Pipeline transportation of tailings slurry is the key to efficient tailings discharge in mines. In order to explore the variation law of flow resistance of mine tailings slurry during pipeline transportation, the influence of multiple factors such as flow velocity, transportation pipe diameter, slurry mass concentration and their interaction on the flow resistance of tailings slurry transportation was analyzed by orthogonal test, and the pipeline transportation parameters of tailings slurry were optimized. The results are as follows. Firstly, the F value of the established regression model is 4 660.66, the P value is less than 0.000 1, and the fitting value R²is 0.997, indicating that the model has high reliability. The model can accurately characterize the change of the resistance along the slurry transportation under different working conditions. Secondly, when the flow rate is low(1−2 m/s), the slope of the curve of the resistance loss along the way is small. When the flow rate is greater than 2 m/s, the curve of the resistance loss along the way is steeper, the resistance is greater, and it increases sharply. Thirdly, the influence degree of multi-factor interaction from large to small is the interaction of tailings slurry flow rate and mass concentration, the interaction of tailings slurry flow rate and transportation pipe diameter, and the interaction of slurry mass concentration and transportation pipe diameter. Fourthly, the optimized pipeline transportation parameters are as follows: the mass concentration of slurry transportation is 56%, the slurry transportation speed is 1.6 m/s, the transportation pipe diameter is 127 mm, and the predicted resistance along the way is 0.242 kPa/m. The measured resistance along the way in the test is 0.233 kPa/m, and the error is only 3.7%, which is within the controllable range. It shows that the accuracy of the prediction model is high, and the resistance along the way is small under this working condition, which meets the requirements of long-distance transportation. The test results provide a reference for the efficient pipeline transportation design of high mass concentration tailings slurry in mines.

In order to quantitatively evaluate the flocculation effect of flocculants and acquire the change law of floc pore structure, fractal theory was introduced to analyze the fractal dimensions and flocculation effect of four flocculants, including ZYD, JYC-1, JYC-2 and ZYZ. The pore structures inside different flocculants samples were analyzed using three-dimensional reconstruction and CT slices, and the mechanism of tailings flocculation and settlement was revealed. Through the static and dynamic flocculation tests, the optimal flocculant type and process parameters were determined, and the relationship between the underflow mass concentration of slurry and the material layer height was deduced. The results show that ZYZ-type flocculant has the largest fractal dimension, which is 1.842 3, with the best flocculation settling effect. ZYZ-type flocculant forms flocs with larger size at a faster settling speed, and forms a better water-conducting channel, achieving a significant thickening effect. Through the static and dynamic flocculation tests, the slurry mass concentration of 20% and the flocculant unit consumption of 20 g/t were determined as the best flocculation process parameters. If the overflow mass concentration of slurry is up to 70%, the height of the sediment layer that needs to be settled is 731 mm. On-site industrial application shows that the average sand mass concentration of the thickener is 70%, the urination rate is no more than 5%, the slurry fluidity is good, and the filling effect is good.

Using high-sulfur tailings and gobi materials as mixed filling aggregates, the mechanical properties of high-sulfur tailings backfill excited by industrial magnesium slag were investigated. A preliminary test was conducted to investigate the effect of different magnesium slag contents on the compressive strength of the backfill. Using NaOH as an alkaline activator, the influence of proportioning parameters of high-sulfur tailings backfill on compressive strength, stress-strain curve characteristics, and failure mode was systematically analyzed through orthogonal tests. The results show that when the ratio of tailings to gobi material is 1:1 and the dosage of NaOH is 1%, the compressive strength of the backfill reaches its maximum value at 28 d, which is 4.54 MPa. NaOH has an optimization effect on the mechanical properties of high-sulfur tailings backfill excited by magnesium slag, and reasonable control of the dosage of high-sulfur tailings, NaOH, and magnesium slag can significantly improve the compressive strength of the backfill. Increasing the dosage of NaOH, magnesium slag, and high-sulfur tailings can enhance the mechanical strength of the backfill, but it will lead to a brittle transition in its failure mode and exacerbate local failure characteristics. The research results can provide a basis for the resource utilization of magnesium slag and high-sulfur tailings, as well as the performance optimization of mine filling materials.

