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Abstract In this paper, based on smoothed‐particle hydrodynamics‐finite element method, numerical models of plunger squeezing water at a sinusoidal velocity were established to simulate self‐excited pulse water jet (SEPWJ). RHT constitutive model was adopted to describe the damage and failure of coal rock impacted by water jet. The morphological evolutions of broken pits and timeliness of rock‐breaking efficiency of SEPWJ and continuous water jet (CWJ) under the conditions with and without stress loadings were obtained and compared. The evolution laws of damage and stress inner coal rock induced by jet impact, and the failure mechanism were revealed. And the influences of different stress loading magnitudes on the fracture characteristics of coal rock were investigated. The results show that the morphologies of broken pits formed by self‐excited pulse jet undergo changes in a semi‐circular, U‐shaped, V‐shaped, and bullet shaped in sequence under the stress‐free loading condition. When applying one‐dimensional (1D) and 2D stress loadings, the shallow but wide broken pits with laminar main cracks along the stress loading direction and the inverted trapezoidal bowl broken pits are formed, respectively. With the increase of 1D stress, the depth and width of broken pits slightly decrease as a quadratic parabolic function and linearly increase, respectively. And the broken pit width and area both show an exponential slow decreasing trends with the increasing 2D stress. SEPWJ can induce higher stresses to cause the earlier occurrence of initial damage and the shorter duration of damage accumulation to coal rock than CWJ, which leads to a better rock‐breaking effect. The surface and deeper coal rock elements are broken mainly due to compressive shear stresses. The 2D stress loading delays the initial damage occurrence and prolongs the damage accumulation duration due to inhibitory effect of stress loading on jet impact.

Abstract Pore difference characteristics and adsorption/desorption experiments of primary coal with different degrees were explored to study the characteristics of pore structure and gas desorption. The results show that the pore volume increases with the increase in the degree of metamorphism. In primary coal with different degrees of metamorphism, the distribution of micropores, small pores, and medium‐sized pores (by volume) is WY PM > CY. According to the pore volume distribution law of different metamorphic primary coals, the average pore diameters of WY, PM, and CY were 68.40, 45.60, and 30.50 nm, respectively. The porosity and specific surface area of primary coal with different degrees of metamorphism show a distribution pattern such that WY > PM > CY. For large pore and visible pore, the pore volumes of WY, PM, and CY coal were 0.0785, 0.0587, and 0.0300 mL/g, respectively, and the proportions were 81.86%, 74.49%, and 55.25%, respectively. The porosity of WY, PM, and CY were 11.57%, 9.03%, and 6.57%, respectively, and the specific surface areas were 6.917, 5.826, and 5.611 cm3/g. According to the desorption characteristics of coal bodies with different metamorphic degrees at different time nodes found that the gas desorption at different time nodes shows similar changes under the same adsorption equilibrium pressure. The higher the degree of coal metamorphism, the greater the amount of gas desorption, and the faster the desorption amplitude. The desorption intensity of WY coal is significantly higher than that of PM and CY coal with the increase of desorption time which also verifies the measurement results of pore structure and adsorption constant.

Stress is one of the main factors influencing coal and gas outbursts. The apparent effects of the crustal stress, the structural stress, and the mining-induced stress increase as the depth of mining increases. At present, there have been few studies of the relationship between the comprehensive analyses of the crustal stress, mining-induced stress, and coal gas. The in situ measurement of the relationship between stress-related behaviors and coal gas under the influence of mining was conducted through experimental analysis of surrounding rock support and coal and gas outburst control and optimization of surrounding rock support materials and system construction. The results showed that the mining-induced stress first increased to a peak value, then gradually decreased, and tended to stabilize as the footage progresses. Stress appears at 96 m ahead due to mining; after 57 m of advancing, there is a large increase until it passes through this area. The stress in front of the working face increases linearly, and the increase range is obviously larger than that of the coal body in a certain range on both sides. The support anchoring force gradually decreased and tended to be stable after rapidly increasing to a maximum value. The deep displacement of the roof increased linearly and tended to be stable after reaching an accumulated displacement which can reach 16-28 mm. The residual gas pressure in front of mining operations decreased rapidly, and beyond 15 m on each side of the roadway, it decreased significantly. The residual gas pressure and gas content were consistent with the gas desorption index of drill cuttings due to the influences of gas predrainage and mining. The stress along the direction of the roadway and the residual gas content, the residual gas pressure, and the gas desorption index of drill cuttings conform to the logarithmic functional relationship. The research results provide a basis for the comprehensive prevention and control of coal and gas outbursts from multiple angles considering stress, coal, and gas.

Coal elasticity and gas expansion are important factors for coal and gas outburst. During the outburst process, the elastic strain energy of coal is mainly released from the stress region, and the gas expansion energy near the working face is larger, and it is not a continuous release process. To reveal the mechanical characteristics and energy evolution of outburst coal seam, uniaxial and triaxial compression tests were carried out on outburst coal seam samples under different loading methods. The experimental results show that the elastic characteristics become more obvious with the increase of loading rate, the peak strain increases, the elastic modulus is linearly related with the loading rate，and the overall degree of fragmentation increases with the increase of loading rate, which is consistent with the severity of macroscopic coal failure. The failure mode of coal under uniaxial compression conditions is often manifest as brittle failure. The strength characteristics of coal under different loading rates comply with the Mohr‐Coulomb criterion, and the peak strength is linearly related to the failure time and loading rate. With the increasing confining pressure causes the failure of coal samples to transition from ductile to brittle, and the failure mode develops from local shear to overall splitting. The elastic energy evolution curve is consistent with its stress‐strain curve. With the increase of confining pressure, the limiting elastic energy and peak total energy increase in a quasi‐linear manner. The accumulated limit elastic energy plays an important role in the failure of coal samples, and the macroscopic manifestation thereof is that the coal samples fail more severely under high confining pressure conditions than under low confining pressure conditions. The research results are of great significance for the comprehensive prevention and control of coal and gas outburst. [ABSTRACT FROM AUTHOR]