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In the process of lower coal roadway support under double gobs in close coal seams, there are problems such as the whole destruction of the lower coal seam and the intermediate strata, the thin roof of the lower coal roadway with the smaller support space, the upper rock of the lower coal roadway without anchor bearing layer, the roof falling, the anchor cable offing, and so on. On the basis of the engineering background of No. 11103 haulage roadway in Fenxi coal mine, the research methods of theoretical modeling analysis, numerical simulation analysis and field engineering test are comprehensively adopted. In this paper, the layout of the roadway with floor insertion in the lower coal seam, the distribution characteristics of surrounding rock deviatoric stress, the control principle of broken surrounding rock and the coupling support method of shed–cable–prop are systematically studied. The study shows that many measures are needed to ensure the stability of the surrounding rock in the strong mining roadway of the lower coal seam. The lower coal seam roadway section was modified from the original rectangular section to adopt an arched section to counteract the broken roof. The staggered position of the lower coal seam roadway layout was modified from the original coal roadway position to a coal–rock roadway, so as to avoid the area of high deviatoric stress in the coal pillar and the area of crushed coal body, and to leave an effective support space for the roof. The support scheme of the lower coal seam roadway was improved to use anchor cable and single prop to support the weak section of U‐type steel shed with local coupling, and realize unequal force coupling by adjusting the size of the force at different coupling points. In the field monitoring the deformation of the roadway is small, and the support effect is good. [ABSTRACT FROM AUTHOR]

The gob‐side entry‐retaining by roof cutting and pressure relief (GERRC) method is an advanced no‐pillar mining method that can significantly improve coal recovery and reduce roadway accidents. However, according to case studies, there is relatively little research on the application of the GERRC method in geological conditions of soft rock formations (especially when the roof and floor are both mudstone), and such research is needed. Therefore, based on the typical soft strata of the Xinyi Coal Mine, this paper analyzes the failure mechanism of the roadways and proposes the GERRC method with coal rib and corner strengthened support (CRCSS). The evaluation indexes of the roof convergence rate and cable safety margin are established, and the effects of roof cutting and CRCSS on roadway deformation control are explored by a numerical comparison test. The results show that roof cutting has a limited effect on roadway deformation. The deformation can be reduced by 75.77% by implementing CRCSS on the basis of GERRC. Then, the evolution characteristics of roadway stress and deformation in the process of mining are further discussed through field contrast tests. The field application results show that compared with that of GERRC with conventional support, the surrounding rock deformation of GERRC with CRCSS is reduced by more than 55%, the safety margin of the cable is improved. The overpressure of the hydraulic prop is diminished, and the stability distance of the roadway is reduced by approximately 30%. Finally, this paper discusses the mechanism to control roadway deformation by CRCSS. This research expounds the rationality of CRCSS, and the results can provide a reference for other situations with similar geological conditions. [ABSTRACT FROM AUTHOR]

This paper is aimed at the problems of multiple factors influencing the roof stability near a heading face and unreasonable values of the unsupported distance leading to a slow excavation speed. The behavior of the direct roof subsidence in the unsupported area is analyzed, and numerical simulation and the response surface method are used to study the key factors affecting the roof deformation and their interaction relations. The results indicate that roof strength is the most critical factor affecting deformation and failure, followed by soft rock thickness, unsupported distance, and support strength. The interaction analysis in Zhaozhuang Coal Mine's 33082 lane shows that it is difficult to control the roof subsidence when the unsupported distance is more than 2 m and the soft rock thickness is more than 4 m. After the unsupported distance is set to about 2 m and the support strength is increased to 0.25 MPa, the excavation speed approximately doubles, and the overall roadway is stable and controllable. By implementing field applications, the rationality of the research approach was confirmed. [ABSTRACT FROM AUTHOR]

Gob‐side entry formation by roof cutting is a new technology for no pillar coal mining, which can maximize coal resources and reduce roadway drivage ratio. However, the mechanical behavior of the formed entry is complex while it is crucial to ensure the stability of the entry for mining safety. This paper proposed a machine learning‐based method for predicting the stability of the formed entry, which combined artificial neural network (ANN) with particle swarm optimization (PSO) algorithm or genetic optimization (GO) algorithm. The data set from 75 coal mining faces from 2009 to 2022 was employed to train and test the models. A descriptive variable of dynamic unstable distance was introduced to evaluate the stability state of the formed entry and six other parameters were chosen as influence parameters. The two intelligent models were compared with each other to have a comprehensive assessment. Model assessment indices such as R2, mean absolute error, mean absolute percentage error, and root mean square error were used to evaluate the accuracy of the models. The results of both developed models are promising, and the predictive accuracy of the PSO‐ANN model is higher than that of the GO‐ANN model. Through sensitivity analyses, it has been found that the coal seam thickness and roof rock hardness are the most important parameters for influencing entry stability. The developed method provides a practical tool for the prediction of entry stability and the optimization of entry design. [ABSTRACT FROM AUTHOR]

