Search Results (76)

In underground mining, the stage pillar (SP) is essential for maintaining stope stability, acting as a load-bearing structure between stages. Determining the minimum safe SP thickness is critical to balancing mineral recovery and operational safety. This study investigates the design and mechanical stability of SP under substantial backfill loads, using a representative Iron ore mine as a case study. Based on the geometry of the overlying backfill and core sampling data, extreme loading conditions were identified, with the stope measuring 85 m in height, 72 m in length, and 18 m in width. A mathematical model incorporating the pressure arch effect and triangular pillar geometry was developed to estimate the backfill-induced load. Safety factors for various SP thicknesses were calculated using thin plate and elastic beam theories. Considering sequential excavation of the first- and second-step stopes, the minimum safe SP thickness was determined to be 6.0 m. This design was evaluated using FLAC^3D numerical simulations. The results reveal that during the first step, stress concentrations occurred mainly at the pillar base, with a maximum displacement of approximately 2.0 cm and peak tensile stress of 0.36 MPa—both within acceptable limits. These findings support improved pillar design for safe, efficient ore recovery in underground metal mining. Full article

Aiming at the problem of surrounding rock instability easily induced by high ground stress in the process of deep-well mining, the optimization of stope structure parameters is studied by combining numerical simulation with theoretical analysis. Firstly, the physical and mechanical properties of rock mass are fully understood using laboratory experiments. Then, six kinds of stope structure parameter schemes are preliminarily designed using the Matthews chart method. According to the geological conditions of the Ruihai Gold Mine, a large three-dimensional numerical model is established. Based on FLAC3D, the follow-filling continuous mining method is used to simulate the six schemes. By analyzing the influence and law of different stope structures on the stress, displacement, and plastic zone evolution of surrounding rock, the most effective mining strategy to balance the safety and economic benefits of the target area is determined. In the area with good rock mass quality, the optimal stope dimensions are 20 m in height, 15 m in width, and 80 m in length. In the rock mass area with fault crossing or relatively developed joint fissures, a reduced configuration of 20 m height, 10 m width, and 70 m length is recommended to enhance stability and stress management. Finally, comparative analysis of mining methods confirms that the follow-filling continuous mining method effectively mitigates ground pressure, offering a theoretical foundation for the safe and efficient extraction of deep mineral resources. Full article

A barrier pillar between the surface and underground mining sections provides a critical buffer zone in the transition from the boxcut highwall to underground sections by isolating stress fields from underground sections and preventing them from affecting the boxcut highwall slope. In this study, an empirical scaled span method and Rocscience RS2 software were used to conduct parametric studies on key parameters for designing barrier pillars and analyzing the room and pillar design for a planned underground mine on the Great Dyke, Zimbabwe. The approach included analyzing the effect of barrier pillar width, assuming a 10° dipping angle of the orebody, with room and pillar dimensions of 7 m and 6 m, respectively. The impact on boxcut slope stability and the roof of the first stope was monitored. The stability of the barrier pillar was analyzed for varying widths (6 m, 10 m, 20 m, 30 m, and 40 m) and orebody dipping angles (0°, 10°, 20°, 30°, and 40°). The effect of deteriorated rock mass conditions, represented by Geological Strength Index (GSI) values from 30 to 50, was assessed. The optimum room and pillar design was evaluated against the planned 6 m pillar sizes. This comprehensive study aims to support the integrity and longevity of the critical structures of the mining operation. Full article

In order to improve the poor prediction effect of current filling body strength design, a support vector machine (SVM) and Lib Toolbox were used to build an optimal match model or strength index of unclassified tailings filling body. Eight main factors were analyzed and screened as condition attributes, and backfill strength as a decision attribute. Next, we selected 72 groups of training samples and 6 groups of calibration samples. Our model adopts a radial basis function (RBF) as the kernel function and uses a grid search method to optimize parameters; it then tests the combination of optimal parameters by cross-validation. Results show that the mean error of regression prediction and verified predictions made by the SVM match model were 1.01%, which were more accurate than the BP neural network model’s predictions. On the premise that stope stability is ensured, the SVM match model may decrease cement consumption and the cost of backfill more effectively, and improve economic efficiency. Full article

