Filling stope double-row medium-length hole blasting method

Abstract

The invention discloses a double-row medium-length hole blasting method for a filling stope. According to the method, a theoretical value range of each blasting parameter is deduced according to field engineering practice, an ore rock blasting crushing mechanism and a blasting parameter theoretical formula. The method comprises the following steps of: constructing a stope blasting model with different hole bottom distances, explosive unit consumption and resistance line combinations by using ANSYS numerical simulation software in combination with physical and mechanical properties of ore rocks and field explosive performance parameters, and preferably selecting double-row medium-length hole blasting parameters according to a chamber effective stress distribution rule and a roof area vibration velocity peak value of the ANSYS blasting model. According to the method, by selecting the double-row optimal medium-length hole blasting parameters, it is guaranteed that ore rocks in the chamber are fully crushed, the effective stress invading the roof and the blasting vibration velocity peak value are reduced, and the roof is prevented from being damaged. And meanwhile, the explosive loading amount is reduced, the economic cost is reduced, and the production efficiency of a mine is improved.

Description

Technical Field

[0001] This invention relates to the field of numerical simulation technology for sector-shaped deep-hole blasting in underground metal mines, specifically a double-row deep-hole blasting method for filling mining areas. Background Technology

[0002] Fan-shaped blast holes, centered on the same drilling tunnel and arranged at fan-shaped angles, can cover a large area of ​​ore body, and are particularly suitable for scenarios with complex ore body shapes such as inclined, branched, and varying thicknesses. In a certain iron mine, after entering deep mining, changes in the mechanical properties of the ore and rock, while the borehole network parameters were still based on old data, led to problems such as roof over-excavation, roof damage, and a high proportion of large blocks during the mining process. Before implementing blasting, a combination of theoretical analysis and numerical simulation was used to preliminarily predict the blasting effect, providing a new double-row blasting scheme for deep mining to guide on-site blasting. Therefore, it is necessary to optimize the various designed blasting schemes based on ANSYS simulation.

[0003] The open-cut backfilling method is characterized by safety, high efficiency, environmental friendliness, and high resource recovery rate. Based on these characteristics, its application in metal mines is continuously expanding, and its application is widely adopted. During open-cut mining, over-excavation of the roof mainly involves two problems: increased risk of roof collapse and inadequate roof connection. Over-excavation leads to irregular spatial shapes between the backfill and the roof, increasing the difficulty of roof connection. Difficulty in achieving a good roof connection prevents the backfill from effectively supporting the roof, affecting the backfilling effect and failing to fully utilize the backfill's support function. The hole mesh parameters determine the distribution of explosive energy during blasting. Uneven energy distribution can easily lead to a series of problems such as large chunks of ore and rock, roof collapse or over-excavation, and under-excavation. Therefore, when optimizing blasting parameters, the impact on the roof should also be considered to prevent excessive stress waves and blasting vibration velocities from damaging the roof.

[0004] ANSYS, a powerful engineering simulation software, is primarily used in the field of blasting to analyze stress fields, strain fields, shock wave propagation, and rock fragmentation mechanisms during blasting processes through numerical simulation technology. This provides a scientific basis for blasting scheme design, parameter optimization, and safety assessment. The software can simulate the propagation of shock waves generated by explosive explosions in media such as rock and soil, analyzing peak pressure, stress distribution, and attenuation patterns at different times.

[0005] An explosive-rock coupling model was established using the LS-DYNA module to calculate the time-varying detonation wave pressure and assess the damage range to the surrounding rock mass. The dynamic stress distribution in the rock mass caused by blasting was studied, identifying areas of concentrated tensile and compressive stress, and predicting crack initiation locations and propagation paths. Optimization of borehole layout (e.g., hole spacing, row spacing) and charge structure was widely applied to avoid excessive fragmentation or under-blasting caused by stress concentration.

[0006] ANSYS's advantages in the field of blasting lie in its support for coupled simulation of multiple processes, including explosive detonation, shock wave propagation, rock mass damage, and structural vibration, realistically recreating the entire blasting process. It includes built-in material models for various materials such as rock, concrete, and metal, and allows for customizable parameters to match the actual engineering material properties. The post-processing interface is intuitive, supporting 3D dynamic effect displays, facilitating engineers' understanding of complex blasting mechanics. Numerical calculations offer high accuracy and can be compared and verified with field monitoring data. Compared to traditional field tests, simulations can rapidly iterate parameters (such as explosive charge and borehole parameters), reducing trial-and-error costs and shortening the design cycle.

