In response to the engineering difficulties of strength design of filling body during deep phosphate mining, a combination of on-site monitoring and numerical simulation was used to monitor the stress of the cemented backfill in the full-time (filling stage, curing stage, bearing stage) of the mining site, so as to obtain the internal stress evolution law of the filling body. Numerical simulation analysis shows that after the first stage of mining and filling is completed, the stress is mainly generated by the reserved pillar bearing the overlying rock mass. After the completion of the second stage of mining and filling, the stress distribution of the filling body in the mining area is uneven, with a stress distribution pattern of “large in the middle and small at both sides” inside the filling body. The average stress on the filling pillars is 3.0 MPa. The on-site measured data analysis shows that the filling body is a passive bearing body, and it gradually begins to bear after blasting and mining the adjacent pillars of the filling body. After the back filling of the side pillar of the filling body, the maximum vertical stress reached 0.714 MPa. The mining area adopts a filling body with a design strength of 3.0 - 3.5 MPa, which can resist the disturbance stress generated by blasting mining, and the filling body is in a stable state.

In order to solve the problems of poor mining safety and high loss and depletion rate of the steeply inclined and soft broken veins in Shanggong Gold Mine, the F1-Ⅲ vein were subjected to on-site industrial experimental research using the quasi-medium-deep hole segmented-open pit method. The results show three aspects. Firstly, all the mining works are arranged in stable rock mass outside the vein, and the safety of the construction process can be ensured by drilling medium-deep holes into the ore body from the roadway outside the vein to control and recover the ore body. Secondly, the non-cutting engineering blasting groove technology, which takes the mine approach as the initial free surface of blasting and the 3-4 rows of holes are simultaneously blasted from bottom to top to form a full section, can significantly reduce safety hazards, reduce mining costs and simplify construction technology. Lastly, the dividing line between the surrounding rock and the vein is accurately determined during the drilling process, the hole is protected by plastic sleeves and the mining is carried out by fine blasting technology. When the stope process parameters are determined as the minimum resistance line of 1.5 m, the row distance of 1.5 m, the hole bottom distance of 2.0 m, and the single explosive consumption of 0.22 kg/t, the blasting effect is the best. The loss rate is reduced to 10.1%, the dilution rate is reduced to 12.4%, the production capacity of the mining site is increased by 54.6 t/d compared to the shallow hole retention method, and the direct mining cost is reduced by 11.06 yuan/t.

In order to realize the smooth transition from caving method to filling method for a high value iron ore body, a combined mining scheme of caving method to filling method was proposed from the perspective of stope stability and full recovery of high value ore body.Using simplified beam method and Rubenieit method, the thickness of isolated top pillar in the transition area of caving method to filling method is 10.6 m and 9.4 m respectively. Based on the theoretical calculation results of two kinds of isolated top pillar thickness, the numerical simulation of the combined mining scheme of the iron ore with caving method to filling method was carried out by FLAC3D software. The thickness of the isolated top pillar in the transition area of the combined mining scheme of the iron ore with caving method to filling method is defined as 10 m. The research results have been applied to the actual production of the iron mine, and good economic benefits have been achieved, which provides a reference for caving method to filling method in the same type of mine.

The waste rock cemented backfill often exhibits stratification during the filling process. Therefore, a study was conducted on the filling effect of layered backfill. Slicing of waste rock cemented backfill was carried out, and the distribution characteristics of aggregates in different layers of the backfill were analyzed. Then, through numerical simulation, the influences of layered and non-layered waste rock cemented backfill on the filling effects of goaf were compared and analyzed. It is found that the stress in the middle of the roof of the goaf filled with layered waste rock cemented backfill and non-layered waste rock cemented backfill is about 3.68 MPa and 3.61 MPa, respectively. The settlement degree of the roof of the non-layered backfill is significantly smaller, which can provide more stable support for the roof. The lateral support of the non-layered backfill is uniformly stressed, with a small displacement and a maximum displacement of 3.76 mm. The lateral displacement of the layered backfill is small in the upper part and gradually increases downwards, with a maximum displacement of 4.46 mm. The distribution of fine aggregates in the upper layer of layered backfill makes it susceptible to pressure deformation, while the waste rocks distribution inside the non-layered backfill is uniform, with a stronger overall bearing capacity and relatively smaller deformation.Therefore, when carrying out the filling material proportion, it is important to control the backfill layering phenomenon as much as possible to provide important guarantees for mine safety production.

