To address the issues of a complex solution space, strong nonlinearity, and easy entrapment in local optima in microseismic localization, the application of swarm intelligence optimization algorithms in microseismic localization was systematically investigated. By comparing the performance of the sparrow search algorithm (SSA), differential evolution (DE) algorithm, and particle swarm optimization (PSO) algorithm in single P-wave localization, it was found that PSO algorithm exhibited the best performance in terms of solution stability. To further increase the localization accuracy and enhance algorithm robustness, S-waves were introduced to construct a dual-wave joint localization model, reducing the inversion space from four dimensions to three dimensions. The experimental results indicate that the dual-wave joint localization significantly outperformed the single P-wave localization, where PSO algorithm achieves the lowest average localization error in the dual-wave joint localization and demonstrated good adaptability. To overcome the shortcomings of the PSO algorithm, such as premature convergence and sensitivity to initial values, an adaptive improved particle swarm optimization (APSO) algorithm integrating opposition-based learning and chaotic mapping was proposed. Opposition-based learning was utilized to generate a high-quality initial population, while chaotic mapping achieved the dynamic adjustment of the inertia weight to balance the global exploration and local exploitation capabilities. The numerical experimental results indicate that the average localization error of APSO algorithm at multiple typical source points is lower than that of the standard PSO algorithm, with the average localization error reduced by approximately 62.4%, and it exhibits stronger stability and convergence efficiency. This method provides an effective tool for high-precision microseismic localization in deep rock mass engineering and has promising prospects for engineering application.

To achieve safe mining in the near-shaft area of steeply inclined thin orebody, a steeply inclined thin gold mine was taken as the engineering background. Combined with the geological conditions of the mine, the applicable mining method was preliminarily selected, and the primary selection scheme was optimized by using the analytic hierarchy process and fuzzy decision theory. FLAC3D numerical simulation was used to analyze the shaft stability during the mining process in the newly added recoverable near-shaft area. The results show that the upward drift filling mining method is the optimal mining scheme for the newly delineated area of the shaft safety pillar. Compared with ascending mining, descending mining reduces the extreme value of the maximum unbalanced force in the stope by 15%-20%, effectively alleviates the local stress concentration in the ore body, and helps restrain ground pressure activities and control shaft displacement. The optimized mining sequence gradually releases stress and creates a favorable stress environment for the stope. The filling body plays a key supporting role, and their synergistic effect significantly enhances the overall stability of the stope and effectively inhibits rock mass failure and propagation. The on-site monitoring results of the shaft indicate that after adopting the descending mining scheme, the maximum compressive deformation of the main shaft is 0.049 mm/m, and the maximum tensile deformation of the blind shaft is 0.151 mm/m, both values meet the requirements specified in safety regulations (2.0 mm/m). The field industrial test results show that the upward drift filling mining method has good safety and economic benefits, the resource recovery rate is increased to 95%, realizing the coordinated optimization of safety, economic benefit and resource recovery.

Based on the occurrence characteristics of the soft, fractured, and thick large orebodies in a gold mine, the mechanized upward drift filling mining method was recommended. Six sets of stope structural parameters were designed through an orthogonal experiment, and the FLAC3D numerical simulation software was utilized to analyse the stope stability. The simulation results indicate that the compressive and tensile stresses in the stope are mainly concentrated at the top and bottom corners of the sidewalls, and both exhibit a symmetrical distribution. The displacement values of the stope roof, floor, and hanging wall surrounding rock increase with the increase of the drift width and drift height, but the overall displacement values are relatively small. On the premise of ensuring the safety and stability of the stope, larger stope structural parameters are preferentially selected to improve the technical and economic indicators. Therefore, Scheme 3 (with both the drift width and drift height being 5 m) is determined as the optimal stope layout scheme. The field application results indicate that when the mechanized upward drift filling mining method is adopted for extraction, the stope production capacity reaches 220 t/d, and the development and cutting ratio is 24.5 m³/kt. Furthermore, the ore loss rate is 6.2%, and the dilution rate is 8.3%, achieving favourable technical and economic benefits.

