According to the well-developed joints and fragmented orebody in certain middle sections of a gold mine, combined with the stability of the ore and surrounding rock, two mining schemes were proposed, including the upward drift cut-and-fill mining method and the sublevel cut-and-fill mining method. Through a comprehensive comparison of the technical and economic indicators, as well as the advantages and disadvantages of each method, the upward drift cut-and-fill mining method was ultimately selected. The Mathews stability graph method was then employed to design the stope dimensions (height × width × length) for the primary and secondary drifts, which were determined to be 3.3 m×8.0 m × 40 m and 3.3 m×7.0 m×40 m, respectively. Based on the pre-support mechanism of anchor cables, a support scheme was designed. Discrete element software was employed to simulate the crack development state in the fractured zone, and an indicator for the surrounding rock cracking zone was proposed to quantitatively characterize the cracking depth, thereby analyzing the stability of the stope. The results show that after anchor cable support, no significant rockfall occurs in the roof of the two-step drift, and no large-scale interconnected crack zones or widespread surrounding rock cracking zones are observed. Considering both support effectiveness and economic requirements, an anchor cable pattern of 3.3 m×4.0 m is identified as the optimal solution. Under this scheme, the roof displacement is approximately 84 mm, with a reduction of 26.96%, the total number of cracks decreases by 37.33%; and the equivalent depth of the cracking zone is 0.64 m, with a reduction of 36.63%, compared to the condition without support. The field industrial application shows that the maximum vertical displacement difference of the roof after support is 91 mm. This solution brings approximately 7.555 million yuan in economic benefits to the mine.

The production of open-pit mines is affected by geological fluctuations, and traditional methods are difficult to effectively coordinate multi-objective ore extraction strategies. Aiming at the uncertainty of ore grade and unit mining cost during the extraction process, a multi-objective uncertainty optimization model was constructed with the objectives of minimizing unit cost, maximizing total extraction amount, and minimizing the standard deviation of grade fluctuation. The Monte Carlo method was used to simulate the random variations of geological parameters. To improve the stability and diversity of the solution set, a dynamic reference point perturbation jumping mechanism and an adaptive scaling strategy for the objective space were designed and embedded into NSGA-Ⅲ, resulting in the DPRS-NSGA-Ⅲ algorithm. A case study was conducted on a large-scale molybdenum mine. The results show that the DPRS-NSGA-Ⅲ algorithm increases the total ore extraction amount by 2.03%, reduces the unit mining cost by 2.17%, and controls the grade deviation within 0.005 percentage points. The obtained Pareto solution set is uniformly distributed in the three-dimensional objective space, without abnormal clustering or collapse. The research results demonstrate good adaptability and practical application value in addressing uncertain factors in the ore extraction strategy of open-pit mines.

To mitigate the impact of variations in cementitious material dosage and slurry mass concentration in backfill batching system on the strength prediction of backfill in the stope, the shrinkage characteristics of unclassified-tailings backfill slurry, which were strongly correlated with two factors, were selected as the research subject. Static shrinkage tests and uniaxial compression tests were conducted with different cement-tailing ratios (1∶4, 1∶8, 1∶20) and mass concentrations (70%, 72%, 74%, 76%). The random forest algorithm was selected to calculate the feature importance scores of different factors, and a compressive strength prediction model was constructed based on the shrinkage ratio. The model was validated by combining in situ measured data from an upward horizontal slice mining stope in a lead-zinc mine. The results are concluded as follows. Firstly, when the cement-tailing ratio is constant, the shrinkage ratio of the backfill slurry decreases gradually with the increase of mass concentration. When the mass concentration is constant, the slurry shrinkage ratio also exhibits a decreasing trend with the increase of cement-tailing ratio. Backfill slurry with low mass concentration and a low cement-tailing ratio is more prone to bleeding and segregation, which leads to the deterioration of backfill quality. Secondly, by using a random forest model to calculate the feature importance scores of each factor, it is identified that the cement-tailing ratio has the greatest impact on the compressive strength of the backfill material (93%), and the shrinkage ratio (5.6%) has a greater impact on the compressive strength than the mass concentration (1.4%). Thirdly, the engineering application results show that compared with the prediction method based on mass concentration, the prediction results of the uniaxial compressive strength of the backfill based on shrinkage ratio have smaller errors in most cases and have a better engineering applicability.