[0007] In summary, existing methods for optimizing sector-shaped deep-hole blasting through numerical simulation mainly rely on theoretical analysis and empirical formulas to calculate the range of blasting parameters, followed by simulation of various schemes using ANSYS. However, current simulations often focus on the blasting effect on the ore and rock, neglecting the protection of the stope's roof. In filled stopes, roof protection is extremely important; over-excavation of the roof often compromises its stability, and irregularities in the roof contact area can prevent the filling material from effectively providing support. Therefore, when selecting the optimal blasting parameters for rock-breaking effect after finite element simulation of the designed blasting scheme, the actual site conditions should also be considered to ensure that the selected parameters do not damage the roof. Based on these requirements, a double-row, medium-deep-hole blasting scheme that can guarantee both rock-breaking effect and roof protection is proposed for the mine. Summary of the Invention

[0008] The purpose of this invention is to solve the problems existing in the prior art and to provide a method for double-row medium-deep hole blasting in filling mining areas.

[0009] In blasting engineering simulation, the multi-hole fan-shaped deep-hole model, compared to the simplified two-hole model, more closely resembles real engineering scenarios, revealing complex blasting mechanics and interaction mechanisms. The limitation of the two-hole model is that it can only simulate stress wave interference between a single pair of holes, such as stress enhancement or cancellation, and cannot reflect the three-dimensional superposition of multi-source stress waves in a multi-hole array. In contrast, the full-size multi-hole model established based on actual engineering conditions can capture the superposition of stress waves at different times under differential blasting of adjacent holes in a fan-shaped layout. The stress wave generated by the first blast hole causes pre-fracture of the rock mass, while the stress wave from the later blast hole attenuates faster as it propagates in the fractured medium, requiring adjustment of charge parameters to avoid energy waste. Furthermore, the simulation cloud map can visually display stress concentration areas (such as the effective stress peak at the midpoint of the line connecting the holes), while the two-hole model can only reflect the stress distribution along a single line direction. Multi-hole sector-shaped deep-hole simulation, by introducing multi-physics coupling, three-dimensional heterogeneous media, and complex detonation processes, solves the problem of "oversimplification leading to distorted results" in two-hole models. It is particularly suitable for practical engineering projects such as sector-shaped hole layout in open-pit mines and deep-hole blasting in underground mining areas. Its core value lies in shifting from "feasibility verification" to "systematic optimization," helping engineers predict the complex effects of multi-hole synergy during the design phase, reducing on-site trial-and-error costs, and improving blasting safety and resource extraction efficiency. Furthermore, in the process of optimizing blasting parameters, the selected parameters must not only ensure the crushing effect on the ore and rock but also be based on the engineering practice of subsequent backfilling in the open area, taking into account the protection of the stope roof. Based on full-size multi-hole sector-shaped deep-hole blasting simulation, a double-row blasting scheme is provided for mines that ensures both the blasting effect in the mining area and the protection of the roof. The specific solution of this invention is: a double-row deep-hole blasting method for backfilling mining areas, comprising the following steps:

[0010] Substituting the rock mass mechanical parameters and on-site explosive parameters into relevant formulas, the required hole bottom distance, explosive consumption, discharge distance, and plugging range for the deep-hole blasting model in the sector are calculated. Multiple research schemes are designed based on the parameter value ranges.

[0011] Based on the obtained hole bottom distance parameter range, the designed charge structure and finite element method are used to establish a finite element hole bottom distance sector-shaped medium-deep hole blasting model.

[0012] Based on field data and theoretical calculations, the material parameters of the ore-rock model for the borehole bottom distance model and the material parameters of the explosive-air-blocking model were determined. Non-reflective boundaries and free surfaces were established, and the equations of state for each material and the constitutive model of the rock were determined.

[0013] Solve the finite element model of borehole bottom distance to simulate the explosive explosion process and obtain the effective stress distribution law and stress time history diagram of the monitoring grid between boreholes in each section. Optimize the borehole bottom distance parameters while ensuring that the ore and rock in the stope are sufficiently broken.