The stability of exposed one-step filling body on both sides is the key to safe mining in the two-step mining of subsequent filling mine. By means of numerical simulation and field monitoring, the stress and displacement states of different positions under the condition of one-step filling body unilateral exposure in 5371 stope of an iron mine were studied. The results show that the vertical stress of the measuring point at the bottom of the exposed filling body changes the most, about 0.48 MPa. The horizontal displacement of the measuring point at the top and closer to the exposed surface is larger, and the maximum horizontal displacement is about 32 mm. The mining rate of the two-step pillar is logarithmically related to the maximum vertical stress and maximum horizontal displacement of the exposed filling body. Different excavation times have little effect on the vertical stress at the bottom of the filling body, which is about 1.92%, and have a great effect on the horizontal displacement of the top of the filling body, which is about 5.65%. The field monitoring results show that during the two-step mining, the vertical stress in the middle of the unilateral exposed filling body changes greatly, about 0.3 MPa, and the horizontal displacement of the top measuring point is the largest, about 30 mm, which is basically consistent with the numerical simulation results.The research results can provide safety guarantee for the two-step mining work of the iron mine.

The evolution law of mechanical properties of tailings materials under freeze-thaw cycles is explored to provide a theoretical basis for the safety and stability evaluation of tailings dams. The tailings samples with different water contents were subjected to 1, 3, 5, 7, 9 freeze-thaw tests, and triaxial shear tests were performed under confining pressure conditions of 100 kPa to 400 kPa. The test results show that with the increase of the number of freeze-thaw cycles, the macroscopic structural deformation of the tailings increases significantly. Especially when the water content is 15%, after 7 freeze-thaw cycles, the pore structure of the tailings changes significantly, and the number of pores at the volume level of 0 − 250 μm³ increases significantly. In addition, it is found that increasing confining pressure can improve the elastic modulus of tailings and inhibit the damage of tailings caused by freeze-thaw cycles. The deformation modulus of the tailings tends to be stable after 5 freeze-thaw cycles, while the tailings samples with high water content show lower cohesion, internal friction angle and deformation modulus.

Aiming at the technical bottleneck of low compressive strength, high filling cost and poor filling effect of cement-based filling body, ultrasonic technology was used to improve the performance of cement-based filling slurry. Through the orthogonal tests of three factors and four levels, the effects of ultrasonic power, ultrasonic time and stirring time on the rheological properties and mechanical properties of cement-based filling slurry were investigated, and the strengthening mechanism of ultrasonic action was further discussed. There search results indicate that the primary and secondary factors affecting the compressive strength of the filling body are ultrasonic power, ultrasonic time and stirring time. Ultrasonic power is a significant factor affecting the yield stress and plastic viscosity of the filling slurry, followed by stirring time, and ultrasonic time has the smallest impact. The compressive strength of the filling body at 3 days, 7 days, 14 days and 28 days shows a trend of first increasing and then decreasing with the increase of ultrasonic power, ultrasonic time and stirring time. When the ultrasonic power, ultrasonic time and stirring time are 70 W, 4 min and 6 min, respectively, the compressive strength of the filling body reaches its optimal value. The yield stress and plastic viscosity of the filling slurry are negatively correlated with ultrasonic power and ultrasonic time, and show a trend of first decreasing and then increasing with the extension of stirring time. This is because the ultrasonic action promotes the dispersion of cement and tailings particles in the liquid phase, improves the uniformity of the slurry, and promotes the cement hydration process. However, excessive ultrasonic power and prolonged ultrasonic time lead to the reunion of the flocculation structure of cement particles, which inhibits the diffusion of water molecules and the compactness of the slurry structure. The appropriate stirring time is beneficial to the dispersion of the slurry and weakens the friction and shear between the particles, while the excessive stirring time leads to the destruction of the slurry structure and weakens the mechanical properties of the filling body.