With the shift of coal resources to deep mining, the occurrence of long‐distance coal seams has increased, and protective layer mining is facing new challenges. This paper attempts to explain the stress evolution law of the upper coal group in the long‐distance mining of the lower coal group in Pingdingshan No. 8 Coal Mine. A simulation model of advance mining of lower‐group coal long‐distance was established. The stress evolution law of the upper coal seam under the influence of advanced mining disturbance of the lower coal seam is studied. The following conclusions were obtained: The advance mining of the lower coal group had a positive or negative impact on the stress distribution of the upper coal seam group. With the recovery of the lower coal group of the F‐21030 working face, the overburden of the F‐21030 goaf finally formed a "Y" type pressure relief area. The pressure relief effect of the E‐21070 working face near the stopping line was obvious. The coal seam of Group E was divided into three areas affected by the advance mining of the lower coal seam. The maximum pressure relief value was 6.6% lower than the initial stress. According to the simulation results, the E‐21070 working face was divided into three regions, namely, the pressure relief region, the stress increase region, and the original stress region. According to the field drainage results of pressure relief gas, the extraction curve could be divided into three parts, namely, the stable area, pressure relief area, and stress recovery area. The maximum pure gas drainage volume could reach seven to eight times of the original area. The pressure relief extraction effect was remarkable, and the optimal extraction range was 22–210 m behind the coal face of the group. [ABSTRACT FROM AUTHOR]

Gas drainage is an important measure to ensure energy safety in coal mines. However, for mines with high gas content and low permeability, the critical question is how to enhance the drainage effect of drilling. By using the self‐unloading pressure effect on the coal body in front of the mine, the permeability of the coal body can be improved, which facilitates gas drainage. Therefore, establishing the fluid–solid model of gas‐bearing coal mass under the influence of mining, this paper simulates the advanced self‐pressure‐relief range in front of the coal working face and the permeability of coal seam at different heights. Through simulations and field tests, the reasonableness of the advanced self‐pressure‐relief gas drainage was verified, and the spatial and temporal characteristics of the gas drainage from the borehole were quantified and analyzed. The results show that the width of the advanced self‐pressure‐relief zone after simulated mining is 25 m, the vertical permeability of the coal seam increases first and then decreases with the increase of the height from the floor, and the maximum permeability of the layer is 2.6 m from the floor, which is 8.62 times higher than the initial permeability; the gas drainage effect of test boreholes has a periodic law with the coal seam recovery, which, divided into the initial drainage stage, the scalar volume of gas drainage increase stage, and the scalar volume of gas drainage decay stage. At a distance from the working face 20–40 m, the boreholes' scalar volume of gas drainage increased by 3.89 times compared with the initial drainage stage, and the scalar drainage volume of the boreholes through the seam increased by 8.01 times compared with the initial drainage stage; in the section of increasing scalar volume of drainage, the gas drainage volume of the boreholes at different heights of the coal seam had obvious "stratification" phenomenon. The test boreholes were optimally placed in the middle of the seam at the lower end. The study provides a reference for self‐pressure‐relief gas drainage technology and optimization of borehole placement. [ABSTRACT FROM AUTHOR]

In northern Shaanxi province, the coal seam deposit conditions are generally characterized by large thickness and shallow depth of burial, the phenomenon of coal pillar sloughing is serious in the process of working face retrieval, which brings great potential danger to mine safety production. In this paper, by monitoring the stress and damage of the coal pillar in the field section, the stress changes and damage degree of the working face in different mine stages are determined. A new mechanical model of coal pillar in a thick coal seam is constructed and the stress changes and structural breakage law of each section are obtained, the stress transfers mechanism of overburden structure and the stress damage to coal pillar during the mining process of upper and lower section comprehensive mining working face is analyzed. The stress, microscopic damage, and the degree of sloughing in the coal pillar were significantly reduced by monitoring through the top plate hydraulic cracking pressure relief control program. The results of the study show that the extent of sloughing coal pillar and internal damage is closely related to the length of key block B of the masonry beam formed by the overlying rock layer and the position of the break line, by changing the orientation of the overlying masonry beam rock layer and the position and length of the tendency to break line, the sloughing coal pillar is effectively controlled and its stress and damage are reduced. We have achieved good results in control of the sloughing in the Hanjiawan coal mine and similar mines in northern Shaanxi. [ABSTRACT FROM AUTHOR]