[0014] A finite element blasting model for explosive consumption per unit area was established based on the optimal hole-to-bottom distance and existing model parameters. An effective stress monitoring grid was set in the edge region of the roof. The effective stress time history diagram of the roof region was output, and the peak effective stress was compared with the dynamic tensile strength of the ore and rock. The optimal explosive consumption per unit area data was selected without damaging the roof region.

[0015] Multiple finite element blasting models with varying row spacing were established based on optimal borehole spacing and explosive consumption. Effective stress monitoring grids were designed on the side adjacent to the free face and between the two rows of explosives, while blasting vibration velocity monitoring grids were set up in the top plate area. According to the national "Blasting Safety Regulations," the blasting vibration velocity should not exceed 25 cm / s, and the effective stress in the monitoring grid on the free face side and between the two rows of explosives should exceed the dynamic tensile strength of the ore and rock.

[0016] Furthermore, the finite element model is established using ANSYS software;

[0017] Furthermore, the parameter assignments and state equations for materials such as ore, rock, explosives, and air were all performed using LSDYNA software;

[0018] Furthermore, rock mechanics parameters are obtained through on-site or laboratory testing. These parameters include: rock density ρ, uniaxial compressive strength σ. c Uniaxial tensile strength σ t Bulk modulus K, shear modulus G, Poisson's ratio μ.

[0019] Furthermore, based on theoretical formulas, the initial radial peak value of the stress wave during blasting and the radial and tangential stresses induced by the stress wave in various particles of the ore and rock were calculated. This provides data for subsequent analysis and calculation of the pulverized zone and fracture zone of the ore and rock.

[0020] Formula for calculating the initial radial peak value of stress wave:

[0021]

[0022] In the formula: P0 is the initial pressure of the shock wave acting on the ore; P cj ρ0 is the detonation pressure of the explosive; ρ0 and ρ1 are the density of the explosive and the density of the ore / rock, respectively; C r γ and γ represent the wave velocity of the ore and the adiabatic index of the explosive, respectively.

[0023] Radial stress calculation formula:

[0024] σ r =P0(r a / r b ) -α

[0025] Where: σ r Radial stress; r a To calculate the distance r from the point to the center of the explosive, bThe radius of the borehole is μ1 = 0.8μ, where μ is the static Poisson's ratio of the ore and rock; α is the load transmission attenuation index in the ore and rock. According to relevant studies, in the shock wave zone, α = 2 + μ1 / (μ1-1), and in the stress wave zone, α = 2 - μ1 / (μ1-1).

[0026] Formula for calculating tangential stress:

[0027] σ θ =-fσ r

[0028] σ is the tangential stress, f is the lateral pressure coefficient, f = μ1 / (1-μ1).

[0029] Furthermore, based on the rock-breaking mechanism of explosives and the theory of columnar explosive action, the radius of the crushed zone and the radius of the fracture zone generated after rock blasting are analyzed. Combined with the calculation formulas for various parameters, the range of hole pattern parameters for medium-deep hole blasting is obtained. This confirms the hole bottom distance, charge structure, explosive consumption per unit area, and minimum resistance line for medium-deep hole blasting.

[0030] Formula for calculating the radius of the crushing zone:

[0031]

[0032] In the formula: r b σ is the radius of the borehole; μ1 = 0.8μ, μ is the static Poisson's ratio of the ore; α is the load transmission attenuation index in the ore. According to relevant studies, α = 2 + μ1 / (μ1-1) in the shock wave zone and α = 2 - μ1 / (μ1-1) in the stress wave zone; σ is the tangential stress, f is the lateral pressure coefficient, f = μ1 / (1-μ1). Where σcd is the dynamic compressive strength of the ore.

[0033] Formula for calculating fractured zone:

[0034]

[0035] Where: σ td It represents the dynamic compressive strength of the ore.

[0036] Formula for calculating the unit consumption of explosives:

[0037]

[0038] In the formula, q represents the explosive consumption per unit volume; ρ0 represents the explosive density; r represents the borehole radius; L 药 ρ represents the total length of the explosive charge in the borehole; a represents the distance from the bottom of the borehole; m represents the density coefficient; S represents the cross-sectional area of ​​the borehole; and ρ represents the density of the ore / rock.

