Taking high-fluidity slag and desulfurization gypsum alkali-activated filling material (HSD - AAM) as the research object, the influence of seawater mass concentration on its setting time, flowability and strength was analyzed. Furthermore, the early hydration mechanism of HSD - AAM was revealed through SEM, XRD and TG - DSC tests.The results show that with the increase of seawater mass concentration, the setting time of HSD - AAM shortened, the flowability decreased, and the flexural and compressive strength increased. Compared with the mixing conditions of pure water, when mixed with seawater of 100% mass concentration, the flow time of HSD - AAM is decreased by about 12%, the initial and final setting time are shortened by about 28% and 27%, the flexural strength and compressive strength are increased by about 18% and 19% respectively after curing for 1 day, and the flexural strength and compressive strength are increased by about 27% and 17% respectively after curing for 7 days. Seawater can accelerate the hydration process of C−(A)−S−H, AFt and Friedel salt, alkali activator can accelerate the hydration reaction between precursor and chloride ion, thus improving the early strength of HSD - AAM.

The existing tunnel boring machines (TBM) construct tunnel surrounding rock is mainly for hard rock, and the open-type TBM construction of water-rich soft rock tunnel is an engineering problem.In view of the difficult geological conditions of water-rich soft rock in a mine project, the feasibility of open-type TBM construction in typical water-rich broken soft rock strata was analyzed from the aspects of geological condition adaptability, advancement by supporting shoes in soft and broken sandstone, and water-induced softening of water-rich broken sandstone using numerical simulation and field construction monitoring data. Graded support measures were put forward. The field monitoring results show that after using an open-type TBM for excavation and graded support, the maximum deformation of the roof is reduced from 9.2 cm to 4.5 cm, and the deformation degree is reduced by 51.5%, which effectively ensures the safety and stability of the surrounding rock of the roadway. At the same time, based on the field investigation, the construction cost and construction period were compared and analyzed, which further verified the feasibility of using open-type TBM construction. The final construction results show that the project takes 396 days to safely and successfully complete the excavation task. The maximum daily excavation is 31.53 m, and the maximum monthly excavation is 521.53 m. The excavation task is completed 6 months ahead of the planned construction period, which has good economic and social benefits.

In order to achieve the goal of cost decreasing and benefit increasing of blasting operation in underground metal mines, the medium-deep hole blasting technology of ore-rock in the middle section of 1 100−1 200 min Taibai Gold Mine was studied. The rock Proctor hardness coefficient, ore-rock bulk density, wave impedance, integrity coefficient and tensile strength were selected as the classification indexes of ore-rock blastability. Based on the game theory combination weighting method, these five indexes were given reasonable weights. The set pair analysis method was used to calculate the maximum connection degree between the measured value of the index to be classified and the common criterion of ore rock blastability level, so as to determine the ore-rock blastability level of the six exploration areas in Taibai Gold Mine.Based on the classification of ore-rock blastability, the original blasting parameters of the mining area were updated. Nine three-factor and three-level orthogonal tests were designed in the Ⅳ-level blasting area where the blasting effect was still not ideal. According to the field test results, the influence degree of each factor on the blasting effect index was analyzed. According to the relationship between the three factors and the blasting effect index, the blasting parameters of the Ⅳ-level blasting area were optimized. The optimal blasting parameters of the Ⅳ-level blasting area in Taibai Gold Mine are as follows: the hole row distance is 1.5 m, the hole bottom distance is 2.8 m, and the proximity coefficient is 1.8. These parameters were applied to the actual production of grade Ⅳ blasting area in Taibai Gold Mine. The bulk rate and explosive unit consumption were significantly reduced, and the amount of one-time ore collapse was significantly increased, which realized the cost decreasing and benefit increasing of medium-deep hole blasting technology in Taibai Gold Mine.

After conducting fan-shaped medium-deep hole blasting in underground metal mines, a series of problems such as roof collapse, hanging roof, and high block rate are prone to occur, which have certain safety hazards and high costs of secondary crushing. It is necessary to optimize the hole network parameters of the fan-shaped medium-deep hole blasting in order to ensure the uniformity of blasting blocks and the safety of the blasting operation surface. Based on the analysis of the blasting funnel test results for a fan-shaped medium-deep hole blasting with a diameter of 89 mm in a Russian mine, the reasonable ranges of hole network parameters were determined as follows, hole bottom distance of 2.8−3.2 m and row spacing of 2.3−2.5 m. Subsequently, ANSYS/LS-DYNA numerical simulation software was used to calculate and analyze the blasting models with different hole bottom distances, and the optimal hole bottom distance was determined to be 2.6−2.8 m. The optimized hole network parameters were used for on-site industrial testing, and after blasting, the fragmentation of the mining area was uniform, the roof stability was good, and the overall blasting effect was good.

