Mining potash deposits at great depths is associated with increasing extraction losses. Conventional room-and-pillar mining systems that leave barrier pillars between panels prevent their subsequent recovery because of progressive stress accumulation and deterioration of excavation stability, which necessitates the development of new technological solutions. This study proposes and substantiates a mining approach for gently dipping potash seams at great depths aimed at reducing ore losses through secondary recovery of barrier pillars between panels while controlling the rock mass stress state by backfilling. The study was conducted using finite element modeling with a Mohr–Coulomb elastoplastic constitutive model. The model was calibrated against field measurements of roof subsidence for mining conditions at a depth of 1100 m. The influence of extraction thickness, chamber filling ratio, and deformation properties of backfill materials (dry fill, hydraulic fill, and cemented backfill) on the stress state of the barrier pillar was evaluated. The results show that in the absence of backfilling, stresses in the barrier pillar at the stage of ground movement stabilization exceed the geostatic stress level by more than six times, which precludes its subsequent extraction. An empirical relationship between the stress concentration factor and the properties of the backfill mass was derived, enabling the prediction of safe mining conditions. A configuration of inter-chamber pillars with variable width, increasing from the center toward the periphery, is proposed to achieve a more uniform load distribution. A method was developed for calculating the width of first-stage pillars, ensuring that the factor of safety remains above the regulatory threshold (>1). Simultaneous mining and backfilling operations limit stress buildup in the barrier pillar. This creates conditions for the safe recovery of the pillar using second-stage chambers. The proposed technology enables additional recovery of mineable reserves, does not require major modifications to the existing development layout, and allows mining waste to be used as backfill material.

This study addresses the growing recognition of zirconium raw materials as a critical mineral resource in most industrialized countries and the need for a comprehensive assessment of the complex global zirconium market. Using statistical, graphical, and analytical methods, the study examines the zirconium resource base, the spatial distribution of zirconium deposits by geological type, global commodity flows (production, imports, exports, and consumption) by country, as well as prices and future production and consumption trends. The analysis shows that global consumption of zirconium raw materials has increased rapidly, from 39 kt in 1950 to 2.191 Mt in 2024. The main demand-side trend is the sharp rise in consumption in China, driven by rapid economic growth: from 84 kt (8.7% of the global market) in 1997 to 1.83 Mt (78%) in 2024. At the same time, growth of the global zirconium raw materials market is constrained by rising demand and prices, the high share of international trade, and conflicting interests between producing countries (Australia, South Africa, Mozambique, Indonesia, and Senegal) and the major consuming countries (China, the European Union, the United States, India, and Japan). Another major challenge is that a significant share of global reserves is located in complex endogenic deposits that are difficult to develop both technologically and economically. In most industrialized countries, zirconium is classified as a critical mineral resource. Global proven reserves of zirconium raw materials in developed deposits are estimated at 95 Mt, while forecast resources amount to 232 Mt. Production is currently concentrated mainly in titanium-zirconium placer deposits; however, zirconium also occurs in complex endogenic deposits in carbonatites and alkaline igneous rocks in the form of zircon, baddeleyite, and eudialyte. Global production of zircon concentrate increased from 537 kt in 1970 to 1.64 Mt in 2024 (+2.4% per year), while cumulative global production for 1950–2024 reached 59.7 Mt. Export supply of zircon concentrate to the global market, including re-exports, increased from 395 kt in 1970 to 1.86 Mt in 2024. In the 2010s, the share of exports in global zirconium raw material production ranged from 61% to 98%. High demand led to the emergence of a new group of producers developing placer deposits in Indonesia, Mozambique, Senegal, Kazakhstan, Madagascar, Kenya, Vietnam, and Sierra Leone. Their share of global exports increased from 0.2% in 1999 to 30% in 2024. Production from placer deposits may increase significantly in Mozambique, Madagascar, and Vietnam, while new mining operations may also emerge in Namibia and Tanzania. In addition to placer deposits, projects are being considered for the development of complex endogenic deposits in which zircon concentrate would be produced as a by-product, including Strange Lake and Thor Lake (Canada), Bear Lodge (USA), Baerzhe, Bozigor, and Tudiling (China), Khalzan-Buregtei (Mongolia), and Katuginskoye, Ulug-Tanzekskoye, and Zashikhinskoye (Russia). Development of deposits representing a new technological type—eudialyte ores, which constitute complex zirconium–rare-earth raw materials, is also possible. These projects include Nechalacho (Canada), Tanbreez–Kvanefjeld (Greenland), Toongi–Dubbo (Australia), Lovozerskoye (Russia), and Saima (China).