[0039] Formula for calculating the minimum resistance line:

[0040]

[0041] In the formula, k is the correction coefficient; d b ρ is the diameter of the medium-deep borehole; ρ0 is the explosive density; L is the depth of the underground mine borehole; m is the borehole density coefficient; q is the explosive consumption per unit volume; H is the stope segment height.

[0042] Furthermore, it is necessary to set the rock constitutive model, rock mechanics parameters, explosive material parameters and explosive equation of state, air material parameters and air equation of state, plugging material parameters and plugging material constitutive model, and the model's non-reflective boundaries and free surfaces for the deep hole model in the finite element sector. Specifically, this includes:

[0043] Based on the actual site conditions, a finite element sector-shaped deep hole model was constructed. The MAT-RHT model was selected for the ore and rock, the Jone-Wilkine-Lee (JWL) equation of state was selected for the explosive, and the MAT-NULL stiffness-free material model and MAT-SOLL-AND-FOAM material constitutive model were selected for the air. The Map function in the mesh module of LSDYNA was used to finely mesh the model, with the maximum mesh size controlled at 10cm.

[0044] Furthermore, the stress distribution characteristics of the mine stope roof under different blasting parameters were studied, and the effective stress distribution law under different hole bottom distances and explosive consumption was obtained, as well as the effective stress peak value of the monitoring grid. The optimization of the hole bottom distance simulation aims to ensure that the ore and rock between holes can be fully broken, while the optimization of the explosive consumption simulation aims to prevent the ore and rock in the roof from being broken by stress waves. The purpose of the row spacing simulation is to provide a new double-row medium-deep hole blasting scheme for the mine, thereby replacing the old single-row medium-deep hole blasting scheme and improving production efficiency. Attached Figure Description

[0045] Figure 1 This is a flowchart illustrating the optimization of blasting parameters according to an embodiment of the present invention;

[0046] Figure 2 This refers to the 1.4m hole bottom distance sector-shaped medium-deep hole blasting model involved in the embodiments of the present invention;

[0047] Figure 3 The hole bottom distance model involved in this embodiment of the invention provides the effective stress monitoring point location for the rock breaking effect between holes;

[0048] Figure 4 This is an effective stress distribution diagram at a hole bottom distance of 1.4m according to an embodiment of the present invention;

[0049] Figure 5 The effective stress time history curves of each monitoring grid at a borehole bottom distance of 1.4m are shown in the embodiments of the present invention.

[0050] Figure 6The 1.98 kg / m³ mentioned in the embodiments of the present invention 3 A fan-shaped medium-deep hole blasting model with low explosive consumption;

[0051] Figure 7 The location of the effective stress monitoring point for top plate protection in the explosive consumption model involved in the embodiments of the present invention;

[0052] Figure 8 The 1.98 kg / m³ mentioned in the embodiments of the present invention 3 Effective stress distribution diagram under explosive unit consumption

[0053] Figure 9 The 1.98 kg / m³ mentioned in the embodiments of the present invention 3 Effective stress time history curve of the monitoring grid under explosive unit consumption;

[0054] Figure 10 This refers to the sector-shaped medium-deep hole blasting model with a 2m resistance line and a 1m row spacing involved in the embodiments of the present invention;

[0055] Figure 11 The location of the effective stress monitoring point for the rock-breaking effect of the stope in the row spacing model involved in the embodiments of the present invention;

[0056] Figure 12 This is an effective stress distribution diagram under a 2m resistance line plus a 1m row spacing, as described in an embodiment of the present invention.

[0057] Figure 13 The effective stress time history curves of each monitoring grid under a 2m resistance line plus a 1m row spacing are shown in the embodiments of the present invention.

[0058] Figure 14 This is a fan-shaped medium-deep hole blasting model for the third (3.1m, 2.0m, 3.1m, 2.0m) plugging scheme involved in the embodiments of the present invention;

[0059] Figure 15 This is a grid diagram for monitoring vibration velocity in a blockage model mine chamber according to an embodiment of the present invention;

[0060] Figure 16 The vibration velocity time history diagrams are for the three blocking schemes (3.1m, 2.0m, 3.1m, 2.0m) involved in the embodiments of the present invention. Detailed Implementation

[0061] The following is in conjunction with the appendix Figure 1-16 The specific embodiments of the present invention will be further described below.