To investigate the rheological properties of filling slurry with steel slag and unclassified tailings under different proportioning conditions, rheological tests were conducted using the controlled shear rate method, and the effects of mass concentration, cement-sand ratio, and steel slag content on the rheological properties of the slurry were analyzed. The interaction among mass concentration, cement-sand ratio, and steel slag content on the rheological properties of slurry was investigated using the central composite design response surface method.The results show that the mass concentration has the most significant effect on the yield stress and viscosity, and with the increase of mass concentration, the yield stress and viscosity show a power function growth law. The steel slag content reduces the viscosity and yield stress of the slurry, significantly improving its flowability. When the steel slag content is increased from 0 to 25%, the yield stress decreases by 85.43% and the viscosity decreases by 53.83%. For the interaction of various influencing factors, the interaction between mass concentration and steel slag content has the most significant effect on yield stress and viscosity. The increase of steel slag content promotes the influence degree of mass concentration on rheological parameters, that is, as the steel slag content increases, the influence of mass concentration on yield stress and viscosity gradually increases. Parameter optimization shows that when the mass concentration is 60%, the steel slag content is 30 %, and the cement-sand ratio is 1:8, the yield stress and viscosity of the slurry are the smallest, and the flowability is the highest.

Advancements in exploration theory and equipment have significantly enhanced mine reserve expansion. However, adapting new resources in the deep periphery of mines to existing production systems requires reassessment. Green stoping-and-filling operations face challenges from varying filling gradients and existing slurry pipeline transportation parameters that no longer meet practical demands. To address these challenges, tests were conducted on the particle size distribution of backfill materials and slurry flow performance at the mine. Results identify that a reasonable range of slurry mass concentration is 70%-72%. Through theoretical analysis and calculation, the matching degree between the design parameters of the existing filling system and the orebody filling gradients was verified. The maximum gravity flow filling gradients of the slurry mass concentration of 70% and 72% are 9.31 and 5.56, respectively. Simulation models were constructed to verify the correctness of theoretical analysis. The results show that a 72% slurry mass concentration is suitable for gravity flow transportation in older mine zones with filling gradients of 2.3-4.9, while reducing concentration to 70% enables gravity filling in high-gradient goafs with filling gradients of 7.2-8.7 in newly developed zones. For the mine to achieve optimal technical and economic performance, two schemes are proposed:implementing additional pumps to transport slurry with a mass concentration of 72% under pressure(first scheme), and utilizing slurry with a mass concentration of 70% for gravity flow (second scheme). A comparative analysis shows that the first scheme has obvious advantages and disadvantages, while the second scheme requires no technical modification, offers a more reliable process, and reduces operational costs by RMB 2 334 500 compared to the first scheme. Therefore, the second scheme is the preferred option.

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.

In order to accurately predict the mine water inflow after exposing the aquifer in the open-pit mine of the main ore body of Kambove mine, the mine field pumping test was carried out in combination with the existing hydrogeological data of the mine, and the mine pit water inflow in different mining years was calculated based on the pressure bearing transfer to no pressure formula in the large well method. Based on field hydrogeological data and pumping test results, the main ore body is divided into strong water-rich zone (Ⅰ), medium water-rich zone (Ⅱ) and an aquifer (Ⅲ). When the mine is pushed to a level of 1 300 min the third year of production, the underground water inflow is the smallest. When the mining level reached 1 270 m in the sixth year of production, the underground water inflow reached its maximum value.Combined with the rainfall in normal rainy season and rainstorm season, the total water inflow of the mine pit can reach 23 152 m³/d and 46 552 m³/d respectively. The research results can provide important data support for the development of subsequent waterproofing and drainage plans for the mine.