To address the challenges associated with deep ground pressure control at the Sanshandao Gold Mine, a pre-controlled top-to-middle and deep-hole upper and lower-wall goaf subsequent filling mining method was proposed. Three distinct downward mining schemes were designed, the excavation procedure is systematically designed with 18 steps, and the temporal and spatial evolution characteristics of stress and displacement were analyzed using FLAC3D. The results revealed that stress concentration occurred during excavation steps 1–3. As excavation progressed to steps 4–9, the stress concentration area shifted primarily to the filling zones of partially excavated and filled sections. By steps 10–12, the stress concentration in these areas was alleviated. Upon completion of all excavation and filling steps, a small plastic zone was observed, accompanied by an alternating distribution of high and low stress within the backfill. Throughout the excavation process, vertical displacement ranged from 4.42 to 22.73 mm, while horizontal displacement ranged from 1.72 to 3.69 mm, indicating that vertical displacement had a more significant impact on stope stability than horizontal displacement. Furthermore, the fuzzy comprehensive evaluation method was applied to optimize the selection among the three schemes, with Scheme 2 identified as the optimal. Field industrial trials subsequently confirmed the technical rationality and practical applicability of Scheme 2 under actual mining conditions. Full article

This study investigates the stability risks associated with a substandard-thickness (42 m) backfill isolation layer in the open-underground coordinated mining system of the Yongping Copper Mine’s eastern panel at the −150 m level. A numerical simulation based on FLAC3D 3.00 was conducted to evaluate the impacts of four mining sequences (south-to-north, north-to-south, center-to-flank, and flank-to-center) on stress redistribution and displacement evolution. A three-dimensional geomechanical model incorporating lithological parameters was established, with 23 monitoring points tracking stress and displacement dynamics. Results indicate that the mining sequence significantly influences the stability of both the isolation layer and the slope. No abrupt displacement occurred during mining, with incremental isolation layer settlement controlled within 3 mm. Post-mining maximum displacement increased to 10–12 mm. The “north-to-south” sequence emerged as the theoretically optimal solution, reducing cumulative displacements in pillars and stopes by 9.1% and 7.8%, respectively, compared to the suboptimal scheme. However, considering the engineering continuity of the existing “south-to-north” sequence at the −100 m level, maintaining consistent directional mining at the −150 m level is recommended to ensure synergistic disturbance control, ventilation system stability, and operational management coherence. Full article

Unplanned dilution is a critical challenge in underground mining, directly affecting operating costs, resource recovery, stope stability and operational safety. This study presents an empirical–statistical framework that integrates the Mathews–Potvin stability graph, the Equivalent Linear Overbreak/Slough (ELOS) metric, and a site-specific linear calibration to improve dilution prediction in sublevel stoping operations. A database of more than 65 stopes from a Brazilian underground zinc mine was analyzed and classified as cable-bolted, non-cable-bolted, or self-supported. Planned dilution derived from the Potvin graph was compared with actual ELOS measured by cavity-monitoring surveys. Results show a strong correlation between cable-bolted/supported stopes (r = 0.918), whereas non-cabled/unsupported and self-supported stopes display lower correlations (r = 0.755 and 0.767). Applying a site-specific linear calibration lowered the mean absolute dilution error from 0.126 m to 0.101 m (≈20%), with the largest improvement (≈29%) occurring in self-supported stopes where the unadjusted graph is least reliable. Because the equation can be embedded in routine stability calculations, mines can obtain more realistic forecasts without abandoning established empirical workflows. Beyond geotechnical accuracy, the calibrated forecasts improve grade-control decisions, reduce unnecessary waste haulage, and extend resource life—thereby enhancing both the efficiency and the accessibility of mineral resources. This research delivers the first Brazilian case study that couples Potvin analysis with ELOS back-analysis to generate an operational calibration tool, offering a practical pathway for other sites to refine dilution estimates while retaining the simplicity of empirical design. Full article

To address the stability control challenges of the “support-surrounding rock” system in fully mechanized top-coal caving faces within steeply dipping coal seams, this study employs an integrated approach combining theoretical analysis and numerical simulations, revealing the three-dimensional boundary morphology of the top-coal limit equilibrium zone and establishing a quantitative framework for boundary delineation. The results show that the boundary spatial morphology of the limit equilibrium zone in the fully mechanized caving stope in steeply dipping coal seams is an “asymmetric arc-shaped ribbon-like curved surface”. Along the inclined direction of the working face, the boundary distribution presents an “asymmetric circular-arc arch”, with the vault located in the middle-upper part of the working face. Along the strike direction of the working face, the distance from the boundary to the longwall face shows a gradually increasing pattern from top to bottom. Upon comparing the results from the numerical simulation, theoretical calculation, and field monitoring, a consistent overall pattern emerges. This consistency validates the rationality of the analytical representation of the boundary of the top-coal limit equilibrium zone. The research findings hold significant importance in predicting the stability of the “support-surrounding rock” system and the top-coal cavability. They can offer a scientific foundation for guiding the stability control practices of the support–surrounding rock within this type of mining stope. Full article

