Search Results (18)

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

As the world’s largest resource production and consumption country, China is rich in mineral resources, but its per capita occupancy is lower than the world average. In this context, mining enterprises need to choose efficient, green, and safe mining methods. The backfilling method has been paid more and more attention in domestic mines because of its remarkable advantages in controlling ground pressure, reducing ore loss and dilution rate, and protecting the surface. Due to the high cost of land acquisition and the difficulty of tailing reservoir construction in Luohe Iron Mine, after careful consideration, the sublevel caving method is abandoned and the suitable filling mining method is explored. By using the finite element analysis software MIDAS GTS NX 2019 to simulate the stress and deformation of the stope roof and backfill under different stope structural parameters, the stope structural parameters that meet the mining technical conditions of Luohe Iron Mine can be determined, and combined with the actual mine conditions, the sublevel open stope subsequent filling method is finally adopted. The practical results show that after the transformation of the mining method, the economic and environmental benefits of the mine exceed expectations. This paper aims to show readers the latest progress in the transformation of mining methods in China and provide a useful reference for similar mining method transformation at home and abroad. Full article

The integrated optimization of stope design and underground mine production scheduling is an approach that has been shown to effectively leverage the synergies among these two underground mine planning components to generate truly optimal stope layouts and extraction sequences. The existing stochastic integrated methods, however, do not include several elements of a mining complex, such as stockpiles, due to the computational complexity and non-linearity that they might add to the optimization of the long-term mine plan. Additionally, sequential simulation methods that rely on two-point statistics and Gaussian distribution assumptions are commonly used to generate the input realizations of the mineral deposit. These methods, however, are not able to properly characterize complex spatial geometries or the high-grade connectivity of non-Gaussian and non-linear natural phenomena. The present work proposes an extension of previous developments on the integrated stope design and underground mine scheduling optimization through an expanded stochastic integer programming formulation that incorporates long-term stockpiling decisions. An application of the proposed method at an operating underground copper mine compares the cases in which the geological simulated orebody models are based on high-order and Gaussian sequential simulation methods. The extraction sequence and related final stope design are shown to be physically different. It is seen that the optimization process takes advantage of the better representation of high-grade connectivity when high-order sequential simulations are used, by targeting the areas with grades that follow the mill’s blending requirements and by making less use of the stockpiles. Overall, a 4% higher copper metal production and a resultant 6% higher net present value are observed when high-order sequential simulations are used. Full article

This study, with the engineering background of the design of a stope involving a sublevel mining method in a certain underground metal mine, explored the application of stress-solving methods based on the complex variable function theory in actual engineering. Three mathematical calculation models based on the functions of a complex variable were established. Through triangle interpolation, mapping functions of a plane with a roadway section and a plane with the stope section were determined. An improved Schwarz alternating method was adopted to study the stability of the roadway and the influence of an adjacent roadway from the perspective of the stress field. In addition, in light of the distribution characteristics of a gangue in the stope, the design parameters of a pillar were optimized, with the pillar’s optimal dimensions determined. The results showed that when the magnitudes of two far-field principal stresses in the rock mass are relatively close, the distribution of the surrounding rock stress around the roadway is more uniform, and tensile stress is less likely to occur. The excavation of a neighboring roadway exacerbates to some extent the side stress of the other roadway, especially the compressive stress concentration on the side closer to the neighboring roadway. However, when the distance between the two roadways is significantly larger than the roadway size, this effect is not pronounced. In the engineering case studied in this research, the thickness of the pillar is approximately linearly positively correlated with the safety factor of the pillar approximately linearly negatively correlated with the recovery rate. Overall, this research explored the application of the complex variable function theory in an underground mine design, demonstrating its accuracy and practicality. Full article

