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To address control of the coal wall rib spalling in a work face with large mining height, an investigation was conducted on the Caojiatan mine 122107 work face as a case study. Using a combination of theoretical analysis, orthogonal experiments, and numerical simulations, we determined the influence of hydraulic support parameters on coal wall rib spalling. Our results show that reducing the end-face distance and the resultant force application point and increasing the vertical force and the horizontal force of a hydraulic support provide increased stability of the coal wall. Extensive verification was made through simulations. The work presented herein suggested a structural zoning of the coal wall rib spalling, it also established the coal wall stability coefficient, and determined the lower limit of the coal wall's nonstructural area stability coefficient. Additionally, this study provided methods for the prevention and control of coal wall rib spalling in a large mining height work face, via improving the nonstructural area mechanical parameters, and optimization of the hydraulic support. [ABSTRACT FROM AUTHOR]

In mining, the roof structure of a working face with a large mining height in a shallow‐buried coal seam is prone to cutting instability in front of the support, which causes support crushing and coal‐wall spalling. This study analyzes 2201 working faces with large mining heights in the Haiwan No. 3 Mine by investigating the mechanical response law of coal walls under the transient excitation of roof structure instability. A mechanical model of coal walls under dynamic and static load coupling is established, revealing the formation mechanism of coal wall spalling and its main influencing factors. Prevention and control technologies of advanced grooving and pressure relief are proposed, and grooving parameters are determined. The results show that: During the mining process of large mining height working face in shallow buried coal seam, the step change of static response of coal wall is induced by the transient instability of roof structure, and the high abutment pressure is formed on the coal wall. At the same time, the strain energy and gravitational potential energy of roof cutting instability are released to form a dynamic load. Under the action of dynamic and static load, the plastic failure dissipation power of the coal wall is greater than the critical power, and the coal wall spalling occurs. The main influencing factors are mining load, mining height, and internal friction angle of the coal wall. The method of advanced grooving pressure relief can change the roof structure and coal force mechanism, by cutting off the connection between the coal wall and immediate roof, reducing the height of the coal wall force, effectively controlling the position of the roof cut off break line, reducing the static load effect of the roof on the coal, and transferring the position of the dynamic load response to the depth of the coal wall. When the grooving height‐depth ratio k is 0.75, the peak value of the abutment pressure moves 5.57 m to the front of the working face, and the pressure relief effect is the best. The above research results provide an effective active protection method for controlling coal wall spalling in fully mechanised working face with large mining height. [ABSTRACT FROM AUTHOR]

Abstract In mining, the roof structure of a working face with a large mining height in a shallow‐buried coal seam is prone to cutting instability in front of the support, which causes support crushing and coal‐wall spalling. This study analyzes 2201 working faces with large mining heights in the Haiwan No. 3 Mine by investigating the mechanical response law of coal walls under the transient excitation of roof structure instability. A mechanical model of coal walls under dynamic and static load coupling is established, revealing the formation mechanism of coal wall spalling and its main influencing factors. Prevention and control technologies of advanced grooving and pressure relief are proposed, and grooving parameters are determined. The results show that: During the mining process of large mining height working face in shallow buried coal seam, the step change of static response of coal wall is induced by the transient instability of roof structure, and the high abutment pressure is formed on the coal wall. At the same time, the strain energy and gravitational potential energy of roof cutting instability are released to form a dynamic load. Under the action of dynamic and static load, the plastic failure dissipation power of the coal wall is greater than the critical power, and the coal wall spalling occurs. The main influencing factors are mining load, mining height, and internal friction angle of the coal wall. The method of advanced grooving pressure relief can change the roof structure and coal force mechanism, by cutting off the connection between the coal wall and immediate roof, reducing the height of the coal wall force, effectively controlling the position of the roof cut off break line, reducing the static load effect of the roof on the coal, and transferring the position of the dynamic load response to the depth of the coal wall. When the grooving height‐depth ratio k is 0.75, the peak value of the abutment pressure moves 5.57 m to the front of the working face, and the pressure relief effect is the best. The above research results provide an effective active protection method for controlling coal wall spalling in fully mechanised working face with large mining height.

Abstract To address the issues of coal rib instability and high occurrence of roof falls in high extraction height fully mechanized coal mining of soft coal seams, this study comprehensively employs theoretical analysis, numerical calculations, and field industrial experiments. By analyzing the distribution characteristics of coal rib displacement and deformation instability zones under different mechanical parameters of soft coal seams, the instability mechanism of coal ribs in high extraction height mining of thick and soft coal seams is revealed, and corresponding stability control techniques for high extraction height coal ribs and supports are developed. The study shows that the stability of coal ribs in high extraction height mining fields of soft and thick coal seams decreases as the values of mechanical parameters decrease. In other words, the smaller the cohesion and deformation parameters of the coal body, the more prone the coal rib is to roof falls. By utilizing advanced deep‐hole static water infusion technology, the frequency of roof falls in extremely soft high extraction height working faces is reduced by 80%, the average depth of roof falls is decreased by 35%, and the average length of roof falls is reduced by 50%. The stability of the coal ribs is effectively improved.

To investigate the evolution characteristics of overlying rock fractures, based on a geological prototype of a large-height comprehensive mining face in Shanxi, a three-dimensional large-scale physical similarity model was established. The experiments were carried out using microseismic monitoring and physical model cutting methods to study the activity and fissure evolution of the overburden rock. Model cutting revealed that, approximately 65 m from the bottom of the coal seam, delamination occurred, marking the top of the overburden rock fissure zone and the bottom of the bending and sinking zone. At 25 m from the coal seam bottom, the rock layer was highly fragmented, forming the collapse zone, which was 4.8 times the mining height. Between 25 and 65 m from the bottom, a fissure zone existed, which was 12.5 times the mining height, with abundant delamination fissures at the top of the fissure zone. Significant microseismic events were observed as the coal face advanced to 45 m, with notable increases in the concentrations and distribution ranges of these events in both the strike and height directions of the coal seam. The subsidence range of the overlying rock layer expanded from the top to the bottom, with the subsidence slope area extending gradually and the central compaction area remaining relatively flat. The overall shape presented an irregular ellipse, with peripheral uplift phenomena observed in the subsidence area. At 39 m from the coal seam bottom, the maximum subsidence of the rock stratum was 4.0 m, with subsidence amounts decreasing with increasing stratum height. Fissure density along the coal seam inclination and direction exhibited a double hump pattern, with fissure areas on both sides showing high densities and the central compaction areas having low densities. Coal seam mining caused stress redistribution in the surrounding rock layer, and the stress in front of the work was divided into the stress reduction zone, dynamic influence zone, mining influence zone, and unaffected zone. Coal rock porosity under high stress was less sensitive to stress changes, resulting in smaller changes in fissure permeability and fissures remaining mostly closed. [ABSTRACT FROM AUTHOR]

