Search Results (49)

The heading face is one of the zones most severely affected by dust pollution in underground coal mines, and dust control becomes even more challenging during roadway excavation with continuous miners. To improve dust mitigation in environments characterized by intense dust generation, high ventilation demand, and large cross-sectional areas, this study integrates numerical simulations, laboratory experiments, and field tests to investigate the physicochemical properties of dust, airflow distribution, dust migration behavior, and a comprehensive dust control strategy combining airflow regulation, foam suppression, and dust extraction fan systems. The results show that dust dispersion patterns differ markedly between the left-side advancement and right-side advancement of the roadway; however, the wind return side of the continuous miner consistently exhibits the highest dust concentrations. The most effective purification of dust-laden airflow is achieved when the dust extraction fan delivers an airflow rate of 500 m³/min and is positioned behind the continuous miner on the return side. After optimization of foam flow rate and coverage based on the cutting head structure of the continuous miner, the dust suppression efficiency reached 78%. With coordinated optimization and on-site implementation of wall-mounted ducted airflow control, foam suppression, and dust extraction fan systems, the total dust reduction rate at the heading face reached 95.2%. These measures substantially enhance dust control effectiveness, improving mine safety and protecting worker health. The resulting reduction in dust concentration also improves visibility for underground intelligent equipment and provides practical guidance for industrial application. Full article

To enhance the understanding of dust diffusion laws in tunnel blasting operations of metal mines and determine optimal ventilation dust removal times, a scaled physical model of a metal mine tunneling face under the China Zijin Mining Group was established based on field measurements. Numerical simulation was employed to investigate airflow movement and dust migration in the tunneling roadway, and the fundamental features of airflow field and dust diffusion laws after tunnel blasting operations in the fully mechanized excavation face were revealed. The effects of three main factors included airflow rate (Q), ventilation distance (S), and tunnel length (L) on the dust removal time after tunnel blasting operations were investigated based on the orthogonal design method. Results indicated that reducing the dust concentration in the roadway to 10 mg/m³ required 53 min. The primary factors influencing dust removal time, in order of significance, were determined to be L, Q, and S. The lowest dust concentration occurs when the ventilation distance was 25 m. A predictive model for dust removal time after tunnel blasting operations was developed, establishing the relationship between dust removal time and the three factors as T = 20.7Q^−0.73S^0.19L^0.86. Subsequent on-site validation confirmed the high accuracy of the predictive model, demonstrating its efficacy for practical applications. This study contributes a novel integration of orthogonal experimental design and validated CFD modeling to predict ventilation dust removal time, offering a practical and theoretically grounded approach for tunnel ventilation optimization. Full article

The temperature of a surrounding rock mass decreases continuously due to the ventilation in its roadway, and the range of the rock mass with the temperature decreasing is called a Heat-Regulating Ring. Considering the steady-state temperature field, a steady-state heat conduction model of the Heat-Regulating Ring is established, and a formula of the radius and temperature of the Heat-Regulating Ring is obtained. It is found that the radius of the Heat-Regulating Ring is related to the thermal conductivity of the rock, the surface heat transfer coefficient of the tunnel, the radius of the ventilation tunnel, the original rock temperature, the rock wall temperature, and the air temperature. As assessed through field experiments and numerical simulation experiments, the error between the theoretical values and the simulation-derived values for the heat conduction model is very small, and the theoretical formula has a universal applicability. After long-term ventilation, the section shape and the radius of the ventilation tunnel have little effect on the Heat-Regulating Ring’s radius. The wind speed increases from 1 m/s to 5 m/s, and the radius of the Heat-Regulating Ring increases from 26.9 m to 28.4 m. When the ventilation wind speed reaches a certain value, although the wind speed is still increasing, the temperature value of the Heat-Regulating Ring is basically unchanged, or the change amplitude is very small. When the wind speed is 5 m/s, after 1800 days of ventilation, the radius of the Heat-Regulating Ring along the roadway is 27.9 m to 28.4 m. Full article