In order to improve the accuracy of the prediction of ore flowability in the ore storage section of the ore pass based on numerical simulation and indoor similarity test, the parameters of the parallel bond model based on the slump and expansion test were introduced in the numerical simulation process. The numerical simulation results obtained by this method were highly consistent with the indoor similarity test. The results indicate a significant decline in ore flowability within the moisture content range of 0 to 9%. Particularly when the moisture content is 7%−9%, the ores are prone to blockages. Therefore, this specific moisture content range should be strictly avoided during actual ore drawing operations. When the moisture content is 3%−5%, the flowability of the ore is the best. It is suggested that the moisture content of the ore should be controlled in this range in the actual project. Furthermore, the study reveals that as the ore storage height increases from 20 m to 40 m, the flowability of the ore improves significantly, reaching a peak mass flow rate at 40 m. However, when the ore storage height is further increased to 60 m, the mass flow rate stabilizes and exhibits a slight decline.

To address the issue of low blasting efficiency in deep buried roadways excavation, the influence of detonation sequence under high stress conditions on the blasting effectiveness of parallel cut was studied. By combining physical model tests with numerical simulations, the dynamic laws of strain field evolution, crack expansion, and groove cavity formation with different detonation sequences were revealed. The metal wire electric explosion technology was used to simulate explosive blasting, combined with a confining pressure loading device to apply static pressure. The digital image correlation method (DIC) and ultra-high-speed camera were used to capture the surface strain field and crack expansion process of the specimens. Based on the 3D scanning results, the differences in groove shape with different detonation sequences were analyzed. The results show that when the first blast hole is located in the direction of the maximum principal stress (detonation sequence two), the circumferential strain of the empty hole is significantly increased, and it is easy to form penetrating cracks. The 3D scanning results show that the groove cavity volume of detonation sequence two is larger than that of detonation sequence one, and the residual rate at the bottom of the hole is reduced. Numerical simulation further verifies that when the first blast hole is along the direction of maximum principal stress, the distribution of damage zone is more uniform, the opening angle of the groove cavity increases, and the energy utilization efficiency of subsequent blast holes is effectively improved. The research has shown that prioritizing detonation in the direction of maximum principal stress can optimize the distribution of damage zones, improve the shape of the groove cavity, and reduce residue at the bottom of the hole.

Aiming at the problem of roadway stability during the collaborative mining of three middle sections of Jinchuan No.2 mining area, the maximum principal stress, roadway convergence rate, rock body integrity(K_v), rock quality (RQD), and support effectiveness were selected as evaluation indexes. Based on the combined weighting-boundary cloud modeling algorithm, the stability of the roadways during the collaborative mining of three middle sections in deep section was comprehensively evaluated and predicted, then compared with on-site working conditions. The results indicate that the stability grades of the roadways in the 1 018 m, 934 m, and 814 m sublevels across different mining phases primarily fall within grades Ⅱ to Ⅲ, but there are differences in the stability of the main roadways in each middle section. In current mining stage, the stability of the roadway in the 1 018 m sublevel is superior to that in the 934 m and 814 m sublevels, with the western and eastern areas of the 1 018 m sublevel exhibiting better stability than the central area. Conversely, the central and eastern areas of the 934 m and 814 m sublevels show greater stability than the western area. In the later stages of mining, the stability of the main roadway in each middle section gradually shifts from more stable to generally stable, resulting in a decrease in stability. The evaluation results align with the actual project, providing a reference for roadway support in the late stages of the collaborative mining of the three middle sections in Jinchuan No.2 mining area.

In order to solve the problem of roof instability and material waste caused by unreasonable value of roof support parameters, the optimization of support parameters was studied based on the research background of the north stope of Zhongguan Iron Mine. Firstly, based on the suspension theory and the loose circle theory, combined with the numerical simulation results without support, the plastic zone was used as the basis for the determination of the loose circle, and the thickness of the loose zone was obtained, and then the length of the bolt was 2.4−4.4 m. Secondly, according to the quality of rock mass, the designed row spacing between bolts was 0.8 m×0.8 m and 1.2 m×1.2 m respectively, the length of bolts was 2.4 m and 4.4 m respectively, and the anchoring force was 17.5 MPa and 20 MPa respectively. The FLAC3D software was used to establish the mining model, and the vertical stress, vertical displacement and plastic zone distribution of the roof of each scheme were compared and analyzed. Finally, based on the simulation results, the index satisfaction method was used to select the roof settlement, the height of the stress release zone, the area of the plastic zone and the total length of the bolt, and the optimization model of the bolt parameters considering the safety and economic benefits was established. The highest comprehensive satisfaction is 0.899, and the optimal scheme is determined to be a row spacing between bolts of 0.8 m×0.8 m, bolt length of 4.4 m and anchoring force of 20 MPa. The simulation results provide theoretical support for the determination of the bolt parameters of the ore body.