Coal seam permeability is a key parameter controlling degassing efficiency, the intensity of methane emission, and the safety of mining operations. As permeability decreases with depth and is critically dependent on the stress-strain state, its reliable prediction requires models capable of adequately describing the interaction between sorption-induced deformation, poroelastic effects, and fracture aperture closure mechanisms. Owing to the absence of a unified approach for permeability assessment under complex stress-strain conditions, the objective of this study was to systematize and compare the principal empirical and analytical models describing this dependence. To this end, an analytical review of models accounting for sorption-elastic deformation, porosity evolution, effective stress effects, thermoelastic behavior, and cleat system parameters was conducted. Model comparison was performed through numerical simulations of permeability variation over an effective stress range of 0–50 MPa and at depths of up to 1500 m. The models incorporated parameters such as the Biot coefficient, deformation modulus, sorption compressibility, initial permeability, and geometric characteristics of fractures. The results of the parametric calculations demonstrate that, despite conceptual differences, all models exhibit a common trend of nonlinear permeability reduction with increasing effective stress. This behavior reflects the physical processes of pore space compression and fracture aperture reduction. It was established that the most intensive permeability decline occurs within the effective stress range of 5–15 MPa, corresponding to active cleat closure, whereas at depths exceeding 1000 m permeability changes tend to stabilize due to exhaustion of the deformation potential of the fracture structure. Overall, the analysis revealed differing model sensitivities to geomechanical parameters, with the influence of sorption-induced deformation being comparable to that of poroelastic effects. Model selection is shown to be condition-dependent: the Seidle (1992) model is most suitable for accounting for sorption-induced deformation, the Palmer & Mansoori (1998) model for deep coal seams with variable porosity, and the Karkashadze & Hautiev (2015) model for describing elastic deformation effects. The derived relationships can be applied to assess the natural permeability of coal seams in undisturbed rock masses.

Certain types of waste from mining and metallurgical industry have the potential to directly absorb carbon dioxide (CO₂) from the atmosphere, as some of them contain minerals capable of carbonation. The paper demonstrates that belite sludges from the Achinsk Alumina Refinery (AAR, RUSAL Achinsk JSC), a byproduct of processing nepheline ores from the Kiya-Shaltyrskoye deposit, possess this property. An investigation of the variability in the mineral composition of the sludges as a function of storage duration in the sludge field (sludge storage facility), across the duration intervals of 0–5, 5–25, 25–50 years revealed a steady decrease in the content of calcium silicates (larnite, wollastonite, merwinite) and an increase in the content of carbonates as the sludge aged from fresh to old. This study examines the factors influencing the rate of sludge carbonation including those observed under conditions typical of an operational sludge storage facility. The electron microscopy (SEM-EDS) analysis of the sludge revealed the porous structure of the silicate particles in the sludge, as well as the extent to which they had been replaced by calcite. An assessment has been conducted of this storage facility’s potential for carbon dioxide deposition through the carbonation of silicate minerals in sludge that are chemically unstable under atmospheric conditions. Based on the results of the analysis and literature data on the CO2 absorption capacity of calcium silicates under atmospheric conditions, it was concluded that larnite has the maximum absorption potential in belite sludges. Based on the conditions at the ARR’s belite sludge storage facility, the maximum volume of CO₂ that can be absorbed during complete interaction between larnite and atmospheric air has been calculated. The absorption capacity of one ton of the sludge solely due to larnite (with its content of 32.6%) is 83.3 kg of CO₂, and taking into account wollastonite and merwinite, the total potential reaches 262 kg/t. The scale and dynamics of the process of converting silicates into carbonates in the sludge storage facility will allow the volume of absorbed carbon dioxide to be taken into account in calculations of the carbon footprint of the enterprise’s end-use product, aluminum produced from nepheline ore.

The limited availability of natural core material and its unsuitability for repeated use in laboratory experiments create a need for alternative ways of producing specimens for geomechanical investigations. Against this background, 3D-printed rock replicas are attracting increasing interest. Among the available approaches, 3D-LCD printing is a readily accessible and precise stereolithographic technology based on the layer-by-layer curing of liquid photopolymers through a liquid-crystal display. The aim of this study was to experimentally evaluate the physical and mechanical properties of materials produced by this method and to compare them with those of natural rocks. To address this aim, a set of tests was performed, including microstructural analysis, nondestructive testing based on elastic-wave velocity measurements, and uniaxial compression tests. The results showed that the printed specimens were characterized by a high degree of isotropy in their elastic properties and by stable mechanical parameters under varying post-curing and storage conditions. The inclination of the printed layers relative to the loading direction was also found to have a significant effect on compressive strength and Young’s modulus: the highest strength values were obtained at an orientation angle of 60° (up to 162 MPa), whereas the lowest were recorded at 30° (up to 120 MPa). Comparison with test data for natural silicites, silicified dolomites, and opoka, a carbonate-siliceous sedimentary rock, showed comparable mechanical properties and similar deformation behavior within the elastic range (up to 20 MPa). Thus, 3D-LCD rock replicas can be used for the physical modeling of geotechnical processes and for laboratory studies of failure behavior; however, differences in failure mechanisms during plastic deformation must be taken into account.

