Thickening and dewatering of tailings is a key link in the tailings filling process. In order to explore the thickening and dewatering performance of tailings in a mine, static and dynamic flocculation settlement tests of tailings were carried out according to the basic physicochemical properties of tailings. The static flocculation settlement tests were designed with the mass concentration of tailings slurry, the flocculant type and the flocculant content as the influencing factors and the settlement rate as the research index. The dynamic flocculation settlement tests were designed with feed rate, feed mass concentration and flocculant content as influencing factors, and solid content of overflow water, underflow mass concentration and settling velocity of tailings as research indexes. Based on response surface analysis method, the dynamic flocculation settlement law of total tailings under the interaction of multiple factors was revealed. Firstly, through the static flocculation settlement tests, the optimal type of flocculant was determined as CKFA02, the optimal content of flocculant was determined as 20 g/t, and the optimal feed mass concentration of tailings slurry was determined as 10%−12.5%. Secondly, according to the response surface model constructed based on the dynamic flocculation settlement tests, the maximum influencing factors on the underflow mass concentration, solid content of overflow water, and settling velocity of tailings are the slurry feed rate. The interaction between slurry feed rate and feed mass concentration has the greatest impact on the underflow mass concentration, while the interaction between slurry feed rate and flocculant content has the greatest impact on the solid content of overflow water and settling velocity of tailings. Thirdly, the optimal settlement parameters are the slurry feed rate of 0.55 t/(m²·h), feed mass concentration of 11.12%, flocculant content of 30 g/t. Under the optimal flocculation settlement parameters, the underflow mass concentration is 60.6%, the solid content of overflow water is 251.4 mg/L, and the settling velocity of tailings is 58.06 mm/s.

There are significant differences in mechanical properties and failure modes among backfill with different sizes. In order to study the size effect of backfill, the uniaxial compression experiments were conducted on four groups of backfill specimens with different sizes, and an acoustic emission meter was used to monitor the energy release during the loading process. Numerical simulation was used to simulate and verify the compression failure of the specimens and study the size effect and damage characteristics of cylindrical backfill specimens with different sizes. Finally, the scanning electron microscope tests were carried out on the cylindrical specimens of gangue cemented backfill with different sizes, and the microscopic morphology difference of different sizes of specimens was observed. The results show that the uniaxial compressive strength of the specimen is inversely proportional to the size of the specimen. The small size specimen is dominated by shear failure. When the size of the specimen increases to a certain extent, it shows a certain tensile failure. With the increase of the size of the specimen, the specimen shows a more obvious size effect. The acoustic emission characteristic of the cylindrical specimen can be divided into the initial preparation stage, the acoustic emission increasing stage and the acoustic emission decreasing stage. The acoustic emission activity of the specimen with small size is more intense than that with the large size. The larger the size of specimen, the more weak points in the specimen, and the worse the cementation ability of the specimen. The research results can provide a scientific basis for the determination of backfill parameters of different size structures in gangue cemented filling.

In order to solve the problems such as high block rate, severe damage to the filling bodies and top pillars on both sides during the two-step mining process, and high ore impoverishment rate encountered during the mining by sublevel drilling and stage open stope with subsequent filling method in Sanguikou lead-zinc mine, optimization research was conducted on the original blasting parameters. Using a diameter of the blast hole of 76 mm and based on the analysis of rock drillability and blastability, through theoretical calculation, it is determined that the row spacing is 1.8 m and the hole bottom spacing is 2.1 m; alternatively, the row spacing is 1.9 m and the hole bottom spacing is 2.0 m. According to on-site industrial tests, compared with the stope using the original blasting parameters, the optimized blasting parameters reduced ore impoverishment rate by 19.8%, ore loss rate by 4.6%, the explosive consumption by 12.2%, and the block rate by 34.8%. The simulation analysis of the blasting effect before and after optimization of blasting parameters was conducted using LS-DYNA finite element software, and the simulation results showed that when using optimized blasting parameters for blasting, the ore block rate is lower, and the damage to the filling body on both sides, top pillar, and rock drilling roadway arch line is minimal.