In a gold mine, the uniaxial compressive strength of ultra-fine unclassified tailings backfill cannot satisfy the requirements of stope filling. To improve the backfill strength, steel fibers, polypropylene fibers and steel-polypropylene hybrid fibers were adopted as reinforcement materials for backfill, and the effects of fiber type, fiber dosage and curing age on the backfill strength were systematically analyzed. Based on the test data, a prediction model for backfill strength integrating Gaussian Process Regression (GPR) and Shapley Additive Explanations (SHAP) was proposed. The research results show that both steel fibers and polypropylene fibers can effectively enhance the backfill strength. Compared with the fiber-free blank control group, all fiber systems achieve the optimal strength enhancement at a fiber dosage of 0.3%. A typical nonlinear relationship of "promotion at low dosage and inhibition at high dosage" exists between fiber dosage and backfill strength. In addition, the established strength prediction model exhibits excellent prediction performance, and the average relative error of strength prediction is controlled within 10%. The SHAP analysis can accurately quantify the influence degree of fiber type, fiber dosage and curing age to backfill strength. The research results provide a quantitative basis for the mix proportion design of fiber-modified backfill and help promote the intelligent development of mine backfilling engineering.

To address the problems of weak research on solidification and slurry properties of classified overflow fine tailings and slow progress in underground backfill treatment of fine tailings, a small spread cylinder was used to measure the spread value to determine the paste backfill concentration. Using slag-based cementitious materials, solidification tests and slurry property studies were carried out on ultrafine tailings from middling overflow of an iron mine, classified overflow fine tailings from some gold mines, and unclassified tailings from several iron mines. The results show that the use of slag-based cementitious powder can effectively solve the problems related to the solidification and strength requirements of classified overflow ultrafine tailings. Ultrafine tailings have high porosity and low bulk density, and the slurry mass concentration can be as low as below 50%. Under the same binder dosage, the cement-tailing ratio of ultrafine tailings is relatively high, but the 28-day strength loss of the backfill is limited, and the increase in backfill cost is not significant. The slurry mass concentration has a significant effect on solidification strength. Under a filling line of 3.5 to 4.0, the ultrafine tailings backfill slurry can achieve gravity flow transportation without pumping.

To effectively solve the problems of severe backbreak damage, high oversize block ratio and poor stability of working face side when using large-diameter long-hole caving method in underground stopes, the optimization of axial charge structure was carried out based on the engineering background of three rows of large-diameter long-hole lateral blasting in underground stope of a copper mine in Anhui Province. LS-DYNA was used to establish five numerical models of axial segmented charge structures with three rows of blast holes. By analyzing the damage distribution law and failure volume of rock mass in different zones, the optimal charge structure for backbreak control in each row of blast holes were determined respectively. The numerical simulation results show that the axial charge structure can alter the distribution of explosion energy, thereby regulating the damage distribution and failure range of the rock mass. Considering both rock breaking effect and backbreak control effect comprehensively, it is recommended that the three rows of blastholes adopt axial charge structures with 1 m charge and 1 m air interval, 0.5 m charge and 1 m air interval, and 0.5 m charge and 1.5 m air interval, respectively. The field industrial test results show that after adopting the optimized charge structure, the number of oversize blocks in the collapsed muck pile decrease significantly, the reduction of sonic wave velocity in the rock mass of the imminent blasting zone is about 8.8%, and the rock damage value is basically less than 0.19. This indicates that the rock mass features favorable integrity and slight damage, and the backbreak control effect is remarkable.

To realize safe and efficient mining in a copper-nickel mine in Xinjiang, a horizontal lateral medium-deep hole blasting mining technology was proposed. Firstly, blasting parameters are preliminarily determined by the empirical analogy method: the blast hole diameter is 65 mm, the minimum resistance line is 1.625-2.275 m, the hole spacing is 1.300-2.275 m, continuous in-hole charging is adopted, the unit explosive consumption is 0.35-0.5 kg/t, and the millisecond delay initiation interval is 25 ms. Interval stemming is adopted for blast holes, with stemming lengths of 1 m and 2 m respectively. Subsequently, the LS-DYNA finite element software was used to carry out numerical simulation analysis on horizontal lateral medium-deep hole blasting. The numerical simulation results show that rock can be effectively broken by blasting and the stability of backfill roof can be guaranteed when the minimum resistance line is 1.5 m and the hole spacing is 1.0-1.5 m. Finally, field verification tests were carried out. The results show that the shape of the drift stope after blasting is relatively regular. The predicted overbreak and underbreak volume is less than 10%, No oversize blocks or fine ore are generated, and the backfill roof remains intact, achieving favorable blasting performance.

