiation stress, crack propagation time and failure pattern of coal-rock composite structure are analyzed. The improved load-unload response ratio considering the damage variable is established, and the stability of the coal-rock composite structure under cyclic loading and unloading is quantitatively analyzed. It is concluded that the larger the coal-rock height ratio is, the greater the load-unload response ratio of the composite structure is, and the more unstable the composite structure is. This conclusion is consistent with the crack initiation stress and fracture propagation time, and can quantitatively reflect the progressive instability process of coal-rock composite structure, and provide a basis for revealing the dynamic disaster of coal pillar and roof composite structure. [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]

In bench blasting, backbreak is the unwanted result that causes instability to the highwall and can lead to safety hazards. Hence, it is utmost necessary to minimize the generation of backbreak to improve mine's safety. It has always been difficult to predict the backbreak because of various parameters involved, i.e. blast design, explosive properties, rock mass, etc. In this study, multivariate regression analysis (MVRA) and genetic programming (GP) techniques were performed on 70 blast data sets of previously published papers. Both the models have been developed and tested with the same mine data set. For validation of the models, a total of 14 trial blasts have been conducted in Indian coal mines with different geological strata and other parameters. The values of R2, RMSE, MAPE and prediction level at 25% and 90% were computed for GP and MVRA techniques. Also, the GP model is compared with the other state-of-the-art techniques. It has been found that the level of prediction for validation data set at 25% using GP is 78.57% and for MVRA is 21.42%. The mean magnitude of relative error (MMRE) value for GP and MVRA is 0.18 and 0.42, respectively. The results show that the GP is a more efficient tool for prediction of backbreak in comparison with MVRA. On performing sensitivity analysis, it has been found that stemming length and powder factor are the most influencing parameters to backbreak. [ABSTRACT FROM AUTHOR]

Determining the strata and surface movement of deep mining can provide a scientific basis for wellbore protection and mine land utilization. In this paper, numerical simulation methods were employed to investigate the stratum and surface movement boundary curves of deep caving and backfilling mining. The findings can be concluded in three aspects. First, the stratum and surface movement boundaries of deep caving and backfilling mining were all curved and conformed to an exponential function, but the influence range of strata and surface movement of deep different mining methods were different. Second, the backfilling rate of deep backfilling mining had an impact on movement boundary boundaries of strata and surface. With the backfilling rate decreasing, the influence range of strata and surface movement boundary were increased. Last, the research results were applied to a case to confirm the new approach of determining the influence range of strata and surface movement of deep mining. The example application shows that it not only can ensure the safe production of the wellbore, but also can enhance the reuse area of the mine field. [ABSTRACT FROM AUTHOR]

A new approach for prediction of face advance rete (FAR) prior to mining operation and determination of the operation efficiency after mining operation in retreat longwall mining panel is presented based upon the concepts of rock engineering system (RES). For this purpose, six longwall panels considered in Parvadeh-I coal mine. Seven major effective parameters on FAR was selected including coal mine roof rating, gas propagation, safety factor of longwall face, ratio of joint spacing to cutting depth at longwall face, longwall face inclination, panel width, floor rock mass rating. To performance evaluation of the presented model, the relationship between the average vulnerability indexes of advance operation with FAR was determined in considered panels with coefficient of determination (R2) equal to 0.884 that indicate relatively acceptable correlation and compatibility. Investigations of the research indicated that it is possible to determine the actual operation efficiency under fair conditions by a RES-based model. The inevitable reduction of FAR for each longwall panel was determined by presented model that the difference amount between the maximum possible practical face advance rate (FARmpp) and recorded actual face advance rate (FARa) indicate the operation efficiency. Applied approach in this paper can be used to prediction of FAR in retreat longwall mining panel for same conditions that can have many benefits, including better and more accurate planning for the sales market and mine operation. Also, presented method in this paper can be applied as a useful tool to determination of actual operation efficiency for other sections and extraction methods in coal mines. [ABSTRACT FROM AUTHOR]

With the increasing tension of current coal resources and the increasing depth of coal mining in China, Non Coal Pillar Gob-side Entry Retaining Technology has become a preferred coal mining method in underground coal mines. In the current study of the non coal pillar gob-side entry retaining mining with roof cutting and pressure releasing, the theory and design of roof cutting under near-level conditions have formed a certain system. However, there is a lack of systematic research on the design of the roof cutting and the connection requirement of the cutting seam and other aspects about the inclined coal seam. Therefore, choosing the concrete representative target mines as the research foundation, using mechanics analysis, geometric analysis, numerical simulation and other means, on the basis of considering the change of mining height, this paper mainly studies about the effect of coal seam inclination on roof slit depth and angle, then get the three steps of the cutting height design, the three considerations of the roof cutting angle design and the connection rate requirement of the roof cutting. At last the key roof cutting parameters of the three target mines are calculated accordingly. Based on the existing research, this paper puts forward the concept and design method of roof cutting connectivity rate. The design system of roof cutting seam parameters of non coal pillar gob-side entry retaining mining with roof cutting and pressure releasing is further improved, and the corresponding calculation method is deduced for the design of the key parameters of the inclined coal seam and the gently inclined coal seam, and proposing the second roof cutting scheme of the upper side gob-side entry retaining. [ABSTRACT FROM AUTHOR]