The traditional anchoring method of bolts has insufficient control over the surrounding rock of the coal roadway. Based on this background, full-stress anchoring technology of bolts was proposed. Firstly, a mechanical relationship model of a bolt-drawing, anchoring interface was established to obtain the equations of the axial force and obtain shear stress distribution as well as the decreasing-load transfer law of the anchoring section of bolts. Through studying the prestress-loading experimental device of bolts, we found that increasing the initial preload could increase the axial force under the same conditions and the retarded anchoring section could control the axial-force loss of bolts in the middle of the anchoring section. Under the full-stress anchoring mode, the effect of applying a pre-tightening force was better than that of applying a pre-tightening force under traditional anchoring methods. Moreover, FLAC3D (Fast Lagrangian Analysis of Continua 3D; ITASCA (Ita sca International Inc), Minnesota, USA) numerical simulation calculation was performed. Under the full-stress anchoring mode of bolts, the increased anchoring length reduced the damage of the anchoring section, with a wider control range of the rock formation and higher strength of the compressive-stress anchoring zone. Based on the above research, four methods for applying the full-stress anchoring technology of bolts in engineering were proposed. The full-stress anchoring technology of bolts in the coal roadway has been applied in the support project of the return-air roadway at working face 3204 of the Taitou Coking Coal Mine of the Xiangning Coking Coal Group, Shanxi. The maximum moving distance of the roof and floor of the roadway was reduced from 200 to 42 mm, and the maximum moving distance on both coal sides was reduced from 330 to 86 mm. The full-stress anchoring technology of bolts was able to control the surrounding rock in the coal roadway.

Based on the engineering background of providing advance support for the working face of mining roadways, this paper studies the reasonable support technology of advance roadway roofs by combining theoretical analysis, numerical simulation, and field tests. Based on the geological conditions of the 1304 working face of Yineng Coal Mine, the FLAC3D numerical simulation software was used to compare and analyze the effects of the original single hydraulic prop advance support and the bolt-mesh-cable support without the single hydraulic prop. The results show that although the deformation of the surrounding rock is reduced under the support of the single hydraulic prop, the convergence of the roof and floor of the roadway and the left and right sides are still as high as 288 mm and 308 mm, respectively, which does not meet the requirements for safe production. Based on this problem, this study proposes full-stress anchoring technology. FLAC3D numerical simulation software is used to simulate and analyze the supporting effect of the full-stress anchoring support technology in advanced mining roadways. The results of numerical simulation experiments show that the convergence of the roof and floor and the convergence of the left and right sides of the roadway surrounding rock are 33 mm and 52 mm, respectively, which have a good control effect on the roadway surrounding rock. The field test of bolt full-stress anchoring support technology was carried out in the return air roadway of the 1304 working face. The deformation of the surrounding rock of the roadway was monitored by setting up stations. The measured results show that the maximum roof and floor convergence of the roadway is 42 mm, and the maximum convergence of the two sides of the roadway is 69 mm, which meets the requirements for safe mining on site. In this study, by comparing with the advance support effect of the original single hydraulic prop, the rationality of the full-stress anchoring technology of the mining roadway in the advance section of the working panel is determined. The use of bolt full-stress anchoring instead of the traditional single hydraulic prop for advanced support has a better surrounding rock control effect and a lower support cost. This is a new technology for advanced support of surrounding rock in mining roadways, which enriches the control technology of roadway surrounding rock and also provides technical reference for other similar engineering cases.

Studies have confirmed the poor stability of layered roofs with weak interlayers, and it is necessary to study the roof-caving mechanism of such roadways. A model of a weak interlayer was established to study the influence of the layer position and horizontal stress of layered roofs with weak interlayers on the stability of roadway roofs. FLAC^3D numerical simulation software was used to study the damage characteristics of weak interlayers and different horizontal stresses on roadway roofs. Based on Proctor’s theory, a maximum caving arch model was proposed to obtain the maximum caving arch height and span range. The pressure measurement coefficient and the vertical displacement change of the roof strata conformed to corresponding linear functions by fitting the pressure measurement coefficient and the maximum vertical displacement. A bolt-while-drilling support method (BWD) was proposed based on team measurement-while-drilling (MWD) to accurately determine the location of weak rock layers and the development range of plastic zones. A maximum caving arch model was proposed to obtain the range of influence of the maximum caving arch and the range of the maximum slope collapse angle. Furthermore, a method of anchor rod support-while-drilling was proposed and tested on-site in Jingu Coal Industry, Guxian County, Shanxi Province, China. According to the site conditions, short anchor cables were used to pass through the weak interlayer of roofs, with a good support effect. The results provide a new method for layered roof support containing weak interlayers.

Coal-accompanying gas is an essential resource, with numerous mining methods. The practice has proved that injecting high-pressure air into the coal seam can replace and flush the gas in the coal seam, effectively solving the problem of inadequate single gas drainage in soft and low permeability coal seams. This paper uses the finite element method to solve the model, simulate and study the gas drainage by high-pressure air injection in the bedding drilling, and establish a fluid-structure coupling model for gas drainage by high-pressure air injection. The competitive adsorption of N₂, O₂, and CH₄, diffusion and migration of CH₄ in coal matrix and fissure, matrix deformation caused by CH₄ adsorption, and desorption and control of coal deformation by applied stress are considered in the model. When the fixed extraction time is 600 days (d), the optimal spacing between the extraction hole and injection hole is 12.5 m. The safe extraction effect and minimum drilling amount can be ensured. It provides a basis for guiding gas drainage by injecting high-pressure air on-site.