The construction of a blasting network is challenging, with high risks and costs associated with blasting equipment. Considering the geological structure, physical and mechanical properties of the excavation target, and the detonation characteristics of the detonating cable, a delayed detonation network was designed to combine a digital electronic detonator and a detonating cord. This design replaced the digital electronic detonator with the detonating cord in auxiliary and peripheral holes within the same section. In contrast, a single digital electronic detonator was used to initiate detonation to reduce the networking complexity and improve the overall reliability of the network. Field tests conducted at Guangshan iron mine and Gemstone phosphate mine verified the applicability of this method in different mineral environments. The results show that the fusion delay detonation network can significantly reduce detonation equipment costs, simplify initiation network construction, improve cutting effectiveness, and reduce safety risks during excavating small-section roadways in iron mines. However, during the underground phosphate rock test, the unique geological structure and the properties of phosphate rock prevented the designed network from achieving the expected results, indicating a need for further research.

To carry out deep coal mining safely and efficiently, the dynamic mechanical characteristics and fracture mechanism of coal rock in deep earth were investigated under the ‘three high and one disturbance’ environment. A dynamic impact test of coal rock was carried out using the self-improved ϕ 50 mm high temperature synchronous split Hopkinson pressure bar (SHPB) test equipment at temperatures between 25~200℃. A ZWT viscoelastic constitutive model was also improved to establish a dynamic constitutive equation considering the temperature effect. The influence of high temperature on crack development law and the dynamic strength of coal rock was investigated based on the coupling of the finite difference and discrete element methods. The results show four stages to the dynamic stress-strain curve of coal rock under high-temperature impact: compaction, elastic, crack propagation, and softening failure. The dynamic compressive strength and dynamic elastic modulus of coal rock significantly decrease as temperature increases. In contrast, the failure strain increases, and the absorbed energy varies in a W-shaped pattern. The fractal dimension increases linearly as the particle size decreases. The degree and complexity of the fragmentation mechanism increase as the compressive strength decreases. Although the improved dynamic constitutive model based on ZWT could adequately express the stress-strain relationship following a high-temperature impact, it does not apply to the compaction stage. According to the simulation and test results, water and adsorbed gas actively escape in the coal rock at 150℃. The coal matrix is also heated and expanded, which induces cracks. There are apparent mesoscopic cracks initially and gradually developed through cracks, mainly shear ones. The crack development of coal rock under dynamic compression at 100℃ develops through the impact surface, and the high temperature deteriorates the strength of coal rock.

There are hundreds of blast holes in a tunnel blasting. Since the traditional manual drawing of the blasting scheme is laborious and depends on the experience of blasting engineering, a digital method for the planar and three-dimensional spatial distribution of blast holes was proposed based on formula derivation to study an intelligent design method for tunnel blasting blast holes. Subsequently, the programming of blast hole parameters was achieved by utilizing computer programming techniques, which can lead to the development of an intelligent blast hole design system. The results show that the parametric expression method of tunnel contour, cut holes, peripheral holes and auxiliary holes can realize their rapid creation and meet the needs of tunnel blasting. By establishing a correlation between the coordinates of the blast hole opening and bottom, a refined expression of the spatial distribution of the blast hole can be realized, and intuitive guidance for on-site drilling operations can be provided. Furthermore, a computer programming method can realize an intelligent and fine design of a blast hole layout. After tunnel blasting, the residual marks of the blast holes are complete, the contour of the tunnel excavation is smooth, and the overall excavation effect is good. The research results can improve the efficiency and intelligence of tunnel blast hole design.

The rock mass joints can affect the propagation of explosive stress waves. The angle between their direction and the surrounding holes and the relative position changes have different effects on the blasting effect. Based on the attenuation law of stress waves at different jointed angles, a method was proposed for zoning the surrounding holes of tunnel blasting in jointed rock masses. The parameters of the surrounding holes are optimized when the angle between the joint and the surrounding hole is 30°, 60°, 90°, and 0° (parallel). The zoning layout method was validated by combining LS-PREPOST numerical simulation and on-site tests regarding rock damage depth and blasting vibration speed. The results show that the rock mass's damage depth and blasting vibration speed under the zoning arrangement of surrounding holes are significantly better than that of the original layout plan of surrounding holes. Based on the geological conditions of the research section of the Bayueshan Tunnel of the Tongliang Anyue Expressway, the angles between the joints and the surrounding holes are set to 30°, 60°, and 90°, respectively. The spacings between the surrounding holes are set to 43 cm, 50 cm, 58 cm, and 60 cm when the joints parallel the surrounding holes. The average over-excavation value can be controlled at 18cm after blasting, and the over-consumption of concrete is controlled within 100% per linear meter.

