• Geomechanics and Engineering
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• 제25권3호
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• pp.233-243
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• 2021

Rib spalling is a major issue affecting the safety of steeply inclined coal seam. And the failure coal face and support system can be affected with each other to generate a vicious cycle along with inducing large-scale collapse of surrounding rock in steeply inclined coal seam. In order to analyze failure mechanism and propose the corresponding prominent control measures of steeply inclined coal working face, mechanical model based on coal face-support-roof system and mechanical model of coal face failure was established to reveal the disaster mechanism of rib spalling and the sensitive analysis of related factors was performed. Furthermore, taking 3402 working face of Chen-man-zhuang coal mine as engineering background, numerical model by using FLAC3D was built to illustrate the propagation of displacement and stress fields in steeply inclined coal seam and verify the theory analysis as mentioned in this study. The results show that the coal face slide body in steeply inclined working face can be observed as the failure height of upper layer smaller than that of lower layer exhibiting with an irregular quadrilateral pyramid shape. Moreover, the cracks were originated from the upper layer of sliding body and gradually developed to the lower layer causing the final rib spalling. The influence factors on the stability of coal face can be ranked as overlying strata pressure (P) > mechanical parameters of coal body (e.g., cohesion (c), internal fraction angle (φ)) > support strength (F) > the support force of protecting piece (F') > the false angle of working face (Θ). Moreover, the corresponding control measures to maintain the stability of the coal face in the steeply inclined working face were proposed.

• Geomechanics and Engineering
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• 제36권5호
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• pp.427-440
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• 2024

This study concentrates on the 301 comprehensive caving working face, notable for its considerable mining height. The roof model is established by integrating prior geological data and the latest borehole rock stratum's physical and mechanical parameters. This comprehensive approach enables the determination of lithology, thickness, and mechanical properties of the roof within 50 m of the primary mining coal seam. Utilizing the transfer rock beam theory and incorporating mining pressure monitoring data, the study delves into the geometric parameters of the direct roof, basic roof movement, and roof pressure during the initial mining process of the 301 comprehensive caving working face. The direct roof of the mining working face is stratified into upper and lower sections. The lower direct roof consists of 6.0 m thick coarse sandstone, while the upper direct roof comprises 9.2 m coarse sandstone, 2.6 m sandy mudstone, and 2.8 m medium sandstone. The basic roof stratum, totaling 22.1 m in thickness, includes layers such as silty sand, medium sandstone, sandy mudstone, and coal. The first pressure step of the basic roof is 61.6 m, with theoretical research indicating a maximum roof pressure of 1.62 MPa during periodic pressure. Extensive simulations and analyses of roof subsidence and advanced abutment pressure under varying working face lengths. Optimal roof control effect is observed when the mining face length falls within the range of 140 m-155 m. This study holds significance as it optimizes the working face length in thick coal seams, enhancing safety and efficiency in coal mining operations.

• Geomechanics and Engineering
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• 제28권3호
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• pp.299-311
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• 2022

Water-resisting key stratum (WKS) between coal seams is an important barrier that prevents water inrush from goaf in roof under multi-seam mining. The occurrence of water inrush can be evaluated effectively by analyzing the fracture of WKS in multi-seam mining. A "long beam" water inrush mechanical model was established using the multi-seam mining of No. 2+3 and No. 8 coal seams in Xiqu Mine as the research basis. The model comprehensively considers the pressure from goaf, the gravity of overburden rock, the gravity of accumulated water, and the constraint conditions. The stress distribution expression of the WKS was obtained under different mining distances in No. 8 coal seam. The criterion of breakage at any point of the WKS was obtained by introducing linear Mohr strength theory. By using the mechanical model, the fracture of the WKS in Xiqu Mine was examined and its breaking position was calculated. And the risk of water inrush was also evaluated. Moreover, breaking process of the WKS was reproduced with Flac3D numerical software, and was analyzed with on-site microseismic monitoring data. The results showed that when the coal face of No. 8 coal seam in Xiqu Mine advances to about 80 m ~ 100 m, the WKS is stretched and broken at the position of 60 m ~ 70 m away from the open-off cut, increasing the risk of water inrush from goaf in roof. This finding matched the result of microseismic analysis, confirming the reliability of the water inrush mechanical model. This study therefore provides a theoretical basis for the prevention of water inrush from goaf in roof in Xiqu Mine. It also provides a method for evaluating and monitoring water inrush from goaf in roof.

