In order to select the best filling mining scheme and to optimize stope structure parameters of a high-value iron deposit, combined with the mining technical conditions of the deposit in the stage of −120 m to −180 m, the selection of mining scheme was regarded as a typical fuzzy decision problem. The analytic hierarchy process was used to obtain the weights of 11 indexes, and the fuzzy mathematics was used for comprehensive evaluation. The mining scheme of this iron deposit was selected as the sublevel drilling stage open stoping with subsequent filling method. Based on numerical simulation, four groups of stope structural parameters of the preferred mining scheme were optimized. The simulation results are concluded as follows. Firstly, in the process of mining, the tensile failure of each scheme is obvious, and the stress concentration occurs at the junction of the pillar and the roof. The maximum horizontal displacement of the central monitoring point at two sides of the stope and the maximum vertical displacement of the central monitoring point at the roof are 9.4 mm and 9.6 mm, respectively. Secondly, in the process of pillar mining, the shear failure of each scheme is obvious, and the maximum vertical displacement of the stope roof center in scheme Ⅱ is 28.0 mm. From the perspective of safety and economy, the optimal stope structure parameters of an iron mine are 15 m room width and 15 m pillar width. The scheme and the optimized stope structure parameters have achieved good economic benefits in mine production.

In underground metal mining, the roof stability of stope is directly related to the overall safety of the high-stage filling stope. Taking 2 types of stopes with different heights in Zhongguan Iron Mine as the engineering background, by testing the range of loose ring of pre-control top roadway without mining influence, the appropriate length of anchor rod was selected, and four kinds of combined support schemes of “prestressed long anchor cable+anchor net spray” with different row spacings were proposed. The roof stability of stope in the process of multiple mining under different anchor cable row spacing conditions was simulated and compared to determine the optimal anchor cable row spacing and other support parameters. The results show that the loose ring range of the side of the pre-controlled top roadway in the 30 m height stope is 1.6−2.0 m, and in the 60 m height stope is 1.4−2.3 m. The selected length of anchor rod is 2.4 m. When the anchor cable row spacing is 3.5 m, 2.5 m and 1.5 m respectively, the support control effect of 30 m height stope is 37.15%, 37.50% and 26.25% higher than that of no anchor cable, anchor cable row spacing of 3.5 m and anchor cable row spacing of 2.5 m respectively. When the anchor cable row spacing is 5 m, 4 m and 3 m respectively, the support control effect of 60 m height stope is 36.65%, 34.00% and 30.35% higher than that of no anchor cable, anchor cable row spacing of 5 m and anchor cable row spacing of 4 m respectively. Considering the economic cost and construction intensity, it is determined that the anchor cable row spacing of 30 m and 60 m height stope pre-control top roadway is 2.5 m and 4.0 m respectively. It is verified by engineering practice that there is no caving in the process of mining, and the stability is good. The recovery rate is 88.25% and the dilution rate is 10.39%, which can be further popularized and applied.

In response to the problem of excessive local support parameters and excessive support material consumption leading to longer support cycles and higher annual support costs in the support of roadway outside the vein in a certain lead-zinc mine, the support of roadway outside the vein in the 12-15 line of 2416 section of this mine was taken as the engineering background. After conducting rock mass quality evaluation and measuring relevant mechanical parameters by test of the rock mass, the optimal support scheme was selected by comparing the support effect and cost through numerical simulation and on-site testing. The results show that the combined support method of anchor rod+ steel plastic mesh + double reinforcement bar + shotcrete can meet the stability of the roadway outside the vein during its service life. Following parameters were selected, including the length of anchor rod with a length of 2.4 m and a diameter of 18 mm, as well as the MSCKa resin anchoring agent, with a spacing of 1.1 m, a steel plastic grid mesh specification of 4.5 m×2.5 m, a hole mesh of 6 cm, a tensile load of 60 kN, a double reinforcement bar diameter of 6 mm, and a sprayed concrete strength of 20 MPa. The support cost can be reduced by 599.46 RMB/m. The research results can provide a reference and guidance for the design of underground roadway support in similar mines.