In order to enhance the shear performance of graphite tailings, based on the indoor direct shear test, the effects of clay content, water content, quicklime content and polypropylene fiber content on the shear strength of graphite tailings were determined by single factor test. The optimum mix ratio of clay, water and quicklime was obtained by response surface method (RSM), and the optimum mix ratio scheme was verified. The microscopic characteristics of the samples were analyzed by scanning electron microscopy. The results show that with the addition of clay, water, quicklime and polypropylene fiber, the shear strength and cohesion of the solidified graphite tailings increase first and then decrease, and the internal friction angle tends to be stable.When the clay content is 29.38%, the water content is 15.82%, and the quicklime content is 3.59%, the cohesion of the solidified graphite tailings reaches the best, which is 9.417 3 kPa. The average cohesion obtained by the test is 9.487 kPa. The relative error between the predicted value and the test value is 0.7%, less than 5%, which verifies the feasibility of the response surface design model.The SEM results show that the agglomeration structure is formed between the modified graphite tailings particles and the clay, which improves the shear strength of the solidified material. The research results are of great significance to the reinforcement and protection of graphite tailings reservoir slope.

Compound portland-sulphoaluminate cement coal-based solid waste grouting material was prepared by using coal gangue and fly ash as raw materials, supplemented by portland cement and sulphoaluminate cement as cementitious materials. By adjusting the mass ratio of portland cement and sulphoaluminate cement, the effects of cement with different mass ratios on the mechanical properties and flow properties of grouting materials were studied. The pore size characteristics of grouting materials were analyzed by nuclear magnetic resonance, and the hydration products of grouting materials were analyzed by scanning electron microscope and energy spectrum. The results show that when the mass ratio of portland cement to sulphoaluminate cement is 3:7, the 28 d compressive strength and flexural strength of the specimen are 17.05 MPa and 5.00 MPa, respectively, the slump is 34.83 cm, the setting time is 174 min, and the comprehensive performance is the best. The main hydration product of the grouting material is C-A-S-H gel, which is mutually reinforced with the skeleton formed by gangue powder aggregate, fly ash, C-S-H gel, which effectively improves the strength of the specimen. The internal pore size of the grouting material is mainly distributed in 0−0.63 μm and 1−4 μm, and the pore size of 0.05−0.4 μm occupies the main position. The research results can provide some theoretical guidance for the development and engineering application of grouting materials.

Conventional ecological restoration surveys in mines have the characteristics of long time, high cost, and difficulty in tracing the past, which have certain limitations. In order to quickly and intuitively obtain the distribution and restoration progress of historical abandoned mines in aimed region, taking northwest Yunnan as the research area, multi temporal remote sensing data was used to summarize the interpretation indicators of distribution in historical abandoned mines, mine ecological restoration models, and specific restoration measures in this area. Through conducting detailed interpretation and analysis and combining with field investigation and verification work, the distribution of historical abandoned mines in this area and the subsequent ecological restoration situation of mines were clarified. The results show that there are a total of 3 060 historical mining spots in this area, with mining and transfer sites as the main types of damage, accounting for 81.8% of the total number of spots. The involved minerals are mainly non-metallic minerals. In terms of mine restoration, 765 mining spots with an area of approximately 1 468 hectares have been repaired and treated, mainly through engineering restoration. The investigation and research results can provide a reference for the investigation and restoration of abandoned mines, as well as important data support for the planning, policy formulation, and deployment of mine ecological restoration and management work.

In order to study the influence of dynamic obstacles on mine wind speed, four influencing factors of roadway inlet wind speed, mine car driving speed, mine car length and blocking ratio were selected when the mine car was driving in the transportation roadway. The dynamic grid technology was used for numerical simulation, and the variation law of the influence of different factors on the roadway wind flow was obtained. The results show that the greater the difference between the inlet wind speed of the roadway and the driving speed of the mine car, the more obvious the average wind speed change of the cross section of the roadway, and the longer the interference distance of the mine car to the wind flow of the roadway, which greatly affects the stability of mine ventilation system. The mine car length and the blocking ratio have little effect on the stability of the mine ventilation system. Using the stepwise linear regression analysis method in SPSS software, the influence of four influencing factors on the change degree of roadway wind speed from strong to weak is mine car driving speed, roadway inlet wind speed, blocking ratio and mine car length. The research results are of great significance to maintaining the stability of mine ventilation system.

