Coal and gas outburst is an extremely complex dynamic phenomenon of mine gas, which is mainly manifested in a very short time. A large amount of coal and rock are thrown out from the coal body to the mining space and accompanied by a large volume of high-pressure gas. In the process of coal and gas outburst, the internal energy consumption of gas is composed of two parts: one is used to throw out broken coal and rock mass, and the other is used to pulverize broken coal. In this article, from the perspective of energy dissipation, the experiment of broken coal ejection with different coal particle sizes, different adsorption gas, and pressure is studied. The characteristics of coal ejection are studied and analyzed, and the proportion of adsorbed gas participating in the outburst work is quantitatively analyzed. The results show that after outburst excitation, residual gas will continue to desorb and work on outburst until the power is insufficient to throw coal body; compared with air, CO2 gas has a stronger ability to work on the outburst, and the outburst coal is thrown far away, and the pulverization effect is stronger. Through the energy analysis in the process of outburst, the results show that when the particle size of the coal sample is consistent, the greater the outburst pressure is, the larger the desorption amount of the adsorbed gas is, and the larger the volume involved in the outburst work is. When the test gas is consistent with the outburst pressure, the gas desorption amount of the small-size coal sample is more, the desorption gas has a stronger ability to work on the outburst, and the proportion of participating in outburst work is higher. The crushing degree of coal plays an important role in the expansion and release of gas internal energy.
Coal and gas burst is one of the significant and catastrophic hazards in underground longwall operations. To date, the protective coal seam mining has been recognized as the most effective mining method for minimizing or even avoiding the effect of the coal and gas burst. In this paper, numerical modelling and field test were carried out for the longwall operation in Qidong Coal Mine in order to investigate the induced stress and coal seam gas drainage performance in the protected coal seam after the complete extraction of the protective coal seam. It was found that four stress zones can be classified in the protected coal seam being the original stress zone, stress concentration zone, stress relief zone, and recompaction zone. In addition, the monitoring data of gas concentration and volume change in the field agree well with the numerical modelling results.
In order to solve the problems of the uneven deformation of Gangue Filled Wall and the difficulty of large-scale promotion of roadway side support, and to achieve the purposes of direct disposal of coal mine waste, reducing costs, and protecting the environment, the failure mechanics model of the bagged gangue was established, and the mechanical action relationship between longitudinal external load and transverse external load of gangue woven bag was deduced. Through the uniaxial compression test of large-scale flexible backfill (coal gangue of different particle sizes), it was obtained that when the strain is 0.2, the bearing capacity of particles with particle sizes between 0 and 10 mm is greater than 5 MPa, and when the strain is 1.27, the bearing capacity of particles with particle thicknesses between 10 and 20 mm is greater than 0 mpa, which meets the requirements of resistance value and resistance growth rate of gob side entry. In the “load deflection” test of backfill (gangue) samples, it was found that the maximum failure load of wet shotcrete is greater than that of dry shotcrete, and the wet shotcrete can withstand greater deformation under the same load conditions. Through the analysis of the experimental results of “flexural strength thickness” and “maximum failure load thickness”, it was finally determined that the thickness of the spray layer with good flexibility and sufficient support force is controlled at about 80 mm.
To study the evolution law of axial force and shear stress of a full-length anchorage bolt in a rectangular roadway during roadway driving and working face mining, based on the stress analysis of the bolt, considering the elastic parameters and geometric size of the bolt, the effect of a bearing plate on surrounding rock, roadway cross-section shape, roadway deformation degree, and roadway elastic parameters, elastic mechanics and mathematical analysis methods were used to establish the mechanical model describing the interaction between the bolt and surrounding rock, and the mechanical formulas for calculating the axial force and shear stress of the bolt were derived. Taking the mining roadway of 1,131(1) working face in the Zhujidong coal mine of the Huainan mining area as the engineering background, the axial force and shear stress of the bolt in the middle of the roof and side of the rectangular roadway with the advance of driving face and working face were analyzed. The mechanical model and theoretical analysis results are verified by installing force measuring bolts with the same mechanical properties as the field and observing the real axial force distribution of the bolts.
