• Geomechanics and Engineering
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• v.17 no.4
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• pp.333-342
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• 2019

Geological dynamic hazards during coal mining can be caused by the failure of a composite system consisting of roof rock and coal layers, subject to different loading rates due to different advancing velocities in the working face. In this paper, the uniaxial compression test simulations on the composite rock-coal layers were performed using $PFC^{2D}$ software and especially the effects of loading rate on the stress-strain behavior, strength characteristics and crack nucleation, propagation and coalescence in a composite layer were analyzed. In addition, considering the composite layer, the mechanisms for the advanced bore decompression in coal to prevent the geological dynamic hazards at a rapid advancing velocity of working face were explored. The uniaxial compressive strength and peak strain are found to increase with the increase of loading rate. After post-peak point, the stress-strain curve shows a steep stepped drop at a low loading rate, while the stress-strain curve exhibits a slowly progressive decrease at a high loading rate. The cracking mainly occurs within coal, and no apparent cracking is observed for rock. While at a high loading rate, the rock near the bedding plane is damaged by rapid crack propagation in coal. The cracking pattern is not a single shear zone, but exhibits as two simultaneously propagating shear zones in a "X" shape. Following this, the coal breaks into many pieces and the fragment size and number increase with loading rate. Whereas a low loading rate promotes the development of tensile crack, the failure pattern shows a V-shaped hybrid shear and tensile failure. The shear failure becomes dominant with an increasing loading rate. Meanwhile, with the increase of loading rate, the width of the main shear failure zone increases. Moreover, the advanced bore decompression changes the physical property and energy accumulation conditions of the composite layer, which increases the strain energy dissipation, and the occurrence possibility of geological dynamic hazards is reduced at a rapid advancing velocity of working face.

• 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.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.

• Journal of Environmental Health Sciences
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• v.17 no.1
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• pp.32-38
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• 1991

Airborne suspended particulate concentration in drilling sites of underground coal mines in Taebaek area was evaluated. And respirable coal dust exposure level was evaluated. Airborne suspended particulate mass include total suspended particle(TSP) and thoracic particle(TPM). TSP (by open-face filter holder) and TPM(by elutriator) concentration were determined by low volume air samplers. Personal air samplers were attached to the coal workers including drillers, coal cutters, and their assistants. Normality and log-normality of TSP, TPM, and respirable dust(RPM) concentration were tested by Kolmogorov-Smirnov one-sample test. Differences of means of TSP, TPM, and RPM concentration were tested by paired t-test. Relation between TSP, TPM, and RPM with pairs were tested by regression test and Pearson's correlation.

• Geomechanics and Engineering
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• v.22 no.2
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• pp.133-142
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• 2020

The evolution of the mining-induced fracture network formed during longwall top coal caving (LTCC) has a great influence on the gas drainage, roof control, top coal recovery ratio and engineering safety of aquifers. To reveal the evolution of the mining-induced stress and fracture network formed during LTCC, the fracture network in front of the working face was observed by borehole video experiments. A discrete element model was established by the universal discrete element code (UDEC) to explore the local stress distribution. The regression relationship between the fractal dimension of the fracture network and mining stress was established. The results revealed the following: (1) The mining disturbance had the most severe impact on the borehole depth range between approximately 10 m and 25 m. (2) The distribution of fractures was related to the lithology and its integrity. The coal seam was mainly microfractures, which formed a complex fracture network. The hard rock stratum was mainly included longitudinal cracks and separated fissures. (3) Through a numerical simulation, the stress distribution in front of the mining face and the development of the fracturing of the overlying rock were obtained. There was a quadratic relationship between the fractal dimension of the fractures and the mining stress. The results obtained herein will provide a reference for engineering projects under similar geological conditions.

• Transactions of the Korean Society of Mechanical Engineers B
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• v.22 no.11
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• pp.1499-1508
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• 1998

A numerical study is performed to investigate pressure drops, particle trajectories and erosion around orifice plates in pulverized coal pipe lines. Particle impaction rates change significantly with orifice shapes and Stokes numbers. At Reynolds number of $5{\times}10^5$, the pulverized coal flows well with streamlines and do not collide at the orifice plates at small sizes (${\sim}20{\mu}m$). However, the large particles (over $70{\mu}m$) impact on the front face of the orifice and erode the orifice surface. The pressure loss coefficients around the erode orifice are largely different from the designed original orifice.

• Geomechanics and Engineering
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• v.1 no.2
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• pp.143-154
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• 2009

Many factors influence occurrences of rock burst in coal mines, such as mining methods, control methods of the coal roof, lithological characteristics of the roof and floor, tectonic stress, groundwater and so on. Among those factors, lithological characteristics in the roof are the intrinsic controlling factors that affect rock burst during coal mining. Tangshan colliery is one of the coal mines that have suffered seriously from rock bursts in China. In this paper, based on the investigating the lithological characteristics of coal roofs and occurrence of rock bursts in Tangshan colliery, a numerical method is used to study the influence of roof lithological characteristics on rock burst potential. The results show that the lithological characteristics in the roof have an important impact on the distributions of stresses and elastic strain energy in coal seams and their surrounding rocks. Occurrences of rock bursts in this colliery have a close correlation with the thick-bedded, medium- to fine-grained sandstones in the roof. Such strata can easily cause severe stress concentration and accumulate enough energy to trigger rock bursts in the working face during mining operations.

• 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.22 no.6
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• pp.541-552
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• 2020

With the increasing tension of current coal resources and the increasing depth of coal mining, the gob-side entry retaining technology has become a preferred coal mining method in underground coal mines. Among them, the technology of the gob-side entry retaining with the high-water filling material can not only improve the recovery rate of coal resources, but also reduce the amount of roadway excavation. In this paper, based on the characteristics of the high-water filling material, the technological process of gob-side entry retaining with the high-water filling material is introduced. The early and late stress states of the filling body formed by the high-water filling materials are analyzed and studied. Taking the 8th floor No.3 working face of Xin'an coal mine as engineering background, the stress and displacement of surrounding rock of roadway with different filling body width are analyzed through the FLAC^3D numerical simulation software. As the filling body width increases, the supporting ability of the filling body increases and the deformation of the surrounding rock decreases. According to the theoretical calculation and numerical simulation of the filling body width, the filling body width is finally determined to be 3.5m. Through the field observation, the deformation of the surrounding rock of the roadway is within the reasonable range. It is concluded that the gob-side entry retaining with the high-water filling material can control the deformation of the surrounding rock, which provides a reference for gob-side entry retaining technology with similar geological conditions.

• Explosives and Blasting
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• v.24 no.1
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• pp.57-62
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• 2006

The typical blasting method adopted in Pasir Coal Mine is a surface blasting technique with a single free face. It means that there is only one free face, which is usually the ground surface. This kind of blasting method is easy to use but inevitably causes enormous ground vibrations, which, in turn, can affect the stability of the slopes comprising the various boundaries of the open pit mine. In addition, the method also has the problem of lowering the blast efficiency compared to other methods such as bench blasting methods or ones with more than two free faces. In this respect, a project was launched to develop a new blasting method that is suitable for controling the ground vibration and enhancing the blast efficiency. As a part of the project, authors investigated the current blasting method as well as the overall pit developing process in the mine, and established some important guidelines that should be observed during the whole development process. This paper presents the details of the typical blasting pattern and the pit developing method in the mine, and suggests the guidelines determined from the results of the observations.