To address the issues of coal rib instability and high occurrence of roof falls in high extraction height fully mechanized coal mining of soft coal seams, this study comprehensively employs theoretical analysis, numerical calculations, and field industrial experiments. By analyzing the distribution characteristics of coal rib displacement and deformation instability zones under different mechanical parameters of soft coal seams, the instability mechanism of coal ribs in high extraction height mining of thick and soft coal seams is revealed, and corresponding stability control techniques for high extraction height coal ribs and supports are developed. The study shows that the stability of coal ribs in high extraction height mining fields of soft and thick coal seams decreases as the values of mechanical parameters decrease. In other words, the smaller the cohesion and deformation parameters of the coal body, the more prone the coal rib is to roof falls. By utilizing advanced deep‐hole static water infusion technology, the frequency of roof falls in extremely soft high extraction height working faces is reduced by 80%, the average depth of roof falls is decreased by 35%, and the average length of roof falls is reduced by 50%. The stability of the coal ribs is effectively improved. [ABSTRACT FROM AUTHOR]

When the coal mining face enters the final stage of mining, the roadway faces the superimposed influence of surrounding rock stress redistribution and roof rotary moment. As affected by the strong disturbance in the coal mining process, the roof plate of the roadway has undergone serious deformation, which seriously affects the stability of the roadway. Taking the 108 working face of the Jinjitan coal mine as the engineering background, a comprehensive study was conducted on the control of the perimeter rock in the retracement of a tunnel in a heavy coal seam with a large mining height. By analyzing the physical properties of the enclosing rock of the retreated roadway, and using theoretical analysis, numerical simulation, on-site monitoring, and other methods, the characteristics of the peripheral rock's movement relationship and mineral pressure manifestation in the final mining stage of the large-height working face have been studied. The structural mechanics model was established, and in the case where the support cannot be solved just by strengthening the support, the design scheme of "blasting roof break + constant resistance anchor cable support" was innovatively tried. FLAC3D simulation results show that the stress release of the surrounding rock is more adequate when the height of roof cutting is 20 m. The stress of the surrounding rock near the roadway is reduced by 30~40%, and the stress state is reasonable. The constant resistance and large deformation anchors can absorb the deformation energy of the rock body, maintain constant working resistance and stable deformation, and have good rock stability control, which is conducive to the stability of the roadway. [ABSTRACT FROM AUTHOR]

Abstract To study the fracture instability characteristics and fracture mechanism of overlying strata in steeply dipping and large mining height stope, through the combination of physical simulation experiment, numerical calculation, theoretical analysis and field measurement, the correlation between the evolution of mining stress field and the fracture of overlying strata in large mining height stope under the effect of dip angle is analyzed, and the stress transfer path and the fracture mechanism of overlying strata are revealed. Finally, the influence mechanism of key strata on the evolution of the mining stress field of overlying strata is explained. The research shows that under the influence of gravity dip angle effect and mining height increase, the mining stress is non-equilibrium transferred and evolved in space, and the principal stress field presents the characteristics of partition evolution. The fracture trajectory of rock strata is multi-step and “八” shaped, and the inclined masonry structure and multi-step key strata form a cooperative bearing structure. The key layer of the ladder is the stress transmission structure between the caving zone and the stress arch, and the high stress in the overhanging area under the main roof is transmitted to the advanced roof. With the increase of dip angle, the height of the caving zone decreases, and the fracture range of ladder and principal stress arch decreases. The cracks in the middle and upper broken rock blocks increase, and the instability of the “high ladder rock layer” is easy to induce the simultaneous fracture of the “inclined masonry structure”, resulting in regional unbalanced instability and impact. The research results can provide some reference and guiding significance for the stability control of surrounding rock in longwall stope with steeply dipping.

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In order to improve rib stability, failure criteria and instability mode of a thick coal seam with inter-band rock layer are analysed in this study. A three-dimensional mechanical model is established for the rib by considering the rock layer. A safety factor is defined foy the rib, and it is observed that the safety factor exhibits a positive correlation with the thickness and strength of the inter-band rock. A calculation method for determining critical parameters of the rock layer is presented to ensure the rib stability. It is revealed that incomplete propagation of the fracture at the hard rock constitutes a fundamental prerequisite for ensuring the rib stability. The influence of the position of the inter-band rock in the coal seam on failure mechanism of the rib was thoroughly investigated by developing a series of physical models for the rib at the face area. The best position for the inter-band rock in the coal seam is at a height of 1.5 m away from the roof line, which tends to provide a good stability state for the rib. For different inter-band rock positions, two ways of controlling rib by increasing supports stiffness and flexible grouting reinforcement are proposed.