To address ventilation challenges in the working face of metal mine excavation, an equal-scale physical model was established with a mine section as the test site, combined with field-measured data and relevant parameters of spent air reuse equipment. Numerical simulations were carried out using Fluent 2020 R2 software to analyse the characteristics of the airflow field, temperature field, and dust distribution in the excavation roadway. The results show that when the cold air outlet temperature (T₀) is 22 °C, the temperature within the cooling zone does not exceed 26.3 °C, thereby demonstrating effective cooling. The equipment parameters significantly impacted cooling and dust removal. When the distance from the cold air outlet to the heading face was set to Z_m = 8 m, the air outlet temperature was T₀ = 22 °C, and the ventilation circulation rate was F = 40%, the working area achieved better cooling and dust removal effects. On-site application showed that within 15 m of the working face, temperatures dropped by 3–3.5 °C, reaching a low of 25.1 °C. The relative humidity at a point 1 m away from the working face decreased from 90.6% to 70.2%, and the average dust removal efficiency was 44.9%, which significantly improved the comfort and safety of the working environment at the heading face. Full article

Gas accumulation triggered by coal and gas outbursts is the core cause of secondary disasters in coal mines. This study focuses on the risk assessment of gas accumulation during disaster scenarios, proposing a multidimensional evaluation method integrating the analytic hierarchy process (AHP), information entropy theory, kernel density estimation, and dynamic risk propagation modeling. A unified intelligent prevention system encompassing “monitoring–prediction–decision making” is established. Leveraging the TFIM3D simulation platform and case studies from the Qunli Coal Mine accident, this research reveals spatiotemporal evolution patterns of gas concentration and explosion risk thresholds. A ventilation optimization strategy based on risk classification is proposed. The results demonstrate that the dynamic risk index (DRI), derived from the coupling of the roadway air volume stability coefficient and gas concentration information entropy, can accurately identify high-risk zones. The findings provide theoretical foundations and practical pathways for dynamic risk management in ventilation systems during coal and gas outburst disasters. Full article

The objective of this study was to understand the characteristics of gas migration in a mine system network domain during a period of reversal ventilation. Combining field experiments with the TF1M3D simulation program, we analyzed gas migration and distribution during reversal ventilation in the JIU LI coal mine. The results showed that, after implementation of the airflow reversal process for the entire mine, the gas in the return roadways flowed back to the working face and accumulated with the gas emitted from the working face to form a gas concentration peak, after which the gas concentration gradually decreased in a stepwise manner and finally reached a stable state that was maintained until the end of the reversal ventilation. The peak gas concentration and the peak areas of the gas concentration curve during the airflow reversal were positively correlated with the time of airflow stoppage operation. The gas concentration peak affected the safety of the mine airflow reversal process; therefore, countermeasures and technical plans should be made in advance. The TF1M3D simulation results were consistent with the field experiment results. Full article

As resources are extracted from the deeper sections of a mine, the ventilation network becomes increasingly complex. Consequently, determining the optimal installation location for speed-measuring equipment that accurately reflects the average wind speed along the roadway remains a challenging task. In this study, two three-dimensional geometric models, smooth and rough, were developed based on field conditions. The cross-sectional widths, heights, and flow velocities of the model channels were processed dimensionlessly. The dimensionless velocity distributions of the smooth and rough models were then analyzed for different Reynolds numbers. It was observed that the dimensionless average velocity ring distributions for the rough model were smaller than those for the smooth model. Additionally, the maximum values of dimensionless flow velocities were negatively correlated with the flow velocities under laminar flow conditions, whereas they largely overlapped under turbulent flow. The dimensionless distances of the average velocity rings from the top and sidewalls of the channel were studied and determined for both models across different flow regimes. Specifically, the dimensionless distance values $d(-)$ were found to be 0.111 for the smooth model and 0.101 for the rough model under the laminar regime. Under the turbulence regime, the corresponding values were 0.106 and 0.108. Likewise, the values of $h(-)$ were 0.135 and 0.135 for the smooth and rough models in the laminar flow regime, while under turbulent flow, the values were 0.131 and 0.162, respectively. The largest dimensionless velocity value was identified at the center of the velocity distribution circle. For corners that did not maintain simple parallelism with the walls, these regions were incorporated into the circle equation using the Least Squares Method, providing a theoretical basis for the placement of velocity-measuring equipment in practical applications. By using the sidewall as the reference coordinate, an appropriate mathematical model was employed to establish the functional relationship between the centerline velocity of the roadway and the dimensionless horizontal coordinate. The fitting results showed good agreement, and this model can be used to back-calculate and expand the potential installation locations for a mine anemometer. Full article