Earthquakes are one of the main causes leading to dynamic instability of tailings dams. Based on the pseudo-dynamic method combined with the Swedish arc method, the evaluation indexes of dynamic stability of tailings dams can be quickly determined. To improve the computational efficiency of the pseudo-dynamic method in the evaluation of slope earthquake stability, a separated critical initial phase search strategy was introduced, and a fast pseudo-dynamic method was proposed. The traditional pseudo-static method, the pseudo-static method considering the amplification factor, and the fast pseudo-dynamic method were applied respectively to evaluate the seismic stability of a tailings dam in Xizang from two angles of safety factor and failure probability. The results of numerical analysis show that the safety factors obtained by above three methods all meet the minimum stability requirements. In the view of the safety factor and the sliding area comprehensively, the results obtained by the pseudo-static method considering the amplification factor tend to be conservative. The influence of P-wave velocity on the calculation results is small. With the increase of S-wave velocity, the safety factor obtained by the fast pseudo-dynamic method gradually decreases, the failure probability gradually increases, the number of instability sliding surfaces gradually increases, and the risk sources of earthquake instability of tailings dams increase. The research results have a guiding significance for the evaluation of earthquake stability of tailings dams.

The traditional theoretical methods mostly use semi-quantitatively to analyze the safety thickness of the goaf roof, which can not fully reflect the stress and failure mode of the roof. Therefore, the cusp mutation theory was introduced to study the safety thickness of the goaf roof, the research model of the safety thickness of the roof was established, and the formula of the minimum safety thickness of the roof was derived. Based on the occurrence status of goaf in Dabaoshan Mine, the theoretical values of safety thickness of goaf roof under different lithology and different span conditions were calculated. Finally, the formula was verified by the traditional theory and the finite element numerical simulation method. The research results show that it is feasible to introduce the cusp mutation theory into the study of the safety thickness of the goaf roof, and the derived formula of the minimum safety thickness of the goaf roof is reasonable and reliable. Under the conditions of the slightly weathered limestone, slightly weathered secondary dacite-porphyry, copper-sulfur ore body, and sandstone-rock group in Dabaoshan Mine, the safety thickness of the goaf roof calculated based on the cusp mutation theory is 7.3 m, 12.5 m, 9.5 m, and 22.1 m, respectively, which can be used as the basis for on-site goaf treatment.

In order to improve the rationality of mechanical parameters of tunnel surrounding rock, a new inversion model of surrounding rock parameters was proposed based on a tunnel project with a super large-section in Zhuhai. After the initial samples were generated based on Latin hypercube sampling (LHS), the parameter sensitivity analysis was carried out to determine the key parameters of the surrounding rock and improve the sample structure. Then, the whale optimization algorithm (WOA) was used to optimize the number of hidden layer nodes, the initial weights and the thresholds of the extreme learning machine (ELM) to form the LHS-WOA-ELM inversion model. The inversion parameters were substituted into FLAC3D to calculate the displacement and compare with the field measured data. The results show that the parameter sensitivity analysis based on LHS can investigate the co-variation of multi-parameters with fewer samples, and determine the main parameters affecting the displacement of surrounding rock as elastic modulus E, cohesion c and internal friction angle φ. Compared with WOA, ELM and BP algorithm models, the difference between the calculated displacement values obtained by LHS-WOA-ELM inversion model and the measured displacement values is smaller, indicating that the inversion analysis method can well reflect the nonlinear and uncertain characteristics between the surrounding rock parameters and deformation, and further improve the accuracy and efficiency of the surrounding rock inversion in tunnels with super-large section, which can provide a reference for determining the design parameters of underground caverns and mining projects.

In order to reveal the influences of super-thick and hard strata on abutment pressure and the evolution law of energy distribution in stope, a typical mining area of Yangliu Coal Mine was taken as the research object, and the variation law of abutment pressure and energy evolution process were simulated by numerical simulation method. The results are concluded as follows. (1) When there are super-thick and hard strata in the overlying strata, the advance abutment pressure in the middle and both sides of the working face is increased by about 9% and 5% respectively compared with that without super-thick and hard strata. This shows that when there are super-thick and hard strata in the overlying strata, the middle and both sides of the stope are in a high stress state for a long time. When the stope is disturbed by external force, it is more likely to lead to the occurrence of dynamic disasters such as mine earthquake and rock burst. (2) When there are super-thick and hard strata in the overlying strata, the change curve of the advance abutment pressure of the stope shows a cliff-like decline after breaking. When there are no super-thick and hard strata in the overlying strata, the change curve of the advance abutment pressure of the stope is gentler than that of the former. (3) With the continuous advancement of the working face distance, the peak value of the abutment pressure of the free side is increased by 3.60 MPa and 3.39 MPa respectively compared with the peak value of the abutment pressure of the mining side when there are or are not super-thick and hard strata in the overlying strata. The change trend of the side abutment pressure of the two is basically the same, but the side abutment pressure of the free side of the roadway is obviously larger than the lateral abutment pressure of the mining side of the roadway. This shows that the existence of super-thick and hard strata makes the stress concentration, roadway mining will be more prone to damage. (4) The super-thick and hard strata will cause great impact on the stope after breaking. Compared with the stope without super-thick and hard strata, the damage degree is more severe, the elastic energy released is more, and the influence range is more extensive.