When the coal mining face enters the final stage of mining, the roadway faces the superimposed influence of surrounding rock stress redistribution and roof rotary moment. As affected by the strong disturbance in the coal mining process, the roof plate of the roadway has undergone serious deformation, which seriously affects the stability of the roadway. Taking the 108 working face of the Jinjitan coal mine as the engineering background, a comprehensive study was conducted on the control of the perimeter rock in the retracement of a tunnel in a heavy coal seam with a large mining height. By analyzing the physical properties of the enclosing rock of the retreated roadway, and using theoretical analysis, numerical simulation, on-site monitoring, and other methods, the characteristics of the peripheral rock’s movement relationship and mineral pressure manifestation in the final mining stage of the large-height working face have been studied. The structural mechanics model was established, and in the case where the support cannot be solved just by strengthening the support, the design scheme of “blasting roof break + constant resistance anchor cable support” was innovatively tried. FLAC3D simulation results show that the stress release of the surrounding rock is more adequate when the height of roof cutting is 20 m. The stress of the surrounding rock near the roadway is reduced by 30~40%, and the stress state is reasonable. The constant resistance and large deformation anchors can absorb the deformation energy of the rock body, maintain constant working resistance and stable deformation, and have good rock stability control, which is conducive to the stability of the roadway.

Coal wall of large mining height face in steeply dipping coal seam has large free height, wide free movement space and poor self-stabilization balance, coal wall is not an isolated body in the face, it can form a load-bearing structure with the surrounding rock, supports and other mediators, and the mining behavior is closely linked. The dip angle effect can easily promote the dissimilation of the coal wall bearing environment, complicate its response behavior, increase the difficulty of coal wall and surrounding rock stability control, and restrict the safe and efficient production of the face. In order to solve the problem of coal wall stability control of large mining height face in steeply dipping coal seam, theoretical analysis and numerical calculation are comprehensively used for research. The stress in the plastic zone of large mining height face in steeply dipping coal seam increases exponentially, there is an asymmetric arch residual stress influence zone in the vicinity of the coal wall, the plastic zone breadth is dissimilated in different regions, and the width from large to small is the upper, middle and lower part, and the distribution pattern is in the shape of stepped arch, coal in the plastic zone bears repeated pressure, and will be enhanced with the expansion of plastic zone. As the mining height increases, the residual abutment pressure near the coal wall decreases, the compressive strength, displacement value and disturbed range of the coal in front of the coal wall will increase. The stress and transport of the coal wall under the dip angle effect develop in a zonal manner, in which the stress distribution is lower> upper> middle, while for the displacement is middle> upper> lower. In addition, the working face inclination under the oblique angle will lead to the transformation of the coal wall instability mode, when the oblique angle is large, the dip angle of face is small, the compression component of abutment pressure is enhanced, and the external convex spalling mainly occurs. On the contrary, it’s the sliding instability under the coupling of mining stress and self weight. The comprehensive analysis shows that the coal wall mining behavior is regionally heterogeneous under the coupling effect of dip angle and mining height, and because the stress unloading and migration in the middle and upper part of the face are larger, therefore the two regions should be the key prevention and control areas for coal wall instability of large mining height face in steeply dipping panel.

AbstractIn this paper, taking a mine in Xinjiang as the engineering background, through the combination of physical simulation experiment, 3DEC numerical calculation, theoretical analysis and field monitoring, the fracture evolution characteristics of overlying strata and its influence on the stability of support under unbalanced loading conditions are systematically studied. The results show that under the action of gravity-dip effect, the non-uniform filling of gangue leads to the typical cross-layer migration characteristics of overburden caving shape and roof bearing arch. The unstable inclined masonry structure is formed in the middle and upper areas, which is prone to rotation and sliding instability, forming impact pressure. The mining stress of the overlying strata changes significantly due to the mining of the coal seam. The lower the overlying strata, the greater the peak stress, the smaller the stress value in the goaf, and the greater the stress change. The amplitude and instability probability of the support migration increase with the decrease of the working resistance of the support, the increase of the friction between the roof and the support, and the increase of the partial load of the support. Finally, the prevention and control measures of roof caving crushing support in steeply dipping and large mining height working face are put forward.

Copyright of Coal Science & Technology (0253-2336) is the property of Coal Science & Technology 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.)

In order to improve the recovery rate of coal resources in general, Chengzhuang Coal Mine has carried out and implemented the technology of gob-side entry retaining in 5309 large height mining face. In view of the complex conditions such as 6 m thick coal seam and hard thick rock stratum in the roof, the application of directional blasting presplitting and flexible formwork concrete technology in the mining face of “thick coal seam and hard roof” is emphatically discussed by means of theoretical analysis and field test. The results show that the deformation of gob side entry retaining roadway develops rapidly within 10-90 m behind the working face; the wall pressure of gob-side entry retaining reaches the maximum within 50-100 m behind the working face; the cooperative application of blasting presplitting, roof cutting and pressure relief and flexible formwork concrete wall not only reduces the influence of periodic pressure on the roof of the roadway reserved along the goaf, but also enhances the bearing capacity of the roadway side, which obviously improves the overall stability and reliability of the roadway.

The coal walls in a caving face with a tall mining height are prone to rib spalling, which leads to the phased cessation of the mining of the working face, causes heavy losses, and endangers the safety of underground workers. In order to prevent serious rib spalling accidents of coal walls in fully mechanized caving faces with a large mining height and to improve the prediction of and ability to control rib spalling, a load-bearing mechanical model of the roof–coal wall–support system was established based on the moment-balance relationship. The expressions for the deformation and stress distribution in a coal wall were calculated. Then, the influences of key factors on the horizontal displacement of the coal wall were investigated. A numerical simulation model of the working face was established, and an orthogonal test design was introduced. On this basis, the influences of four factors: cutting height, breaking position of the main roof, support strength, and sidewall protecting force of the support on the horizontal displacement and volume of a plastic zone of coal wall, were analyzed. Moreover, their order of importance was ranked on the basis of sensitivity. Based on the engineering conditions and production practices in the Cuncaota II Coal Mine, key parameters for controlling and measures for preventing the rib spalling of the coal wall are proposed to guide practical actions. [ABSTRACT FROM AUTHOR]