Gas explosions are the most serious type of accident in coal mines in China. This study analyzed 125 gas explosion accidents that occurred between 2010 and 2020. The results showed that the number of gas explosion accidents and deaths in 2010–2020 was stable and decreasing. The number of larger gas explosion accidents in 2010–2020 is the largest, but the death toll from major accidents was much greater. Coal faces, headings, and roadways are the main locations where gas explosions are initiated. The coal mines in which gas explosions occur in coal faces and headings are mainly “township” enterprises and private mines, all of which engage in illegal operations. The main cause of gas accumulations in roadways is ventilation system failure; these failures can be reduced with improved ventilations system management. The number of gas explosion accidents and related deaths in the Sichuan, Guizhou, and Heilongjiang provinces are very high. The annual change in the frequency of gas explosion accidents, the quarterly distribution of gas explosion accidents, and time during a mining shift when gas explosion accidents occur are closely related to national policies and regulations, company annual production goals, and the mental status of miners, respectively. Full article

Close-distance coal seams are common in underground mining, and their spacing is short, which produces strong mining disturbance. In instances where the upper seam has been mined and a goaf has formed, a notable issue arises during the lower seam’s mining, characterized by substantial deformation of the roadway along the goaf. Field exploration and three-dimensional geological modeling have revealed that the fourth and sixth working faces and pillar of seam No. 5 are all under seam No. 2’s goaf, with an average distance of 16.70 m. Simultaneously, the double compression effect of the pillar, induced by the linkage rotation of key blocks of the lower and upper seams, is analyzed. The induction mechanism and path of the large deformation are expounded. It is thus proposed that the pillar’s width should be determined by gob-side entry, driving beneath the goaf, with the roof near the pillar being cut off in advance to realize the path of cutting off the compressed pillar. Through the simulation comparison of five kinds of pillar width combined with engineering practice, it has been determined that the best width is 8 m, and the abutment pressure is distributed in a double-peak saddle shape, with the result that the load-bearing ability is notably significant. Through the comparative simulation of roof-cutting, it was found that roof-cutting helps the roof to collapse near the pillar-side and decreases the vertical stress peak to 16.46 MPa, the shear stress peak to 5.93 MPa, and the J₂ peak to 7.23 × 10¹³ Pa, which further alleviates the pressure on the pillar. In the field, the haulage roadway’s roof was cut by two-way shaped-charge blasting, and the sandy mudstone (5.90 m) was successfully cut off. Concurrently, anchor cable reinforcement was implemented on the roof and two ribs of the ventilation roadway in proximity to the pillar, thereby ensuring stabilization and mitigating the mining effect. The engineering research provides a case and scheme reference for the operation of gob-side entry driving beneath close-distance goafs worldwide. Full article

To solve the problem of the inability to achieve Y-shaped ventilation in the boundary coal mining of high-gas mines and the problem of gas accumulation in the upper corner of a fully mechanized mining face, non-pillar coal mining technology is proposed by a driving roadway during the mining period. A high-gas working face requires Y-shaped ventilation to achieve upper corner gas control, but Y-shaped ventilation conditions are not available at the boundary coal body. In order to handle this challenge, studies have suggested non-pillar coal mining technology, which involves excavating roadways while mining in order to achieve non-pillar coal extraction and use recoverable wide coal pillars. During the simultaneous excavation of a working face and roadway, studies analyzed the distribution characteristics of the complicated stress environment. Following an evaluation of the impact of coal pillar width on the quality of an excavation roadway, this study’s development is in terms of an effective technique for retaining coal pillars as established. During the mining period of a working face, in the goaf of the working face, the research analyzed the distribution properties of the gas flow field, and findings from the study indicate that the width of the recovered coal pillar influences the distribution of gas. Finally, the width of the coal pillar was comprehensively determined, forming non-pillar coal mining technology by a driving roadway during the mining period. The on-site practice has shown that using a wide coal pillar with a width of 70 m to protect the roadway significantly reduces the deformation of the surrounding rock in the mining roadway, the gas concentration at the return airway is lower than the safety production standard, and by decreasing the mining succession time by 15 months, studies achieved improving the working face’s coal extraction rate by 12.6%. Full article

