With the gradual depletion of shallow resources in mines, deep mining has become an important way to obtain resources. A domestic metal mine under construction with an ultra-deep well under large-load and high-speed lifting system was taken as the research object. Simulation analysis of the aerodynamic pressure distribution laws of lifting containers was carried out by comparing the dynamic changes of air flow fields under different arrangement modes of shaft lifting containers and operating conditions. The standard k-ε turbulence model in fluid mechanics was adopted, and numerical simulations of the air flow behavior in the shaft during the operation of lifting containers were conducted using Fluent software. The aerodynamic pressure distribution around the containers during the lifting process was analyzed, and the container offset was calculated, thereby revealing the aerodynamic characteristics and lateral swinging patterns of the lifting containers in the shaft. The research results show that when two lifting containers running in opposite directions intersect in the shaft, obvious aerodynamic impact force and impact torque will be generated. Moreover, the lateral aerodynamic force of two-container in the shaft is greater than that of four-container, and the piston effect of the two-container shaft layout is more pronounced than that of the four-container shaft layout. From an aerodynamic perspective, the four-container layout should be prioritized for ultra-deep well lifting systems, with the two-container layout considered as a secondary option.

Take the high-sulfur ore in Gaofeng mine as the research object to address the problems of resource waste and potential safety hazards caused by oxidation and consolidation during the mining process of deep high-sulfur ore. The consolidation characteristics of the ore was analyzed through the weight gain method from the quantitatively perspective, and the anti-consolidation test of sulfur ore was carried out. The anti-consolidation measures of sulfur ore mining were put forward according to the test results. Based on the test results, corresponding anti-consolidation measures for high-sulfur ore mining were proposed. The test results indicate that the consolidation degree of the ore increases with time, and the pressure on the ore can promote its consolidation, and the maximum weight gain rate of the ore sample in 12 d is 0.460%. The composite anti-consolidation agent prepared by sodium bicarbonate (NaHCO₃) and hydroxyethyl cellulose (HEC) can significantly inhibit the consolidation of ore samples. When the mass ratio of NaCO₃ and NEC is 1:2 and the mass fraction of solute is 2%, the anti-consolidation coefficient is 85.5% and the weight gain rate of ore sample in 27 d is 0.215%. Combined with the test results, the anti-consolidation measures such as optimizing the blasting parameters to control the particle size of ore, continuous ore drawing in stages to shorten the exposure time and spraying anti-consolidation agent were put forward. The research results can provide theoretical basis and practical guidance for safe and efficient mining of deep high-sulfur mines.

A set of intelligent mining system solutions has been proposed to address the problems of numerous small and medium-sized mines, complex ore bodies, and difficult safety control in China's underground metal mining resources. Taking the complex and fragmented ore body of Tongkeng Mine as the research background, an intelligent system based on the concept of "spatio-temporal coordination" was constructed based on segmented airfield subsequent filling method. In this system, a global wireless positioning network was adopted to achieve unmanned and less-manned operations for key processes such as rock drilling, loading, and shovel transportation, as well as cross temporal and spatial collaborative scheduling of multi-equipment. Ore flow and equipment flow were innovatively integrated in depth. In the practical application of Tongkeng Mine, the system has run stably for one year, and has built three sets of standardized intelligent mining processes, one set of multi process and multi equipment collaborative support system, one complete intelligent equipment chain, and one scheduling platform. It has safely released resources in hidden danger areas, completed over 4 000 meters of perforation, and extracted 80 000 tons of ore, significantly improving operational efficiency and safety. The research results can provide a replicable technological path for intelligent mining of complex ore bodies, and provide a reference for promoting unmanned mining operations and intelligent equipment in mines.

There are many residual pillars and top pillars under the condition of filling body enveloping in a high-grade iron mine. In order to recover high-grade iron ore resources as much as possible, a recovery scheme of residual pillar and top pillar was studied. Comparative analysis of the technical and economic indexes between the downward drift filling mining method and the sublevel filling mining method was conducted. And the sublevel filling mining method with simple mining process, high mining intensity and low production cost was selected as the best mining method for residual pillar recovery. Based on the selected mining method, numerical simulation was used to analyze the displacement, stress, and plastic zone distribution of the surrounding rocks and filling materials in the mining area under different structural parameters of mining room. Optimization research was conducted on the structural parameters of mining room under the mining of the 606^# inter pillar and the top pillar at －525 m sublevel. The simulation calculation results show that the optimal structural parameter scheme for the mining room under the recovery of 606^# inter pillar is a mining room length of 30 m without leaving a top pillar. The maximum displacement of the mining room roof in this scheme is 50.00 mm, and the volume of plastic zone of roof is the minimum, which is 1 534.55 m³. The optimal structural parameter scheme for the mining room under the recovery of top pillar is that the spans of the first and second step mining rooms are 8 m and 10 m, respectively, and the length is 25 m. The maximum displacement of the mining room roof in this scheme is 57.64 mm, and the volume of the plastic zone per unit area of the roof is small, which is 1.02 m³. There is no obvious connection in the plastic zone of the filling body inside the mining area.

