• Geomechanics and Engineering
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• v.22 no.3
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• pp.255-264
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• 2020

With the ongoing development of deep mining of coal resources, some coal mine dynamic disasters have exhibited characteristics of both coal-gas outbursts and rockbursts. Therefore, research is required on the mechanism of rockburst-outburst coupling disaster. In this study, the failure characteristics of coal-rock combination structures were investigated using lab-scale physical simulation experiments. The energy criterion of the rockburst-outburst coupling disaster was obtained, and the mechanism of the disaster induced by the gas-solid coupling instability of the coal-rock combination structure was determined. The experimental results indicate that the damage of the coal-rock structure is significantly different from that of a coal body. The influence of the coal-rock structure should be considered in the study of rockburst-outburst coupling disaster. The deformation degree of the roof is controlled by the more significant main role of the gas pressure and the difference in the strength between the rock body and the coal body. The outburst holes and spall characteristics of the coal body after the failure of the coal-rock structure are strongly affected by the difference in strength between the roof and the coal body. The research results provide an in-depth understanding of the mechanism of rockburst-outburst coupling disasters in deep mining.

• Geomechanics and Engineering
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• v.28 no.6
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• pp.613-624
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• 2022

A study of the evolution of overburden fractures under the solid-fluid coupling state was conducted based on the geological and mining characteristics of the coal seam depth, weak strata cementation, and high-intensity mining in the mining areas of West China. These mining characteristics are key to achieving water conservation during mining or establishing groundwater reservoirs in coal mines. Based on the engineering background of the Daliuta Coal Mine, a non-hydrophilic simulation material suitable for simulating the weakly cemented rock masses in this area was developed, and a physical simulation test was carried out using a water-sand gushing test system. The study explored the spatial distribution and dynamic evolution of the fractured zone in the mining overburden under the coupling of stress and seepage. The experimental results show that the mining overburden can be vertically divided into the overall migration zone, the fracture extension zone and the collapse zone; additionally, in the horizontal direction, the mining overburden can be divided into the primary fracture zone, periodic fracture zone, and stop-fracture zone. The scope of groundwater flow in the overburden gradually expands with the mining of coal seams. When a stable water inrush channel is formed, other areas no longer generate new channels, and the unstable water inrush channels gradually close. Finally, the primary fracture area becomes the main water inrush channel for coal mines. The numerical simulation results indicate that the overlying rock breaking above the middle of the mined-out area allows the formation of the water-conducting channel. The water body will flow into the fracture extension zone with the shortest path, resulting in the occurrence of water bursting accidents in the mining face. The experimental research results provide a theoretical basis for the implementation of water conservation mining or the establishment of groundwater reservoirs in western mining areas, and this theoretical basis has considerable application and promotion value.

• Geomechanics and Engineering
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• v.36 no.5
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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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• v.14 no.5
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• pp.421-428
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• 2018

In deep mining, the lateral deformation of strip coal pillar appears to be a new characteristic. In order to study the lateral deformation of coal-mass, a monitoring method and monitoring instrument were designed to investigate the lateral deformation of strip coal pillar in Tangkou Coalmine with the mining depth of over 1000 m. Because of without influence of repeated mining, the bedding sandstone roof is easy to break and the angle between maximum horizontal stress and the roadway is small, the maximum lateral deformation is only about 287 mm lower than the other pillars in the same coalmine. In deep mining, the energy accumulation and release cause a discontinuous damage in the heterogeneous coal-mass, and the lateral deformation of coal pillar shows discontinuity, step and mutation characters. These coal-masses not only show a higher plasticity but also the high brittleness at the same time, and its burst tendency is more obvious. According to the monitoring results and theoretical calculations, the yield zone of the coal pillar width is determined as 15.6 m. The monitoring results presented through this study are of great significance to the stability analysis and design of coal pillar.

