Yongle Liu

To elucidate the impact of mine water on weakening mechanisms of coal under dynamic compressive loading conditions, this investigation employed experimental analyses utilizing Split Hopkinson Pressure Bar (SHPB) apparatus. Parameters including moisture content, impact gas pressure, and confining pressure were systematically varied to clarify the weakening properties. The results indicate a significant decrease in both the strength and stiffness of submerged coal under dynamic loading conditions. Furthermore, moisture infiltration accelerated strength degradation. A damage model is derived based on elastic damage mechanics. The model demonstrated strong concordance with the experimental data through parameter testing. This research provides quantitative evidence for the comprehension of the weakening mechanism of submerged coal.

The mechanical properties of the surrounding rock at the island working face have significantly altered due to extended exposure to high-salinity mine water, raising concerns over the safety of mining operations. This research seeks to conduct a thorough analysis of the mechanical properties of water-saturated roof rock in deep underground mining circumstances, with particular emphasis on sandstone samples typically found in these settings. A series of dynamic loading tests are undertaken on sandstone specimens with differing water contents to examine the effect of high-salinity water on rock strength. These tests replicate the circumstances encountered by the rock during mining operations and facilitate the assessment of the rock material's degradation behavior while in contact with saltwater. Through systematic analysis of the test findings, accounting for fluctuations in water saturation levels and the corresponding stress-strain responses, we may extract essential insights about the mechanical properties of the roof rock. The research underscores the significance of comprehending these mechanical properties to anticipate problems in mining. Subsequent to the testing phase, a comprehensive evaluation of the mechanical properties of the roof rock is conducted, resulting in the development of a damage constitutive model specifically designed for the engineering rock mass. This model integrates the observed weakening events, offering a more precise depiction of rock behavior under saltwater conditions, thereby improving our capacity to plan safer and more efficient underground mining operations in analogous high-salinity situations.

This study aims to comprehensively analyze the stability of rock mass around the gob side driving roadway in a significant inclined mining face. By considering the geological parameters and engineering conditions of specific cases, we carefully explored many factors affecting the stability of the surrounding rock mass, such as the mechanical properties of rocks, formation inclination, and the existence of goaf. Based on this, we construct a numerical model for analyzing the stability of the rock mass around the gob side driving roadway. Subsequently, we made a detailed investigation of the stress distribution and the characteristics of the plastic zone. In addition, through the analysis of the destruction mode and stress distribution characteristics of the rock mass, we establish the appropriate width of the coal pillar and thus provide a scientific basis for the formulation of support countermeasures; finally, the support countermeasures proposed in this study have achieved remarkable results in practical application and verify the feasibility and practicability of the research method. This support scheme ensures the stability of the gob side entry over its service life. It is hoped that this scheme can be promoted in similar projects to help maintain mine safety.

In coal mine paste filling technology, geomaterials like coal gangue and fly ash are used as the main component, and cement is applied as the cementing material. In the mining production, mining-and-filling is a cyclic work, where the filling immediately after mining and mining immediately after filling. Long solidification time after filling will affect mining; consequently, the paste should have early strength. In addition, the prepared paste will be conveyed to goaf through the pipeline. The paste flow characteristics will change to some extent in the conveying process, and there is uncertainty about whether the paste can meet the requirements of pumpability and strength. Therefore, the influence of pipeline conveying on flow characteristics of paste before filling the goaf should be taken into consideration. Based on the above two points, this paper studies the paste strength, backfill strength, and pumpability parameters in coal mine paste filling and determines the early and later strength of coal mine paste, as well as the pumpability parameters such as slump degree, segregation degree, setting time, and paste gradation. With the determined mass proportion of coal gangue, fly ash, and silicate cement, the orthogonal test was carried out with three factors including gypsum content, the content of early strength agent (Na₂SO₄), and the mass concentration, and at three levels. The factors affecting paste flow characteristics were determined by range analysis, and the factors affecting the pastes early strength were determined by the XRD test and SEM test on its microstructure. With paste proportioning and pipeline conveying simulation system, taking slump, segregation degree, backfill strength, and other parameters as indicators, we obtain the influence law of pipeline conveying on the flow characteristics of paste. The research has great theoretical and practical significance for developing coal paste with early strength and its flow characteristics.