During the process of mining underground coal, the coal emits a large amount of methane into the mining space, which may lead to methane accumulation and exceed explosion safety limits When the methane encounters a fire source, a methane explosion may occur. The forceful impact caused by a methane explosion in an underground roadway can cause serious damage to the roadway structures and even lead to the collapse of the ventilation system. At the same time, the explosion impact may result in the death of workers and cause physical injury to the surviving workers. Therefore, it is necessary to study the damage effect and injury range of methane explosions. On the basis of the damage criteria and damage characteristics of methane explosions, according to the overpressure distribution of shock waves in the propagation process of a methane explosion, the explosion hazard range is divided into four ranges (from inside to outside): death range, serious injury range, minor injury range, and safety range. Four injury degrees of shock wave overpressure to personal body (slight, medium, serious injury, death), and seven damage degrees of overpressure to structures are also analyzed. The thresholds of their damage (destruction) are determined. On this basis, an experimental system and numerical simulation are constructed to measure damage characteristics, the overpressure value, and the range distance of a methane explosion with different initial explosion intensities. According to the experimental and numerical results, the attenuation formula of a methane explosion shock wave in the propagation process is derived. The research results show that the overpressure and impulse of shock waves are selected as the damage criteria for comprehensive evaluation, and the overpressure criterion is suitable of determining the injury (failure) range over long distances. The four injury ranges are in line with the actual situation and are reasonable. The injury degree also conforms to the medical results, which can be used to guide the injury degree of mine methane explosions. The injury range caused by methane explosions with different initial explosion intensities is reasonable and is basically consistent with the on-site situation. The derived attenuation formula and calculated safety distance are in good agreement with the experimental and numerical results. The research results can provide guidance and help in the escape, rescue, and protection of coal mine underground person. Full article

In order to solve the problem of gas overrun in the fully mechanized caving face and the upper corner of high gas and extra-thick coal seam, the fracture and caving process of the roof in the goaf is analyzed and studied by using the relevant theories of fracture mechanics and seepage mechanics. The mathematical model of fracture and caving of the immediate roof and main roof in the goaf is established. Combined with ANSYS Fluent 6.3.26, the seepage process of gas in coal and rock accumulation in the goaf under different ventilation modes is simulated. The distribution law of gas concentration in the goaf is obtained, and the application scope of different ventilation modes is determined. In addition, the influence of the tail roadway application and the wind speed size on the gas concentration in the goaf and the upper corner of the fully mechanized caving face is also explored. The results show that, affected by wind speed and rock porosity, along the strike of the goaf, about 30 m near the working face, the gas concentration is low and growth is slow. In the range of 30~160 m, the gas concentration increases rapidly and reaches a higher value. After 160 m, the gas concentration tends to be stable. Along with the tendency of the working face, the gas concentration in the goaf increases gradually from the inlet side to the return side, and the gas concentration increases noticeably near the return air roadway. Along the vertical direction of the goaf, the gas concentration gradually increases, and the concentration of the fracture zone basically reaches 100%. Different ventilation modes have different application scopes. The U-type ventilation mode is suitable for the scenario of less desorption gas in the coal seam, while U + I and U + L-type ventilation modes are suitable for the scenario of more desorption gas in coal seam or higher mining intensity. The application of the tail roadway can reduce the gas concentration in the upper corner to a certain extent, but it has limited influence on the overall gas concentration distribution in the goaf. In addition, when the wind speed of the working face should be controlled at 2.0~3.5 m/s, it is more conducive to the discharge of gas, the method of reducing the gas concentration in the upper corner by increasing the wind speed of the working face is more suitable for the case where the absolute gas emission of the fully mechanized caving face is low, and the effect is limited when the absolute gas emission is high. The above conclusions provide a reference for solving the problem of gas overrun in the goaf and the upper corner of a fully mechanized caving face. Full article