Most autonomous navigation systems used in underground mining vehicles such as load–haul–dump (LHD) vehicles and trucks use 2D light detection and ranging (LIDAR) sensors and 2D representations/maps of the environment. In this article, we propose the use of 3D LIDARs and existing 3D simultaneous localization and mapping (SLAM) jointly with 2D mapping methods to produce or update 2D grid maps of underground tunnels that may have significant elevation changes. Existing mapping methods that only use 2D LIDARs are shown to fail to produce accurate 2D grid maps of the environment. These maps can be used for robust localization and navigation in different mine types (e.g., sublevel stoping, block/panel caving, room and pillar), using only 2D LIDAR sensors. The proposed methodology was tested in the Werra Potash Mine located at Philippsthal, Germany, under real operational conditions. The obtained results show that the enhanced 2D map-building method produces a superior mapping performance compared with a 2D map generated without the use of the 3D LIDAR-based mapping solution. The 2D map generated enables robust 2D localization, which was tested during the operation of an autonomous LHD, performing autonomous navigation and autonomous loading over extended periods of time. Full article

In order to study the surface movement and deformation law of deep continuous mining of steeply inclined orebodies in high-stress areas, the surface movement and deformation law of deep continuous mining by caving method in the Shizishan mining area was studied based on the field fissures investigation, GPS monitoring, and large-scale geotechnical engineering numerical simulation software FLAC3D 5.0. The results show that with deep continuous mining of the orebody, surface fissures, and monitoring displacement are rapidly increasing. After the stoping of different sublevel orebodies, there will be an obvious settlement center on the surface, and the horizontal surface displacement also shows a trend of gradual increase. The results indicate that surface subsidence at the mine site is in an active development phase. The research results are of great significance to the prevention and control of surface rock movement disasters in mining areas. Full article

The sublevel open stoping with backfill method has recently been widely used in underground metal mines. The primary CPB stope is frequently affected by blasting in the secondary ore stope, leading to stope collapse and ore dilution, which has become a common problem and has received widespread attention. Numerical simulations are carried out in the present work, and a 1/4 numeral model consisting of a primary CPB stope and a secondary ore stope is built. The secondary ore stope is divided into four layers on average in the simulation model, and the incident stress induced by each blasting at the interface of the CPB and ore is simulated. The results show that the CPB stope in the range within the height of the explosive charge induced horizontal compressive stress and tensile stress induced from the explosive charge height, while the mined section under the charge height has no obvious blasting impact. The maximum incident compressive stress is close to 1.2 MPa and occurs in the area closest to the blast hole The maximum induced tensile stress occurs in the range above the charge height, which is about 0.2 MPa. The stress ratios of the four-layered lift blasts are 3.6%, 3.8%, 4.0%, and 4.8%, respectively, showing a slight cumulative effect of layered blasting. In addition, the positive correlation between incident stress and the stress ratio is studied in the present work, and the results show that the greater the incident stress is, the greater the incident ratio is. Full article

Underground mining is one of the riskiest industries. It is well established that the investigation of geomechanical parameters at the design stage of an underground mine provides the approximate rock mass characteristics, which are associated with some risks in the design. From a realistic risk assessment point of view, it is essential to classify risky design parameters as relevant to risk groups and determine a suitable weighting strategy for risk-prone elements aiming at risk assessment. Therefore, a realistic weighting procedure is an essential step in making realistic design decisions to increase the safety of mining operations and economic vitality. This study aimed to develop a realistic weighting procedure to assess and compare various geomechanical parameters that pose a risk to opening stability. In this research, sub-level stoping mining methods, which are commonly used in the Kazakhstan mining industry, were selected to test the developed weighting algorithm. In this study, the risk-prone geomechanical parameters for the chosen mining method were defined, and a weighting procedure was developed using the Fuzzy Analytic Hierarchy Process (FAHP) method. The proposed methodology was verified against available data from the Ridder–Sokolny underground mine, and the analysis results showed good agreement with actual observations in the mine. The obtained preliminary results show that FAHP is a reliable method for weighting geomechanical parameters and can be used as an input in any geomechanical risk assessment practice. Full article