As ventilation networks in modern mines expand, airflow distribution control becomes increasingly complicated due to three factors: insufficient control depth combined with unsynchronized operating schedules of individual mine sectors; increasingly complex aerodynamic interactions between working areas and control devices; and growing system inertia. This highlights the need for a unified approach to designing ventilation control systems for complex networks, so that the conditions under which their implementation is technically feasible and economically justified can be assessed in advance. To achieve this objective, two key tasks were addressed: identifying the appropriate spatial depth and temporal scale of ventilation control. The methodology was based on a mathematical framework for analyzing aerodynamic interactions in branched networks with numerous fans, shafts, levels, and diagonal connections. The framework includes the use of aerodynamic influence matrices, their graphical analysis, clustering, and decomposition of the network into subsystems. Dimensional analysis was also applied to estimate the characteristic times of various dynamic processes in mines. As a result, principles for designing control systems were proposed, including the selection of the control level with regard to the number of air consumers and their operating schedules, as well as matching the control time scale to the characteristic times of ventilation and mining processes. It was shown that the duration of production cycles permits ventilation control to be applied at that level, opening a promising direction for further research. It was also established that, in complex ventilation networks, control algorithms should be designed primarily to maintain the required airflow rate, whereas control based on gas concentration is less effective.

A mathematical model describing the formation of dynamic load components acting on the working units of mining machines during rock cutting is an essential component of the digital twin of a mining shearer and is used for engineering analysis, design calculations, process simulation, and machine-parameter optimization. The application of various numerical methods to cutting-force modeling, including FEM and DEM, is constrained by the need to identify a large number of parameters, typically 10 to 20, many of which are difficult to determine either analytically or experimentally. A stochastic mathematical model of the rock cutting process has been developed by representing the process as a flow of random events, namely elementary loading events and fracture events associated with the failure of a certain volume of rock mass, that is, chip formation events. The interval between elementary loading events in time or space is treated as a random variable. The closest agreement with experimental data obtained from tests of a shearer cutting a full-scale coal-cement block was achieved with a model based on a truncated exponential distribution of the interval between successive fracture events. For each elementary loading event, the maximum cutting force at which chip formation occurs is determined analytically from the known expected value of the cutting force. For the cutting force acting on an individual pick, the modeling error does not exceed 7% for the expected value and standard deviation and 15% for the maximum value. Good agreement was also confirmed between the histograms of the force distribution and the spectral density plots obtained from full-scale and computational experimental data. The proposed model contains no more than three parameters requiring identification and can be used as a component of the digital twin of a mining shearer. The same approach is also applicable to mathematical modeling of the cutting of hard soils using the working tools of earthmoving machines and to modeling the operating processes of crushing machines.

Comprehensive consideration of the geomechanical and structural characteristics of the rock mass is a prerequisite for ensuring the safety and efficiency of underground mining operations. The key parameters that must be incorporated into computational models include the physical, mechanical and strength properties of rocks, their degree of fracturing, as well as the initial and mining-induced stress state of the rock mass. This study investigates geomechanical criteria for determining support parameters of development workings under conditions of plastic deformation. The methodological framework is based on a combination of theoretical analysis, finite element numerical modeling using the RS2 software package (Rocscience), and generalization of experimental data on the properties of host rocks. This approach enabled a detailed analysis of the stress-strain state of the rock mass at various stages of mining operations, including the roadway junction where a new roadway is driven from an existing excavation and its subsequent development. The modeling results established the spatial boundaries of deformation zones: the depth of developed inelastic deformation reaches 0.6–0.7 m from the excavation boundary, while the elastoplastic deformation zone extends to 1.8–1.9 m. Stress analysis showed that in the abutment pressure zone ahead of the excavation face, stresses reach values of 20.48 MPa, which is about 25% higher than the in-situ stress level at a depth of 600 m. Analysis of the Factor of Safety (FoS) revealed local zones with FoS < 1 in the junction zone, indicating the need for reinforcement of the support system. The results provide a scientifically substantiated basis for predicting the geomechanical state of the rock mass and for selecting rational support parameters, thereby improving excavation stability, reducing maintenance requirements, and enhancing industrial safety in underground coal mining.