The mine ventilation system is the important subsystem in the underground engineering of the mine. The accurate control of the air volume is the basis for the optimization of the ventilation system, and the local resistance is one of the important factors affecting the air volume. In order to study the local flow field characteristics of the double-frame louver windshield, the Computational Fluid Dynamics (CFD) numerical simulation software was used to analyze the dynamic characteristics of the local flow field of the louver windshield under different inlet wind speeds and different blade opening angles. The relationship functions between the effective opening area of the windshield with the air volume and local resistance were established and compared with the measured results. The results show that when the blade opening angle is 45°, a high-speed flow field is gradually formed at the inlet of the windshield. With the increase of the inlet wind speed, the peak wind speed and the flow field area increase. When the blade opening angle increases from 60° to 90°, the local resistance at the windshield decreases and the distribution characteristics of the rear flow field change. During the process of the wind flow passing through the windshield, the static pressure energy and kinetic energy are converted to each other. The pressurization deceleration area and the vortex area formed on the leeward side of the air door become the key factors affecting the local resistance and energy loss in the flow field behind the double-frame louver windshield. The local resistance coefficient of the windshield depends on the effective wind area, which has a power function relationship with the blade opening angle. The maximum relative error between the local resistance of the windshield calculated by the formula and the measured results is 5.57%, which shows that the local resistance calculated by the formula is more accurate.

In response to the complexity and ambiguity of safety evaluation for coal mine electromechanical systems, interval approximation theory was introduced to improve the objectivity and reliability of evaluated results, and a comprehensive safety evaluation model for coal mine electromechanical systems based on G1-FDEMATEL-interval approximation was established. Firstly, according to the theory of accident causation intersection, 17 indicators were filtered from the four dimensions of human, machine, management, and environment to construct an evaluation system. Secondly, the G1-FDEMATEL method was coupled to obtain the comprehensive weights of indicators. Based on the principle of interval approximation, the interval estimation method was used to determine the evaluation data of indicators, which compensated for the limitation of discontinuous point estimation values and information missing in traditional models. By calculating the comprehensive distance, the safety level of coal mine electromechanical systems was determined. Finally, taking 5 coal mines as examples, the G1-FDEMATEL-interval approximation model was applied to analyze the safety status of the electromechanical systems. The results show that the safety levels of the electromechanical systems in all five coal mines were relatively safe, consistent with the expert conclusions. Compared with traditional VIKOR and TOPSIS models, it can effectively reflect the true safety status of the electromechanical systems, verify the adaptability and scientificity of the models, and provide a theoretical support for preventing and handling electrical system safety accidents.

In order to objectively evaluate the low-carbon ecological construction effect of mines and promote the coordinated development of economic construction and environmental protection of mining enterprises, an evaluation index system was established based on the framework of DPSIR. The index weight algorithm of DPSIR-DEMATEL-AHP was constructed using the DEMATEL method and AHP method. Using the interval number optimal algorithm and the binary connection number optimal algorithm, a method for evaluating the low-carbon ecological construction effect in mines based on interval number goodness was established to objectively reveal the close degree connection among the driving force of mining enterprise development, ecological environment pressure, environmental quality status of mining areas, the impact of mining on the environment and the response of mining enterprises. It provided a basis for dynamic evalution of the low-carbon ecological construction effect in mines and analysis of the key factors affecting low-carbon ecological construction. The model method was used to evaluate the low-carbon ecological construction effect of an open-pit mine and analyze its key influencing factors. The results show that the low-carbon ecological construction effect of this mine is the best in 2021 and the worst in 2020. Important measures to improve the effectiveness of low-carbon ecological construction in mines include enhancing the mining profit margin, saving energy and increasing efficiency in mines and reducing the CO₂ emissions, enhancing the solid waste utilization per unit of industrial output value, increasing the greening coverage rate in mining, implementing effective dust-reducing measures, strengthening the management of key work processes in mine production, as well as enhancing the coverage rate of energy-saving and emission-reduction campaigns and training and so on.