Dilution in underground mining refers to the unintended incorporation of waste material into the ore, reducing ore grade, revenue, and operational safety. Unplanned dilution, specifically, occurs due to overbreak caused by blasting inefficiencies or poor rock stability. To mitigate this issue, various factors related to rock quality, stope geometry, drilling, and blasting must be carefully considered. This study introduces a statistically rigorous methodology for the prediction of dilution in underground mining operations. The proposed framework incorporates a 10-fold cross-validation procedure with 30 repetitions, along with the application of nonparametric statistical tests. A total of eight supervised machine learning algorithms were investigated, with their hyperparameters systematically optimized using two distinct genetic algorithm (GA) strategies evaluated under varying population sizes. The models include support vector machines, neural networks, and tree-based approaches. Using a dataset of 120 samples, the results indicate that the GA-ANN model outperforms other approaches, achieving $MAE$, $R^2$, and $RMSE$ values of 0.2986, 0.8457, and 0.3928 for the training dataset, and 0.1882, 0.9508, and 0.2283 for the testing dataset, respectively. Furthermore, four stacking models were constructed by aggregating the top-performing base learners, giving rise to ensemble metamodels applied, for the first time, to the task of dilution prediction in underground mining. Full article

To explore the influence of the arching effect on stress distribution in jointed backfill structures, this study employs three-dimensional numerical modeling to systematically analyze the mechanical behavior of backfill materials. A finite-difference approach was adopted to establish a representative stope model incorporating interface elements to simulate rock–backfill interactions. The methodology involved parametric studies examining key material properties (internal friction angle, cohesion, elastic modulus, Poisson’s ratio) and geometric configurations, with boundary conditions derived from typical mining scenarios. The results demonstrate that stress distribution follows nonlinear relationships with all investigated parameters. Increasing the internal friction angle and the cohesion reduce internal stresses, though the arch effect exhibits a distinct upper limit. Mechanical properties significantly influence stress transfer characteristics, with the elastic modulus governing stiffness response and the Poisson’s ratio affecting lateral stress development. Geometric parameters control the spatial extent of arching, with larger dimensions modifying the stress redistribution pattern. This research quantitatively establishes the operational limits of arching in backfill structures, providing crucial thresholds to prevent stability risks from overestimating its benefits. The findings offer practical guidelines for optimizing backfill design in deep mining and paste filling applications, contributing both technical solutions for mine safety and fundamental insights for geomechanical theory. The developed methodology serves as a robust framework for future studies on complex backfill behavior under various loading conditions. Full article

This study proposes a mechanical model for evaluating the stability of thickened structural walls (TSWs) under complex local loading conditions. The model allows for the calculation of stress distribution and yielding status of TSWs based on the Drucker–Prager (D–P) yield criterion. Compared with two existing theoretical models, the proposed model improves calculation accuracy by approximately 5% and 53%, respectively. The analysis results indicate that the maximum principal stress of TSWs primarily occurs at the midpoint of the left boundary (0, h/2), the center of lateral loading on the bottom boundary (LP, 0), or the center of lateral loading (LP, h/2). As the lateral load position (LP) and width (LW) increase, both the maximum principal stress and the yielding area increase. Increasing the sidewall thickness (ST) and length (SL), while reducing the sidewall height (SH), significantly enhances the overall stability of TSWs. To meet residual ore recovery requirements, it is recommended to increase SL and reduce LP, LW, and SH. In the residual ore recovery project of the Jubankeng tungsten mine, the critical thicknesses of four TSWs were calculated using the proposed mechanical model, yielding values of 4.6 m, 4.2 m, 2.6 m, and 11.9 m. Based on field validation conducted in stopes V3412 and V3301, the discrepancy in maximum principal stress (MPS) between the mechanical model and numerical simulations was within 4% for both cases, further confirming the accuracy and applicability of the proposed model in engineering practice. Full article