Rock burst is the main technical challenge that troubles the safe mining of phosphate mines in the western Hubei region. In order to reduce the rock burst risk of brittle direct roof of phosphate mines in western Hubei, taking the +80 m working face of Shaqihe Phosphate Mine as the research object, the applicability of different roadway section shapes in mines with brittle direct roof was analyzed, and an inverted trapezoidal section scheme conducive to prevention and control of rock burst was proposed. Numerical simulation method was used to study the distribution characteristics of stress, elastic energy, and displacement of surrounding rocks under different section shapes, the optimal section shape was selected, and the design of relatively safe and economical section parameters were optimized. Finally, engineering verification was carried out on the working face site. The results show that for mines with brittle direct roof, the stress concentration of the roof with rectangular section is the most significant, the roof subsidence of the positive trapezoidal section is large, while the roof stress distribution of the inverted trapezoidal section is relatively uniform, and the roof stress and elastic energy are obviously released, thus reducing the risk of rock burst. Thus, the inverted trapezoidal section is the optimal section shape, with corresponding parameters of 3.4 meters for the upper bottom, 3 meters for the lower bottom, and 4 meters for the height. After optimizing the roadway section, the rock burst of direct roof on the working face has been significantly reduced. The research results can provide a successful example for the prevention and control of rock burst in similar mine roadways in western Hubei.

In order to optimize the performance of filling materials while reducing filling costs, the research on filling materials proportion was carried out. Based on the RSM-CCD method, an experiment with 4 factors and 5 levels was designed to construct a response surface regression model with factors of fly ash mass fraction A, cement mass fraction B, calcium carbonate mass fraction C, and slurry mass fraction D. The slump of the filling slurry and the uniaxial compressive strength of the filling body at 7 d and 28 d were the response values, respectively. The influence of single factor and the interaction of various factors on material properties were analyzed, and the lowest cost filling material proportion scheme was sought on the basis of meeting performance requirements. The results show that the slump of high-concentration filling slurry is only affected by a linear single factor, and the interaction has no significant effect on it. The slump increases firstly and then decreases with the increase of A, B, and C, and decreases with the increase of D. Linear and interactive effects have a significant impact on the strength of high-concentration filling body at 7 d and 28 d. There is no significant relationship between the interaction effect AC and the strength of the filling body. The increase of single factor in AB or CD will promote the strengthening effect of the other factors on the compressive strength of the filling material at 28 d, while the increase of single factor in AD, BC, and BD will inhibit the enhancement effect of the other factors on the compressive strength of the filling materials at 28 d. The optimal proportion of filling materials considering cost-effectiveness is 18.0% of fly ash mass fraction, 10.9% of cement mass fraction, 1.3% of calcium carbonate mass fraction, and 84.0% of slurry mass fraction. This proportion scheme meets the design standards for paste filling in engineering examples.

In order to meet the requirements of the two-step stope mining in the Peki Copper-Gold Mine of Serbia for the backfill strength and reduce the backfill cost, the backfill material mix ratio was optimized. The backfill strength in different parts of the stope was preliminarily determined by theoretical calculation, and several groups of indoor tests of backfill material mix ratio were carried out. The test shows that the uniaxial compressive strength of the backfill increases with the increase of the mass concentration of the backfill material and the cement-sand ratio. When the mass concentration is 68%, the cement-sand ratio is 1:6 and 1:8, the backfill strength can meet the strength requirements of the theoretical calculation. On this basis, the industrial test was carried out. The results show that the backfill strength requirements of the two-step stope mining can be met using 1:6 cement-sand ratio at the bottom of the first-step stope and 1:8 cement-sand ratio at the top and middle of the first-step stope. At the same time, compared with the original backfill ratio scheme, the backfill cost can be reduced by 2.53 yuan/t.