Traditionally, longwall top coal caving (LTCC) panels adopt the single‐window drawing mode and their drawing efficiency is relatively low. The synchronous multi‐window drawing (SMWD) mode is a remarkable approach to resolve the incompatibility between shearer coal mining and top‐coal drawing. The present paper analyzes the caving characteristics of top coal under SMWD in LTCC mining of extra‐thick coal seam. The caving and migration of top‐coal fragments were simulated by using the continuum–discontinuum element method (CDEM). The results show that the SMWD mode can effectively improve the drawing smoothness with larger caving channels compared to the traditional single‐window drawing mode. The main caving channel for top‐coal fragments was directly above the assembled multi‐window (AMW), and the size of the AMW had a great influence on the distribution of the velocity field. Additionally, the simulated draw body of the initial drawing adopted an elliptical shape, which changed into a variation ellipse with the increasing size of the AMW. The boundary of the simulated draw body of the initial drawing presented an irregular shape that deviated from the results predicted by the Bergmark–Roos model. Furthermore, the drawing efficiency increased with the expansion of the AMW, whereas the recovery ratio first increased and then decreased with the expansion of the AMW. Finally, field tests revealed that the SMWD mode can greatly improve the drawing efficiency. Thus, the recovery ratio under A2 and A3 SMWD modes was higher than that of the traditional single‐window drawing mode. Based on the results of field tests and CDEM simulation, it is recommended to adopt A2 and A3 SMWD modes to enhance the top‐coal recovery ratio and drawing efficiency. These research results could constitute a technical reference for automated highly efficient top‐coal drawing and high‐yield LTCC panel extracting extra‐thick coal seam. [ABSTRACT FROM AUTHOR]

Rational estimation of the mechanical state of a competent rock layer is of prime importance for structural and support design in longwall coal mining. This paper analytically examines the stability of a voussoir‐structured stiff rock by considering it acting as a symmetric flat arch with discontinuities at the abutments and midspan. Three typical failure modes including snap through, midspan crush and sliding failure are inspected. The solution assumes that the shape of the compressive arch obeys a parabolic variation as the voussoir beam deflects downwards with large‐deformation. The crush failure due to stress concentration in the beam hinge is predicted by relating the squeezing area with the beam rotation, which is absent in previous studies. The proposed approach can examine the voussoir beam stability without cumbersome iteration. Primarily, this work facilitates the evaluation of the support requirements to reinforce roof beds in longwall panels. After accounting for the total loads of the overlying rock layers, the stability assessment developed in this paper can further be utilized to predict the span limit of the main roof above a longwall panel and to estimate the critical width which governs the magnitude of mining subsidence. This manuscript provides an analytical solution for the stability of a competent rock stratum in voussoir structure. Typical failure modes including snap through, crush at midspan and sliding failure have been examined. The proposed approach is capable of predicting the elastic buckling when the voussoir beam bends downward with large deflection. Moreover, the relationship between beam rotation and contact area alteration due to crush at midspan hinge is first considered. This work improves the reliability of stability analysis for underground longwall panels. Furthermore, the proposed guidelines can be utilized to predict the occurrence of dynamic ground pressure, overlying strata movement and subsequent surface subsidence. [ABSTRACT FROM AUTHOR]

To control roadway under small abandoned mines in ultrathick coal seam, this paper analyzed its deformation and failure characteristics. As a result, the roadway occurred roof leakage and right rib extrusion. Based on above characteristics, the secondary support schemes were put forward. To predict the control effects of the roadway, a numerical model was built to analyze its deformation under the secondary support schemes. Besides, the industrial test was carried out. From the test and numerical simulation, a significant increase could be found in vertical stress and bearing capacity on ribs after the secondary support. Moreover, under the secondary support, the roof separations of roadway decreased more than 90%, while the lateral displacements of the roadway dropped over 85%. Based on the above, it was indicated that the deformation of the roadway could be well controlled. [ABSTRACT FROM AUTHOR]