The safety issues posed by blasting vibration to surrounding buildings urgently need to be addressed, while traditional vibration-reduction methods rely on empirical parameters and lack statistical support based on small sample sizes. By conducting blast vibration monitoring at a certain mine, the peak vibration frequency was extracted using the fast Fourier transformation (FFT). The characteristics of the peak vibration frequency were then analyzed using a comparative approach that incorporated the Bootstrap method, Bayesian analysis, and maximum likelihood estimation. Based on this, the inter-hole delay time was adjusted to half a cycle and verified through field tests to confirm the accuracy of this method. The results show that as the distance from the explosion source increases, the vibration velocity of the particle gradually decreases, and the attenuation characteristics of vibration velocity vary with direction. The Bootstrap method does not require the assumption that the population follows a specific distribution, instead drawing samples directly from the original sample. This approach better reflects the characteristics and distribution of the samples, resulting in statistical outcomes that are more representative of reality. An 8-ms half-cycle is determined to be the optimal inter-hole delay time. Field tests show that, compared to a 20-ms inter-hole delay time, an 8-ms inter-hole delay time reduces peak vibration velocity in the X, Y, and Z directions by 14.81%, 12.68%, and 36.51%, respectively, while the peak vibration velocity in the combined velocity direction decreases by 22.55%, demonstrating a significant vibration reduction effect. The research findings provide a basis for optimizing parameters in medium-deep holes underground blasting, effectively addressing safety issues caused by blasting vibrations.

Slopes containing steeply dipping weak interlayers often fail due to the development of sliding surfaces along the weak interlayer. To address the sensitivity analysis of factors affecting the stability of such slopes, a grey relational-orthogonal test composite algorithm was proposed. The algorithm combines orthogonal test design, range analysis and grey relational method, a slope in an open-pit mine in Yunnan Province was taken as an engineering case for study. The strength reduction method was used to calculate the slope safety factor. An orthogonal test L₁₆(4⁵) with four levels and five factors was designed to analyze the influence of interlayer thickness, interlayer dip angle, distance from interlayer to slope surface, interlayer internal friction angle and interlayer cohesion on the stability of the open-pit slope. The results show that the interlayer dip angle has the most significant influence on slope stability, with its range proportion reaching 51.89%, followed by the internal friction angle of the weak interlayer, with a range proportion of 24.51%, while the other factors have weaker influences. The analysis results of the composite algorithm show a consistent trend with the traditional range analysis results, but the sensitivity ranking is clearer, verifying the reliability of this method. The study suggests that in slope design, the slope angle should be controlled to be smaller than the interlayer dip angle, and the slope stability should be enhanced by increasing the internal friction angle of the weak interlayer. By quantifying the sensitivity weights of key factors, theoretical support is provided for disaster warning, support design and emergency response.

To achieve accurate prediction of roadway deformation, a parallel prediction model that combines a Temporal Convolutional Network (TCN) and a Transformer, optimized using the Alpha evolution (AE) algorithm was proposed. This model integrated TCN and Transformer architectures through a parallel framework. The TCN utilized dilated causal convolutions to extract local temporal features, while the Transformer employed a multi-head self-attention mechanism to enhance the capture of long-term temporal patterns. Finally, the model output results through a dynamic multi-model feature fusion mechanism guided by temporal attention. At the same time, the AE algorithm was introduced to optimize hyper parameters, leveraging adaptive evolutionary paths and a two-step mechanism to enhance the model's robustness as handling complex time-series data. Experimental validation was conducted using roadway displacement monitoring data from three different monitoring points in the middle levels of an iron ore in Yunnan. The results show that when the prediction step size ranges from 6 to 48, the AE-TCN-Transformer model maintains a root mean square error (RMSE) below 0.013 2 mm, a symmetric mean absolute percentage error (SMAPE) below 0.09%, and a coefficient of determination (R²) above 0.97. These performance metrics significantly outperform those of single models such as TCN, Transformer, LSTM, and other single-model approaches. This fully validates the accuracy and reliability of the AE-TCN-Transformer model in roadway deformation prediction and provides a new approach for precise prediction of roadway deformation.