Aiming at the problem of large loss and dilution in the ore drawing process of non-pillar sublevel caving method, physical ore drawing tests under different sublevel heights and end wall inclination angles were carried out, and the variation laws of the drawn-out body shape and the variation characteristics of the ore dilution rate and the recovery rate were obtained. The results indicate that when the sublevel height is constant, with the increase of the end wall inclination angle, the horizontal thickness of the drawn-out body along the route direction increases, the axial deflection angle increases, and the width and vertical height along the end wall decrease. During the process of ore drawing, and with the increase of the sublevel height, the time of waste rock release is later, the cumulative ore dilution rate is smaller. For different sublevel heights, when the end wall inclination angle is 90°, the cumulative ore dilution rate at each sublevel height is the smallest after reaching the cut-off grade. With the increase of the end wall inclination angle, the ore recovery rate at each sublevel height increases first and then decreases, and the maximum value will be obtained when the end wall inclination angle is 85°. When the end wall inclination angle are 85°, 90° and 95°, with the increase of the sublevel height, the ore recovery decreases first and then increases, and the minimum values can be obtained at the sublevel heights of 18.8 m,19.17 m and 18.73 m, respectively.

The open stope mining with subsequent filling method is adopted in the open-pit to underground mining in Sijiaying Iron Mine, and horizontal isolation pillars are required to realize the isolation of open-pit and underground mining system to ensure the safety of underground mining. In order to improve the overall stability of the mining area, based on the gravity increase method, the effects of gravity acceleration on the distribution of stress field, displacement field and plastic zone of the open-pit bottom pillar structure system were studied when the thickness of open-pit bottom boundary pillar was 45 m and the width of panel pillar was 20 m. The results show that under the initial gravity acceleration, the pillar structure system is basically in a stable state, and with the increase of the gravity acceleration, the stress concentration phenomenon of the pillar structure system continues to increase. When the gravity acceleration increases to 2 times of the initial one, the maximum principal stress reaches 56.98 MPa. The distribution range of tensile stress also expands with the increase of gravity acceleration, and the value of tensile stress finally reaches the tensile strength of rock mass. Under the initial gravity acceleration, the maximum displacement of the pillar structure system is 259 mm and the distribution range is very small and the pillar system is in a stable state. However, the displacement increases exponentially with the gravity acceleration. The pit bottom boundary pillar and panel pillar within the range of N22 to N26 exploration lines in the mining area are high-risk areas prone to instability, which is likely to cause multiple pillar chain instability and failure. It is suggested that in the process of design and mining, measures should be taken to improve the bearing capacity of the panel pillars and the filling quality of the goaf in the range of N22 to N26 exploration lines, so as to ensure the safety and reliability of the open-pit to underground mining of Sijiaying Iron Mine.

The large blocks of ore rocks with a size exceeding 1.2 m generated during the natural caving mining method of Pulang Copper Mine have caused problems of equipment blockage and increased costs. To address these problems, the block size distribution and energy distribution characteristics of secondary crushing for three typical large ore rocks under different numbers of blast holes were analyzed in detail using finite element numerical simulation. Firstly, the block size distribution data of ore rocks were obtained through image processing and field measurement, and the ore rocks were classified. Secondly, finite element software was used to simulate the evolution of blasting stress, crack propagation, and energy dissipation. The results show that the number of blast holes has a significant effect on the crushing efficiency of large ore rocks. With the increase in the number of blast holes, the stress distribution within the rock becomes more uniform, and the crushing effect is enhanced. When the number of blast holes is the same, the three typical large ore rocks are subjected to similar blasting load pressures at the same distance from the blast holes. The blasting load pressure decays rapidly with the increase of the distance from the blast holes. During multi-hole blasting, the stress value on the central line of two-hole distribution is the maximum, while that on the central line of three-hole distribution is the minimum. With the increase in the number of blast holes, the number of cracks correspondingly increases, forming longer cracks or a more extensive crack network in the ore rock. Moreover, multi-hole blasting creates directional cracks or crack zones. After multi-hole blasting, the block size distribution characteristics of the ore rock are uniform and well-crushed. The number of blast holes is directly proportional to the maximum values of kinetic energy and internal energy. Under the condition of the same number of blast holes, the internal energy of ore rocks of various sizes remains stable, and the ratio of kinetic energy to internal energy of the rock decreases with the increase of the size of large ore rocks. Therefore, in the secondary crushing construction of rocks at Pulang Copper Mine, single-hole blasting can be used for small-sized large ore rocks, while double-hole blasting is suitable for medium-sized and large-sized large ore rocks.

Reasonable mining methods and stope structural parameters are crucial for ensuring the safety and efficient production of mines. According to the characteristics of large thickness variation of a copper ore body, combined with the analysis of rock mass joints and fissures, the evaluation of rock mass integrity, the allowable exposed area of ore rock and the self-stability ability of rock mass, the mining scheme of ore body was optimized. The stress, displacement, and failure laws within the stope mined by the upward horizontal filling mining method with reserved retaining wall were analyzed by FLAC3D numerical simulation software. The results show that the maximum tensile stress during the mining process does not exceed the tensile strength of the ore body. As the mining progresses, the maximum compressive stress migrates from the lower left tip of the goaf to the roof near the upper right tip, and as the stratification increases, the compressive stress at the tip gradually decreases. Under the condition that the exposed area is determined, the smaller the short-side span is, the greater the roof subsidence is. The mining process leads to the intensification of bottom stress and deformation, and the floor heave of the filling body floor and the surrounding rock of the footwall is significant. The shear failure is mainly concentrated at the tip of the goaf and extends to the deep part of the surrounding rock. Due to the presence of the hanging retaining wall, the sidewalls effectively block the expansion of the surrounding rock failure zone into the goaf, thereby enhancing the stability of the stope.