Protective coal layer (PCL) mining is a preferred regional method for preventing coal and gas outbursts; stress‐relief of PCL mining is the prerequisite of pressure‐relief gas drainage. In order to study the stress‐relief effect of PCL mining on protected coal seams and determine the protection scope of PCL mining under different inclinations, this paper uses physical simulations to study the stress‐relief law and deformation characteristics of the strata during upper PCL mining with a gentle inclination (25°), an intermediate inclination (45°), and a steep inclination (65°). The results show that as the coal seam dip increases, the degree of deformation and failure of the overburden strata gradually decreases, the severe subsidence and expansion area of the overburden strata tends to gradually develop in the uphill direction of the working face, and the floor strata underlying the protective seam significantly expand toward the goaf. However, the absolute value of the upward movement of the floor strata is much smaller than that of the roof strata; and the greater the dip angle, the greater the ratio of floor displacement to roof subsidence. The stress‐relief curve of the protected layer is approximately U‐shaped. As the inclination of the coal seam increases, the stress‐relief range decreases gradually. The maximum relief‐stress and the stress release coefficient of the protected layer initially increase and then decrease. The stress concentration coefficient of the protected coal seam initially increases first and then decreases in both uphill and downhill of the working face in the PCL. As the inclination increases, the stress‐relief angle of the downhill boundary initially increases and then decreases, while the stress‐relief angle of the uphill boundary decreases gradually. Our research results provide guidance for the delimitation of the protected scope of PCLs with different dip angles. [ABSTRACT FROM AUTHOR]

The revelation of the distribution characteristics of stress and fractures in the floor strata is of great significance to improve the regional gas drainage efficiency and to eliminate or reduce the risk of coal and gas outburst in adjacent coal seams during multi‐seam mining. In this paper, the stress distribution in the floor under mining condition was derived, and the characteristics of stress zones in the floor were elaborated while mining the 22201 working face in Shaqu coal mine in Shanxi province, China. Besides, the propagation laws of fractures in different stress zones were theoretically analyzed, and the distribution law of fracture angles in the floor strata under different mining‐induced stresses was obtained. The research results demonstrate that the compression and transition zones show high vertical stress and less developed fractures and are dominated by fractures with a large angle (>60°), which is unfavorable for gas drainage. The expansion zone in the floor within 2‐40 m in the rear of the working face is characterized by high unloading degree of vertical stress and developed fractures. Moreover, fractures with a small angle (<60°) dominate the spatial distribution of the fractures, which is favorable for gas drainage. The research results can not only be a beneficial supplement to the theory of coal and gas co‐exploitation in multi‐seam mining but also provide theoretical support for optimizing the design parameters of cross‐measure boreholes under similar conditions. [ABSTRACT FROM AUTHOR]

Nonpillar mining (NPM) with an automatically formed roadway is an emerging mining technology that does not use coal pillar retention and drivage. This technology can be used to actively control the caving height of the roof by directional cutting technology and can promote the rapid equilibrium of a fractured high roof under the support of the caved gangue. Therefore, when analyzing the stability of a high roof under NPM, not only the interaction force of the fractured rock blocks but also the effects of the support provided by the gangue, and the coal body should be considered. In addition, if necessary, the roof support in the roadway should also be considered. Based on this idea, in this paper, we analyzed the stability characteristics of a fractured high roof and suggested the stability conditions of the fractured rock of a high roof, including actual stability, pseudostability, and instability and provided the criteria for determining these three stability conditions. Furthermore, a visualized approach for the determination of roof stability was proposed. The core idea of the proposed approach is to transform the parameters that cannot be measured in the abovementioned conditions into visualized data that can be directly or indirectly obtained in the field, allowing for the detection and occasional assessment of the stability condition of the fractured rock of a high roof. This visualized determination approach can be used to assess the roof stability condition in a timely manner during the production process and to ensure production safety. The abovementioned research results were successfully used in the S1201‐II mining face of the Ningtiaota Coal Mine, indicating that the proposed approach is feasible and can be used under similar conditions. [ABSTRACT FROM AUTHOR]

The coal production exploited from extra‐thick coal seams accounts for 45% of China's total coal production. With large coal output, the extra‐thick coal seams require that the chamber section be more than 100 m2. In case of parting (the noncoal rock, usually mudstone, sandwiched in the coal seam) in the extra‐thick coal seams, the integrity of the coal seams will be damaged and the large‐section chamber is extremely easy to occur rib spalling and roof leakage, which will increase the chamber's supporting difficulty. To solve such problems, this paper took the electromechanical chamber in Tashan Coal Mine as the engineering background and analyzed the deformation and failure characteristics of this chamber surrounding rocks by the numerical simulation method during the excavation process. The results indicated that the failure areas of surrounding rocks were mainly in partings, chamber ribs, and chamber dome with the failure depths of 15 m, 10 m, and 5 m, respectively. Based on the obtained deformation and failure characteristics, the supporting method of grouting integrated with high‐strength bolts and anchor cables was put forward. The grouting method was firstly used to improve the integrity of the chamber surrounding rocks, and the high‐strength bolts and anchor cables were used for supporting. Then, the supporting parameters of the bolts and anchor cables and the grouting parameters were optimized by the numerical simulation method. The optimized parameters were applied in the electromechanical chamber in Tashan Coal Mine, and the on‐site chamber ribs lateral displacement and roof separation amounts were only 0.018 m and 0.012 m, respectively, indicating that the surrounding rocks were controlled effectively. [ABSTRACT FROM AUTHOR]

