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Energy is an internal variable during rock deformation and failure, and its dissipation and conversion law can reflect the rock’s internal damage and deterioration state. Analysis of rock deformation and failure process from the perspective of energy is helpful to deeply understand the mechanism of rock damage, fracture and instability failure, and has important theoretical and practical significance for the stability evaluation and support control of surrounding rock. In this study, through single cyclic loading and unloading (SCLU) experiments, cyclic triaxial loading and unloading (CTLU) experiments and conventional triaxial compression (CTC) experiments, the equivalent elastic modulus method based on elastic strain energy is proposed to analyze the energy conversion of rock. The results show that the error of the elastic strain energy calculated by the strain energy formula method is generally higher than 10% with the secant and tangent modulus of the loading and unloading curve as input parameters. Taking the equivalent elastic modulus proposed in this study as an input parameter, more accurate elastic strain energy can be obtained by the strain energy formula. During the rock failure process, the equivalent elastic modulus shows a three-stage characteristic of increase, steady and decrease. The equivalent elastic modulus can be estimated by the quadratic function between the equivalent elastic modulus and confining pressure and axial strain. Under the same deformation and deviatoric stress, the elastic strain energy stored in rock increases with increasing confining pressure. The local maximum energy dissipation rate corresponds to stress drop, and the peak energy dissipation rate appears near the peak strength. High energy dissipation mainly occurs in a short time after peak strength, and energy release and dissipation are more sudden and severe under high confining pressure.

The surrounding rock of a coal mine roof fractures with the excavation of the underground working face in coal mining. These mining-induced fractures are connected and extended upward to form water inrush channels. A water inrush accident may occur when there is a sufficiently large water source. To elucidate the formation mechanisms of the roof water inrush channel and the characteristics of water inrush in goafs, we performed a case study of No.18401 Panel of Xiqu Coal Mine in China and determined whether the roof water inrush channel is connected by theoretical calculation and microseismic monitoring. The modified mechanical parameters of rock masses were embedded into the numerical model based on microseismic data. Microseismic monitoring and numerical simulation were organically combined to analyze the connection process of the water inrush channel, after which the roof water inrush channel is calibrated on the No.18401 Panel. We established a non-Darcy flow model for water inrush in the water-conducting fractured zone of mines by coupling the Darcy, Forchheimer, and Navier–Stokes equations. Finite element language and its compiler (FELAC) was used to study the water inrush mechanism of non-Darcy seepage. The results show that the pressure, velocity, and porosity in the water-conducting fractured zone are non-uniform in water inrush occurrence and development, and the mixed fluid mainly passes through the “dominant channel.” The development of water inrush is accompanied by the release of the hydrostatic pressure in the aquifer, the sudden increase of the velocity at the water inrush position, and the increase in fluid concentration. Hence, the underground water inrush can be predicted and prevented by monitoring the aforementioned indicators. This research is of great significance for the calibration of the water inrush channel of the roof and the prediction of water inrush disasters.

After coal mining enters the deep, the mining environment changes dramatically, and engineering disasters become increasingly prominent, which are mostly related to rock instability and failure. As traditional support is difficult to meet production needs, it is necessary to improve the support system. Based on the engineering background of the Pinggang mining roadway, this work studies the migration law of overlying strata in deep goaf by theoretical analysis and numerical simulation. The results show that the vertical stress and plastic failure range of the surrounding rock in front of the working face increase with the advance distance and when the working face advances to the first square, reaching the maximum. A stope spatial model considering the influence of horizontal stress is established. Combined with the theory of key strata, the stress transfer characteristics of overlying strata are analyzed. It can be seen that 0~30 m in front of the coal wall of the working face is the influence range of advanced abutment pressure, and the dynamic mining pressure in this range has a great influence. The inclined direction of the working face, 0~20 m away from the coal wall of the roadway, is the influence range of the solid coal abutment pressure. On this basis, the “migration- transfer- control” technical system of surrounding rock in deep stope face is put forward, i.e., the stress transfer of surrounding rock is caused by overlying rock migration, and the large deformation of surrounding rock is controlled by supporting means. Based on the original support scheme of the roadway, three reinforcement schemes are designed for the roof, the sidewalls, and both the roof and sides. The deformation control effect of the reinforcement scheme is far greater than that of the single factor, and the field monitoring effect is good. The research results aim to provide theoretical and technical support for the deformation control of mining roadways in the deep mining process.