[0062] As shown in the figure, this invention proposes a method for optimizing blasting parameters in medium-deep holes based on ANSYS simulation, which specifically includes the following steps:

[0063] Substituting the rock mass mechanical parameters and on-site explosive parameters into relevant formulas, the required hole bottom distance, explosive consumption, and minimum resistance line range for the deep-hole blasting model in the sector are calculated. Multiple research schemes are designed based on the parameter value ranges.

[0064] Based on the obtained hole bottom distance parameter range, the designed charge structure and finite element method are used to establish a finite element hole bottom distance sector-shaped medium-deep hole blasting model.

[0065] Based on field data and theoretical calculations, the material parameters of the ore and rock model for the borehole bottom distance model were determined;

[0066] Based on on-site data and theoretical calculations, the material parameters of the explosive-air-blocking model were determined.

[0067] Set up non-reflective boundaries and free surfaces, and determine the equations of state for each material and the constitutive model for the rock;

[0068] Solve the finite element model of borehole bottom distance to simulate the explosive explosion process and obtain the effective stress distribution law and stress time history diagram of the monitoring grid between boreholes in each section. Optimize the borehole bottom distance parameters while ensuring that the ore and rock in the stope are sufficiently broken.

[0069] A finite element blasting model for explosive consumption per unit charge was established based on the optimal hole bottom distance and existing model parameters.

[0070] An effective stress monitoring grid is set up in the edge region of the roof. The effective stress time history diagram of the roof region is output, and the peak effective stress is compared with the dynamic tensile strength of the ore and rock. Explosive consumption data is optimized without damaging the roof region.

[0071] Multiple sets of minimum resistance line finite element blasting models were established based on the optimal hole bottom distance and explosive unit consumption.

[0072] An effective stress monitoring grid is designed on the side adjacent to the free surface;

[0073] A blasting vibration velocity monitoring grid was installed in the roof area. According to the national "Blasting Safety Regulations", the blasting vibration velocity should not exceed 25 cm / s, and the effective stress of the monitoring grid on the free side should exceed the dynamic tensile strength of the ore and rock.

[0074] Furthermore, the finite element model is established using ANSYS software;

[0075] Furthermore, the parameters and equations of state for materials such as ore, rock, explosives, and air are all input using LSDYNA software;

[0076] The specific experimental data and calculation formulas are as follows:

[0077] The rock mechanics parameters of the ore were obtained through field or laboratory testing.

[0078] Rock mass mechanical parameters include: ore density ρ, uniaxial compressive strength σ c Uniaxial tensile strength σ t Bulk modulus K, shear modulus G, Poisson's ratio μ;

[0079] The parameters of emulsion explosives used in underground mines are obtained from the inspection reports provided by the Explosives Quality Supervision and Inspection Center;

[0080] The existing sector-shaped deep-hole blasting scheme is provided by the mine;

[0081] Furthermore, based on theoretical formulas, the initial radial peak value of the stress wave during blasting and the radial and tangential stresses induced by the stress wave in various particles of the ore and rock were calculated. This provides data for subsequent analysis and calculation of the pulverized zone and fracture zone of the ore and rock.

[0082] Formula for calculating the initial radial peak value of stress wave:

[0083]

[0084] In the formula: P0 is the initial pressure of the shock wave acting on the ore; P cj ρ0 is the detonation pressure of the explosive; ρ0 and ρ1 are the density of the explosive and the density of the ore / rock, respectively; C r γ and γ represent the wave velocity of the ore and the adiabatic index of the explosive, respectively.

[0085] Radial stress calculation formula:

[0086] σ r =P0(r a / r b ) -α (2)

[0087] Where: σ r Radial stress; r a To calculate the distance r from the point to the center of the explosive, b The radius of the borehole is μ1 = 0.8μ, where μ is the static Poisson's ratio of the ore and rock; α is the load transmission attenuation index in the ore and rock. According to relevant studies, in the shock wave zone, α = 2 + μ1 / (μ1-1), and in the stress wave zone, α = 2 - μ1 / (μ1-1).

[0088] Formula for calculating tangential stress:

[0089] σ θ =-fσ r (3)

[0090] σ is the tangential stress, f is the lateral pressure coefficient, f = μ1 / (1-μ1).