In order to reduce the energy consumption of ball mill, ZQM250×100 ball mill was used to study the grinding energy consumption of an iron ore after high-pressure roller grinding, and the energy required for the newly generated －0.074 mm grade product after grinding was used as the evaluation standard of grinding energy consumption. The effects of medium filling rate, ore addition quantity, water addition quantity and mill frequency on the energy consumption of the newly generated －0.074 mm product, the significance of the interaction between the factors and the optimal test conditions were analyzed and predicted by the response surface methodology, and the results were compared with the single factor test results. The optimum conditions obtained by single factor test are as follows: medium filling rate is 40%, ore addition quantity is 700 g, water addition quantity is 300 mL (mass concentration is 70%), mill frequency is 50 Hz, and energy consumption of the newly generated －0.074 mm grade product is 169.54 (kW·h)/t. The response surface methodology analysis show that the mill frequency has the greatest influence on the energy consumption of the newly generated －0.074 mm grade product, followed by the medium filling rate, water addition quantity, and ore addition quantity. The interaction between medium filling rate and mill frequency is the most obvious, followed by water addition quantity and mill frequency, ore addition quantity and water addition quantity, medium filling rate and ore addition quantity, ore addition quantity and mill frequency, medium filling rate and water addition quantity. The model prediction based on the response surface methodology shows that the optimum test conditions are: 39.71% medium filling rate, 698.15 gore addition quantity, 339.65 mL (mass concentration 67.27%) water addition quantity, and 49.73 Hz mill frequency. Under this optimal condition, the predicted energy consumption of the newly generated －0.074 mm product is 120.53 (kW·h)/t, and the average energy consumption of the verification test is 118.97 (kW·h)/t, which is very close to the predicted value, indicating that the response surface prediction model is reasonable and effective.

In order to solve the problems of high energy consumption, material accumulation in the air duct, and local wear in the swing mill, an analysis was conducted on the existing structures of the intake air duct, including volute and air duct blades. Based on fluid dynamics, multiphase flow, and discrete element coupling simulation technology, numerical simulations were conducted on the internal wind and particle fields of a pendulum mill. Response surface simulation tests were conducted on the volute shape, the blades angle of air duct, and the blades number of air duct. The results show that optimizing the volute shape is beneficial for reducing pressure loss and equipment energy consumption. Reasonable selection of blade angle and number is beneficial for reducing the uneven coefficient of air flow velocity in the air duct and regulating the flow field. Based on multi-objective optimization of pressure loss and non-uniformity coefficient, the optimal solution for the air duct structure was determined as volute shape of 2, blades numbers of 18, and blade angle of 55°. The collection efficiency of finished particles in the optimized scheme is as high as 95.3%, which is 13.5% higher than that of the original scheme. The simulation results are consistent with the field test data, which can provide a guidance for the structural optimization design of the pendulum mill.

Aiming at the problem that the convolutional neural network applied in the existing copper ore grade classification lacked the relationship between the construction and the induction of long-distance features, a copper ore grade classification method based on the integrated model of Swin Transformer and EfficientNet was proposed. This method made full use of the inductive ability of Swin Transformer V2-t architecture for long-distance feature relations and the advantage of EfficientNet V2-s in capturing subtle local features. By adding a linear layer to integrate the output results of the two models, and dynamically adjusting the weight of the linear layer according to the output of a single model itself, the mapping relationship was optimized, and the classification performance was significantly improved. The experimental results show that the accuracy rate of the fusion model in the classification task is 92.891%, the accuracy rate is 93.095%, and the recall rate is 92.654%. Compared with the single model before integration, the integrated comprehensive model has improved the classification accuracy by 1.3 percentage points, the accuracy rate by 1.9 percentage points and 2.186 percentage points respectively, and the recall rate by 0.474 percentage points and 0.237 percentage points respectively.