Intense cutting-induced dust production in fully mechanized mining faces (FMMFs) with large mining heights produces a high amount of dust that is difficult to capture and severely affects the working environment, threatening the health of occupational staff. The effective spray range and atomization performance of tracking sprays are counteracted by the influences of the mine's height and ventilation airflow in FMMFs. Thus, optimizing the spray's parameters and relationship between the effective spray range and atomization performance to reduce dust levels is the main priority of dust-control techniques. In this study, a new swirl-core atomization nozzle is developed based on fluid mechanics and the solid–liquid coalescence mechanism. The liquid generates a circumferential velocity when passing through the swirl core, which considerably increases the droplet breaking power and reduces the droplet cohesion factor, achieving a remarkable atomization effect. The spray angle of the new nozzle is 57°, which is 80.9% greater than the GZPW-16 mine-use nozzle (31.5°); the effective spray range increases from 5.2 to 5.9 m; and the spray's mist saturation is significantly better than the GZPW-16 mine-use nozzle. Under different test pressures, the particle size range of the droplets produced by the new nozzle and dust particles on site satisfied the best synergy of droplet–dust coalescence. The total and respirable dust-reduction rates were 78% and 75.1%, respectively, which were 42% and 65% higher than those of the original nozzle. The new nozzle effectively improves the efficiency of the single dust-control technique of the tracking spray, which is significant for the dust-prevention and -control technology of FMMFs with large mining heights. [ABSTRACT FROM AUTHOR]

Double-roadway tunneling could mitigate the contradiction between mining production needs and tunneling speed, which is pivotal to the sustainable development of underground mines. However, it is very difficult to control the stability of a mining roadway on an adjacent working face suffering from strong mining disturbance due to double-roadway tunneling, especially at a large mining height working face. In order to control the stability of the return air roadway (RAR) 23205 of a strong mining roadway at working face 23205 in the Zhuanlongwan Coal Mine in Inner Mongolia, we carried out field monitoring, theoretical analysis, numerical simulations, and engineering practice to identify the main factors influencing the deformations and the stress distribution law of the surrounding rock in order to propose countermeasures for strong mining roadways. The results show the factors influencing the large deformation of strong mining roadways include large mining height, repeated mining, stress concentration due to the large coal pillar, and a small thickness of the anchorage layer in the roof. The stress peak in the central coal pillar caused by the first and second mining is 23.19 MPa and 27.49 MPa, respectively, and the stress concentration coefficients are 4.538 and 5.379, respectively. Countermeasures (pressure relief via large-diameter boreholes in the large coal pillar and long anchorage for roof reinforcement) were created to control the stability of a strong mining roadway, i.e., RAR 23205. Field measurements indicated that deformations in RAR 23205 could be efficiently controlled. The maximum deformation of the surrounding rock was 50 mm, which meets the safety and efficient production requirements of the coal mine. In addition, new roadway layout optimization and control countermeasures are put forward to control the stability of mining roadways. [ABSTRACT FROM AUTHOR]

With the increasing coal cutting height of a longwall panel with a large mining height, joints and fractures have a more obvious influence on coal wall stability, which intensifies rib spalling. To investigate the influence of the fracture distribution on the stability of a coal wall with a large mining height, this paper utilizes a large panel in the Zhaogu No. 2 coal mine as the research object. The coal strength and fracture occurrences are measured to obtain the probability distribution function of the fracture spacing and density. A 3D numerical model of the panel with a non-through fracture is established by using 3DEC software based on the Poisson’s disk model and Monte Carlo principle, and numerical experiments of rib spalling and the optimization of flexible reinforcement control technology are conducted. The method of prefabricating non-through fractures and physical experiments are proposed to analyze the stability and damage degree of the coal wall under different fracture tendencies. The results show that the existence of fractures changes the direction and magnitude of the stress field, thus making the stress and displacement of the coal wall show anisotropy, revealing a high correlation between the densities and sizes of joints and the coal strength. When the joints are parallel to the coal wall with a 45° dip angle, the coal wall stability is the most unfavorable with the largest damage range. The optimal grouting position and flexible material diameter are determined; i.e., a grouting position at 60–70% of the coal wall height and a material diameter of 20 mm exhibit the best control effect. The research results provide a new approach for the control of rib spalling in fractured coal. [ABSTRACT FROM AUTHOR]

In order to solve the problems of vicious development of surrounding rock fissures, severe mining disturbance, weak bearing capacity of narrow coal pillar and damage and failure of roadway support system of gob-side entry driving with large section of large mining height in deep coal mine, taking the small coal pillar roadway excavation of 20203 working face of Dahaize Coal Mine as the engineering background, the internal and external stress field, limit equilibrium theory, numerical calculation simulation, engineering practice are used. It is comprehensively determined that the reasonable width of narrow coal pillar of gob-side entry driving with large section of 6.8 m mining height in deep coal mine is 6.5 m, and the surrounding rock comprehensive control technology of “high-strength concrete bedding + high prestressed bolt support of roadway sides + roof flexible bolt strong support + diamond metal mesh” is proposed. The research shows that: the width of internal stress field under this geological condition is 14.4 m, the width of coal pillar crushing area after mining previous working face is 3.1 m, and the reasonable width range of narrow coal pillar is 5.8 to 8.2 m; when the coal pillar size is 6.5 to 9.5 m, the stress of the coal pillar is improved, and the elastic zone appears above the coal pillar and the roadway roof, which can give full play to the anchoring effect of bolt (cable) to ensure the stability of the roof and sides; the convergence deformation of the roadway tends to be stable 32 days after the support is completed. The maximum deformation of the solid coal wall, coal pillar wall, solid coal side roof and coal pillar side roof are 52 mm, 60 mm, 62 mm and 67 mm respectively. During the mining of this working face, the overall deformation of the roadway is in a controllable range. The 6.5 m narrow coal pillar and comprehensive control technology of surrounding rock can realize the stable bearing of narrow coal pillar and long-term effective control of surrounding rock of gob-side driving with large section of large mining height in deep coal mine.