Taking the loose and weak strata in the freezing section of the main well of a mine in Laizhou as the research object, the temperature field distribution law of the double-ring freezing hole was compared and analyzed by means of field measurement and ANSYS numerical simulation. It was found that the freezing effect of the double-ring freezing hole was not ideal. Based on the comprehensive study of the formation and development law of the freezing wall for the double-ring freezing hole, a new optimization scheme of the three-ring freezing hole was proposed, that is, the single-ring freezing hole on the inner wall of the well and the double-ring freezing hole on the outer wall were arranged, and the ANSYS finite element software was used to analyze and compare the freezing effect of the temperature field before and after optimization. The results show that the optimization scheme shortens the circulation time by more than 10 days, effectively reduces the internal temperature of the wellbore and shortens the circulation time of the freezing wall. After freezing for 60 days, the thickness of the freezing wall of the optimization scheme has reached about 4.8 m, which is much larger than that of the original scheme, and the temperature is about 4−5 ℃ lower than that of the original scheme. It shows that adding a circle of freezing holes in the well is an effective method, which can quickly reduce the temperature around the well to the lowest, and play a certain guiding role for well excavation with complicated geological conditions such as safe construction of well excavation and difficult freezing.

The rational design of stope filling body is very important to ensure the safe and efficient mining of mines. Based on the engineering background of a stope mined by large-diameter deep hole open stoping with subsequent filling mining method, the reasonable strength of filling body in stope was studied by combining theoretical calculation and numerical simulation. The stress, displacement and surface subsidence deformation law inside the filling body of the stope were explored. The results show that the maximum principal stress inside the filling body gradually increases with the progress of mining. The stress is mainly compressive stress, which is more obvious at the roof and floor. The strength change of the filling body has little effect on the maximum and minimum principal stress peaks, but the increase of the strength of the filling body reduces the stress radiation range. The increase of the strength of the filling body reduces the roof subsidence space and limits the roof deformation, and its constraint effect on the roof is better than that of the surface. Under the influence of excavation disturbance, the displacement inside the filling body presents a step-like change rule. Under the working condition that the strength of the filling body in stope is 2.5 MPa at the 10 m of top and bottom and 1.5 MPa in the middle part, the maximum tilt deformation, curvature, maximum horizontal deformation in X-direction, and maximum horizontal deformation in Y-direction of the surface monitoring line are 0.109, 0.181×10^－3,－0.024, and 0.059 mm/m, respectively. All these values meet the requirements of the safety regulations for the impact of surface subsidence on structures.

In order to solve the problems of serious accumulation of solid waste and low overall strength of backfill in a mine, tailings and waste rocks were used as filling aggregate, and polypropylene fiber was added to prepare filling material. Through orthogonal test, the effects of cement-sand ratio, waste rock gradation index and fiber content on the slump of filling slurry and the strength of backfill were investigated, and the optimum ratio of filling material was determined. The results show that the primary and secondary factors affecting the slump of filling slurry are waste rock gradation index, fiber content and cement-sand ratio, while the primary and secondary factors affecting the 28 days compressive and tensile strength of backfill are cement-sand ratio, fiber content and waste rock gradation index. The decrease of cement-sand ratio and the increase of waste rock gradation index will weaken the strength of backfill. When the fiber content is 0.6%, the strength of backfill is the largest. The fitting function of the strength of backfill and the slump under the influence of the coupling of the main factors was constructed. According to the fitting function, the three-dimensional model diagram of the influence relationship between the main factors and the performance index of backfill was established, which conformed to the relationship of Z=Ax²+By²+Cx+Dy+Exy+F. Based on the orthogonal test and the actual needs of the mine, the optimal ratio parameters of the filling material are obtained: cement-sand ratio of 1:5, waste rock gradation index of 0.8, and fiber content of 0.6%. Under this condition, the fluidity of filling slurry and the strength of backfill can meet the filling requirements of the mine, and the cost of the filling material is relatively low.

Calcium carbonate can be used as a retarder to increase the flowable time of the filling slurry. In order to study the influence of calcium carbonate on the mechanical properties of paste filling body, the single factor and orthogonal method were used to design the filling body material ratio scheme, and the filling body was prepared. The uniaxial compression test was carried out on the filling body specimen, and the acoustic emission change of the filling body during the loading process was monitored in real time. The microstructure of the crack of the specimen was observed by SEM. The influence law and weight order of each factor on the compressive strength of the filling body were studied, and the failure mechanism of the filling body was analyzed from multiple perspectives. The results show that with the increase of calcium carbonate mass fraction, the strength of filling body decreases first and then increases. After adding calcium carbonate, the influence of mass fraction of fly ash, cement and solid material on the strength of filling body at each curing age (3 d, 7 d, 28 d) is basically the same as that before adding calcium carbonate. Calcium carbonate has a weak degradation effect on the compressive strength of filling body, but it has a price advantage and can be used as a retarder for filling materials. Based on the range analysis, the weight order of the influence of each factor on the 28 d strength of the filling body is cement mass fraction>calcium carbonate mass fraction>solid mass fraction>fly ash mass fraction. According to the failure characteristics of filling body, with the increase of curing age, the failure mode of filling body changes from ductile splitting to brittle shear failure. The acoustic emission ringing count changed abruptly before and after the initial compaction of the filling body approaching the elastic deformation stage and the peak stress, which revealed the internal damage evolution process of the filling body.

