In the transition process of open-pit to underground caving mining, with the deformation and fracture of the overlying surrounding rocks, its permeability increases sharply, which provides channels for heavy rainfall infiltration and easily induces disasters such as underground water inrush. In order to accurately evaluate the runoff infiltration in the subsidence area under heavy rainfall conditions, taking Shilu Iron Mine as the engineering background, a calculation method of rainfall infiltration based on the accurate division of surface catchment infiltration area was proposed. Hydrologic Engineering Center's River Analysis System (HEC-RAS) was used to delineate the catchment area outside the subsidence area. Combined with the numerical simulation results of surface subsidence deformation and surface catchment in the mining area, the infiltration areas such as caving zone, water-conducting fracture zone and runoff zone under different mining conditions were divided, and the rainfall infiltration under rainstorm conditions was calculated in different areas. The results show that the runoff infiltration in the subsidence area of the North No.1 stope increased from 182 874 m³/d under the current condition to 218 118 m³/d at the end of mining, with an increase of 19.3%. Three anti-seepage grade areas were further divided, and anti-seepage measures were proposed for different areas. Among them, the area where the caving zone and the water-conducting fracture zone were submerged by water was the first-level anti-seepage area, and the comprehensive anti-seepage measures of layered filling ground cracks, laying anti-seepage film and pumping were adopted. The research results can provide scientific basis for the design of mine drainage system and the prevention of water inrush disaster.

Aiming at the problem that the slope is easy to lose stability in the mining of ionic rare earth mines, and the traditional stability criterion method is difficult to quantify the uncertainty of dynamic multi-field coupling, a slope stability evaluation model based on game theory and cloud model was constructed. According to the four principles of uniqueness, purpose, feasibility and observability, a multi-dimensional evaluation index system covering geomechanical parameters, leaching process, slope geological conditions, hydrometeorological characteristics and regional geological characteristics was constructed. The game theory was used to optimize the subjective weight determined by the fuzzy analytic hierarchy process (FAHP) and the objective weight determined by the entropy weight method, so as to realize the collaborative weighting of subjective and objective weights. The cloud model was introduced to quantify the membership degree of slope stability, and the expectation, entropy and super entropy were used to characterize the random fuzziness of parameter values. An ionic rare earth mining area in southern Jiangxi was taken as an example for industrial application. The results show that the comprehensive stability scores of 1^#, 2^# and 3^# slopes in the mining area are 74.8 (sub-unstable), 58.9 (relatively stable) and 99.3 (unstable), respectively, which are consistent with the actual observation results and the calculation results of the improved AHP method. Compared with the traditional limit equilibrium method, the potential risks can be found in advance. It is verified that the game theory-cloud model can provide reliable discriminant results and has high-risk sensitivity. The research results provide a theoretical basis for the dynamic control of slope stability of ionic rare earth mines.

A rare earth mining area in Longnan County, Jiangxi Province, is affected by long-term mining operations and complex natural environments, and the slope stability problem is becoming increasingly prominent. Based on field investigation and data analysis, the geological and environmental conditions and slope stability characteristics of the study area were systematically evaluated. Physical and mechanical parameters of rock and soil mass were obtained through drilling, sampling and laboratory tests. Typical slope profile models were established by using Lizheng Geotechnical Software. On this basis, the circular slip method and broken-line slip method were adopted for computational analysis, and the slope stability under natural working conditions and rainstorm working conditions were simulated respectively. The calculation results show that under natural working conditions, the slope safety factor is 1.59, and the slope is in a stable state. Under rainstorm working conditions, the slope safety factor decreases to 0.913, and the slope is in an unstable state. Rainfall conditions significantly reduce the shear strength of soil mass and increase pore water pressure. Rainfall is the key factor triggering slope instability. The research results can provide a scientific basis for the formulation of slope protection measures in mining areas.

In view of the fact that Sijiaying Iron Mine does not have the condition of plane boundary expansion, it is necessary to set up the slope boundary pillar after the transition from open-pit to underground mining. The gravity increase method was adopted to study the stability of the slope boundary pillar after the transition from open-pit to underground mining. The evolution characteristics and laws of the stress field, displacement field and plastic zone of the slope boundary pillar under different gravity accelerations were obtained. The results show that under the condition of initial gravity acceleration of 1.0g, the 40 m thick slope boundary pillar can ensure the safety of underground mining and avoid instability of open-pit slope, such as collapse, caving and landslide. With the increase of gravity acceleration, the progressive instability failure of the slope boundary pillar can be divided into four stages: stable, slight deformation, serious deformation and instability failure. Slope boundary pillar is affected by the unloading of underground mining. The maximum principal stress is basically within 30 MPa, and the minimum principal stress and its distribution range increase greatly. The tensile stress value reaches more than 3 MPa, which is prone to tensile failure. The displacement field and plastic zone expansion characteristics of the slope boundary pillar are highly consistent with the tensile stress, expanding from the slope toe to the slope top on the west slope of the open pit. The research results have a certain guiding role in preventing and controlling slope instability after the transition from open-pit to underground mining.