This study presents a comprehensive approach to solve the problem of low ore recovery caused by the difficulty in separating small-particle size ore from soil after blasting in a limestone building stone mine. Firstly, a correlation model between blasting fragmentation and dynamic damage of rock mass was established based on field measurement data and numerical simulation results, which can determine dynamic damage thresholds corresponding to various rock particle sizes. Secondly, the numerical simulation test of bench blasting in a three-dimensional fractured rock mass was carried out by using different air-decked charging stages and borehole distribution parameters, which can improve the particle size yield of 0.3~0.9 m and control the bulk ratio to obtain the best blasting parameters. Finally, the field blasting tests were conducted to optimize the charge structure and borehole distribution parameters based on numerical simulation results. The results show a negative exponential function relationship between the blasting block size and the dynamic damage value of the limestone. Specifically, the dynamic damage thresholds corresponding to the blasting size of 0.3 m and 0.9 m are 0.793 and 0.286, respectively. Using only an air-decked charging structure alone can increase the particle size ratio of 0.3~0.9 m and significantly raise the bulk rate. Conversely, combining an air-decked charging structure with a reduced hole spacing markedly enhances the particle size ratio of 0.3~0.9 m while maintaining a stable bulk rate. Optimal blasting results are achieved using a two-stage air interval charging structure and a strategic reduction in hole distribution parameters. The field application results show a 20.09 percentage point increase in the 0.3~0.9 m particle size ratio, with the bulk rate remaining virtually unchanged. Additionally, the unit consumption of explosives decreased by 10.29%.

Taking the explosion of a shipyard as an example, the investigative techniques had been comprehensively utilized (such as on-site investigation, interview, numerical simulation, theoretical calculation, trace analysis, etc.) to conduct in-depth research on the development of the accident, consequences and mechanism of the explosion. An on-site survey of the involved hull revealed fresh welding slag remaining on the starboard main deck in the area of the No. 7 empty cabin utility hole. In addition, the empty compartment No. 7 on the starboard side was identified as the explosion's origin based on the ignition source traces, the extent of the hull damage, and the cracking direction. The root of the explosion accident was restored according to results from a comprehensive on-site investigation. The paint and thinner would volatilize and produce massive organic combustibles during painting operations, forming explosive mixtures when mixed with air and accumulating in the empty compartment's limited space. The explosive gas mixture contacted spattered weld slag at the utility hole, causing an explosion in the empty compartment No. 7 on the starboard side, which triggered explosions in the remaining compartments. As a result of the breeding-development-dissipation of the shock wave, the explosion first intensified and then gradually reduced damage to the surrounding of the NO. 7 empty cabin as the center. The simulation with the CFD analysis software FLUENT revealed that the explosion mixture diffused 10% outward through the utility hole in 12 h. According to the diffusivity analysis results, the gas mixture's volumetric concentration was calculated to be 7.3%, which is sufficient for a combustion explosion. The equivalent amount of TNT for the explosion was 188 kg, which was inverted to 203 kg depending on the extent of damage to the buildings at the accident site after the explosion. It is in good agreement with the TNT equivalent of the explosion obtained from calculations based on the physical parameters of the gas mixture, which proves the practicability of the calculation.

A parameterized hole placement design method for medium-depth holes was proposed to solve the problems of large subjectivity of manual interaction, cumbersome and complex adjustment of hole placement and difficulty in ensuring the uniformity of hole bottom distance. Firstly, a mathematical model in underground mines was constructed using the parameterized hole layout idea combined with the blasting boundary space constraints and blasting parameter requirements. Furthermore, the mathematical model was solved using the operation research method to get the optimal hole design on a digital mining software platform. Finally, the model was used in an underground mine. The result shows that the standard deviations of the bottom distance for the manually laid holes are 0.08 m, 0.10 m, 0.07 m and 0.07 m, respectively, while those for the automatically laid holes with the parameterized method are 0.02 m, 0.03 m, 0.02 m and 0.02 m, respectively. The hole laying time is shortened from 4 hours to 5 min. The proposed method significantly reduces the workload of the designing technicians and maximizes the guarantee of inter-hole laying between holes. It maximizes the uniformity of the bottom distance and avoids the randomness and errorprone nature.