Aiming at the problem of the deformation of the roadway floor plate during the laneway during the retention period, the mechanical model of the roadway floor is established, and the deformation characteristics of the roadway floor and the change law of the bottom drum are studied and analyzed through theoretical calculation and calculus simulation, revealing the instability mechanism of the surrounding rock of the roadway under the stress disturbance environment, and when not affected by the adoption, the roadway forms a certain stress concentration area within the effective range of support. During mining, under the comprehensive action of the original peripheral stress field and the mining stress field, the cliffhanger is unstable under the comprehensive action of the original peripheral stress field and the mining stress field, and the extrusion and stretching effect of the unflapped part of the rock layer above the goaf section of the coal seam is set up along the air, resulting in violent deformation such as the bottom drum, and the rotational sinking of this part of the unflinted rock layer further aggravates the transfer of the overburden load to the surrounding rock of the lane, so that the surrounding rock along the empty lane is subjected to a large additional stress, and the mining stress field plays a leading role, and the mining stress “far field” is the compound stress field, of which the tensile stress is the leading destructive factor. The deformation of the surrounding rock is mainly based on the bottom, and the horizontal stress on the bottom plate along the empty lane is mainly generated by the horizontal strain that occurs after the lower rock layer of the filling body and the coal gang is subjected to the supporting pressure transmitted by the top plate. With the mining of the working surface, the roof of the goaf area is broken and collapsed to form the characteristics of “vertical three belts,” which is affected by the “large support” of the coal body of the working surface and the “small support” of the surrounding rock along the empty roadway, and the pressure relief of the cut roof can make the roof plate along the empty lane change from the “long arm beam” structure when the roof is not cut into the “short arm beam” structure, blocking the lateral stress of the goaf area to the roof plate of the alley and significantly reducing the degree of stress superposition of the roof plate of the alley. The technical means of blasting cutting roof active pressure relief and protective lane are used to block the transmission of lateral support pressure, the roof slate layer is precracked in advance, the sinking of the rock layer is accelerated, the disturbance time is reduced, the vertical stress of the rock layer and the rock layer above it along the empty roadway is reduced, the vertical stress concentration of the roadway is reduced, the stress concentration coefficient is reduced, the degree of damage of the surrounding rock after the top is weakened, the damage range is reduced, and the technical problem of large deformation prevention and control along the bottom drum of the empty alley can be solved. Constructing the mechanical structure model of the top plate of the cut top pressure relief and the uncut top pressure relief along the empty lane, the stress change characteristics of the active protective rock surrounding rock along the hollow top of the cut top pressure relief were calculated, and after the technical scheme of the blasting cut top active pressure relief and protection lane was adopted, the deformation along the empty roadway was significantly weakened, the stability of the surrounding rock of the roadway after the blasting of the cut roof was significantly improved, the maintenance state along the section of the empty roadway was good, and the cross-sectional convergence rate was reduced by 37.3% compared with the original section. Cutting the roof active pressure relief and protective lane can effectively improve the stability of the surrounding rock.
In order to reduce the risk of coal and rock dynamic disasters in the coal mine production process, the coupling mechanics characteristics of coal and rock produced in the process of coal mining in the Dingji Coal Mine are taken as the research object, and the experimental study on the deformation characteristics and the variation rule of mechanical parameters of raw coal under multifield coupling (temperature, gas, and stress coupling) was carried out. The results show that the elastic modulus, peak strain, and peak stress of raw coal samples under the thermal-hydraulic-mechanical coupling have the same change law in the test temperature range and all of them show a linear decreasing law as the temperature increases. Under the same temperature gradient increasing condition, the elastic modulus, peak strain, and peak stress show a nongradient decreasing trend as the temperature increases. Both the deformation modulus and the lateral expansion coefficient show a linear increase as the temperature increases, while the deformation modulus and the lateral expansion coefficient show a nongradient increase trend as the temperature increases under the same temperature gradient increasing condition. Under the action of the thermal-hydraulic-mechanical coupling, unloading confining pressure obviously accelerated the yield process of the coal sample, and the confining capacity of confining pressure on transverse strain of the coal sample decreased. To prevent the occurrence of coal and gas outburst, it is necessary to take specific prevention measures according to the change law of triaxial compression mechanics of a raw coal specimen under the action of the thermal-hydraulic-mechanical coupling.
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