In traditional sequential single-wing mining practices, one-entry longwall mining systems make it challenging to efficiently and smoothly transfer mining equipment during a continuous mining sequence. In two-entry longwall systems, the headgate of the current panel and the tailgate of the next panel are excavated parallel to one another, effectively creating space for the transfer of mining equipment. The tailgate of the panel, however, is subjected to high-mining-induced stresses, causing severe floor heave, which seriously affects the efficiency of coal production. In this paper, field measurements and numerical simulation methods are used to reveal the mechanism of floor heave induced by the rupture and instability of a competent roof. The results show that the positional relationship between the adjacent tailgate and the longwall face is divided into three stages. Throughout the three stages, the area in which the coal pillar is not horizontally displaced moves from the center of the pillar to the goaf, and the area of peak vertical stress within the coal pillar shifts from the center of the pillar to the side nearest to the tailgate. Field studies suggest that the proposed technologies can effectively control floor heave in the tailgates of two-entry longwall mining systems.

Blasting has been widely used in coal mining as a method of reducing the integrity of rock mass. Its low controllability often leads to the unsatisfactory fragmentation effect of rock mass. The empty borehole effect has great significance in avoiding the excessive breakage of rock mass and guiding the directional propagation of blasting-induced fractures. Nevertheless, the rock mass damage type evolution induced by the empty borehole has been rarely studied. A tension-compression constitutive model of rock mass damage is established in this paper. The model is incorporated into the numerical modeling code LS-DYNA as a user-defined material model. Then, LS-DYNA is used to investigate the evolution mechanism of rock mass damage under the empty borehole effect. The damage types of rock mass caused by the empty borehole effect are also studied. The Fortran language is utilized to monitor the number variation of the tensile damaged elements and the compressive damaged elements in the rock mass around the empty borehole. The results indicate that existence of the empty borehole significantly enhances the tensile stress and the stress concentration factor in the rock mass nearby the empty borehole. Meanwhile, the rock mass nearby the empty borehole mainly damages in tension. Both the number of the tensile damaged elements and the tensile stresses in the elements increases as the empty borehole diameter increases. The number of the compressive damaged elements decreases with increasing empty borehole diameter.Energies 2020, 13, 756 2 of 21 mass. Yue et al. [4] found that the empty borehole is able to change the stress distribution in the rock mass and induce the initiation and directional propagation of the blasting-induced fractures.The existence of the empty borehole in the rock can change the stress distribution and the damage zone in the rock mass [5]. The empty borehole concept has been successfully applied to rock fracturing, tunnel excavation, and coal seam gas extraction [6,7]. The dynamic evolution of the blasting-induced stress and the directional propagation of the blasting-induced fractures have been widely studied by many researchers based on theoretical analysis and laboratory experiments. Cho et al. [8] investigated the influence of the empty borehole on fracture propagation, using polymethyl methacrylate (PMMA) and revealed the law of fracture propagation from the aspect of stress concentration. Based on the evolution characteristics of the fractures between blasting boreholes, He and Yang [9] studied the law of fracture propagation between adjacent boreholes, using a high-speed camera and digital image processing. Ma and An [10] simulated the stress evolution and fracture propagation around an empty borehole in the blasting process by numerical modeling and provided a theoretical basis for the formation of blasting-induced fractures. Chen et al. [11] used LS-DYNA (a numerical modeling code) to simulate the effective stress distribution in the rock under the superimposition effect of stress waves and refl...

Aiming at the broken failure of anchor cable in the mining roadway roof during the mining process, the lagging support scheme of anchor cable is proposed. Based on the results of indoor anchor cable pull-out test, the Cable element in FLAC3D is modified to realize the extension breaking of anchor cable in the calculation process. Furthermore, the minimum principal stress and volume strain rate mutation point are used as the failure criteria of the anchor cable. Through the comparative analysis of five anchor cable lagging support schemes of 6208 transport tunnel in Wangzhuang Mine Coal, the results demonstrate that the lagging support reduces the initial support resistance of the supporting structure. With the increase of lagging time, the ability of anchor cable to adapt to deformation increases gradually. When the lagging time reaches the gentle area of roadway deformation, its ability to adapt to deformation remains stable. Finally, it was determined that the support should start at 10–15 m of the anchor cable lagging head of the 6208 transport tunnel. Industrial tests show that the lagging support scheme ensures that the anchor cable can withstand a certain deformation, and the support body has no broken failure, which effectively controls the large mining-induced deformation of surrounding rock.

Deep horizontal high stress and high permeability geological factors appear when coal mines are converted to deep horizontal mining. When the roadway is damaged by the mining face, and the supporting components are mismatched, the deep roadways necessitate extensive repair work, which has a negative impact on the coal mining economy and sustainability. This paper carried out a series of field tests on the roadways deformation, crack distribution, and loose rock zone of the deep roadways. Furthermore, a numerical calculation model was established using the discrete element method (DEM) and calibrated with laboratory tests and RQD methods. Both the stress and crack distribution in the surrounding rock of the deep roadway were simulated. The field test and the corrected numerical model showed consistency. A FISH function was used to document the propagation of shear and tensile cracks around the roadway in three periods, and a damage parameter was adopted to evaluate the failure mechanism of the deep roadways under the dynamic stress disturbance. The matching of specifications of anchor cables, rock bolts, and anchoring agent is the primary point in the control of deep roadways, and revealing the stress evolution, crack propagation, and damage distribution caused by mining effects is another key point in deep roadway controlling. The field test and DEM in this paper provide a reference for the design of surrounding rock control of deep roadways and the sustainable development of coal mines.