To study the mechanical properties of high water filling materials, achieving the purpose of filling the side of the roadway, through uniaxial compression tests on high water materials with different water-cement ratios, we analyze the relationship of uniaxial stress-strain and the relationship of strength and maintenance at different stages. Taking Xishan Coal and Electricity Tunlan Mine as the engineering background, the deformation characteristics of the surrounding and the stress evolution law were investigated by numerical analysis. The results show that with the increase of water-cement ratio, the strength of high-water filling material decreases obviously, the uniaxial compressive strength of the material is 6.62 MPa when the water-cement ratio is 1.5∶1, 4.08 MPa when the water-cement ratio is 2∶1, and 2.85 MPa when the water-cement ratio is 2.5∶1, which are reduced by 38% and 60%, respectively; in the tunneling stage, the maximum roof and floor movement is 45.6 mm, and the maximum roof and floor movement is 203.0 mm during mining. During the period of roadway retention, the roadway is greatly affected by mining within 30 m of the advance working face, and the roof obviously sinks.

The mining technology of gob-side entry retaining is widely adopted by China’s coal mining enterprises because of its high economic significance. However, gob-side entry retaining also has some problems, such as long service time, many times of mining influence, and difficult surrounding rock support. In order to explore the key technology of surrounding rock control of gob side entry retaining in 3304 working face, this paper studies the influence of roadway side filling parameters on surrounding rock control of gob side entry retaining by using research methods such as theoretical analysis, numerical simulation and field measurement, and proposed the bolt asymmetric support scheme of gob side entry-retaining through the “three high and one low” bolt support design concept, through the comparison of numerical simulation test and field practice test, it is concluded that the reasonable parameters of roadway side support body and the asymmetric support scheme of anchor bolt have obvious effect on the control of surrounding rock of gob-side entry retaining roadway.

In recent years, with the rapid development of equipment manufacturing and advances in the research on mine pressure and rock mechanics, large-mining-height comprehensive mechanized coal mining technology has become increasingly widely used in thick-seam mining. On this basis, because the free space extracted is several times that of layered mining, the occurrence of strong ground pressure caused by the violent movement of the basic roof and the low recovery rate of the top coal have become the main problems. In this study, Dong Liang Coal Mine was taken as the engineering background by which to study the three typical movement forms of the basic roof, and the critical fracture conditions of the three forms were obtained through theoretical calculations. The movement result state of the basic roof strata was taken as the different overlying boundary conditions of the top coal to simulate its impact on the recovery rate, to explore the interaction mechanism between the basic roof strata and the top coal, and for use as a theoretical basis to guide engineering practice to improve energy recovery rates and safety efficiency. [ABSTRACT FROM AUTHOR]

The stability of the roof above the filling body is a key factor affecting the effect of gob-side entry retaining. In view of the problem of roof separation control of gob-side entry retaining at 22301 working face with large mining height in Tunlan Coal Mine, by analyzing the stress characteristics and roof separation mechanism of each stage during the roadway retaining, the mechanical model of roof separation above the filling body was established, the partition dynamic reinforcement control technology of the straight roof above the gob-side entry retaining filling body in thick coal seam was proposed, that is “anchor cable”+ “high pre-stressed anchor cable”+ “high resistance hydraulic prop” stage support for the entry retaining. The field test results show that the maximum subsidence of the roof is 140 mm, and the floor heave is 660 mm. Outside the 150 m behind the working face, the cumulative separation of 0 m to 6 m is 16 mm, and the cumulative separation of 0 m to 2.5 m is 14 mm. The surrounding rock deformation is dominated by floor heave. The overall stability of the roof supported by sub-regional dynamic support reinforcement technology is significantly improved.

The roof in a fully-mechanized face of a shallow coal seam with large mining height is prone to form a combined cantilever—articulated rock beam structure. When the support resistance is insufficient, the articulated rock beam will sink. This will make the combined cantilever beam rotate and fracture. It is then easy to induce the sliding and instability of the articulated rock beam, which results in large-scale roof cutting and the support crushing. Taking the combined cantilever beam structure as the main research object, and considering the mining damaged characteristics of cantilever beam rock stratum, the rock beam was regarded as a finite plate model with an edge crack of arbitrary dip angle. In addition, a fracture mechanics model controlled by a set of structural planes was established, the instability conditions of rock beam and the main control factors were analyzed, and the method of determining the support resistance were discussed. The results show that the cantilever beam rotates and fractures. This causes a chain reaction of the rock beam that leads to fracture. The combined cantilever beam then loses stability with the increase of the length of the crack length and the crack dip angle, and therefore it is easier to penetrate the cantilever beam and cause roof instability. The necessary condition for rock beam instability was crack activation, and the sufficient condition was the airfoil branch crack propagate through the rock beam. The influence degree of each parameter on the support resistance was thus determined: crack length a > crack dip angle β > rock thickness h > weighting interval l. The theoretical analysis results were proven to be reasonable by an in situ monitoring example of no. 22,310 working face in the Daliuta coal mine, China. On this basis, the reasonable value of support resistance was obtained. The conclusions of this research provide a new method for researching the roof instability mechanism. They are also conducive to the green and sustainable development of mines.

As an important technology of thick coal seam mining, fully mechanized mining with a large mining height has high mining efficiency. In order to study the roof safety control of large mining height working face, the 122106 working face of Caojiatan coal mine is taken as the engineering background. The numerical simulation method is used to analyze the control ability of roof subsidence when the support strength is 1.2 MPa, 1.4 MPa, 1.6 MPa, 1.8 MPa, 2.0 MPa, and 2.2 MPa. The results show that the support strength of hydraulic support is negatively correlated with roof subsidence. Through theoretical analysis of the mechanical model of the support and surrounding rock under the filling condition, it is shown that the height of the gap between the filling body and roof is the main influencing factor of roof subsidence: the smaller the height of the gap between the filling body and roof, the better the control effect on the roof. Through numerical simulation, the roof subsidence and surface subsidence under different filling rates are analyzed. The results show that when the filling rate increases to 80% the control of roof subsidence achieves better results. Taking production safety and economic benefits into consideration, when the reasonable support strength of the working face is determined to be 2.0 MPa and the filling rate is 80%, the safety control of the working face roof can be ensured. [ABSTRACT FROM AUTHOR]