The calculation of abutment stress has always been the most important topic in safe underground mining and design. In this paper, based on the formation mechanism of abutment stress, combined with the mechanical properties of actual strata and the theory of elastic foundation beams, a structure model of overlying strata with an elastic foundation is established, and a new theoretical calculation method of abutment stress is proposed. The new calculation method takes into account the interaction of the overlying key strata and the effect on the transmission of the front abutment stress and then analyzes the stress patterns of multilayered key strata under the combined action of the bending moment, shear force, and nonuniform load and its transmission to the underlying strata. In addition, based on field observations of panel 10 305 in the Xinglongzhuang coalmine, the influence range, peak position, and variation trend of the front abutment stress before and after breaking of the overlying key strata are studied and verified. The results show that the influence range of the abutment stress remains unchanged after KS1 breaks (eg, from 0 to 180 m), and its peak position shifts forward from 8 to 19 m, while the influence range (0‐180 m) and its peak position (6‐8 m in front of coal wall) remain unchanged after KS2 breaks; the abutment stress decreases, and the maximum decrease is 3.9 MPa in the range of 0‐22 m ahead of the coal wall after KS1 breaks, while the maximum decrease in the abutment stress is 2.1 MPa in the range of 0‐71 m ahead of the coal wall after KS2 breaks. The results of the new calculation method are basically consistent with the field observations and the existing theoretical results, showing that the method has good adaptability and reliability and can provide a new theoretical basis for on‐site design and construction practices. [ABSTRACT FROM AUTHOR]

The mechanism and fracture aperture of thick hard roof failure were theoretically analyzed from the perspective of elastoplastic mechanics. A uniform load cantilever beam was established by considering the voussoir beam structure and stress condition. According to the stress inverse principle in solving the plane strain problem, the internal stress component analytic solution was deduced. Considering the lateral thrust and shear force and by applying the Mohr‐Coulomb failure criterion, the plastic zone boundary equation was obtained. Further analysis of the equivalent stress, equivalent strain, elastic energy density, distortion energy density, and strain energy density was studied with blocks of four different lengths, combined with the rotation angles after failure, to analyze the fissure aperture between straight and curved fracture traces. The theoretical analysis reveals that the deviatoric effect plays the lead role in rupture initiation and propagation; the cracks will deflect during failure, and the fracture trace presents a curved shape. The variation of fracture aperture exhibits a tendency to increase first, then decrease, and then increase along the fracture trace. The analytical approach presented in this paper provides a theoretical basis for determining the main seepage channel flow in underground mining. [ABSTRACT FROM AUTHOR]

Due to the complex geostress and mining conditions in the coal seam with depth of 800 m, stability of surrounding rock for gob‐side entry retaining is very difficult to achieve. In this paper, we firstly propose an innovative bolt‐grouting controlled roof‐cutting for gob‐side entry retaining (BCR‐GER) approach for deep coal mines. Secondly, a mechanical model of "surrounding rock‐supporting body" for BCR‐GER is constructed, which consists of coal wall, roadside props, and gangues in gob (the whole supporting body). Thirdly, the key parameters (ie, cutting height, cutting angle, grouting cable length, and row of roadside props) are designed. Finally, field practice was applied at the No. 31120 haulage roadway of the Suncun coal mine in China, and in situ investigations were conducted for verification. Field measurement results show that maximum convergences of roof‐to‐floor and side‐to‐side were 264 mm and 113 mm, respectively. What is more, the maximum support resistance of roadside props was reduced by approximately 58%. The deformation and failure of surrounding rock were effectively controlled, and the pressure on roadside props was greatly reduced. This research fully considers the bearing properties of gangues in gob, eliminates the secondary disasters caused by borehole blasting, and provides guidance and reference for deep surrounding rock control of the same or similar gob‐side entry. [ABSTRACT FROM AUTHOR]