[0091] Effective stress σ i The expression:

[0092] σ i =σ r [f 2 +f+1-2μ1(1-f) 2 ] 1 / 2 (4)

[0093] Furthermore, based on the rock-breaking mechanism of explosives and the theory of columnar explosive action, the radius of the crushed zone and the radius of the fracture zone generated after rock blasting are analyzed. Combined with the calculation formulas for various parameters, the range of hole pattern parameters for medium-deep hole blasting is obtained. This confirms the hole bottom distance, charge structure, explosive consumption per unit area, and minimum resistance line for medium-deep hole blasting.

[0094] Formula for calculating the radius of the crushing zone:

[0095]

[0096] In the formula: r b σ is the radius of the borehole; μ1 = 0.8μ, μ is the static Poisson's ratio of the ore; α is the load transmission attenuation index in the ore. According to relevant studies, α = 2 + μ1 / (μ1-1) in the shock wave zone and α = 2 - μ1 / (μ1-1) in the stress wave zone; σ is the tangential stress, f is the lateral pressure coefficient, f = μ1 / (1-μ1). Where σ cd It represents the dynamic compressive strength of the ore.

[0097] Formula for calculating fractured zone:

[0098]

[0099] Where: σ td It represents the dynamic compressive strength of the ore.

[0100] The range of values ​​for the hole bottom distance is determined according to equation (7):

[0101] R²≤a≤2R² (7)

[0102] In the formula: R2 is the radius of the fracture zone of the ore and rock, and a is the bottom distance of the deep hole in the sector.

[0103] Formula for calculating the unit consumption of explosives:

[0104]

[0105] In the formula, q represents the explosive consumption per unit volume; ρ0 represents the explosive density; r represents the borehole radius; L 药 ρ represents the total length of the explosive charge in the borehole; a represents the distance from the bottom of the borehole; m represents the density coefficient; S represents the cross-sectional area of ​​the borehole; and ρ represents the density of the ore / rock.

[0106] Formula for calculating the minimum resistance line:

[0107]

[0108] In the formula, k is the correction coefficient; d b ρ is the diameter of the medium-deep borehole; ρ0 is the explosive density; L is the depth of the underground mine borehole; m is the borehole density coefficient; q is the explosive consumption per unit volume; H is the stope segment height.

[0109] Furthermore, it is necessary to set the rock constitutive model, rock mechanics parameters, explosive material parameters and explosive equation of state, air material parameters and air equation of state, plugging material parameters and plugging material constitutive model, and the model's non-reflective boundaries and free surfaces for the deep hole model in the finite element sector. Specifically, this includes:

[0110] Based on the actual site conditions, a finite element sector-shaped deep hole model was constructed. The MAT-RHT model was selected for the ore and rock, the Jone-Wilkine-Lee (JWL) equation of state was selected for the explosive, and the MAT-NULL stiffness-free material model and MAT-SOLL-AND-FOAM material constitutive model were selected for the air. The Map function in the mesh module of LSDYNA was used to finely mesh the model, with the maximum mesh size controlled at 10cm.

[0111] Furthermore, the stress distribution characteristics of the mine stope roof under different blasting parameters were studied, and the effective stress distribution law under different hole bottom distances and explosive consumption was obtained, as well as the effective stress peak value of the monitoring grid. The optimization of the hole bottom distance simulation aims to ensure that the ore and rock between holes can be fully broken, while the optimization of the explosive consumption simulation aims to prevent the ore and rock in the roof from being broken by stress waves. The maximum effective stress of the monitoring grid on the side near the free surface under different resistance lines and the peak value of blasting vibration in the roof area were studied to achieve the goal of fully breaking the ore and rock while ensuring good roof stability without exceeding the allowable standard value for blasting vibration safety.

[0112] See Figure 1 Based on the theoretically calculated parameters for deep-hole blasting in a sector shape, a thin-block blasting model including boreholes was first established using ANSYS. The Map function in the software's built-in mesh module was then used to mesh the model, and the initial mesh model was output. This mesh model was then expanded using LS-DYNA. Material parameters and equations of state for the ore, explosives, air, and plugging were then set.