To accurately identify the key factors influencing the mining safety of underground metal mines and clarify the internal interaction logic among the influencing factors, an integrated analysis model for factors affecting mine mining safety based on Grey-DEMATEL-ISM-MICMAC was constructed. Through literature review and case analysis, 17 influencing factors were screened and identified from five dimensions: personnel, equipment, environment, management and technology. The Grey-DEMATEL method was employed to quantitatively analyze the importance of each factor. Combined with the ISM method, a multi-level hierarchical structure model was constructed to reveal the interrelationships among the factors. Finally, the MICMAC method was used to analyze the attribute characteristics of the factors. The research results indicate that the 17 influencing factors can be divided into 6 levels and 3 types. Employee training and education, safety production management system, safety hazard identification and rectification, and rationality of mining technology are the core influencing factors of mine mining safety, possessing characteristics of high driving force and low dependency. The reliability of safety protection devices, daily equipment renewal and maintenance, environmental hazard source control, working environment conditions, key technology application, and technological innovation are direct influencing factors, exhibiting high dependency characteristics. In response, combined with the actual engineering situation, targeted optimization measures are proposed, guided by the modernization and practicability of the management system, centered on the comprehensive improvement of personnel quality, based on the improvement of the working environment, and supported by the upgrading of technical equipment. The research results can provide theoretical basis and decision-making reference for enhancing the level of mine safety management.

To address the severe water inflow issue in the underground workings of the Mingniwan Phosphate Mine in Guizhou, after confirming substantial water loss from the surface river, water samples were collected from six underground water inflow points, the dolomite aquifer, and the Yangshui River for hydrochemical analysis. Six hydrochemical components, Ca²⁺, Mg²⁺, Na⁺+K⁺, , Cl^- and , were selected as discrimination indicators. The Piper trilinear diagram combined with Grey Relational Analysis (GRA) was applied to identify the sources of underground water inflow. The results show that the hydrochemical types of water samples from the Yangshui River, the 368 m grit chamber, the 490 m water sump, the 440 m goaf, the 524 m water sump, and the 520 m drift are all type, while those from the 650 m main slope surface and the dolomite aquifer are type. The water inflow source in the 440 m goaf is most likely from the Yangshui River, with the highest correlation degree of 0.94. The 650 m main slope surface shows the lowest correlation with the Yangshui River, but its correlation degree with the dolomite aquifer reaches 0.86, indicating a strong correlation, suggesting that the dolomite aquifer is a highly probable water inflow source. The Piper-GRA combined analysis method provides an effective technical solution for identifying water inflow sources and can serve as a scientific basis for subsequent formulation of water prevention and control measures.

The simulation method of drainage hole unit for finite element calculation of seepage field was summarized. On this basis, the direct simulation method of drainage unit for seepage well of tailings pond was proposed. This method establishes a solid mesh model containing optimization parameters of seepage wells all at once at the initial stage of calculation. Afterwards, the optimized elements of seepage wells were grouped in advance by setting different internal boundary conditions. During the finite element solution of the seepage field, the condensation of permeability conduction matrix only needs to be performed once. In the calculation process, the node information of each group was processed according to the first kind of known head boundary conditions. The whole process is easy to program and features fast solving speed. Compared with commonly used engineering methods such as the implicit composite element method and the improved drainage substructure method, the proposed method significantly improves the efficiency of modeling and solution through optimized parameter settings. Meanwhile, it greatly enhances computational accuracy by finely simulating the size effect of seepage wells. Finally, the calculation accuracy of the method is verified to meet the engineering requirements through a fully penetrating well example and a specific engineering case.