Most of the backfill near the mining working face are in the environment of cycle temperature. In order to study the influence of cycle temperature on the damage evolution and failure characteristics of cemented tailings backfill, digital image correlation (DIC) and acoustic emission (AE) techniques were used to conduct uniaxial compression tests on cemented tailings backfill after curing at six different cycle temperatures. The fracture process and acoustic emission behavior characteristics of cemented tailings backfill at different cycle temperatures were analyzed, and the damage model and constitutive equation of cemented tailings backfill at different cycle temperatures were established. The results show that with the increase of cycle temperature, the compressive strength, elastic modulus and peak strain of cemented tailings backfill increase first and then decrease, and all reach the maximum value when the cycle temperature is 25−40 ℃. With the increase of cycle temperature, the failure pattern of cemented tailings backfill changes from shear failure to tensile-shear composite failure, and then to tensile failure. The intensity of the acoustic emission ringing count signal in the initial compaction stage and the post-peak failure stage is different due to the different cycle temperature. The sudden change of the acoustic emission ringing count can be used as a precursor to the failure of the backfill. The reliability of the model is verified by comparing the theoretical calculation value of the constitutive model of the backfill parameters with the experimental measured value. The research results can provide reference for the design of backfill near the working face.

Aggregate size distribution of waste rock and the cement-sand ratio of backfill slurry have a significant effect on the mechanical properties of waste rock cemented backfill. In order to explore the relationship between the two and the mechanical properties of the backfill, the backfill ratio tests were carried out. The particle flow model was constructed. The laws of crack development, force chain evolution and particle failure in the process of uniaxial compression of waste rock cemented backfill were analyzed. The results show that the higher the cement-sand ratio, the higher the sensitivity of the strength of waste rock cemented backfill to the change of fractal dimension of aggregate size distribution. When the cement-sand ratio increases from 1:8 to 1:4, the strength range of samples with different fractal dimensions increases from 0.40 MPa to 1.54 MPa. In the cemented backfill, the interface between the aggregate and the cemented matrix is easy to be destroyed along the aggregate boundary and develops along the interface. With the increase of fractal dimension, the proportion of fine aggregate in the sample increases, and the proportion of shear crack increases. In addition, the higher the fractal dimension, the more conducive to the uniform distribution of the cemented matrix, which can better participate in the bearing capacity. In the design of backfill scheme, the strength of backfill can be flexibly adjusted to meet the actual needs by changing the fractal dimension of waste rock aggregate.

The porous and equal discharge technology for filling pipelines can eliminate the unattached goaf area formed by the self-flow gradient of the paste, improve the roof-contact effect of the mining site, and avoid the problem of roof subsidence. Firstly, through theoretical analysis, the head loss along the porous and equal flow discharge pipeline and the pressure head at each discharge hole were preliminarily determined, and on this basis, the diameter of the discharge hole was obtained. Then, using a self-developed porous and equal discharge test device, the influences of inlet flow rate, the number of holes, and hole spacing on the equal discharge effect were studied through experiments. Finally, numerical simulation methods were used to analyze the pressure and flow rate variations of the discharge pipeline. The results indicate that the spacing between the discharge holes has the most significant impact on the uniformity of discharge, followed by the number of holes. When the hole spacing is 14.5 cm, the flow rate at the inlet is 330 mL/h, and the number of holes is 6, the uniformity of discharge can reach up to 96%. When the paste flows in a horizontal pipe section, the static pressure of the pipeline decreases along the flow direction, and negative pressure appears at the discharge holes. The closer the discharge hole is to the inlet of the paste, the greater the pressure difference. After the paste reaches a stable state, the flow rate of each discharge hole is basically the same, and the uniformity coefficient reaches 98.8%, which means that the paste can achieve porous and equal discharge.

In the process of protective layer mining, hard rock faults and hard nodules are often encountered, which causes that it is difficult for the shearer to cut, the cutting parts are seriously worn, and the advancing speed of the working face is seriously restricted. In order to solve the above problems, based on the engineering background of Ji₁₅₋₁₇−13070 soft rock protective layer working face in Pingmei Thirteenth Mine, the mechanism of ultra-deep hole loose pre-splitting blasting was clarified. The radius range of the fracture zone in the fault hard rock was calculated to be 0.7−1.2 m. The influences of aperture and hole spacing on blasting effect were studied by numerical simulation method, and aperture and hole spacing were determined to be 94 mm and 2.0 m respectively. The technical scheme of “advance drilling in two roadways+ ultra-deep hole loose pre-splitting blasting in working face”, which can realize the parallel operation of propulsion and blasting, was put forward and the industrial test was carried out. The test results show that after adopting the technical scheme, the fault hard rock is fully broken, the wear of the shearer pick is reduced by more than 40%, the advancing speed of the working face is increased by about 25%, and the blasting pre-splitting effect is remarkable.