The transport performance of paste filling slurry and the backfill strength are important contents of filling process. In order to explore the transport performance of paste filling slurry with coarse aggregate and the evolution law of backfill strength in a mine, the response surface method was used to carry out the fluidity test of filling slurry and strength ratio experiment of backfill. Based on the response surface test results, the effects of single factor and multiple factors interaction such as proportion of coarse aggregate, cement-sand ratio and mass concentration on the transport performance of the filling slurry and the strength of the backfill were analyzed. The results are shown as follows. Firstly, under the influence of a single factor, the sensitivity factors of backfill strength from large to small are cement-sand ratio, mass concentration and proportion of coarse aggregate; for the slump of filling slurry, the sensitivity factors from large to small are the proportion of coarse aggregate, mass concentration and cement-sand ratio. Secondly, based on the response surface method, the response regression model of the slump of filling slurry and backfill strength was established. The predicted value of the model has a good linear relationship with the test value, and the model has high fitting degree and high reliability. Thirdly, the interaction between mass concentration and cement-sand ratio has the most significant influence on backfill strength, the interaction of proportion of slump and mass concentration has the most significant effect on the slump of filling slurry. Finally, the parameters of backfill strength and slump of filling slurry has been optimized as follows, proportion of slump of 30.7%, cement-sand ratio of 1:6, and mass concentration of 82%. Under this condition, the errors between the experimental values and the predicted values of the backfill strength and the slump of filling slurry are within 5%, which verifies that the application of the response surface analysis method in optimizing backfill strength and filling slurry slump is reliable.

Pipeline transportation is a crucial component of mine filling systems, and the rheological parameters of the filling slurry are essential for pipeline transport design. Taking the unclassified tailings filling slurry of Daye Iron Mine as the research object, the rheological parameters and pipeline transportation characteristics were studied by rheological test and Fluent numerical simulation. The results indicate that when the slurry mass concentration is between 60% and 70%, the fitted relationship curve between shear stress and shear rate conforms to the change rule of Herschel-Bulkley (H-B) model. With the increase of slurry mass concentration, the yield stress and viscosity coefficient increase exponentially, and the viscosity coefficient increases significantly when the slurry mass concentration reaches 70%. When the mass concentration is 60%−65%, the slurry flow is unstable and the pressure distribution in the pipeline is uneven. When the mass concentration is between 68% and 70%, the slurry flow state is relatively stable, and the pressure is symmetrically distributed along the pipe diameter. The resistance loss of the pipeline is positively correlated with the slurry mass concentration and the slurry flow rate. When the slurry mass concentration reaches 68%, the resistance loss begins to increase significantly. It is suggested that the mass concentration of filling slurry in Daye Iron Mine should be controlled between 65% and 68%. Considering that the maximum flow velocity of slurry appears at the bottom of vertical elbow (inflection point D), where wear-resistant materials should be used. The research results can provide a theoretical basis for the design of pipeline transportation parameters of mine filling system.

Aiming at the influence of surface subsidence induced by open stoping with subsequent filling mining on surface industrial site in a phosphate mine, a three-dimensional model of top-down stoping with subsequent filling of ore body was established by FLAC3D. Numerical analysis was conducted on the deformation of overlying strata and surface subsidence in different mining stages, and the distribution laws of surface subsidence under different mining stages were obtained. As the advancement of mining, the cumulative goaf volume increases, and the maximum surface subsidence value gradually increases, showing a characteristic of larger subsidence values in the early and middle stages and smaller subsidence values in the later stage. The maximum surface subsidence value is 72.94 mm. Combining with the calculation results of displacement, inclination, and curvature near the industrial site on the surface, it can be concluded that the deformation values of the selected measuring points meet the protection level requirements of level Ⅰ and level Ⅱ, and the safety impact of designed underground mining scheme on the surface industrial site meets the regulatory requirements. The research results can provide useful references for the study of surface subsidence and ground pressure control during the top-down stage open stoping with subsequent filling process in similar large-scale mines.

The existing old landslide body is a major safety hazard in the open-pit mining field, and its stability research is of great significance to ensure the safe production of the open-pit mining field. The slope stability of an existing old landslide body in an open-pit mine was studied by using multiple geophysical methods + pattern search method. Firstly, through the seismic imaging method, supplemented by transient electromagnetic method and other geophysical methods, the sliding surface of the existing landslide body was investigated to obtain the position of the existing sliding surface. Then, the pattern search method was used to search the most dangerous sliding surface of the whole slope, and the Bishop method was used to analyze its stability. The results show that the seismic imaging method and the transient electromagnetic method can be used as an effective method to investigate the sliding surface of the existing landslide body after mutual verification. The safety factor of the existing old sliding surface is greater than 1.15, indicating that the sliding surface is relatively stable as a whole. The potential most dangerous sliding surface outside the sliding surface of the existing landslide body has been searched by the pattern search method. A total of eight unsafe sliding surfaces have been found in the P1 slope under the condition of self-weight and earthquake, which are mainly located outside the existing old landslide body, at the top of the slope and below the toe of the existing old landslide body, that is, the position where the slope is steep and varied, and the position with local outburst. A total of three unsafe sliding surfaces have been found in the P1 slope under the self-weight and rainfall conditions, which are mainly located at the lower part of the exit of the existing old sliding surface. P2 slope is basically stable. The research results can provide a reference for slope monitoring and geological disaster prevention.