Obtaining the stability level of rock mass is of great significance to the deep mining of ore body, surrounding rock support and filling of goaf. Based on the engineering geological conditions of Zhongguan Iron Mine, the basic quality BQ values of rock mass at the measuring points were determined through laboratory rock mechanics basic property test, field rock mechanics test, field rock joint investigation and in-situ stress test. Using the data processing and spatial analysis function of ArcGIS software, the spatial distribution prediction model of rock mass quality in the study area was established through Kriging interpolation technology. The rock mass quality of –230 m horizontal ore body was predicted and analyzed using the model, and the field measurement verification was carried out. The results show that the prediction results of the spatial distribution of rock mass quality are in good agreement with the actual situation. The basic quality BQ value of the –230 m level ore body in Zhongguan Iron Mine is 57.843 - 347.573, which is in the fourth and fifth grades. The ore body quality is poor, easy to break and unstable. The spatial distribution prediction model of rock mass quality can provide a method for studying the stability level of rock mass under the condition that rock mass engineering is not fully exposed, and also can build a theoretical basis for the safe mining of Zhongguan Iron Mine in the future.

The underground rock body is often affected by external factors, after repeated cycles of loading and unloading, it will lead to the increase of internal cracks in rock and the change of local stress environment. Therefore, the triaxial graded cyclic loading and unloading tests with different upper limit stresses were carried out on sandstone, and the damage mechanical properties and the laws of energy evolution of sandstone were obtained. The test results show that under the conditions of graded cyclic loading and unloading, the stress-strain curve of sandstone is characterized by hysteresis loops and downward bending, and the hysteresis loops are gradually enlarged with the increase of stress. Under the cyclic loading and unloading with different upper limit stresses, as the number of cycles increasing, the axial elastic strain and axial plastic strain of sandstone show linear relationships with the number of cycles, while the sudden drop of axial plastic strain indicates that the internal pore cracks of sandstone are compacted rapidly. Meanwhile, the input energy and elastic energy curves are very similar during the cycling process, showing an upward trend of nearly exponential curve. Dissipation energy grows relatively slowly in the initial cycling stage, which indicates that the original cracks in the sandstone are compressed and closed; while in the mid-cycling stage, the dissipation energy shows a nearly linear growth trend, which indicates that the rock samples are in the stage of stable expansion of the cracks. The dissipation energy ratio shows a sudden decrease in the initial cycling stage; while in the mid-cycling stage, it maintains an approximate horizontal trend. In addition, under the graded cyclic loading and unloading, there is a linear relationship between the degree of damage to sandstone and the number of cycles, indicating that the number of load cycles has a significant impact on the degree of damage to sandstone.

To study the mechanical properties of marble after damage caused by dynamic and static combined loads, the improved split Hopkinson pressure bar device was used to carry out dynamic and static combined load damage tests on marble at four axial compression levels (0, 18 MPa, 36 MPa, 54 MPa) and three impact velocity levels (8 m/s, 9 m/s, 10 m/s). The uniaxial compression acoustic emission tests on the damaged marble was carried out. The results show that under the action of static and dynamic combination, with the increase of axial static pressure, the dynamic peak stress, dynamic elastic modulus and wave velocity of marble increase firstly and then decrease. The absorption energy is closely related to the axial static pressure. When the axial static pressure is low, the energy required for the closure of microcracks in the rock is mainly obtained from the impact bar. With the increase of axial static pressure, the proportion of energy obtained from the impact bar gradually decreases. With the increase of axial static pressure, the uniaxial compressive strength and elastic modulus of rock after damage caused by dynamic and static combined load increase firstly and then decrease. When the axial static pressure is 18 MPa, the uniaxial compressive strength and elastic modulus of rock reach the maximum value. Appropriate axial static pressure can effectively improve the ability of rock to resist dynamic load and reduce rock damage. In the middle and later stages of the uniaxial loading failure process, the less the number of high-frequency acoustic emission signals (>400 kHz) of the rock with small damage degree, the more the number of medium and low-frequency acoustic emission signals (165 - 205 kHz). The proportion of acoustic emission signals at 165−205 kHz increases firstly and then decreases during the whole loading process, which can be used as a precursory feature of rock failure.