• Geomechanics and Engineering
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• v.19 no.3
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• pp.255-268
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• 2019

Taking top-coal caving mining face (TCCMF) as research object, this paper considers the combination of top-coal and immediate roof as cushion layer to build the solution model of support resistance based on the theory of elastic foundation beam. Meanwhile, the physical and mechanical properties of coal-rock combination influencing on strata behaviors is explored. The results illustrate that the subsidence of main roof in coal wall increases and the first weighting interval decreases with the increase of top-coal and immediate roof thicknesses as well as the decrease of top-coal and immediate roof elastic modulus. Moreover, the overlying strata reflecting on support has negative and positive relationship with top-coal thickness and immediate roof thickness, respectively. However, elastic modulus has limit influence on the dead weight of top-coal and immediate roof. As a result, it has similar roles on the increase of total support resistance and overlying strata reflecting on support in the limit range of roof control distance. In view of sensitive analysis causing the change of total support resistance, it can be regards as the rank of three components as immediate roof weight > overlying strata reflecting on support > top coal weight. Finally, combined with the monitoring data of support resistance in Qingdong 828, the validity of support resistance determined based on elastic foundation beam is demonstrated, and this method can be recommended to adopt for support type selecting in TCCMF.

• Geomechanics and Engineering
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• v.29 no.4
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• pp.451-461
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• 2022

The roadways surrounded by rock and coal will lose their stability or even collapse under rock burst. Rock burst mainly involves an evolution of dynamic loading which behaves quite differently from static or quasi-static loading. To compare the dynamic response of coal and rocks with different static strengths, three different rocks and bituminous coal were selected for testing at three different dynamic loadings. It's found that the dynamic compression strength of rocks and bituminous coal is much greater than the static compression strength. The dynamic compression strength and dynamic increase factor of the rocks both increase linearly with the increase of the strain rate, while those of the bituminous coal are irregular due to the characteristics of multi-fracture and heterogeneity. Moreover, the absorbed energy of the rocks and bituminous coal both increase linearly with an increase in the strain rate. And the ratio of absorbed energy to the total energy of bituminous coal is greater than that of rocks. With the increase of dynamic loading, the failure degree of the sample increases, with the increase of the static compressive strength, the damage degree also increases. The static compassion strength of the bituminous coal is lower than that of rocks, so the number of small-scale fragments was the largest after bituminous coal rupture.

• Geomechanics and Engineering
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• v.37 no.3
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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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• v.16 no.2
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• pp.177-185
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• 2018

The subsidence mechanism of ground surface is a complex phenomenon when multiple seam coal mining operations are carried out. Particularly, the coal mining beneath karstic formations causes a very special form of subsidence. The subsidence causes elasto-plastic deformation of the karstic layers and the collapse of cavities leads to dolinization and/or sinkhole formation. In this study, a sinkhole with a depth of 90 m and a width of 25 m formed in Gelik district within the coal-basin of Zonguldak (NW, Turkey) induced by multiple seam coal mining operations in the past has been presented as a case-history together with two-dimensional numerical simulations and InSAR monitoring. The computational results proved that the sinkhole was formed as a result of severe yielding in the close vicinity of the faults in contact with karstic formation due to multiple seam longwall mining at different levels.

• Geomechanics and Engineering
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• v.29 no.6
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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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• v.11 no.5
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• pp.607-624
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• 2016

Rockburst is becoming more serious in Chinese coal mine. One of the effective methods to control rockburst is blasting. In the paper, we monitored and analyzed the blasting waves at different blast center distances by the Hilbert-Huang transform (HHT) in a coal mine. Results show that with the increase of blast center distance, the main frequency and amplitude of blasting waves show the decreasing trend. The attenuation of blasting waves is slower in the near blast field (10-75 m), compared with the far blast field (75-230 m). Besides, the frequency superposition phenomenon aggravates in the far field. A majority of the blasting waves energy at different blast center distances is concentrated around the IMF components 1-3. The instantaneous energy peak shows attenuation trend with the blast center distance increase, there are two obvious energy peaks in the near blast field (10-75 m), the energy spectrum appears "fat", and the total energy is greater. By contrast, there is only an energy peak in the far blast field, the energy spectrum is "thin", and the total energy is lesser. The HHT three dimensional spectrum shows that the wave energy accumulates in the time and frequency with the increasing of blast center distance.