In this study, the transfer and dissipation of strain energy in the surrounding rock of a deep roadway were analyzed, considering the objective strain softening and dilatancy behaviors. The strain energy increment was decomposed, and its variation was analyzed based on the incremental plastic flow theory; then, a numerical simulation was conducted for verification and further analysis. The results were verified by the field monitoring data of a coal mine gateway. The results show that in the elastic stage, the volumetric elastic strain energy density Uev decreases, while the shear elastic strain energy density Ues increases. In the plastic stage, both Uev and Ues decrease. The volumetric plastic strain energy density Upv is negative, and its absolute value increases, leading to strain energy accumulation. In contrast, the shear plastic strain energy density Ups is positive and increases, leading to strain energy dissipation. Considering strain softening, the elastic strain energy decreases, the plastic strain energy increases, and the region of strain energy dissipation expands. Considering dilatancy, the plastic strain energy varies more significantly, and the effect of strain softening is amplified. The strain energy is transferred from the deep part to the shallow part of the elastic zone and then to the plastic zone. The preexisting and input strain energies in the plastic zone are transformed into considerable amounts of plastic strain energy and then dissipated. Thereafter, a significant plastic strain appears, leading to the large deformation of the surrounding rock.
Numerical simulations have often been used in close-distance coal seam studies. However, numerical simulations can contain certain subjective and objective limitations, such as high randomness and excessively simplified models. In this study, close-distance coal seams were mechanically modeled based on the half-plane theory. An analytical solution of the floor stress distribution was derived and visualized using Mathematica software. The principal stress difference was regarded as a stability criterion for the rock surrounding the roadway. Then, the evolution laws of the floor principal stress difference under different factors that influence stability were further examined. Finally, stability control measures for the rock surrounding the roadway in the lower coal seam were proposed. The results indicated the following: (1) The principal stress difference of the floor considers the centerline of the upper coal pillar as a symmetry axis and transmits radially downward. The principal stress difference in the rock surrounding the roadway gradually decreases as the distance from the upper coal pillar increases and can be ranked in the following order: left rib > roof > right rib. (2) The minimum principal stress difference zones are located at the center of the left and right “spirals,” which are obliquely below the edge of the upper coal pillar. This is an ideal position for the lower coal seam roadway. (3) The shallowness of the roadway, a small stress concentration coefficient, high level of coal cohesion, large coal internal friction angle, and appropriate lengthening of the working face of the upper coal seam are conducive to the stability of the lower coal seam roadway. (4) Through bolt (cable) support, borehole pressure relief, and pregrouting measures, the roof-to-floor and rib-to-rib convergence of the 13313 return airway is significantly reduced, and the stability of the rock surrounding the roadway is substantially improved. This research provides a theoretical basis and field experience for stabilizing the lower coal seam roadways in close-distance coal seams.
Because both faults and cleats exist in coal, sharp stress drops occur during loading when coal is deformed. These drops occur during the pre-peak stage and are accompanied by sudden energy releases. After a stress drop, the stress climbs slowly following a zigzag path and the energy accumulated during the pre-peak stage is unstable. A stress–strain curve is the basic tool used to evaluate the bursting liability of coal. Based on energy accumulation in an unsteady state, the pre-peak stress–strain curve is divided into three stages: pre-extreme, stress drop, and re-rising stage. The energy evolution of the specimen during each stage is analyzed. In this paper, an index called the effective elastic strain energy release rate (EESERR) index is proposed and used to evaluate the coal’s bursting liability. The paper shows that the propagation and coalescence of cracks is accompanied by energy release. The stress climb following a zigzag path prolongs the plastic deformation stage. This causes a significant difference between the work done by a hydraulic press during a laboratory uniaxial compression experiment and the elastic strain energy stored in the specimen during the experiment, so the evaluation result of the burst energy index would be too high. The determination of bursting liability is a comprehensive evaluation of the elastic strain energy accumulated in coal that is released when the specimen is damaged. The index proposed in this paper fully integrates the energy evolution of coal samples being damaged by loading, the amount of elastic strain energy released during the sample failure divided by the failure time is the energy release rate. The calculation method is simplified so that the uniaxial compressive strength and elastic modulus are included which makes the new index more universal and comprehensive. Theoretical analysis and physical compression experiments validate the reliability of the evaluation.
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