In order to reveal the deformation and failure laws of surrounding rock in jointed roadway under dynamic load, 3DEC software was used to simulate and analyze the influences of different burial depths, peak stresses, and load frequencies on the deformation and instability characteristics of surrounding rock in jointed roadway under blasting load in a certain mine. The results show that with the increase of blasting load disturbance time, the displacement of the roof, floor, and right wall of jointed roadway firstly increases, then decreases, and finally stabilizes. The displacement of the roadway surrounding rock increases linearly with the increase of burial depth and peak stress, and the plastic zone area of the roadway surrounding rock also increases accordingly. The deformation and failure degree of the roadway surrounding rock deepen, and the most serious location is on the floor.The displacement of the roadway surrounding rock decreases linearly with the increase of load frequency, and the area of the plastic zone around the roadway slightly decreases. The influences of three factors on the deformation of roadway surrounding rock were analyzed by response surface methodology. It was found that the burial depth had the most significant impact, followed by the peak stress, and the load frequency was the smallest. The research results can provide a theoretical guidance for mine safety production.

In order to study the mechanism of debris flow in the underground development project of weak cemented aquifer, the weak cemented sandstone of a mine in Xinjiang was taken as the research object, and its cementation characteristics and water characteristics were analyzed. The mathematical models of porosity, expansion rate and permeability were used to obtain the permeability of weak cemented aquifer under different expansion rates. On this basis, the fluid-solid coupling model was used to analyze the changes of surrounding plastic zone and seepage field in the process of water expansion of surrounding rock of roadway loose circle under different buried depth conditions.The research shows that the water expansion of surrounding rock will increase the permeability of aquifer and improve its permeability. When the expansion rate of surrounding rock increases to 30%, the unit water inflow after roadway excavation will increase by 1.5 to 2.0 times, and with the increase of development depth, the unit water inflow will increase greatly, which will provide sufficient water source for the occurrence of underground debris flow. The strength of weak cemented siltstone is low, and it is easy to form a plastic failure zone on the tunnel face of the development project. With the increase of the development depth, the plastic failure of the tunnel face is greater, and it is easy to collapse and decompose into a source under the condition of groundwater erosion and scour. The stability of surrounding rock in the underground fracture development area is poor, the groundwater seepage channel is smooth and the supply is sufficient. It is easy to expand, collapse and decompose under the action of groundwater, and it is carried by groundwater to form debris flow, which is the key area of underground debris flow. The research results are helpful to further reveal the formation mechanism of debris flow in the underground development project of weak cemented aquifer.

In view of the deficiency of the research on the prevention and control of rock burst in hard rock, a method of joint prevention and control of rock burst by cutting seam and bolt anchoring was proposed.The bolt anchoring samples without cutting seam, the bolt anchoring samples with cutting seam and the bolt anchoring samples with cutting seam grouting were prepared. The pre-tightening force relaxation tests and biaxial compression tests of bolts with different anchorage body samples were carried out. The strength of composite surrounding rock and safety factor with cutting seam and bolt anchoring were discussed. The results show that compared with the simple bolt anchoring, the stress concentration is transferred from the bolt anchoring zone to the end of the cutting seam by cutting seam, and the fracture evolution is transferred from the surrounding of the bolt anchoring zone to the distal end of the cutting seam. The bearing capacity, damage severity and crushing degree of the sample are reduced. The stress concentration area is enlarged by the cutting seam grouting, and the key layer of flattening is formed. The area is extended from the bolt anchoring zone to the end of the cutting seam, and the fracture evolution is extended from the surrounding of the bolt anchoring zone to the the distal end of the cutting seam. The overall bearing capacity of the sample is improved.The relaxation law of bolt pre-tightening force under the three anchoring conditions before and after loading is basically the same, showing three stages of rapid decay relaxation, slow decay relaxation and stable decay relaxation. The relaxation process of bolt pre-tightening force can reflect the internal fracture evolution of anchorage body. A simplified formula for calculating the strength of composite surrounding rock under different anchoring conditions is established, the energy consumption coefficient is defined, and a formula for calculating the safety factor of composite surrounding rock after excavation considering the energy consumption is derived, three methods of rock burst prevention and control are given, which can be used in engineering. The research results can provide reference for the local danger-relief of rock burst, bolt anchoring, prediction and prevention and control of hard rock.