In view of the current situation where research on the dust diffusion laws of different dust source points is limited and the gap with the actual field situation is too large; this study employs an innovative gas–liquid–solid triphase coupling method to investigate how dust moves and spreads in the fully mechanized excavation face 431305 at the Liangshuijing Mine; focusing on both the dust field and the dust–fog coupled field. The results indicate that using the long-pressure short-suction ventilation method; dust movement in the roadway is primarily influenced by the airflow; which can be classified into vortex; jet; and return flow regions. The analysis reveals that different dust source points affect dust distribution patterns. Dust source 1 generates the highest dust concentration; primarily accumulating on the duct side and return air side of the roadway. By contrast; dust source 2’s dust mainly gathers at the heading and the front of the cutting head. Dust sources 3 and 4 show lower dust concentrations near the top of the roadway. Dust source 5 achieves the most effective dust removal; aided by airflow and a suction fan; showcasing superior dust performance. A comprehensive comparison indicates that dust source 1 has the highest overall dust concentration. Therefore; further simulation of the distribution law of dust generated at dust source 1 under the action of water mist reveals that the dust concentration near the heading face is reduced from 2000 mg/m³ under the action of single air flow to about 1100 mg/m³. At t = 5 s; the spray droplets almost cover the entire tunneling face; leading to a significant decrease in dust concentration within 10–25 m from the tunneling face. Within 40 s; both coal dust and spray droplets are significantly reduced. The field measurement results verify the accuracy of the simulation results and provide certain guidance for promoting the sustainable development of the coal industry. Full article

In ventilation systems of metal mines, the real-time measurement of the airflow field and a reduction in pollutants are necessary for clean environmental management and human health. However, the limited quantitative data and expensive detection technology hinder the accurate assessment of mine ventilation effectiveness and safety status. Therefore, we propose a new method for constructing a mine intelligent ventilation system with a global scheme, which can realize the intelligent prediction of unknown points in the mine ventilation system by measuring the airflow parameters of multiple known points. Firstly, the nodal wind pressure method combined with the Hardy–Cross iterative algorithm is used to solve the mine ventilation network, and the airflow parameters under normal operation and extreme working conditions are simulated, based on which an intelligent ventilation training database is established. Secondly, we compared the airflow parameter prediction ability of three different machine learning models with different neural network models based on the collected small-sample airflow field dataset of a mine roadway. Finally, the depth learning method is optimized to build the intelligent algorithm model of the mine ventilation system, and a large number of three-dimensional simulation data and field measurement data of the mine ventilation system are used to train the model repeatedly to realize the intelligent perception of air flow parameters of a metal mine ventilation network and the construction of an intelligent ventilation system. The results show that the maximum error of a single airflow measurement point is 1.24%, the maximum overall error is 3.25%, and the overall average error is 0.51%. The intelligent algorithm has a good model training effect and high precision and can meet the requirements of the research and application of this project. Through case analysis, this method can predict the airflow parameters of any position underground and realize the real-time control of mine safety. Full article

Coal dust is one of the environmental factors that seriously affect the health of workers as well as the mining equipment in underground coal mines. At present, coal dust is commonly generated during drilling, blasting, excavation, and transportation processes in mining operations. During mining blasting processes, coal dust is generated with varying particle sizes and high concentration levels. High concentrations of dust will affect mining operations and increase the ventilation time required for mining faces. In addition, coal dust exists in suspended form in the roadway and is harmful to human health, especially fine dust particles that have a negative impact on work efficiency. To improve ventilation efficiency and eliminate coal dust, this article presents a CFD-DPM numerical modeling method that integrates a DEM collision model based on the finite element method to analyze the motion characteristics of airflow and dust particles in the mine tunnel, while considering collisions between particles and between particles and walls. The article analyzes the distribution of wind speed, the dispersion of dust in the space around the roadway, and dust concentrations at distances of 1 m, 3 m, and 6 m from the working personnel and at a position 1.5 m above the roadway floor, corresponding to the breathing zone of the workers, with varying parameters such as velocity and duct position. The results indicate that with a wind velocity of V = 18 m/s and an air duct height h = 3.0 m, the best dust reduction results are achieved, and they provide theoretical guidance for selecting and optimizing ventilation parameters in dust control. Full article