Aiming at the problem that the impact vibration caused by roof caving threatens the stability of stope in caving mining, a physical simulation experiment was designed based on the mechanism of caving impact vibration and the similarity law of physical simulation. The relationship between the caving impact vibration velocity and the falling height of the caving body, the amount of one caving, the porosity of the cushion, the thickness of the cushion and the distance of the measuring point was systematically studied. The results show that the caving impact vibration velocity is positively correlated with the falling height of the caving body and the amount of one caving, and negatively correlated with the thickness of the cushion and the distance of the measuring point. With the increase of the porosity of the cushion, the caving impact vibration velocity decreases first and then increases. The correlation between the caving impact vibration velocity and each factor from high to low is the distance of the measuring point, the amount of one caving, the falling height of the caving body, the thickness of the cushion, and the porosity of the cushion. Based on the test results and the calculation method of blasting vibration velocity, an empirical formula for calculating the caving impact vibration velocity was constructed and applied to the west No.2 mining area of Jinchuan Longshou Mine. The maximum vibration velocities of roof caving at 1 595 m, 1 554 m and 1 494 m are 6.4 cm/s, 3.7 cm/s and 3.0 cm/s, respectively. The maximum vibration velocities measured in the field are 1.21 cm/s, 0.26 cm/s and 0.09 cm/s, respectively, which verifies the reliability of the measurement method. The research results can provide theoretical basis and practical guidance for the prevention and control of roof caving impact disasters in underground mines.

Research on slope deformation mechanism forms the foundation and key to formulating reasonable slope treatment measures. After slope cutting and load reduction combined with micro-pile treatment, a submerged slide zone in an open-pit mine in the Democratic Republic of the Congo experienced renewed deformation during subsequent deeper mining operations. To develop a treatment scheme for this submerged slide zone, field investigations and borehole surveys were conducted within the deformed area. The formation process of the sliding surface in the submerged slide zone was analyzed. Based on the investigation results, the deformation mechanism of the submerged slide zone was analyzed in detail. Ultimately, a treatment scheme centered on comprehensive slope cutting and localized reinforcement, comprising slope cutting and load reduction, large-diameter anti-slide piles, and soil nail walls, was formulated. The research results indicate that the formation of the sliding surface is predicated on geological tectonic processes. Under the disturbance induced by open-pit mining, continuous deformation occurs, and the sliding surface extends progressively downward, potentially coalescing to form an integral sliding surface in the final stage. The formation mechanism of the sliding surface in the submerged slide zone can be divided into three stages: the advancing-type sliding stage induced by unloading rebound, the retrogressive-type sliding stage triggered by excavation at the lower section, and the integral advancing-type sliding stage during later development. Based on the research findings regarding the formation mechanism of the submerged slide zone, a combined treatment measure incorporating slope cutting and load reduction, soil nail walls, and anti-slide piles was formulated. The rationality of the treatment scheme was verified through computational analysis. The research results can provide a reference for the study of treatment measures for similar soft rock slopes in open-pit mines.

To address the issue that the dust dispersion mechanism in open-pit mines under the influence of multiple factors remains unclear, this study takes the Qidashan open-pit mine as the engineering background. A numerical simulation method was employed to analyze the effects of multiple factors, including wind speed, wind direction, dust particle size and dust source location, on the dust dispersivity. Furthermore, the spatiotemporal distribution characteristics of dust under multiple dust source conditions were investigated. A three-dimensional flow field model of the Qidashan stope was constructed based on on-site meteorological data. Simulation analyses of the dust dispersivity were conducted under various working conditions. The results indicate that dust particle size is the primary factor influencing the dust dispersivity; the smaller the particle size, the higher the dispersivity. Wind speed ranks second, exerting a significant effect on dust particles larger than 30 μm. In contrast, dust source location and wind direction have relatively minor effects on the dust dispersivity. Moreover, under multiple dust source conditions, dust accumulates within the stope and disperses along the wind direction as the duration of continuous dust production extends. After dust production ceases, the dust concentration gradually decreases. The dust dispersion velocities from different source points exhibit significant differences due to variations in their locations. Affected jointly by terrain and airflow, the PM_2.5 concentration exhibits a variation pattern of "increase-decrease-increase" as the stope depth increases. Based on the research findings, relevant measures and recommendations for dust prevention in production at the Qidashan open-pit stope are proposed, with a focus on dust source control, suppression of the transmission process, and differentiated precise prevention and control.