After grouting reinforcement of fractured rock masses, both filled and unfilled fractures often coexist internally. To investigate the shear characteristics of rock masses under fracture differential filling, a numerical model of grouting-bonded particle flow was constructed based on the indoor direct shear tests of intact and fractured grouting-bonded red sandstone, and the meso-mechanical parameters of grouting rock mass were obtained. The numerical simulation of normal stress direct shear was carried out for red sandstone samples with prefabricated 45° parallel double fractures under unfilled, half-filled (left-filled, right-filled) and full-filled conditions. The results show that under the half-filled condition, the tensile-shear fractures are first generated at both ends of the open fractures, and gradually expand to both ends of the rock until failure. After failure, the proportion of tensile-shear, compression-shear and tensile fractures remains basically unchanged, among which the tensile-shear fractures account for the largest proportion, about 80%. Compared with the left-filled condition, the right-filled condition enters the failure stage faster, but the number of fractures is less than that of the left-filled condition. Filling fractures with epoxy resin slurry can weaken the guiding effect of open fractures on fracture initiation when rock mass is damaged, and improve shear strength. The shear strength under left-filled and right-filled conditions is not much different, and the shear strength increases best under full-filled conditions. The research results have reference significance for the evaluation of grouting effect in fractured rock mass.s.

In order to explore the effect of freeze-thaw action on the strength and acoustic emission characteristic parameters of banded magnetite quartzite, the mechanical and acoustic emission tests were carried out on the banded magnetite quartzite with the freeze-thaw interval of[－20, 20]℃ and the longest freeze-thaw cycle of 280 times. The results show that the uniaxial compressive strength of the rock sample decreases from 200.93 MPa during drying to 106.64 MPa after 280 freeze-thaw cycles, with a decrease of 46.94%. Before and after 70 freeze-thaw cycles, the acoustic emission impact rate, cumulative impact number, energy rate and cumulative energy are significantly different in the compaction stage, elastic deformation stage and microcrack evolution stage. The damage of rock samples caused by high (period≥70 times) and low (period≤40 times) freeze-thaw cycles is obviously different. The cracks of rock samples in high freeze-thaw cycles are prone to friction, slip and extension under load, which leads to the stress redistribution of rock samples and the high degree of energy dissipation. Compared with the rock samples with low freeze-thaw cycle, the cumulative impact number, ringing count and energy of acoustic emission of rock samples with high freeze-thaw cycle in the accelerated expansion stage of micro-cracks decrease by 35.61%, 41.44% and 17.25%, respectively. The intensity of fracture of rock samples with high freeze-thaw cycles is significantly weaker than that of rock samples with low freeze-thaw cycles. With the increase of freeze-thaw cycles, the fracture characteristics of rock samples gradually change from brittle to ductile fracture, and the single b value of rock samples decreases first and then increases. The deterioration of rock samples under low freeze-thaw cycles is mainly caused by the increase of the number and scale of microcracks, while the deterioration of rock samples under high freeze-thaw cycles is caused by the penetration and merging of local fracture networks.

In order to achieve precise control of the process of rock breaking, the characteristics and evolution laws of coda wave signals indifferent rock fracturing stages were studied. The coda wave signal test of supercritical CO₂ fracturing sandstone under triaxial confining pressure was carried out, and the evolution characteristics and signal parameter changing law of coda wave signals during CO₂ fracturing process were analyzed. The sensitive nuclei function of coda wave extension based on probability density was established, and the characteristics and evolution law of coda wave signal in different fracturing stages were revealed. The results indicate that supercritical CO₂ fracturing rock can be divided into four stages: early stage, middle stage, late stage and post-failure stage. The relative change rate of coda wave velocity shows a downward trend as a whole, reflecting the rapid accumulation process of rock damage. The diffusion sensitive nuclei reveals that the rock damage gradually accumulates during the fracturing process, resulting in enhanced wave scattering inside the rock, increased energy loss and rapid development of coda waves.