Aiming at optimization of bench height for open-pit mines in high-altitude cold regions, a multi-objective decision-making method considering the influence of freeze-thaw cycles was proposed. Based on parameters including shoveling equipment selection and toe burden, the bench height was preliminarily determined to be within the range of 12-18 m, and seven alternative schemes were proposed. Numerical simulation of slope stability after 25 freeze-thaw cycles was conducted by adopting GeoStudio software, and the safety factors of each scheme ranged from 1.196 to 1.214. Models for ore dilution and loss, drilling and blasting cost, and ore-rock transportation work were then established to calculate the performance values of each scheme in terms of dilution rate, loss rate, drilling cost, blast cost, and transportation work. On this basis, the entropy-weight method was used to determine the weight of each index, and the comprehensive scores of alternative schemes were calculated. The results show that the 15 m bench scheme achieves the highest comprehensive score (0.516 6), and is identified as the optimal solution. Incorporating the mechanical effects of freeze-thaw cycles into the multi-objective evaluation framework, this method provides a quantitative decision-making basis for bench height design of open-pit mines in high-altitude cold regions.

As for underground engineering projects in shallow-buried soft rock, extensive blast vibration monitoring was carried out along the axial direction on the tunnel ground surface. Based on the monitoring data, probability distribution analysis confirmed that the peak particle velocity (PPV) followed the lognormal distribution. Considering the randomness of PPV, the lognormal distribution theory and confidenc-level analysis method were introduced. A blast vibration safety prediction and control model was established by combining the PPV attenuation model with the probability distribution function, and then applied to the sensitive sections of the project. The results show that in a blasting scheme with the maximum charge per delay at 43.2 kg, the horizontal distance between sensitive point 2^# and the blast source should be greater than 35.044 m to satisfy the safety criterion of a 95% confidence; while in a scheme with the maximum charge per delay at 18.6 kg, that horizontal distance shall exceed 25.660 m to meet the 95% confidence. On the basis of this model, differentiated blasting control schemes are formulated for sensitive tunnel sections. This proposed method provides theoretical support for blasting operation for similar underground softrock projects.

To understand the splitting-tensile behavior of internal cross-cracks in rock, discrete element splitting simulation tests were carried out by using a particle flow code (PFC^2D) based on the physical and mechanical parameters of rock obtained from laboratory tests. The results show that the tensile strength of specimens with pre-fabricated cracks is lower than that of intact specimens. When the angle between the main crack and the loading direction ranges from 0° to 60°, the peak strength of specimens decreases with the increase of the secondary crack angle; when the angle ranges from 60° to 90°, the peak strength increases with the increase of the secondary crack angle. In general, the tensile strength of specimens follows the rule that the larger the angle difference between the main and secondary cracks, the smaller the tensile strength. Under splitting loading, pre-fabricated cracks do not change the loading mode of specimens but only reduce their bearing strength. Micro-cracks in cross-crack specimens are dominated by tensile micro-cracks, which initiate near the tips of pre-fabricated cracks, propagate toward the loading ends and coalesce with increasing load, resulting in specimen failure. Irreversible transformation of internal energy occurs during rock failure. When the loading reaches the storage limit of elastic strain energy, the elastic strain energy is rapidly converted into dissipated energy for micro-crack initiation and propagation as well as particle movement. A large number of acoustic emission events occur simultaneously, and the number of micro-crack increases exponentially.

When drilling and blasting are adopted for excavation of high slopes, blasting-induced vibration signals are susceptible to noise interference and their effective features are difficult to extract accurately, a noise reduction method based on complete ensemble empirical mode decomposition with adaptive noise (CEEMDAN) combined with wavelet threshold is proposed. Blasting vibration signals are reprocessed based on mean ratio (MR), sample entropy, and correlation coefficient. The proposed method is verified by simulated and field measured blasting vibration signals. The simulation results show that the signal-to-noise ratio (SNR) of the proposed method reaches 27.80 dB and the root-mean-square error (RMSE) is 0.209 1, presenting a superior noise-reduction performance compared to the other four methods. Engineering application results indicate that this proposed method can effectively suppress high-frequency noise and improve the resolution and stability of time-frequency analysis for blasting vibration signals. Further analysis reveals that the dominant frequency of blasting vibration signals is mainly within the range of 10-60 Hz with energy concentrated within 16-125 Hz; the dominant frequency of rockfall impact signals is mostly in the range of 40-720 Hz with more complex spectral structures, and its energy is mainly distributed in 62-1 000 Hz.

Aiming at the inconsistency of grading results and difficulty in determining credibility for rock mass quality classification methods such as the RMR method, Q-system method and BQ method, a credibility evaluation method for rock mass quality classification based on analytic hierarchy process (AHP), fuzzy comprehensive evaluation (FCE) and D-S evidence theory is proposed. Firstly, the evaluation vector of rock mass quality grade is obtained via the AHP-FCE model. The evaluation vectors obtained by the three classification methods are taken as independent bodies of evidence. The D-S evidence theory and credibility weighting model are introduced to fuse and revise multi-source evidence, so as to obtain an optimal rock mass quality classification result integrating multi-source information. On this basis, a calculation platform for the credibility of rock mass quality classification is established with MATLAB GUIDE. Verification with 16 groups of measured data from No.2 Ping' an Phosphate Mine shows that the evaluation results obtained by the proposed method are more consistent with engineering practice than those from any single classification method. After revision by the credibility weighting model, the conflict coefficients of 16 groups of original evidence decrease from 0.60-0.82 to less than 0.01. The conflict problem of multi-source evidence is effectively solved, and the credibility of rock mass quality classification results is improved.