[0113] See Figure 2 , Figure 2This is a multi-hole, fan-shaped, medium-deep hole blasting model built in LS-DYNA with a hole bottom distance of 1.4m. The blasting parameters are as follows: The iron ore has medium hardness, but well-developed joints and fissures cause it to fracture. Currently, a Sandvik DL2710 drilling rig is used to drill 76mm boreholes, with a measured charge return rate of 10%–15%, indicating significant return. Increasing both the hole depth and diameter would not achieve better charging results with the current medium-deep hole charging equipment. Therefore, considering the influence of the medium-deep hole charging equipment, the diameter should not exceed the currently used 76mm. Furthermore, considering that the hole diameter range of the Sandvik DL2710 drilling rig is 76–89mm, a borehole diameter of 76mm is chosen. The model width is 15.6m, with 15m representing the stope, and the sides thickened by 0.3m to monitor the impact on the sidewalls. The height is 15.3m, with the top 0.3m reserved for the pre-existing rock roof. Four hole bottom distance models were established based on on-site blasting parameters and empirical formulas. Taking a 1.4m hole bottom distance as an example, the model thickness is 0.3m, and the total number of grid cells is 2,010,044.

[0114] See Figure 3 To determine whether the deep hole blasting effect in the sector is good, an effective stress monitoring grid needs to be set at the boundary of the stope. Taking a 1.4m hole bottom distance and 14 blast hole model as an example, five stress monitoring grids A to F are set at 35cm above the center of the line connecting the bottoms of two adjacent blast holes.

[0115] See Figure 4 The effective stress cloud diagrams at representative time steps are shown when the hole bottom distance is 1.4m.

[0116] See Figure 5 The diagrams show the effective stress time history for borehole bottom distances of 1.3m, 1.4m, 1.5m, and 1.6m. The effective stresses at 1.5m and 1.6m are not considered because the average effective stress does not exceed the dynamic tensile strength of the ore. At 1.3m and 1.4m, the effective stress exceeds 42.58MPa, indicating sufficient ore fracturing. Therefore, a higher charge rate is chosen for a 1.3m borehole bottom distance compared to a 1.4m borehole bottom distance. Considering all factors, a 1.4m borehole bottom distance is selected. (See reference...) Figure 6 The hole bottom distance is 1.4m and the weight is 1.98kg / m. 3 The fan-shaped medium-deep hole blasting model under explosive unit consumption remains consistent with the hole bottom distance model in terms of all other blasting parameters, material parameters and state equations.

[0117] Refer to 7. Six monitoring grids G to L are set at the edge of the top plate to monitor the effective stress change in this area. Since the LSDYNA output cloud map cannot change the letter number, A to F in the figure are the monitoring grids G to L.

[0118] See Figure 8 It is 1.98 kg / m 3Representative effective stress cloud diagrams at various time steps under explosive consumption;

[0119] See Figure 9 It is 2.06 kg / m 3 2.02kg / m 3 1.98kg / m 3 1.94kg / m 3 Time history diagrams of maximum equivalent stress at each monitoring grid under four different unit consumption levels. 2.06 kg / m 3 2.02kg / m 3 Under the given unit consumption, the stress waves acting on the roof all exceed the dynamic tensile strength of the roof rock, posing a risk of roof damage. 1.98 kg / m 3 Neither 1.94 kg / m³ nor 1.94 kg / m³ will compromise roof stability, but considering mine production efficiency, 1.98 kg / m³ is preferred. 3 ;

[0120] See Figure 10 The hole bottom distance is 1.4m, and the weight is 1.98kg / m. 3 The fan-shaped medium-deep hole blasting model under the combination of explosive consumption, 2m resistance line and 1m row spacing, and other blasting parameters, material parameters and state equations are still consistent with the hole bottom distance model.

[0121] See Figure 11 This is a diagram showing the effective stress monitoring grid layout. Six recording cells (a to f) are set up on the double-row blasting numerical model to monitor the equivalent stress change at this location. In the diagram, A to F represent the six monitoring grid cells (a to f).