This paper analyzes the large deformation of roadway in three‐soft coal seam under the influence of tectonic stress. Taking the auxiliary conveying uphill roadway in Yili No. 1 Coal Mine as the engineering background, the deformation and failure mechanism of the surrounding rock and the supporting technology were studied. First, the characteristics of stress field and the surrounding rock properties of deep mining area were investigated through geostress measurement and rock mechanical test. Then, the roadway deformation and the loose circle of the supporting structure were obtained. Based on the results from measurement and theoretical analysis, we proposed a concept, i.e., “Stress adjustment‐Strengthening‐Grouting‐Secondary support.” A numerical model was established to analyze the stress distribution and the state of plasticity in the surrounding rock. According to the results of the geostress measurement and the numerical simulation, a combined support scheme was proposed, i.e., “Yielding bolt & Shotcreting” as the primary support and “Prestressed grouting anchor cable & U‐shaped shed” as the secondary support. Finally, the underground practice showed that the proposed support scheme can effectively control the large deformation and maintain the long‐term stability of the roadway in deep and three‐soft coal seam. The proposed technology has guiding significance for the support design under similar mining conditions.

In longwall mining, the deformation and destruction of overlying strata always lag behind coal extraction. The overlying strata characteristics at the lateral boundary of the stope can be classified into four categories, i.e., Hard‐Soft, Soft‐Hard, Hard‐Hard, and Soft‐Soft. In order to analyze the effect of the above four structures, we adopt viscoelastic theory to the finite element method (FEM) and define the point safety factor to evaluate the rock damage. The accuracy of programming is verified through example verification. A modified viscoelastic‐plastic FEM model is applied to analyze the performance of four overburden structures. The numerical computation results show the following: From the rupture of overburden rock to its stabilization, the duration time of four typical structures can be sorted as “Soft‐Soft < Hard‐Soft < Soft‐Hard < Hard‐Hard”. The fracture direction and dip angle of each structure vary as well. The fracture zone of the H‐S structure is inclined toward the goaf, while that of the S‐H structure is inclined to the lateral boundary of the stope. The fracture zone of the H‐H structure is also inclined toward the lateral boundary, with a greater angle than the S‐H structure, while the fracture zone of the S‐S structure is inclined to goaf, with a greater angle than the H‐S structure.

The low early strength of materials for paste filling in mines caused by low early strength of composite cementing material has been a severe issue. In this study, the effects of sulphoaluminate cement and gypsum on strengths of composite cementing material were investigated experimentally by employing the constrained formulation uniform design. With the content of the sulphoaluminate cement below 14% and the content of the gypsum below 16%, the compressive strengths of composite cementing materials increased, especially early strength. However, the initial and final setting time does not meet the engineering requirements in this case. Optimization tests of composite additives demonstrated that H₂BO₃(0.3%) + Na₂SO₄(0.1%) and H₂BO₃(0.3%) + NaNO₂(0.1%) were ideal setting retarding and early strengthening composite additives as they can both reduce the initial and final setting time and enhance compressive strengths of composite cementing material. Investigations by XRD and SEM revealed that the hydration products of composite cementing material were dominated by AFt (ettringite) at the early stage and by C-S-H (hydrated calcium silicate) gel + CH (calcium hydroxide) gel at the middle and late stages. The hydration products of ratio-optimized composite cementing material do not restrain each other due to the generation sequence. Instead, they grew interactively and were coupled, thus facilitating the growth of the hardened body. This study can provide references for optimization of composite cementing material for paste filling in coal mines.