In this paper, through a similar simulation experiment, the roof rock breaking situation of horizontal sublevel caving method in extremely thick and steep inclined coal seam is studied, and the response characteristics of coal and rock mass under different dynamic load strength are analyzed. The mechanical response mechanisms of different mining positions under the influence of dynamic load and the law characteristics of the surrounding rock from elastic deformation to impact failure are revealed. The study shows that with the increase in the horizontal sublevel mining depth, the roof is gradually broken from the suspended state, and the broken block fills the goaf. The thin rock strata form a granular structure, and the thick rock strata form a block splicing and occlusion structure. On the same side of the coal seam floor, the concentrated stress gradually increases, and the closer the distance from the goaf is, the more obvious the stress change in the coal body below is. With the increase in dynamic load energy, the dynamic instability and failure of the dynamic load side of the roadway occur, and the stress of the bottom coal in the intake roadway of the working face increases due to the influence of the lateral abutment pressure of the goaf, while the stress of the bottom coal in the middle of the working face and the return airway decreases due to the mining of the upper section, which reveals that the dynamic load-induced dynamic behavior requires the critical energy. Due to vibration waves, the dynamic load effect is short, and the reflective stretching is generated on the surface of the roadway, resulting in the failure of the roadway. The peak acceleration increases linearly with the increase in source energy, indicating that the stronger the dynamic load energy is, the higher the impact risk is. When the dynamic load intensity is constant, the peak acceleration decreases with the increase in propagation distance, indicating that the vibration wave has the dominant propagation direction, and there are certain directional differences in the effect of coal and rock. When the dynamic load is applied, the impact failure of the coal body has a critical displacement. When the displacement of the roadway surface is less than this critical displacement, the impact failure will not occur. Full article

The presence of loose medium backfill above the horizontal pillar will technically hinder the efficient recovery of the pillar since the improper design and preserved roof protection layer height will potentially lead to casualties and equipment damage caused by large area collapse of filled tailings as well as roof fall accidents. In this study, a safe and efficient technique for the recovery of isolated pillars under loose tailings backfill was carried out via field investigation, theoretical analysis, numerical simulation, and analytic hierarchy process using the isolated pillars in the 855 middle sublevel of Hongling Zinc-Lead Mine, Chifeng, Inner Mongolia, as a practical engineering background. Current studies have revealed that the optimal scheme for an isolated horizontal pillar recovered via the cut-and-fill stoping of a drift vertical to ore body strike involves preserving a 1.0-m roof protection layer above the crown pillar combined with a spaced mining extraction sequence. This design minimizes ore dilution and losses during the pillar extraction process under safe operation. Our research results provide theoretical and technical support for the safe and efficient recovery of isolated pillars under loose tailings backfill in similar mines. Full article

During the mining process of the near-vertical seam, there will be movement and collapse of the “roof side” rock layer and the “overlying coal seam,” as well as the emergence of the “floor side” rock layer roof which is more complicated than the inclined and gently inclined coal seams, which causes problems with slippage or overturning damage. With the increase of the inclination of the coal seam, the impact of the destruction of the immediate roof on the stope and roadway gradually becomes prominent, while the impact of the destruction of the basic roof on the stope and the roadway gradually weakens. The destruction of the immediate roof of the near-vertical coal seam will cause a large area of coal and rock mass to suddenly rush to the working face and the two lanes, resulting in rapid deformation of the roadway, overturning of equipment, overturning of personnel, and even severe rock pressure disaster accidents, all of which pose a serious threat to coal mine safety and production. It is necessary to carry out research on the mechanical response mechanism of the immediate roof of near-upright coal seams, to analyse the weighting process of steeply inclined thick coal seam sub-level mining. A four fixed support plate model and top three clamped edges simply supported plate model for roof stress distribution are established before the first weighting of the roof during the upper and lower level mining process. The bottom three clamped edges simply supported plate model and two adjacent edges clamped on the edge of a simply supported plate model are established for roof stress distribution before periodic weighting of the roof during the upper and lower level mining process. The Galerkin method is used to make an approximate solution of deflection equation under the effect of sheet normal stress, and then roof failure criterion is established based on the maximum tensile stress strength criterion and generalized Hooke law. This paper utilizes FLAC^3D finite element numerical simulation software, considering the characteristics of steeply inclined thick coal seam sub-level mining. It undertakes orthogonal numerical simulation experiment in three levels with different depths, coal seam angles, lateral pressure coefficient, and orientation of maximum horizontal principal stress, and translates roof stress of corresponding 9 simulation experiment into steeply inclined roof normal stress. We conclude that the distribution law of normal stress along dip and dip direction of a roof under the circumstance of different advancing distances and different sub-levels. The caving pace of first weight and periodical weight were counted under the effect of the roof uniform normal stress. It can better predict the weighting situation of the working face and ensure the safe, efficient, and sustainable mining of coal mines. Full article