Ground fissures accelerate land degradation and soil erosion, and induce secondary geological disasters such as landslides and collapses. To reveal the development characteristics of ground fissures under high-intensity mining conditions, relevant research was carried out with the 615 working face of a mine as the engineering background, and multiple methods including field measurement, experimental simulation, theoretical analysis and engineering application were comprehensively utilized. The results show that ground fissures are mainly divided into tensile, step and collapse types according to different deformation properties. During the advance of the working face, the propagation period of fissures within the working face is shortened and the closure degree decreases. The development degree of ground fissures is more significant than that under conventional geological and mining conditions. The number of fissures with widths of 0-100 mm, 100-500 mm and greater than 500 mm accounts for 38%, 56% and 6%, respectively. Overburden fissures develop synchronously with the breaking of thick and hard rock strata and eventually penetrate the ground surface. Based on the critical deformation values of ground fissures, the fissure angle along the strike, the fissure angle on the dip side, and the fissure angle on the rise side are predicted to be 79.8°, 74.1°, and 79.5°, respectively, with a maximum relative error of 7.1%. In addition, the calculation formula for the critical depth-thickness ratio of through fissures was proposed. The research results are of important reference significance for land remediation and geological disaster prevention in mining areas.

To investigate the fracture distribution law of rock strata under hydraulic fracturing in gently inclined hard roofs and to determine the critical initiation stress range of hydraulic fracturing, this study takes the I010206 working face of Kuangou Coal Mine as the engineering background. Based on the mechanical properties of the roof surrounding rock, similar material specimens were prepared. Using a true triaxial hydraulic fracturing experimental apparatus, physical simulation experiments under different stresses and initiation pressures were conducted under stress conditions of σ_H=12.8-13.9 MPa, σ_h=6.5-8.1 MPa, and σ_v=5 MPa, with a water injection flow rate of 40 mL/min. Furthermore, numerical simulation software was employed to analyze the three-dimensional spatial distribution characteristics of the fractures. The results show that the horizontal stress difference has a significant influence on the fracture propagation morphology. Under low stress difference conditions, fractures exhibit unilateral dominant propagation, with the scaled-up propagation area reaching 4.25 m². Under high stress difference conditions, fractures are symmetrically double-winged, with increased propagation resistance and a reduced area of 3.84 m². The number of acoustic emission events ranges from 3 698 to 4 702, and the peak amplitude ranges from 83.2 to 121.1 dB. The evolution characteristics are related to the height of fracture initiation and propagation. The initiation pressure range for hydraulic fracturing in the hard roof of this mine is determined to be 10.80-16.79 MPa, and a critical initiation pressure of 16.79 MPa is recommended as the lower engineering limit. Engineering application results demonstrate that these fracturing parameters can significantly reduce the suspended roof area and markedly lower the risk of rock bursts. The research results can provide theoretical support and engineering reference for the design of roof fracturing schemes and disaster prevention and control in deep mines.

Aiming at the failure mechanism and control problem of surrounding rock of deep buried solid coal rectangular roadway under thick loose layer and thin bedrock overburden structure, the influence of horizontal stress on the surrounding rock load of this kind of roadway was studied through engineering field investigation, numerical simulation, theoretical derivation and case application. Taking Guotun Coal Mine as the engineering background, the failure characteristics of surrounding rock of solid coal rectangular roadway under thick loose layer and thin bedrock overburden structure were revealed. According to the ground stress characteristics of deep buried roadway, the arch axis equation of balanced arch considering horizontal stress was deduced based on the balance of force and moment. The calculation method of surrounding rock load of rectangular roadway based on the improved model of balanced arch was proposed and programmed, which improved the analysis theory of balanced arch. The effectiveness of the improved balanced arch model was verified by engineering examples and two research cases, and the influencing factors of the failure height of the roadway vault and the roof load were analysed. The results show that the failure of the surrounding rock at the top of the rectangular roadway presents three obvious zones, namely, the broken caving zone, the fracture affected zone and the non-affected zone. The error between the damage range of the vault and the side calculated by the proposed theoretical method and the numerical simulation results is 7.4% and 14.0% respectively, and the error between the proposed theoretical method and the field detection results of similar roadway is 25.0% and 18.4% respectively. The comparative analysis results of the two engineering cases verify the rationality of the improved model of the balanced arch. The failure height of the roadway vault and the roof load decrease nonlinearly with the increase of the lateral pressure coefficient, and decrease linearly with the increase of the internal friction angle of the fragmentation. The research results have important guiding significance for the design of bolt/anchor cable support in this kind of roadway.