Aiming at the problem of dynamic disaster induced by multi-level combined mining in deep mines, the stress evolution law and deformation behavior characteristics of surrounding rock in multi-level mining area were studied. By means of geological survey, theoretical analysis and numerical simulation, the stability of the multi-level mining area of 03-011 line in the eastern part of 2 402-2 522 meter of Xitieshan Lead-Zinc Mine was analyzed, and the deformation and failure behaviors of surrounding rock and ore body in the mining process was demonstrated. The results show that the deformation and failure of the multi-level mining area in this mine mainly occur in the thick and large stope of individual ore bodies and the middle part of the regional mining. The maximum settlement of the top pillar position in the level of the mining status is close to 4 cm, and there is no greater risk. In the subsequent mining of 2 462 m level, the displacement of 14003-3 stope is the largest, and the settlement value exceeds 6 cm. The displacement of multiple stopes on both sides is about 5 cm, and there may be some stope collapse. Combined with the spatial and temporal evolution characteristics of deformation and failure in multi-level mining, aiming at the potential ground pressure in the mining process, the mining planning optimization has been carried out for the mining design, which can provide a theoretical support for the safe and efficient mining of multi-level in the subsequent mining area.

Aiming at the requirement of high accuracy of roof deformation monitoring in mine working face under complex environment, a set of foof deformation of working face monitoring system based on LiDAR was designed, and the matching deformation monitoring data algorithm was developed, and the deformation risk was evaluated. The monitoring system employed time-lapse scanning to acquire two phase point cloud data of the mine working face. Following preprocessing steps such as angular transformation and ICP registration, high-precision triaxial tilt sensors was used to real-time calibrate the LiDAR attitude, enabling precise extraction of the point cloud morphology of roof. Subsequently, the Delaunay triangulation mesh division was applied to process the two phase point clouds, and innovatively incorporated a point-to-plane monitoring algorithm to calculate the distance between the central points of the second phase and the triangular facets of the first phase, thereby enhancing the accuracy of deformation monitoring. Furthermore, based on the DBSCAN clustering algorithm, the deformation point cloud density clustering filtering algorithm was proposed to identify and mark the deformation area, and on this basis, the risk level assessment was conducted to realize the intelligent monitoring of roof deformation. Field research results demonstrate that the deformation displacement range of 0−5 mm in April and July 2024 is in the middle risk level, which needs to be paid attention to and corresponding monitoring and management measures should be taken. The deformation displacement range of 5−10 mm is in the low risk level, which needs to be continuously monitored to prevent the risk from escalating. The monitoring system achieve a monitoring accuracy range from 1 mm to 5 mm, which is significantly improved compared to traditional monitoring methods, and can achieve visual and accurate marking of deformation areas. The research results provide strong technical support for mine safety production, further enhancing the safety and reliability of mining operations.

Monitoring and early warning of dynamic roof hazards in mines is crucial for mining production and personnel safety. Traditional multi-point displacement monitoring and early warning methods rely on a single absolute threshold based on monitoring data, resulting in poor sensitivity and reliability of the warning. In response to this issue, a new early warning method for dynamic roof hazards in mines was proposed based on multi-point displacement data and the Autoregressive Integrated Moving Average (ARIMA) model. In this method, superposition techniques were employed to enhance the signal-to-noise ratio of multi-point displacement time series data, and a graded early warning method combining enhanced windowed waveform characteristic function ratio and kurtosis was introduced, which addressed the inadequacies of traditional threshold warnings. Furthermore, a dynamic weighted parameter ARIMA model was constructed to predict displacement data, enabling the integration of monitoring and prediction data for early warning of dynamic roof hazards in mines. The theoretical testing of this method shows a false alarm rate of 13.3%, with an average error of 15.4% for secondary or higher warning characteristic values. The test results of engineering actual data of kurtosis over two periods show that the warning false alarm rates are 11.8% and 6.7%, respectively, and the average errors of the secondary or higher warning characteristic values are 15.1% and 14.4%, respectively. In summary, the proposed early warning method for dynamic roof hazards in mine has a high application value.

In order to reveal the dynamic variation law of seepage field and stability of the dump slope under long duration rainfall, a case study of the dump in the Democratic Republic of the Congo was conducted. Based on the saturated-unsaturated soil mechanics theory, Geo-Studio software was used to simulate the seepage field of groundwater and perform the limit equilibrium analysis of the dump slope after long duration rainfall. The results show that the rainfall is mainly vertical infiltration in the initial stage, and it is mainly lateral infiltration in the later stage. The initial water level rises rapidly, and gradually decreases after 2 d of rainfall, and tends to be stable after 25 d. Under the condition of short duration rainfall, the safety factor of slope fluctuates whereas under prolonged rainfall, it initially decreases and then increases, and the minimum safety factor appears about 4 d after rainfall. The impact of rainfall on the stability of dump slope has a certain degree of persistence and hysteresis, and safety prevention and control measures should be taken for the slope with poor initial stability.