In order to study the deformation evolution law of fault activation caused by deep mining in Shizishan Copper Mine, China, a monitoring system for fault activation slip is designed and implemented on the basis of the field investigation of footwall fault activation of the main orebody in the mining area. The displacement and stress of the fault are monitored by the multipoint displacement meter, bolt stress meter, and borehole stress meter. According to the measured results, the activation deformation laws of fault F2, fault F3, and fault F4 during deep continuous mining are analyzed in detail. The results show that, when the influence range of underground mining spreads to the fault, the increase in the additional tensile stress on the fault plane will reduce the shear strength of the fault and increase the slip of the fault. When the shear stress exceeds the shear strength of the fault plane, the shear failure of the fault plane occurs, the rock mass on both sides of the fault loses stability, and the fault becomes active; when the orebody in the deep sublevel 14 and sublevel 15 were continuously stoped, the development of the mining influence area to fault F2 leads to fault F2’s activation. When stoping the orebody in sublevel 16, fault F3 also activates. With the continuous downward mining of the deep part, the slip amount increases continuously. The fault activation sequence is from fault F2 to fault F3, and then to fault F4. Full article

This study analyzes the stability and optimizes the parameters of the bottom structure in sublevel stoping with the delayed backfilling method, improves production efficiency, and increases the ore recovery ratio under the premise of ensuring safe production. Theoretical formulas are used to calculate the stability of the pillar with the bottom structure. Numerical simulation is used to study the stability of muck slash during excavation. Finally, the optimization parameters of the bottom structure are obtained by combining a similar physical experimental model and numerical simulation. The results show that the excavation of the muck slash caused different degrees of deformation at the roof and floor of the roadway. The largest stress occurred at the roadway crossing, whereas the smallest stress was in the middle area. The excavation also caused the secondary stress concentration at the adjacent bottom structure but did not significantly impact its stability. During the mining process, the largest displacement deformation occurred at the roadway crossing, and the influence of mining disturbance on the stability of the bottom structure involves timeliness and periodicity. Considering the recovery ratio, dilution ratio, and stability, the spacing of the extracted ore drift is recommended to be 9 m. This study ensures the stability of the bottom structure in the mining process and obtained reasonable parameters of the extracted ore drift, which provides a scientific way for the mines that use sublevel stoping with the delayed backfilling method. Full article

In Canada, many mines have adopted the sublevel stoping method, such a blasthole stoping (BHS), to extract steeply deposited minerals. Sill pillars are usually kept in place in this mining method to support the weight of the overburden in underground mining. To prolong the mine’s life, sill pillars will be recovered, and sill pillar recovery could cause failures, fatality, and equipment loss in the stopes. In this paper, three sill pillar recovery schemes—SBS, SS1, and SS2—were proposed and conducted to assess the feasibility of recovering two sill pillars in a hard rock mine by developing a full-sized three-dimensional (3D) analysis model employing the finite element method (FEM). The numerical model was calibrated by comparing the model computed ground settlement with the in situ monitored ground settlement data. The rockburst tendency of the stope accesses caused by the sill pillar recovery was assessed by employing the tangential stress (Ts) criterion and burst potential index (BPI) criterion. All three proposed sill pillar recovery schemes were feasible and safe to recover the sill pillars in this hard rock mine, and the scheme SBS was the optimum one among the three schemes. Full article

The pillar stability factor (PSF) is calculated in three different mining stages for a sublevel open stoping mining project located in northern Botswana. Several three-dimensional finite element models were developed by varying the stope span. Pillar strength was estimated using the Lunder and Pakalnis equation and pillar stress was obtained from the numerical models. As mining progresses, both the first and second mining stages meet the rib pillar stability factor requirement for safe extraction. Geometrical improvements are suggested in the mining layout for the third mining stage to achieve the required PSF, which is based on international practices. Full article