Copyright of Coal Science & Technology (0253-2336) is the property of Coal Science & Technology 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 coal walls in a caving face with a tall mining height are prone to rib spalling, which leads to the phased cessation of the mining of the working face, causes heavy losses, and endangers the safety of underground workers. In order to prevent serious rib spalling accidents of coal walls in fully mechanized caving faces with a large mining height and to improve the prediction of and ability to control rib spalling, a load-bearing mechanical model of the roof–coal wall–support system was established based on the moment-balance relationship. The expressions for the deformation and stress distribution in a coal wall were calculated. Then, the influences of key factors on the horizontal displacement of the coal wall were investigated. A numerical simulation model of the working face was established, and an orthogonal test design was introduced. On this basis, the influences of four factors: cutting height, breaking position of the main roof, support strength, and sidewall protecting force of the support on the horizontal displacement and volume of a plastic zone of coal wall, were analyzed. Moreover, their order of importance was ranked on the basis of sensitivity. Based on the engineering conditions and production practices in the Cuncaota II Coal Mine, key parameters for controlling and measures for preventing the rib spalling of the coal wall are proposed to guide practical actions.

The gob-side retaining can save coal pillar of mining area and improve coal recovery rate of working face. Meanwhile, this technology helps relieve the tension of coal mining, and reduce the problem of gas accumulation in the tracking corner of working face. Therefore, the gob-side entry retaining ranks in the main development goals of efficient coal mining. The thickness, dip and buried depth of coal seam in the mentioned 3304 working face are large, so it is difficult to control the surrounding rock of the retaining roadway. This paper mainly studies the safe and reliable supporting parameters of the gob-side entry retaining in the 3304 mining area. Firstly, the width and strength of roadway wall filling body and bolt support scheme were designed by orthogonal test method. Then, numerical simulation software is used to comprehensively analyze the influence of the strength and width of roadside support body on the surrounding rock pressure. And the deformation law of surrounding rock under different reinforcing support parameters is further analyzed. Finally, the reasonable support is determined. The width of roadway side filling body is 1.5 m; The strength of the concrete used is C25; The practice proves that the roadway retention scheme and support parameter adopted are effective, which have reference value for similar projects. [ABSTRACT FROM AUTHOR]

In order to study the law of ground pressure appearance in the face with large mining height in shallow coal seam and determine its reasonable working resistance of support, the 5-2 coal 503 working face of Daliuta Coal Mine in Shendong is selected as the research object. The law of ground pressure appearance in the face with large mining height in shallow coal seam is studied through the thin plate theory and the field measured ground pressure data, and the support resistance is calculated by combining two different theoretical calculation formulas force analysis. The results show that the initial step distance of 52503 working face is 64.4 m, which is basically consistent with 61.7 m calculated based on thin plate theory; the periodic weighting step fluctuates between 7.65 m and 29.75 m during the normal mining period, with an average of 19.16 m, a dynamic load coefficient of 1.45-1.57, an average of 1.50, which shows a strong mine pressure; during the weighting period, the support resistance along the inclined direction of the working face presents a “convex” type among them, the lower support resistance of the working face is about 69.8% of the middle part, and the upper support resistance of the working face is about 66.5% of the middle part; the working resistance calculated based on the two support strength theories is 17 871 kN and 19 890 kN respectively, which is smaller than the ZY21000/36.5/70D hydraulic support with the rated working resistance of 21 000 kN, which can meet the support requirements of the roof pressure above the working face.

This paper systematically summarizes the development process of fully mechanized top coal caving technology and equipment in Yankuang Group in recent 30 years. On the basis of the introduction test in China and supported by the national scientific research in the ninth “five-year” plan and the tenth “five-year” plan, the annual output of fully mechanized top coal caving mining arrived 6 Mt for the first time; through the innovation of technology and equipment such as hydraulic support frame type, coal caving theory and technology, the export of automatic coal caving and fully mechanized caving mining technology has been realized, aiming at the problem of fully mechanized top coal caving mining in thick and extra thick coal seam and hard coal seam under complex conditions, a complete set of technology of “mining, caving, supporting and transportation” was developed, the coupling relationship between hydraulic support and surrounding rock in fully mechanized caving mining with large mining height and super large mining height is studied, a complete set of technology and equipment for fully mechanized top coal caving mining with 5.5 m four-pillar and 7 m two-pillar with super large mining height is designed. Then the adaptability, mining efficiency and intelligent level of fully mechanized top coal caving mining are further improved.

Based on the theoretical analysis, numerical simulation and laboratory test, the instability mechanism and characteristics of the coal wall under the action of dynamic load are studied.A PFC2D biaxial compression numerical test model is established, and the fracture is planted in the model to study the scale effect of coal body and coal sample under static load.It is concluded that the number, continuity and distribution of natural defects in coal seam have obvious influence on size effect.The larger the sample is, the lower the strength is, and the change trend is logarithmic curve.Further, the mechanical properties of coal samples under static and dynamic loads are studied by Hopkinson test.It is considered that there is a good consistency between the peak strength obtained from dynamic load test and the stability of coal wall, which reflects the influence of fine joint structure and defects on coal wall stability under the influence of mining.Taking Yangquan mining area as an example, the dynamic and static peak strength of coal samples No.15 from mine No.1, mine No.2, mine No.5 and Xinjing mine are compared with the actual height of coal wall stability.It is concluded that the dynamic peak strength of coal sample can better reflect the influence degree of mining stress.on the stability of coal wall, which provides a theoretical basis for the definition of scientific mining height under strong disturbance.

The timing failure of basic roof of large mining height stope in steeply dipping coal seam results in the sequential pressure of the working face, the bearing state of supports in different regions is dissimilar, and the difficulty of surrounding rock control and production risk increase. In order to find out the failure mechanism of the sequential failure of the basic roof, the temporal failure characteristics of the basic roof, bearing structures in different regions of the working face, and the periodic collapse length of the basic roof were studied by field measurement and theoretical analysis. The results show that the timing failure characteristics of the basic roof of the high-inclination working face are the first middle part, the second upper part, the next lower part, and the broken rock mass (gangue) asymmetrically filling the goaf, which results in the different mechanical performance of gangue to overburden in different areas of the working face, forming a partitioned mechanical bearing structure; the basic roof is in the mechanical environment of different strength, and the action of different strength force leads to different falling step distance of the basic roof, thus causing the timing failure.