Predicting and analyzing the stability of underground stopes is critical for ensuring worker safety, reducing dilution, and maintaining operational efficiency in mining. Traditional stability graphs are widely used but often criticized for oversimplifying the stability phenomenon and relying on subjective classifications. Additionally, the imbalanced nature of stope stability datasets poses challenges for traditional machine learning and statistical models, which often bias predictions toward the majority class. This study proposes a novel methodology for developing site-specific stability graphs using probabilistic modeling and machine learning techniques, addressing the limitations of traditional graphs and the challenges of imbalanced datasets. The approach includes rebalancing of the dataset using the Synthetic Minority Over-Sampling Technique (SMOTE) and feature selection using permutation importance to identify key features that impact instability, using those to construct a bi-dimensional stability graph that provides both improved performance and interpretability. The results indicate that the proposed graph outperforms traditional stability graphs, particularly in identifying unstable stopes, even under highly imbalanced data conditions, highlighting the importance of operational and geometric variables in stope stability, providing actionable insights for mine planners. Conclusively, this study demonstrates the potential for integrating modern probabilistic techniques into mining geotechnics, paving the way for more accurate and adaptive stability assessment tools. Future work includes extending the methodology to multi-mine datasets and exploring dynamic stability graph frameworks. Full article

The width of the pillar is an important factor in the stability of the underground space and the efficiency of resource recovery. This study aims to model the performance of retained walls in panel barrier pillar stopes. By simplifying the three-dimensional problem based on the mining operation, a two-dimensional mechanical model of non-equal-width retained walls was established, and the stress and deflection were solved analytically. The calculated deformation characteristics of equal-width and non-equal-width retained walls were analyzed and compared with numerical simulations. The results indicated that the deformation of retained walls is mainly influenced by the roof loads, the uniaxial compressive strength, and the internal friction angle of backfill materials. For equal-width retained wall design, corresponding to the areas of pillar stopes where the uniaxial compressive strength and internal friction angle of backfill materials are low, great lateral pressure will be created on the retained walls. This results in significant flexural wall deformations in this area, increasing the risk of wall collapses. In comparison, for non-equal-width retained walls, the width is defined based on the surrounding backfill materials, which could greatly reduce the risk of potential damage. For the mining operation at the actual mine, the non-equal-width design with 2.5 m and 4.0 m intervals was adopted for the panel barrier pillar stopes, and the final displacement of the roof of the stope after the completion of the mining is 34 mm, and the two sides of the mine wall remain in good integrity with no significant peeling or cracking identified. This design improves the recovery rate of mineral resources and the stability of mining. Full article

A comprehensive understanding of the mechanical behavior of backfill under compression is crucial for optimizing its design, improving stope stability and enhancing resource recovery. Laboratory testing and numerical simulation were conducted to study the mechanical properties and damage mechanism of cemented tailings backfill (CTB) with different cement-to-tailings (c/t) ratios under uniaxial compression. Laboratory testing was used to investigate the strength and deformation characteristics, macroscopic failure modes, and energy evolution patterns of CTB, while simulation with Particle Flow Code (PFC) was employed to explore the distribution of microcracks and mesoscopic damage mechanisms. A constitutive model accounting for the initial compaction stage was proposed, validated, and applied to practical engineering. The results show that as the c/t ratio decreases, the failure mode of CTB transforms from shear failure to combined tensile–shear failure, and tensile failure. Mesoscopically, a higher c/t ratio leads to more bond contacts, which increases the bearing capacity and consequently causes more cracks to damage CTB. From an energy standpoint, the damage mechanism of CTB is further analyzed and the development of energy is characterized by four stages. Moreover, to explore the failure mechanism of CTB, an innovative constitutive model was proposed and verified through experiments. The matching coefficients, based on the novel constitutive model, indicate that CTB with a c/t ratio of 1:6 is qualified for all current mining depths, and a c/t ratio of 1:10 is sufficient to depths below 300 m. Full article

Stope structural parameters are of great significance for the safe production of mines. To efficiently and safely mine steeply dipping ultra-thick ore bodies, the K. Kegel strength design formula and limit analysis method were used to calculate a reasonable range of stope parameters. Considering the actual mining conditions, the mechanical responses under different structural parameters were obtained through numerical simulations based on a central composite experimental design. A regression model for maximum tensile stress, maximum compressive stress, and maximum vertical displacement was established using the second-order response surface method. The regression model was then used as the objective function, and multi-objective optimization was performed using a simulated annealing algorithm to obtain the Pareto optimal solution set. Based on practical engineering needs, a stope span of 15.0 m, a pillar width of 10.0 m, and a roof thickness of 11.9 m were determined as the optimal structural parameters, achieving a balance between safety and economic efficiency. Full article