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To solve these problems of poor supporting effect and serious deformation and failure of surrounding rock of mining roadway under deep mining stress, a FLAC-3D numerical calculation model is established with −800 m level no. 2424 upper roadway in the Suncun Coal Mine as the background to compare the stress, deformation, and failure law of surrounding rock of mining roadway under once support and multi-level support with the same support strength. It is found that the multi-level support technology has obvious advantages in the surrounding rock of the horizontal roadway on the 2424 working face. From this, the key parameters of multi-level support are determined, and the field industrial test is carried out. The results show that the overall deformation of the surrounding rock is obviously reduced after multi-level support. The displacement of the two sides is reduced by about 40%, the displacement of the roof and floor is reduced by about 30%, and the plastic zone of the roadway is reduced by about 75%. The peak value of concentrated stress decreases from 98.7 MPa to 95.8 MPa, which decreases slightly. The integrity and stability of the surrounding rock are excellent, and the support effect is satisfactory. The research can provide reference and technical support for surrounding rock control of deep high-stress mining roadways. [ABSTRACT FROM AUTHOR]

The stability control of mining roadways is crucial in ensuring the safe and efficient mining of deep coal resources. Given the effect of mining stress on the working face, the support scheme of such roadways, which is designed based on the original in-situ stress parameters, often presents support-related problems. A mining roadway on the 1131 working face of Zhujidong Coal Mine in the Huainan mining area, China was taken as the engineering background of this study. To explore the influence of mining stress on the stability control of this roadway, the 2-dimensional (2D) Particle Flow Code (PFC) numerical simulation software was used to perform a simulation study on the stress distribution characteristics and deformation failure laws of the surrounding rocks under the change in the roadway lateral pressure coefficient caused by the mining stress. An improved support scheme that considers the influence of varying lateral pressure coefficient on the mining roadway was then proposed. Results show that when the lateral pressure coefficient increased from 1 to 1.4, the maximum principal stress (61.2 MPa) is observed at 2.9 m inside the roof of the surrounding rocks in the roadway. When the lateral pressure coefficient decreases from 1 to 0.4, the maximum principal stress (46.2 MPa) is observed at 2.2 m inside the surrounding rock of sidewalls of the roadway, and failure occurs. These findings suggest that the deformation and failure of surrounding rocks are affected regardless of the lateral pressure coefficient increase or decrease. On this basis, the lengths of the anchor bolts in the roof and sidewalls in the original support method are increased from 2,200 mm to 3,000 mm and 2,500 mm, respectively. The field monitoring results indicate that the improved support method mitigates the deformation and realizes the stability control of the roadway surrounding rocks. The findings of this study could provide a scientific basis for the parameter design of roadway support. [ABSTRACT FROM AUTHOR]

This study considered the mining roadway with coal pillar protection in the fully mechanized caving face of the Dananhu No.1 Coal Mine, China. Theoretical analysis, numerical simulation, and field tests were conducted, and the stress environment, deformation, and failure characteristics of the mining roadway in the fully mechanized caving face were analyzed. The results revealed that the intrinsic cause for the large asymmetrical floor deformation in the mining roadway is the asymmetrical phenomenon of the surrounding rock's stress environment, caused by mining. This also results in the non-uniform distribution of the mining roadway floor's plastic zone. The degree of asymmetrical floor heave is internally related to the thickness of the caving coal. When the thickness of the caving coal was in the range of 5.9 m, the deformation of the asymmetrical floor heave, caused by the plastic failure in the floor, became more obvious as certain parameters increased. As the rotation angle of the principal stress direction increased, the maximum plastic failure depth position of the floor gradually moved toward the middle of the roadway. This caused a different distribution for the maximum deformation position. The control of the floor heave deformation was poor, and it was not feasible to use high-strength support under the existing engineering conditions. Hence, the control should mainly be applied to the floor heave deformation. When the thickness of the caving coal was more than 5.9 m, the main roof strata was prone to instability and being cut along the edge of the coal pillar; the rock stress environment surrounding the roadway tended to revert back to the initial geostress state. The proposed floor heave control strategy achieved good results, and as the deformation of the floor heave decreased, the workload of the floor heave was also greatly reduced. [ABSTRACT FROM AUTHOR]