Intense cutting-induced dust production in fully mechanized mining faces (FMMFs) with large mining heights produces a high amount of dust that is difficult to capture and severely affects the working environment, threatening the health of occupational staff. The effective spray range and atomization performance of tracking sprays are counteracted by the influences of the mine’s height and ventilation airflow in FMMFs. Thus, optimizing the spray’s parameters and relationship between the effective spray range and atomization performance to reduce dust levels is the main priority of dust-control techniques. In this study, a new swirl-core atomization nozzle is developed based on fluid mechanics and the solid–liquid coalescence mechanism. The liquid generates a circumferential velocity when passing through the swirl core, which considerably increases the droplet breaking power and reduces the droplet cohesion factor, achieving a remarkable atomization effect. The spray angle of the new nozzle is 57°, which is 80.9% greater than the GZPW-16 mine-use nozzle (31.5°); the effective spray range increases from 5.2 to 5.9 m; and the spray’s mist saturation is significantly better than the GZPW-16 mine-use nozzle. Under different test pressures, the particle size range of the droplets produced by the new nozzle and dust particles on site satisfied the best synergy of droplet–dust coalescence. The total and respirable dust-reduction rates were 78% and 75.1%, respectively, which were 42% and 65% higher than those of the original nozzle. The new nozzle effectively improves the efficiency of the single dust-control technique of the tracking spray, which is significant for the dust-prevention and -control technology of FMMFs with large mining heights.

In order to master the roof failure structure and mine pressure development law of large mining height working face, the roof collapse laws of different mining heights are calculated and analyzed, and the reasonable working resistance of support is determined by dynamic load method. Principal results are as follows: the overhang area of the main roof of the large mining height face increases, and the mine pressure appears obviously; when the mining height is small and the direct roof is thick, the lower direct roof can fill the goaf after the collapse of the lower direct roof, while the upper direct roof is broken and acts as the bearing layer; the mine pressure on the working face is not obvious during the first and periodic pressure; by comparing the distribution law of plastic zone of surrounding rock with different mining heights, the lower direct roof caving is full of goaf and the upper direct roof cannot bear the weight of overlying strata bending and yielding under the action of tensile stress while the mining height is small; when the mining height increases to the point that all the caving of the direct roof is filled with goaf, the old roof forms a bearing structure, and the surrounding rock has fewer failure surfaces. In large mining height, the increase of the overhang area of the old roof causes the increase of abutment pressure in front of the coal wall of the working face, and there are many direct roof failure surfaces. The height of caving zone and fracture zone increased linearly with the increase of mining height; the ZY15000/33/72D hydraulic support was introduced to meet the production requirements according to the calculated working resistance of support.

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Fracture of a critical stratum on a coal seam roof may cause strong strata pressure behaviours. In particular, a large-scale sudden fracture of the critical stratum, which is formed by hard thick sandstones on a mining face with a large mining height, can easily cause dynamic disasters of wells. In this study, a simulation analysis of the deformation and failure laws of a hard thick roof in an island mining face with a large mining height in Xinji 2# Mine 210106 after the mining of Coal bed A1 was performed in accordance with the occurrence characteristics of sandstones. Results demonstrate that large-scale fracture suddenly occurs at a gob roof after the working face is advanced by 120 m. A mechanical analysis model of the island working face was constructed. This model comprises 17.9 m of fracture on the 16 m thick sandstone roof. Hydraulic support provides 16854.1 kN of supporting resistance to maintain the stability of the rock beam and thus prevent threats to the supporting devices of the working face. The hard thick roof was weakened every 15 m in the transport gateway and air return in advance through an advanced deep-hole pre-splitting blasting technique. In accordance with the measurement results of underground pressure, the forced caving of thick sandstones relieves the influence of roof loads on the support and thus realizes safe mining.

In order to obtain the dust distribution of PM5 and PM10 in large mining height fully-mechanized face, the dust concentration around the shearer drum and along the downwind side of the shearer are measured under the conditions of the following wind and upwind respectively. The results show that when coal is cut in following wind, the initial concentration of PM5 and PM10 dust around the rear drum is high, then gradually decreases and becomes stable. The concentration of the front drum is low at first, and rises sharply after 70 to 80 s. When coal is cut in upwind, the initial concentration of the front drum fluctuates greatly, and the concentration of the rear drum rises about 10 s earlier than that in following wind. When coal is cut in following wind, the PM5, PM10 and total coal dust concentration reach the highest at the downwind of shearer around 10 m, then gradually reduce and tend to be stable at 40 m. The PM5 proportion is higher at first, and falls to the lowest at about 15 m, then rises again under the effect of wind. The PM10 proportion is on the decline overall. When coal is cut in upwind, the dust concentration reaches the highest at the downwind of shearer 6 m or so, and is stable about 20 m. PM5 accounts for 60% to 80%, while PM10 accounts for 65% to 95% under both conditions.