The intrinsic mode confusion of empirical mode decomposition (EMD) and the ensemble empirical mode decomposition (EEMD) can only suppress mode confusion to a limited extent, as the white noise added by EEMD cannot be fully neutralized, which compromises the completeness of the original signal. Additionally, both methods fail to avoid interference from endpoint effects. Modal confusion and endpoint effects lead to distortions in the time-frequency analysis results obtained from the Hilbert transforms of EMD and EEMD. A complete ensemble empirical mode decomposition with adaptive noise and endpoint processing (EP-CEEMDAN) is proposed to address these issues. Simulation experiments were conducted to compare EMD, EEMD, and EP-CEEMDAN decomposition results on simulated vibration signals. Through multiscale permutation entropy detection and marginal spectral analysis, it was verified that EP-CEEMDAN has better control over endpoint effects and mode confusion, proving that EP-CEEMDAN is a more effective adaptive algorithm than EMD and EEMD. Finally, EP-CEEMDAN was applied to the processing of measured non-stationary vibration signals, where adaptive white noise was added at the endpoints of the vibration signals during each stage of decomposition. The method successfully generated various intrinsic mode functions (IMF) by calculating a unique residual signal. The EP-CEEMDAN algorithm effectively suppresses IMF endpoint divergence and modal confusion, while the time-frequency spectrum obtained through the Hilbert transform offers high resolution in both time and frequency domains. This result can be used for vibration feature recognition in non-stationary vibration signals.

The amplification of blasting vibration on rock slopes significantly impacts the accuracy of vibration monitoring and slope safety evaluation. This study investigates the phenomenon through numerical simulation and explores the amplification mechanism based on structural dynamics and vibration mode analysis. The simulation results show that the vibration amplification phenomenon primarily occurs in the vicinity of the bench crest. Influenced by the geometric dimensions of the bench crest and the physical and mechanical parameters of the rock mass, higher peak vibration velocities occur at the bench crest than at the bench toe, due to an increase in platform width, a decrease in bench height, a reduction in the slope ratio and a lower rock mass quality. Conversely, the distribution of the first principal stress exhibits an opposite trend to that of the peak vibration velocity. To improve the accuracy of safety assessments, it is recommended that monitoring points be placed at the bottom line of the bench. The vibration mode analysis further confirms that the amplification effect is predominantly governed by the low-order vibration modes, determined by geometric dimensions and mechanical parameters of the rock mass. The agreement between the mode analysis and numerical simulation results highlights the critical role of low-order vibration modes in controlling the slope's vibration amplification.

Rock's mechanical parameters and fragmentation characteristics significantly change under the freezing and thawing environment in high-altitude cold regions, and it is difficult to directly apply traditional blasting parameters for excavation. Therefore, researching blasting design parameters in freeze-thaw environments is of great importance. This study analyzed the impact of freeze-thaw cycles on rock mechanical properties and conducted the crater experiments of single-hole and double-hole simultaneous blasts in ore rocks under freeze-thaw conditions based on Jurong Copper mine. Futhermore, the geometric parameters and block size distribution of the crater were measured after blasting, and the reasonable parameters for blasting design were determined using the mathematical fitting methods. Additionally, the changes in the blasting crater parameters of the mine were also compared and analyzed under four different rock conditions. The results show that the mechanical properties of rock mass significantly deteriorate with a decrease in uniaxial compressive strength and elastic modulus of up to 40.6% and 54.0% under freeze-thaw cycles, respectively. The optimal burial depth ratio for single-hole blasting of freeze-thaw ore rocks is 0.678~0.789 under different lithological conditions, and the ratio of the optimal charge burial depth to crater radius is distributed in the range of 0.875~1.076. For Chibula mining area, the hole diameter is 152 mm, the diorite hole net parameter is 4.5 m×3 m, the corresponding explosives consumption is 0.56 kg/m³, and the tuff hole net parameter is 5 m×4 m with a 0.63 kg/m³ explosives consumption. For Jurong mining area, the hole diameter is 310 mm, the tuff blasting hole net parameter and explosive consumption is respectively 7 m×5 m and 0.61 kg/m³, and the granite porphyry hole net parameter and explosive consumption is respectively 8 m×5 m and 0.64 kg/m³.