With the slopes of an open-pit mine in Inner Mongolia as the research background, the creep characteristics of soft rock in the slopes were investigated by means of field monitoring, laboratory test and theoretical analysis. Real-time monitoring results of the slopes from radar and GNSS found that the creep deformation is relatively larger on the eastern side of the northern slope, which requires key protection. When the axial stress exceeds the long-term strength, the soft rock goes through four deformation stages:instantaneous deformation, decelerating creep, steady-state creep and accelerated creep. Creep deformation and creep rate of soft rock are positively correlated with axial stress. A nonlinear creep model for slope soft rock is established by combining elastic elements, fractional-order element and nonlinear viscoplastic body (NVPB). This model can well describe the whole creep process of slope soft rock, with higher fitting accuracy than that of traditional Burgers model, and also demonstrate favorable applicability.

To reveal the damage and degradation mechanisms of mudstone in tailings reservoir slopes under complex environmental conditions, dynamic uniaxial compression tests, CT scanning, and scanning electron microscopy (SEM) were conducted to analyze the evolution of dynamic uniaxial compressive strength and structural damage mechanisms of mudstone under three typical environmental conditions, i.e., wet-dry cycles, freeze-thaw cycles, as well as wet-dry and freeze-thaw cycles. The results show that the dynamic compressive strength and dynamic elastic modulus of strongly weathered mudstone continuously decay with the increase of cycle number, and the deterioration rate gradually stabilizes after 80 cycles. Among those three environmental damage conditions, the coupling effect of wet-dry and freeze-thaw cycles brings the most significant deterioration to mudstone. After 100 wet-dry and freeze-thaw cycles, the dynamic compressive strength and dynamic elastic modulus of mudstone decrease, and the connectivity of pore networks increases remarkably, presenting a strong linear relationship (negative correlation) between porosity and dynamic compressive strength. Complex environments can reduce the effective bearing area and stress transfer efficiency of rock mass by accelerating pore expansion and crack propagation, thus eventually lead to structural damage and strength deterioration of mudstone. Relevant suggestions for slope reinforcement engineering of tailings reservoirs are finally put forward based on the research results.

With a domestic super-large iron mine taken as research background, FLAC^3D was adopted to calculate and analyze the rock movement law by mining and backfilling operations of underground super-large area, and the deformation degrees of the ground surface and shafts were also evaluated according to relevant standards. It is found that the hanging wall and foot wall in the mining disturbed zone are concentrated within rock movement zones. The peak displacements of subsidence and uplift are approximately 9 cm and 6 cm respectively, which decrease gradually in an arch-shaped pattern toward the ground surface or deep footwall strata. Meanwhile, deep mining activities tend to aggravate the deformation of rock masses in middle sections. The maximum absolute values of surface inclination, curvature and horizontal deformation are 0.22 mm/m, 0.002×10^-3/m and 0.12 mm/m, respectively, all less than the maximum allowable deformation values specified in relevant standards. It indicates that deep mining is unlikely to induce tilting deformation of surface buildings (structures), whereas the southern and northern service shafts are located within the movement zones. Further shaft stability analysis reveals that the southern and northern service shafts are subjected to mining disturbances, and the peak values of inclination, curvature and horizontal deformation of shaft wall are 0.52 mm/m, 0.09 × 10^-3/m and 0.18 mm/m respectively, which are also within the permissible limits according to relevant standards. The maximum cumulative subsidence over time at each monitoring point of the main and auxiliary shafts to the surface is less than 4.6 mm, which indicates that the underground mining of super-large area exerts slight influences on the stability of the main and auxiliary shafts, as well as other buildings (structures).

Aiming at the problems of low blasthole utilization rate, high specific explosive consumption, high detonator consumption, and irregular roadway forming during the excavation of roadways with fractured surrounding rock in a gold mine, an optimization method for burn cut parameters by adopting ultra-deep central hole is proposed. The optimized rock-breaking mechanism is revealed by triple synergistic effect, including enhanced rock throwing via central hole charging, stress wave guidance through large empty-hole spacing, and bottom pre-fragmentation. A three-dimensional cut blasting model is established using ANSYS/LS-DYNA software to simulate four working conditions featuring a 200 mm ultra-deep central hole and enhanced bottom charge, with empty-hole spacings of 130 mm, 160 mm, 190 mm, and 220 mm. By analyzing the effective stress distribution, cut-cavity dimensions, rock-throwing process, and energy variation under different working conditions, the optimal empty-hole spacing is finally determined to be 160 mm. Field test results show that the optimized blasting scheme increases the blasthole utilization rate from 87% to 95.8%-97.8%, reduces explosive consumption by 13.3% and detonator consumption by 14.29%, and controls the over-excavation of two roadway sidewalls within 100 mm. Such findings can provide a theoretical basis and technical reference for similar roadway excavation engineering in fractured surrounding rock.

The shear mechanical behavior of red sandstone specimens with prefabricated V-shaped fractures was systematically investigated via laboratory shear tests and numerical simulations, from the perspectives of strength parameters, crack-initiation locations, coalescence modes and stress-field evolution. The results show that the deterioration of internal friction angle and cohesion of specimens is the most significant when the crack opening angle α=110° and azimuth angle β=0°. The finally formed fracture surfaces with coalescence in specimens are parallel to the shear direction, whereas the crack-initiation locations are strongly controlled by the azimuth angle β. The stress-field distribution shows clear zoning behavior:the low-stress zone presents an inclined S shape, while high-stress and shear-stress zones concentrate at fracture tips and end boundaries. From the mesoscopic perspective, principal stresses dominate the development of wing fractures, and the coupled compression-shear effect is the fundamental cause for the generation of secondary fractures.