• Geomechanics and Engineering
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• 제14권4호
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• pp.367-376
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• 2018

Samples composed of coal and rock show different mechanical properties of the pure coal or rock mass. For the same coal seam with different surrounding rocks, the frequency and intensity of rock burst can be significantly different in. First, a method of measuring the strain variation of coal in the coal-rock combined sample was proposed. Second, laboratory tests have been conducted to investigate the influences of rock lithologies, combined forms and coal-rock height ratios on the deformation and failure characteristics of the coal section using this method. Third, a new bursting liability index named combined coal-rock impact energy speed index (CRIES) was proposed. This index considers not only the time effect of energy, but also the influence of surrounding rocks. At last, a new approach considering the influences of roof and/or floor was proposed to evaluate the impact capability of coal seam. Results show that the strength and elastic modulus of coal section increase significantly with the coal-rock height ratio decreasing. In addition, the values of bursting liability indexes of the same coal seam vary greatly when using the new approach. This study not only provides a new approach to measuring the strain of the coal section in coal-rock combined sample, but also improves the evaluation system for evaluating the impact capability of coal.

• Geomechanics and Engineering
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• 제29권6호
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• pp.697-706
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• 2022

In order to explore the stable anchoring conditions of coal side under the mining disturbance of soft section coal pillar in Wangcun Coal Mine of Chenghe Mining Area, the distribution model of the anchoring support pressure at the coal pillar side was established, using the strain-softening characteristics of the coal to study the distribution law of anchoring coal side support pressure. The analytical solution for the reinforcement anchorage stress in the coal pillar side was derived with the inelastic state mechanical model. The results show that the deformation angle of the roadway side and roof increases with the roof subsidence due to the mining influence at the adjacent working face, the plastic deformation zone extends to the depth of the coal side, and the increase of anchorage stress can effectively control the roof subsidence and further deterioration of plastic zone. The roadway height and the peak support pressure have a certain influence on the anchorage stress, the required anchorage stress of the coal side rises with the roadway height and the peak support pressure. The required anchorage stress of the coal pillar side decreases as the cohesion between the coal seam and the roof and floor and the anchor length increases. Then, applied the research result to Wangcun coal mine in Chenghe mining area, the design of anchor cable reinforcement support was proposed for the section of coal pillars side that has been anchored and deformed, which achieved great results and effectively controlled the convergence and deformation of the side, providing a safety guarantee for the roadway excavation and mining.

• Geomechanics and Engineering
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• 제22권1호
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• pp.1-9
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• 2020

This paper presents an investigation of the ground response of a gob-side gateroad suffering mining stress induced by a 21 m-thick coal seam extraction. A field observation, including entry convergence and stress changes monitoring, was first conducted in the tailgate 8209. The observation results of entry convergence showed that, during the adjacent panel 8210 retreating period, the deformation of the gob-side gateroad experienced a continuous increase stage, subsequently, an accelerating increase stage, and finally, a slow increase stage. However, strong ground response, including roof bending deflection, rib extrusion and floor heave, occurred during the current panel 8209 retreating period, and the maximum floor heave reached 1530 mm. The stress changes within coal mass of the two ribs demonstrated that the gateroad was always located in the stress concentrated area, which responsible for the strong response of the tailgate 8209. Subsequently, a hydraulic fracture technique was proposed to pre-fracture the two hard roofs above the tailgate 8209, thus decreasing the induced disturbance on the tailgate. The validity of the above roof treatment was verified via field application. The finding of this study could be a reference for understanding the stability control of the gob-side gateroad in extra thick coal seams mining.

• Geomechanics and Engineering
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• 제15권6호
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• pp.1173-1182
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• 2018

The coal wall, gob-side backfill, and gangues in goaf, constitute the support system for Gob-side entry retaining (GER) in coal mines. Reasonably allocating and utilizing their bearing capacities are key scientific and technical issues for the safety and economic benefits of the GER technology. At first, a mechanical model of GER was established and a governing equation for coordinated bearing of the coal-backfill-gangue support system was derived to reveal the coordinated bearing mechanism. Then, considering the bearing characteristics of the coal wall, gob-side backfill and gangues in goaf, their quantitative design methods were proposed, respectively. Next, taking the No. 2201 haulage roadway serving the No. 7 coal seam in Jiangjiawan Mine, China, as an example, the design calculations showed that the strains of both the coal wall and gob-side backfill were larger than their allowable strains and the rotational angle of the lateral main roof was larger than its allowable rotational angle. Finally, flexible-rigid composite supporting technology and roof cutting technology were designed and used. In situ investigations showed that the deformation and failure of surrounding rocks were well controlled and both the coal wall and gob-side backfill remained stable. Taking the coal wall, gob-side backfill and gangues in goaf as a whole system, this research takes full consideration of their bearing properties and provides a quantitative basis for design of the support system.