To solve the problems of difficult management of the mudstone composite roof and poor roadway support effect in a coal mine in Shanxi Province, taking a typical working face of the coal mine as the engineering background, an in-depth study was conducted on the support scheme and support effect of the mining roadway with mudstone composite roof by means of field investigation, laboratory test, theoretical analysis and numerical simulation. First, the mudstone composite roof was classified according to the engineering geological conditions. Then, through physical and mechanical tests, the physical and mechanical parameters of each rock layer of the roof were obtained, which provides a reduction basis for the rock mass parameters in the numerical simulation. The corresponding roadway support schemes were formulated, and the support effect was evaluated using numerical simulation technology. The results show that based on the occurrence characteristics of the mudstone layer, the mudstone composite roof can be divided into five types. The main components of mudstone are clay minerals and quartz, with an internal friction angle of 30.78° and a cohesion of 75.362 kPa. A differentiated support scheme is proposed, that is, bolt-mesh-cable support is adopted for type Ⅰ, Ⅲ and Ⅳ roof roadways, bolt-mesh support is adopted for type Ⅱ roof roadway, and I-beam shed support is adopted for type Ⅴ roof roadway. The numerical simulation and field measurement results show that the designed support scheme can effectively control the deformation of the surrounding rock of the roadway and ensure the stability of the roadway with mudstone composite roof. The research results can provide a reference for roadway support design under similar geological conditions.

The open stoping method is mostly used in small and medium-sized iron mines with medium stability of ore and rock. In practical application, it is easy to induce large-scale caving of roof and surrounding rocks in goaf, which seriously threatens the safety production of mines. The complex and irregular goaf group left over by the open stoping method in Zhuba Iron Mine in Nanjiang city was taken as the research object. Through surface point cloud data obtained by high-precision UAV aerial photography, combined with the mine production plan maps, a refined three-dimensional numerical model of the ore and rock was constructed, and the dynamic evolution laws of the mechanical characteristics of the surrounding rock of the goaf groups were studied. The following conclusions are obtained. The height, volume, area, and quantity of the goaf at the 1 318 m middle section are all the maximum values, which are the key factors restricting the safe mining of the lower ore body. The residual ore body at the 1 318 m middle section and the residual horizontal pillar directly facing the goafs at the 1 258 m middle section are not suitable for mining. To ensure the safe mining of the lower ore body, a systematic protection scheme for the caving hazards and impact airwave hazard of the goaf group was proposed. The engineering practice results show that approximately 188 900 tons of ore were recovered from the 1 258 m middle section, and no large-scale caving occurred in the goafs of the 1 318 m middle section during the mining period; during the mining of the central orebody of the 1 208 m middle section, caving of varying degrees occurred in the goafs of the 1 318 m and 1 258 m middle sections, but the caving hazards did not affect the production area, verifying the safety and practicality of the proposed caving hazard protection scheme.

Taking Dahongshan Iron Mine as the engineering case, the study on the treatment of hidden goafs in open-pit slopes was carried out by combining theoretical analysis with numerical calculation. According to the relationship between goaf morphology and open-pit boundary, a treatment scheme of "blasting of goafs inside the boundary and filling of goafs outside the boundary" was proposed. The safety thickness of the goaf roof was determined, the progressive failure mode of the slope above the goaf was revealed, the treatment scope of goaf filling and the compressive strength threshold of the filling body were obtained, and the stability of the open-pit slope after goaf treatment was evaluated. The results indicate that the safety thicknesses of the goaf roofs inside and outside the open-pit boundary are 20-30 m and 20-25 m, respectively. The slope containing hidden goafs has successively experienced the failure processes of tensile cracking at the slope rear edge, progressive failure of goafs, local slope collapse, and large-scale slope instability. The filling treatment range for Layer Ⅱ and Layer Ⅲ goafs is 30 m from the slope side, and the compressive strength of the filling body should reach 20 MPa. After the filling treatment of goafs outside the open-pit boundary, the shear strain of the slope rock mass is reduced and transferred to the interior of the slope, and the safety factor of the slope can meet the stability requirements.

A newly built open-pit mine has power tower foundations of Yiquanyi Line and Jinzhong DC transmission lines distributed within about 300 m on both sides, and slope excavation of the mine may cause the power tower foundations to topple, sink, or be damaged by rockfall impact. Stereographic projection, limit equilibrium method and finite element numerical simulation methods were adopted, and the disturbance range caused by open-pit slope excavation, slope stability and the impact of rockfalls at the slope top on power tower foundations were analyzed from qualitative and quantitative perspectives. The results show that when the open-pit slope is excavated to the final boundary, the slope is in a stable state. The distances from the power tower foundations adjacent to both sides of the stope to the disturbance range are 69.2-192.9 m, and the tower foundations are outside the scope of potential slope sliding surfaces corresponding to the minimum safety factor. The rockfalls at the slope top of the mining area mostly roll into the surrounding depressions. The power tower foundations on both sides of the open-pit stope are located at a high position and there are no rockfalls nearby. The mining of the new open-pit mine will not cause impacts on the power tower foundations on both sides of the stope.