Coal gangue sorting is an essential process in coal mine production, and automatic identification of coal gangue is a prerequisite for intelligent sorting. Coal gangue object detection based on deep learning can make full use of image features, and the upper limit of accuracy is high, but there are problems such as large amount of model calculation and difficulty in deploying real-time sorting sites. Aiming at the above problems, a lightweight DVBS-YOLOv8s coal gangue object detection method was proposed. To enhance the inference speed of the model, DSConv and Vanillanet modules were introduced into the backbone of YOLOv8s, and Slim-neck was incorporated into the neck of the network to achieve lightweight design. Additionally, to improve the feature capture ability of the model for conveyor belt background and coal gangue objects, BoT and Slim-neck were introduced. Finally, WIoU and Soft-NMS algorithms were used to enhance the training effect. An image acquisition test platform was set up, and a dataset of 14 790 coal gangue images was self-constructed for testing. The results show that in the coal gangue object detection task, compared with YOLOv8s, the improved DVBS-YOLOv8s-WS network model pays more attention to global features, increases the receptive field, and inference speed increase by 32%. Compared to the YOLOv5 series, YOLOv7 series, SSD, Faster-RCNN, and YOLOv8s, both detection accuracy and speed are improved, providing reference for achieving high-precision real-time sorting of coal gangue based on deep learning.

Aiming at the problem of excessive arsenic content in zinc concentrate and low grade of arsenic concentrate in a zinc polymetallic mine in Inner Mongolia, the flotation process of copper, zinc and arsenic was systematically studied. The combination of lime, sodium humate and zinc sulfate was used to inhibit arsenic-sulfur minerals, ethyl xanthate and Z-200 were used as combined collectors to enhance the selectivity of copper flotation. During the zinc flotation stage, the lime and NY were used as inhibitors to inhibit arsenic, the copper sulfate was used to activate sphalerite, and the collector ZC was used to improve the recovery rate of zinc. In the arsenic flotation stage, the copper-zinc flotation tailings were further finely ground(－0.074 mm accounts for 68.10%), and the arsenic recovery rate was improved by reagents combined with sulfuric acid, sodium silicate and copper sulfate. Copper concentrate with a grade of 29.35%, recovery rate of 82.17% and arsenic content of 0.46% are obtained through preferential flotation process. The zinc concentrate grade is 46.85%, the recovery rate is 86.34% and the arsenic content is 0.52%. The arsenic concentrate contains arsenic grade of 32.20% and recovery of 87.94%. When the grinding fineness is optimized to－0.074 mm accounts for 80.13%, the dissociation degree of copper-zinc minerals is significantly improved. Compared with the original beneficiation process, the arsenic content of copper-zinc concentrate is greatly reduced. The research results provide technical reference for solving the difficult problem of beneficiation of high arsenic-zinc polymetallic ores.

The eclogite-type rutile ore resources in China are widely distributed and abundant in reserves. However, most of the rutile ore belongs to primary ore, with low grade, fine embedded particle size, complex ore properties and small difference in surface properties between the rutile and gangue, which brings many challenges to mineral selection. At present, the process of rutile beneficiation is complex, with high production costs and long recovery cycles, resulting in difficulty in effectively developing and utilizing China's rutile resources. Regarding a low-grade eclogite-type rutile ore in Qinghai Province, an experimental study on short process technology for full particle flotation recovery was carried out. The results indicate that the short process technology for full particle flotation recovery has good adaptability to this mine. Under the conditions that the grinding fineness of－0.074 mm accounts for 59.52%, screening particle size of－0.013 mm, the dosage of sodium fluosilicate 700 g/t, the dosage of lead nitrate 300 g/t and the dosage of EM-OL 600 g/t, rutile concentrate with yield of 3.21%, TiO₂grade of 61.70% and TiO₂ recovery rate of 63.14% can be obtained by the flotation process of two roughing and four cleaning to separate the screened desliming raw ore, and rutile can be fully recovered and utilized. This process realizes the short-term recovery of rutile minerals, which can simplify the beneficiation process, save production costs, and be also conducive to subsequent re-election and upgrading.