In order to study the dynamic response characteristics of adjacent roadway under blasting action, taking the adjacent roadway of a mine as the engineering background, the 3DEC discrete element simulation software was used to establish a bonded block model considering the discontinuous characteristics of rock mass. The vibration velocity field, stress wave propagation and fracture evolution law of surrounding rock of adjacent roadway under blasting load were simulated, and the response characteristics of the surrounding rock under different blasting rising time was analyzed. The results show that the short rising time makes the energy of blasting wave release intensively, which leads to the high peak vibration velocity of surrounding rock particles, but the attenuation is rapid. The long rising time makes the energy distribution more uniform, but the peak vibration velocity of surrounding rock particles decreases significantly, and the cumulative damage to surrounding rock is more significant. The evolution of fracture fractal dimension presents two-stage characteristics, which increases sharply in the early stage of blasting and increases slowly in the later stage. The long rising time has a stronger promoting effect on the expansion of surrounding rock damage range. Based on the stress wave propagation and the distribution law of the peak particle velocity of the surrounding rock, the short rising time is suitable for efficient blasting scenarios, while the long rising time needs to be combined with charge control to reduce the damage of the surrounding rock. The research results provide a theoretical basis for the safety design and construction control of blasting excavation in adjacent roadways.

In order to solve the problem of foundation cracking and deformation of the ore transportation conveyor belt system in S2—S3 section, which are located in the southern open-pit of the MS Mine of Serbia Zijin Copper Doo, a survey on the current status of foundation cracks, geological investigation, and monitoring of foundation deformation was conducted. It is found that poor rock properties are the main cause of cracking and deformation in the belt corridor foundation. The self-load and operation vibration of the belt corridor, mining blasting vibration, rainfall and snowfall are also important factors that induce foundation cracking and deformation. To ensure the stability of the belt corridor foundation, a treatment scheme was proposed to reinforce the foundation using double-row micro-piles and frame structures. The monitoring results indicate that before taking reinforcement measures, the belt corridor foundation of the S2—S3 section is continuously declining, and the west slope adjacent to the belt corridor shows a trend of sliding. After implementing reinforcement measures, good reinforcement effects is achieved, and the cracking and deformation trend of the belt corridor foundation is suppressed, effectively controlling the development and expansion of cracks, improving the overall integrity and bearing capacity of the foundation, and effectively controlling the propagation and transmission of cracks. The research results can provide a reference for similar geotechnical engineering reinforcement and treatment.

The prediction of mine pit inflow and the spatio-temporal variation of groundwater seepage fields in underground mines is of great significance for mine water control engineering, which is directly related to mine safety production and environmental protection. A copper-cobalt mine in the Democratic Republic of Congo (DRC) is located in a small nappe within the Zambia-DRC metallogenic belt, featuring complex hydrogeological conditions. If the traditional drainage scheme is adopted, the predicted mine pit inflow will exceed 46 000 m³/d. Through on-site hydrogeological condition investigation, a single-layer control drainage scheme was proposed by utilizing the vertical multi-layer structure of the mining area's aquifer and the principle of control drainage. By applying the theory of groundwater flow and numerical simulation techniques, a three-dimensional groundwater numerical model was established to accurately reflect the hydrogeological characteristics of the mining area. Combined with the optimization of drainage borehole layout scheme based on water gushing points during underground development, the prediction shows that the mine pit inflow during the mining period of the 940 m level can be reduced to 25 260 m³/d. The mine has been constructed according to the optimized control drainage scheme, achieving pressure-controlled mining in the first mining section. Compared with the traditional drainage scheme, it saves approximately 2.34 million US dollars in drainage costs annually. The single-layer control drainage method can provide a reference for similar high-water mines in Africa, thereby achieving the reduction of mine pit inflow, protection of mining area water resources, and efficient pressure-controlled mining.

In order to study the effect of heat treatment on the physical and mechanical properties and strength change mechanism of clay-rich sandstone, the clay-rich sandstone specimens were heat treated at 25 ℃, 200 ℃, 400 ℃, 600 ℃ and 800 ℃, respectively, and the physical parameter measurement test and uniaxial compression test were carried out on the specimens after heat treatment.The results show that with the increase of heat treatment temperature, the mass of sandstone specimens decreases continuously, the wave velocity and elastic modulus increase first and then decrease, the porosity and peak strain decrease first and then increase. The strength of the sandstone specimens shows strengthening characteristics at 25 -600 ℃ heat treatment temperature, and shows deterioration characteristics at 600 - 800 ℃ heat treatment temperature. Finally, the strength change mechanism of clay-rich sandstone is revealed from three aspects: particle thermal expansion, mineral dehydration, mineral phase change and thermal decomposition. The research results provide a reference for the study of the thermal characteristics of clay-rich sandstone, and lay a theoretical foundation for the repair and reinforcement of underground space rock engineering under high temperature.