• Geomechanics and Engineering
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• 제37권3호
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• pp.293-306
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• 2024

Influenced by the alternating effects of dynamic and static pressure during the mining process of close range coal seams, the surrounding rock support of cross mining roadway is difficult and the deformation mechanism is complex, which has become an important problem affecting the safe and efficient production of coal mines. The paper takes the inclined longwall mining of the 10304 working face of Zhongheng coal mine as the engineering background, analyzes the key strata fracture mechanism of the large inclined right-angle trapezoidal mining field, explores the stress distribution characteristics and transmission law of the surrounding rock of the roadway affected by the mining of the inclined coal seam, and proposes a segmented and hierarchical support method for the cross mining roadway affected by the mining of the close range coal seam group. The research results indicate that based on the derived expressions for shear and tensile fracture of key strata, the ultimate pushing distance and ultimate suspended area of a right angle trapezoidal mining area can be calculated and obtained. Within the cross mining section, along the horizontal direction of the coal wall of the working face, the peak shear stress is located near the middle of the boundary. The cracks on the floor of the cross mining roadway gradually develop in an elliptical funnel shape from the shallow to the deep. The dual coupling support system composed of active anchor rod support and passive U-shaped steel shed support proposed in this article achieves effective control of the stability of cross mining roadways, which achieves effective control of floor by coupling active support and preventive passive support to improve the strength of the surrounding rock itself. The research results are of great significance for guiding the layout, support control, and safe mining of cross mining roadways, and to some extent, can further enrich and improve the relevant theories of roof movement and control.

• Geomechanics and Engineering
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• 제12권4호
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• pp.657-673
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• 2017

Preventing water seepage and inrush into mines where close multiple-seam longwall mining is practiced is a challenging issue in the coal-rich Ordos region, China. To better protect surface (or ground) water and safely extract coal from seams beneath an aquifer, it is necessary to determine the height of the mining-induced fractured zone in the overburden strata. In situ investigations were carried out in panels 20107 (seam No. $2-2^{upper}$) and 20307 (seam No. $2-2^{middle}$) in the Gaojialiang colliery, Shendong Coalfield, China. Longwall mining-induced strata movement and overburden failure were monitored in boreholes using digital panoramic imaging and a deep hole multi-position extensometer. Our results indicate that after mining of the 20107 working face, the overburden of the failure zone can be divided into seven rock groups. The first group lies above the immediate roof (12.9 m above the top of the coal seam), and falls into the gob after the mining. The strata of the second group to the fifth group form the fractured zone (12.9-102.04 m above the coal seam) and the continuous deformation zone extends from the fifth group to the ground surface. After mining Panel 20307, a gap forms between the fifth rock group and the continuous deformation zone, widening rapidly. Then, the lower portion of the continuous deformation zone cracks and collapses into the fractured zone, extending the height of the failure zone to 87.1 m. Based on field data, a statistical formula for predicting the maximum height of overburden failure induced by close multiple seam mining is presented.

• Geomechanics and Engineering
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• 제17권4호
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• pp.375-383
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• 2019

In underground retreating longwall coal mining, hard roof collapse is one of the most challenging safety problems for mined-out areas. Identifying precursors for hard roof collapse is of great importance for the development of warning systems related to collapse geohazards and ground control. In this case study, the Xinhe mine was chosen because it is a standard mine and the minable coal seam usually lies beneath hard strata. Real-time monitoring of hard roof collapse was performed in longwall face 5301 of the Xinhe mine using support resistance and microseismic (MS) monitoring; five hard roof collapse cases were identified. To reveal the characteristics of MS activity during hard roof collapse development and to identify its precursors, the change in MS parameters, such as MS event rate, energy release, bursting strain energy, b value and the relationships with hard roof collapse, were studied. This research indicates that some MS parameters showed irregularity before hard roof collapse. For the Xinhe coalmine, a substantial decrease in b value and a rapid increase in MS event rate were reliable hard roof collapse precursors. It is suggested that the b value has the highest predictive sensitivity, and the MS event rate has the second highest.