A high-sulfur and high-silicon bauxite mine in Guizhou was treated with reverse flotation for desulphurization and desilication.The effects of grinding fineness, pH adjuster, activator, collector, foaming agent and inhibitor on flotation desulphurization and desilication were analyzed, and the optimal experimental conditions for flotation desulphurization and desilication of the bauxite mine were determined. The research results show that when the grinding fineness of the raw ore is－0.074 mm accounting for 75.00%, the dosage of sodium carbonate is 1 500 g/t, the dosage of copper sulfate is 100 g/t, the dosage of mixed collector is 300 g/t, and the dosage of foaming agent is 84 g/t, the raw ore is subjected to a closed-circuit test of reverse flotation desulphurization with “one coarse, one fine and two sweeps”, and the desulphurized aluminum concentrate with a yield of 79.27%, a desulphurization rate of 85.58% and a sulfur content of 0.31% is obtained. When the grinding fineness of desulphurized aluminum concentrate is－0.074 mm accounting for 88.00%, the dosage of sodium carbonate is 600 g/t, the dosage of soluble starch is 200 g/t, and the dosage of Y28 collector is 525 g/t, the reverse flotation desilication closed-circuit test of “one coarse, one fine and two sweeps” is carried out, and the aluminum concentrate yield is 64.21%, the Al₂O₃ recovery rate is 68.46%, the Al₂O₃ grade is 59.37%, the aluminum silicon ratio is 5.83, and the sulfur content is 0.32%.

In order to explore the effect of particle-particle interaction on slurry rheology under different electrolyte solutions, quartz and calcite were used as research objects. The effects of Ca²⁺/Mg²⁺ concentration on the apparent viscosity and yield stress of single quartz/calcite slurry and mixed slurry were studied by rotational rheometer. Combined with Zeta potential, DLVO theory calculation and suspension turbidity test, the surface electrical properties, particle interaction and coagulation behavior of mineral particles under different Ca²⁺/Mg²⁺ concentrations were revealed, and the theoretical relationship between particle microscopic action and slurry rheology was established. The results show that the increase of Ca²⁺/Mg²⁺ concentration can reduce the thickness of the electric double layer on the particle surface, enhance the attraction between the particles, promote the in-phase condensation of the particles, and result in the increase of the apparent viscosity and yield stress of the single quartz and calcite slurry, and reduce the turbidity of suspension. The increase of Ca²⁺/Mg²⁺ concentration can weaken the attraction between quartz and calcite, weaken the heterogeneous condensation between them, lead to the decrease of apparent viscosity and yield stress of mixed slurry, and increase the turbidity of suspension. The yield stress of quartz + calcite mixed slurry has a logarithmic relationship with the Debye length of the solution, and has a positive correlation with the square of the surface potential difference between the two different particles.

Mastering the characteristics and functional suitability of mine water is a prerequisite for achieving resource utilization and graded utilization of mine water sources. Taking the mine water in Yushuling Coal Mine as the research object, water samples were collected from different areas for water testing, the types of mine water resource utilization were divided, an evaluation system was established, and the functional suitability of different utilization directions of mine water was comprehensively analyzed. Research results show that the chemical type of mine water is mainly , with a membership degree of 0.812 - 0.991 for Class V water. The content of hydrochemical components is high, and there are three potential utilization pathways for drinking, irrigation, and industrial water. The suitability evaluation results of various types of mine water for drinking are grade Ⅲ and Ⅳ, which are not suitable for drinking.In terms of irrigation and industrial water suitability, there are significant differences in the evaluation results of different types of mine water. The sodium adsorption ratio, Razner stability index of water accumulation in goaf and fissure water are 3 - 10.8 (mmol·L)^1/2and 6.1 - 6.5, respectively, indicating that they have little impacts on soil physical and chemical properties and are basically non-corrosive, making them suitable for irrigation and industrial water use. The saturation index of accumulated water in the burnt area is 0.8 - 1.2, exceeding 0.6, indicating a tendency towards scaling. Generally, it cannot be directly used as irrigation water, but can be used for industrial water use. The water quality of the water tank is severely polluted, with indicators such as soluble sodium percentage (78.4% - 84.5%), magnesium hazard ratio (62.9% - 72.0%), Razner stability index(7.2 - 7.7), and Larson ratio (8.0 - 12.0), indicating that this type of water poses a high risk to soil alkalinity and permeability, and has severe corrosiveness, with poor potential for functional utilization.