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In view of the characteristics of large overburden movement space, strong mining disturbance and difficult to determine the initial pressure step distance of the roof in fully mechanized top-coal caving face with large mining height, the working face 101 of a coal mine is taken as the engineering background, through theoretical analysis, numerical simulation and field measurement methods are used to analyze the roof breakage characteristics and mining pressure appearance law of fully mechanized top-coal caving face with large mining height. Firstly, the mining conditions of the working face are systematically introduced, and the influence of the fully mechanized top-coal caving face with large mining height mining conditions on the roof is analyzed. It is obtained that the direct roof collapses and forms an irregular upper direct roof and a regular lower direct roof. Large massive fine-grained sandstone above the working face are extruded into each other to form a semi-arched "rock-gangue" structure. The UDEC software is used to conduct numerical simulation to verify the initial caving step distance of the direct roof of the top coal, mudstone and fine-grained sandstone, and the broken structure of the roof of the fully mechanized top-coal caving face with large mining height under different propulsion lengths and the plastic zone and stress distribution of the roof are obtained. The research shows that: the first caving step distance of top coal is 14.52 m, the first caving step distance of mudstone is 24.06 m, and the first caving step distance of fine-grained sandstone is 47.83 m; in the special case of failure and instability of "rock-gangue" structure, the reasonable support strength PT of the support shall be no less than 1.13 MPa to meet the requirements of roof support. Through the online monitoring of the working resistance of the hydraulic support through the support resistance monitoring system of the fully mechanized top-coal caving face 101 with large mining height, and through the analysis of the working resistance of the hydraulic support, the breaking characteristics of the roof of the fully mechanized top-coal caving face with large mining height and the regularity of the rock pressure are verified. [ABSTRACT FROM AUTHOR]

To analyze the feasibility of fully mechanized large mining height long wall top-coal caving mining (LLTCC) in ultra-thick (20–30 m), parting-rich coal seams, a case study on No. 6 coal seam of the Laosangou mining field was conducted. First, the distribution pattern of the coal-seam structure was mastered and generalized based on the coal-seam thickness, parting thickness, and parting horizon. On this basis, physical simulation and UDEC simulation were performed to analyze the top-coal fracture characteristic and the impact of partings. The top-coal caving of No. 6 coal seam was systematically evaluated with statistical results of coal-seam recovery height (h re) and mining–caving ratio (R mc). Finally, the range of mining height (h) at LLTCC faces was determined as an indicator for the division of LLTCC feasible regions in the Laosangou mining field, which is based on the R mc, h re, coal-wall stability, and equipment selection. The results show that the top-coal caving structure of LLTCC faces can be divided into four layers which were effected by top-coal thickness and upper-horizon partings. LLTCC in ultra-thick (20–30 m), parting-rich coal seams was highly feasible. The top-coal cavability of No. 6 coal seam was good overall, and the R mc was about 1:2.89–1:3.91. Direct mining regions (h = 4.0–5.3 m) and buffer mining regions (h = 5.3–6.0 m) of LLTCC were 52.66% and 42.96%, respectively. This study can be used as a reference for the design of LLTCC in similar coal mines. [ABSTRACT FROM AUTHOR]

In order to solve the problems of large deformations of gob-side entry and the fracture of coal pillars at large-height mining faces in deep mines, a case study of the width design of coal pillars between a slice mining face and a large height mining face is studied by theoretical calculation, numerical analysis, and site engineering measurement. In this paper, the occurrence of the Ji16-17 coal seam and its mechanical properties are analyzed. In addition, driving disturbances on roadway deformations, and the effect of plastic area distributions under conditions of different widths of coal pillars and different falls in heights are studied when mining at a slice face turned to a large height face in a deep mine. The results show that the reasonable width of a coal pillar is 8 m with a fall in mining height of 2.5 m between the Ji16-17–24011 slice face and the Ji16-17 −24030 large mining height face (at a 5.5 m mining height). The monitored results show that at a mining height of 5.5 m, an 8 m panel pillar can effectively maintain the stability of a pillar and gob-side entry, and can meet the requirements for the safe production of a stable transition from the slice mining face to the large mining height face. [ABSTRACT FROM AUTHOR]

Roof sliding and instability along the coal wall usually occur in the working face at large mining heights during the process of the first weighting, which causes roof cutting and support crushing. A mechanical model consists of the main roof, immediate roof, and support based on the nonlinear characteristics of the failure and instability of the immediate roof under the abutment pressure, which we constructed to study the step sinking of the main roof, as well as to assign the reasonable value of the support resistance during the first weighting. The instability mechanism of the system was studied by the catastrophe theory and the principle of energy conservation. A conclusion was drawn that the combined cantilever beam structure for the immediate roof will form with the increase of the mining height, and the instability of the immediate roof causes the catastrophic instability of the system. The system instability was found to be related to the stiffness ratio K, material parameters, the load Q, and the first weighting interval of the main roof by analyzing the necessary and sufficient conditions for system instability. The influence degree of each parameter on the stiffness ratio K was as follows: elastic modulus E > support stiffness k1 > cross-section area a > immediate roof thickness H. The calculation equations of support resistance and subsidence of roof step were obtained. The method of judging the roof instability using catastrophe theory was proved as reasonable on the basis of the monitoring example of no. 12401 working face in Shendong mining area, China. On this basis, a reasonable value of support resistance was further calculated, and the working face was maintained safely when the support resistance exceeded 19,232 kN. [ABSTRACT FROM AUTHOR]

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The dust control basis of fully mechanized mining face is clear dust source characteristics. In view of the problem of unclear dust source and unknown distribution characteristics in the working area of 7.8 m large fully mechanized mining face, the fully mechanized mining face of Shendong Bulianta Coal Mine is selected as the The research object adopts the method of combining numerical simulation and field measurement to study the dust migration law based on the single action and joint action of the two main dust sources of shearer cutting coal and column moving frame, and obtains the dust migration law. According to the characteristics, the distribution of respirable dust concentration in the breathing zone of personnel in the vicinity of the shearer was further analyzed. The results show that the dust produced by dropping the column and moving the frame has the greatest impact on the dust concentration in the operator’s area. The maximum value of the respirable dust concentration in the operator’s area when cutting coal downwind and upwind is in the middle of the shearer, and originates from the shearer cutting coal. The proportions of dust source and column-moving dust source were 44.50% and 76.69%, respectively. The dust traceability research can provide a theoretical basis for the targeted management of dust in fully mechanized mining face with large mining height.

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Copyright of Coal Science & Technology (0253-2336) is the property of Coal Science & Technology 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.)

Copyright of Journal of Heilongjiang University of Science & Technology is the property of Journal of Heilongjiang University of Science & Technology Editorial Department 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.)