Weak interlayer is usually the main internal factor of rock slope instability, and rainfall is the main external factor of slope instability. For rock slope with weak interlayer, it is of great significance to study its instability mechanism and seepage characteristics to prevent slope instability. Taking the limestone open-pit slope with multiple weak interlayers in Nanshan, Chibi, Hubei Province as the research object, the instability characteristics of the slope were deeply explored by numerical simulation method, and the unsaturated seepage characteristics of the slope were studied based on the unsaturated seepage theory. Finally, the Morgenstern-price method was used to calculate the safety factor of each potential slip surface under the influence of seepage field. The results show that the weak interlayer is the main reason for the instability and failure of the slope, whether it is the overall instability and failure of the slope along the weak interlayer or the local instability and failure of the step. Under the influence of rainfall, the change of pore water pressure in the slope gradually develops from the surface to the depth, but the influence depth is limited. For the feature points with a certain distance from the slope surface, the increase of pore water pressure and volumetric water content has a “lag” phenomenon. The safety factor of each potential slip surface of the slope decreases, which decreases greatly in the early stage of rainfall, and tends to be stable in the later stage of rainfall.

Both rock slope instability and tunnel collapse disasters in cold regions are related to the frozen seepage flow in rock fractures. In order to explore the evolution mechanism of the frozen seepage flow in rock fractures at low temperature, a visualized experiment platform for frozen seepage flow in rock fractures under low temperature conditions was developed, and experiments of low temperature frozen seepage flow in rock fractures under different flow conditions were carried out. Based on the image-processing technology, the dynamic development process of liquid-ice phase transition was quantified, then the relationships between the liquid-ice phase transition and the frozen seepage flow properties were analyzed. The results show that due to the difference in heat exchange intensity between liquid phase and fracture surfaces, the liquid phase is frozen firstly in the upper and lower regions of fracture outlet in the frozen seepage flow process. With the development of frozen seepage flow, the areas of ice phase increase gradually, and the seepage channel decreases accordingly, which results in the increase of liquid phase velocity and an intensification of thermal convection intensity under constant flow conditions, so that the ablation phenomenon occurs at the convex places of the ice phase on both sides of the fracture. When the frozen seepage flow reaches a steady state, the Reynolds number of liquid phase flow is obviously greater than its initial value, and the pressure gradient is roughly linear with the flow. Furthermore, the larger flow has an obvious inhibitory effect on the freezing rate of the ice phase.

Chambishi Copper Mine is a typical high-water-inrush mine in Africa, and the hydrogeochemical study of its pit water inrush is crucial for groundwater prevention and control engineering. A total of 2 groups of surface river water samples and 16 groups of pit water inrush samples were collected. Using a combination of Piper trilinear diagrams, ionic ratio coefficients, and Phreeqc modeling, the hydrochemical characteristics and main controlling factors were analyzed, the water inrush sources of the main and western ore bodies were revealed, as well as the hydrochemical evolution mechanism. The results show that the hydrochemical type of pit water inrush is mainly SO₄−Ca⋅Mg type, and the hydrochemical composition is mainly controlled by the dissolution of gypsum and dolomite, with a small amount derived from the dissolution of feldspar. The chert dolomite aquifer, serving as the main water filling source, is continuously distributed within the regional scope. Within the mining area, groundwater flows from the western ore body area to the main ore body area. Affected by the caving mining method in the main ore body area, the chert dolomite aquifer directly recharges the ore body aquifer. The quartzite aquiclude is unevenly distributed, with thinning or absence in some sections. Under natural conditions, there is a hydraulic connection between the chert dolomite aquifer and the ore body aquifer. Combining the current situation of the mining area and the hydrochemical characteristics of water inrush, the prevention and control strategy of “blocking as the main approach and drainage as the auxiliary” has been proposed, providing a basis for underground water prevention and safety production.

The influence of graphene oxide (GO) on the properties of sulfoaluminate cement-based grouting materials was investigated. The macroscopic performance tests reveal that with the increase of GO dosage, the fluidity of the slurry decreases, but the mechanical properties of sulfoaluminate cement-based grouting materials are significantly improved. When the GO dosage is 0.01%, 0.03%, 0.05% and 0.07%, compared with the control group, the 7-day compressive strengths are increased by 4.7%, 14.1%, 13.3% and 11.7% respectively. The mechanical strength enhancement effect of the material is the most significant when the GO dosage is 0.03%, indicating that GO has a remarkable promoting effect on the hydration reaction of sulfoaluminate cement-based grouting materials, and 0.03% is the optimal dosage. XRD and SEM analyses reveal that GO optimizes the microstructure through a bridging effect. Although no new products are formed, the stability of sulfoaluminate cement-based grouting materials and their durability in underground environments are significantly enhanced. The study indicates that GO can improve the early strength and long-term durability of sulfoaluminate cement-based materials, which is of great value for engineering applications.

Aiming at the ecological degradation and damage mechanism of open-pit limestone mines in Jinan, the evaluation index system of ecological vulnerability of open-pit limestone mines in Jinan was constructed based on the "pressure-state-response" (PSR) model framework, including the ecological and environmental elements of the mining area. The engineering application analysis was carried out in Lijiawo abandoned open-pit mine in Jinan. The research shows that the four indicators of landslide prone degree, degree of surface subsidence in the mining area, vegetation coverage and soil erosion can be restored in a short time, which can be used for the acceptance of ecological restoration projects and the evaluation of short-term restoration effects in mine. The measures taken to achieve the short-term restoration effect of ecological system in mines can be regarded as the starting point of the artificial promotion of natural restoration mode. The inverse method based on the limit value of EVI can be used to quantify the degree of artificial interference, and it is considered that the vegetation coverage can achieve ideal natural restoration effect when EVI=3.75. The research puts forward the ecological restoration strategy of "ecological remodeling, regional renovation, landscape reconstruction", which can provide theoretical basis and practical guidance for the ecological restoration work in similar mining areas.

In explosive hazardous environments, transient energy still transfers to the load when the semiconductor switching element of the current-intercepting intrinsically safe power supply cuts off the overcurrent main circuit. In view of this problem, a scheme was proposed to add transient suppression components in the overcurrent protection circuit to reduce the rising rate of short-circuit current and thereby suppress transient energy. Through the analysis of the short-circuit discharge process, a mathematical model of short-circuit discharge was established. Multisim was used to simulate and analyze protection circuits with different inductance values, and the nonlinear relationship between inductance parameters and transient energy was obtained. Increasing inductance can suppress the amplitude of transient current, but excessive inductance will prolong the protection response time, leading to energy accumulation. Based on the short-circuit discharge model, an intrinsically safe power supply was designed and fabricated. The test results show that increasing inductance can effectively suppress short-circuit current and transient energy. A 47 μH inductance can reduce the short-circuit current to 0.988 A and the transient energy to 1.17 μJ, which are 87.3% and 33.9% lower than those of the inductance-free circuit, respectively. Combined with the parallel freewheeling diode, the peak transient voltage has been limited within the safe threshold, which provides a theoretical and test basis for the transient suppression design of intrinsically safe power supplies.

As an auxiliary collector, tributyl phosphate has lots of shortcomings, such as high sensitivity to process conditions, unstable foam, low recovery rate and high reagent consumption in cassiterite flotation of a concentrator in Yunnan. In order to solve these problems, a new type of auxiliary collector SN-F was developed. The adaptability of two auxiliary collectors under different slurry pH, flotation concentration, grinding fineness and aeration rate was compared and investigated through conditional and industrial experiments. The results show that SN-F performs the best when the slurry pH is 6–8 and the slurry mass concentration is 45%. The recovery effect of cassiterite adding SN-F under the condition of grinding fineness of －37 μm accounting for 72.11% is significantly better than that adding tributyl phosphate. The industrial test results show that SN-F can increase the cassiterite recovery rate by 10.69 percentage points, improve the concentrate grade by 0.63 percentage points, and reduce the reagent cost by 34.31%. The auxiliary collection capacity and foam stability are good, and the adaptability to complex slurry is stronger, providing an economical and feasible solution for the efficient separation of fine-grained cassiterite.

The main valuable elements of a copper-molybdenum ore in Pakistan are copper, molybdenum, gold and silver, with copper content of 0.55% and molybdenum content of 0.01% in the raw ore. The contents of associated gold and silver are 0.1 g/t and1.15 g/t respectively. The content of easily mudding gangue minerals such as sericite and chlorite in the ore is high, resulting in a low grade of copper concentrate. In order to solve the above problems, a new-type inhibitor CD-B was developed to inhibit mudding gangue minerals. The process flow of “rough grinding, three-stage roughing with tailings discharge, rough concentrate regrinding and recleaning” was adopted. A copper-molybdenum mixed concentrate with a copper grade of 26.76% and a molybdenum grade of 0.451% can be obtained. The copper recovery rate is 94.49% and the molybdenum recovery rate is 77.48%. The copper-molybdenum mixed concentrate contains 2.06 g/t gold and 32.66 g/t silver, with gold recovery at 40.17% and silver recovery at 55.38%, demonstrating good beneficiation indicators. The new-type inhibitor CD-B is a small-molecule organic compound, featuring zero toxicity, zero pollution, low reagent consumption, and belongs to a high-efficiency green environmental protection reagent. Compared with the conventional reagent of sodium silicate, CD-B can significantly improve the grade of copper concentrate. The consumption of CD-B is only 1/15 of that of sodium silicate, making it convenient for procurement, storage and use.

In order to solve the incompatibility between the green development capability indicators of non-ferrous metal enterprises and the lack of full consideration of subjectivity and objectivity in the process of index weighting, an evaluation model of green development capability of non-ferrous metal enterprises based on game theory combined weighting-improved matter-element extension model was proposed. Firstly, through literature analysis and consulting experts, the evaluation index system of green development capability of non-ferrous metal enterprises was constructed. Secondly, the subjective weight and objective weight of the evaluation index were calculated by the genetic algorithm-analytic hierarchy process and the CRITIC method respectively, and the comprehensive weight was calculated according to the game theory. Finally, the asymmetric closeness degree criterion was used to replace the maximum correlation degree criterion, and the evaluation model of green development capability of non-ferrous metal enterprises based on the improved matter-element extension model was established, and the model was studied by an example. The research results show that the evaluation grade is consistent with the bias degree of the previous year, which verifies that the green development capability evaluation of non-ferrous metal enterprises based on the improved matter-element extension model has certain reliability and accuracy.