[0122] See Figure 12 The effective stress cloud diagrams at various time steps are representative of the 2m resistance line plus the 1m resistance line;

[0123] See Figure 13The time history diagram shows the maximum equivalent stress of each monitoring grid at a 1m spacing. The resistance line was set to a constant 2m during the grid spacing simulation. At a 1m spacing, the maximum effective stress of the three monitoring grids (a to c) exceeds 42.58MPa, indicating effective fracturing of the ore and rock within the range from the first row of boreholes to the free surface. The maximum effective stress of the three monitoring grids (d to f) all exceed the dynamic tensile strength of the ore and rock, suggesting that the ore and rock between the two rows of boreholes can be sufficiently fracturing under the action of double-row explosives; a 1m spacing scheme is suitable. At a 1.2m spacing, the maximum effective stress of the three monitoring grids (a to c) does not exceed 42.58MPa, with the maximum effective stress at point b being only 40.62MPa. Therefore, the ore and rock on the free surface side from the first row of boreholes cannot be guaranteed to be sufficiently fracturing. The values ​​of the three monitoring grids from e to f did not exceed the maximum effective stress of the ore and rock, and the ore and rock between the two rows of blast holes could be broken; when the row spacing was 1.4m and 1.6m, the maximum effective stress of the six monitoring grids from a to f all exceeded 42.58MPa, and neither the ore and rock near the free surface nor the ore and rock between the two blast holes could be fully broken.

[0124] See Figure 14 The hole bottom distance is 1.4m, and the weight is 1.98kg / m. 3 The fan-shaped medium-deep hole blasting model under the conditions of explosive consumption, 2m resistance line plus 1m row spacing and plugging scheme three (3.1m, 2.0m, 3.1m, 2.0m) has all other blasting parameters, material parameters and state equations consistent with the row spacing model.

[0125] See Figure 15 This is a diagram showing the vibration velocity grid layout. To determine whether the roof area was damaged by blasting vibration, three blasting vibration monitoring grids (g, h, i) were set up in the roof area of ​​the stope, and three vibration velocity monitoring grids (j, k, l) were set up in the side area of ​​the stope to record the damage to adjacent stops. A to F in the diagram represent monitoring grids g to l.

[0126] See Figure 16The following are the vibration velocity time history diagrams for each monitoring grid under Blocking Scheme 3 (3.1m, 2.0m, 3.1m, 2.0m). In Scheme 1, the peak values ​​of grids g to i all exceed 25cm / s, posing a risk of roof damage; this scheme is not advisable. In Scheme 2, only grid h's peak value exceeds 25cm / s; the other two grids are close to the safe peak value, making this scheme also undesirable. In Scheme 3, the peak values ​​of grids g to i do not exceed 25cm / s, indicating good safety, and this scheme is acceptable. The peak vibration values ​​of grids j to l remain essentially unchanged, indicating that modifying the blocking length will not significantly affect roof stability. Meanwhile, the maximum vibration velocity in the roof area is 20.85cm / s, which does not exceed the permissible vibration safety standard value of 25cm / s for roadway-type structures. In summary, under all three blocking schemes, the peak vibration velocity in the roof area does not exceed 25cm / s, indicating good safety. Under Scheme 1, the peak vibration velocity of adjacent mine blocks exceeds the safe peak value, making this scheme unsuitable. Under the blocking scheme of Option 2, only some monitoring grids exceed the safe peak value, and the mine still faces the risk of damage; therefore, Option 2 is not an option. Under the blocking scheme of Option 3, none of the monitoring grids exceed 25 cm / s, so Option 3 is an option. In summary, the rock blasting simulation method based on FEM-DEM co-simulation provided by this invention has the following advantages: Rock mass mechanical parameters are obtained through indoor experiments, and explosive parameters are obtained through field surveys. Rock mass mechanical parameters include: rock density, uniaxial compressive strength, uniaxial tensile strength, bulk modulus, shear modulus, Poisson's ratio, etc. Explosive parameters include: detonation velocity, detonation pressure, explosive density, etc. The obtained data are substituted into relevant formulas to calculate the required hole bottom distance, explosive consumption, and minimum resistance line range for the sector-shaped deep-hole blasting model. Multiple research schemes are designed based on the parameter value ranges; focusing on the roof failure mechanism of sector-shaped deep-hole blasting in subsequent filling of the stope, a systematic analysis is conducted on the multivariate combination of key parameters such as hole bottom distance, explosive consumption, and resistance line. Simulations reveal the effective Von Mises stress distribution and peak vibration velocity characteristics of the roof and the mine under different parameter combinations.

[0127] Compared to a simplified two-hole model, the multi-hole sector-shaped deep-hole model more closely resembles real-world engineering scenarios, revealing complex blasting mechanics and interaction mechanisms. The limitation of the two-hole model is that it can only simulate stress wave interference between single pairs of holes, such as stress enhancement or cancellation, and cannot reflect the three-dimensional superposition of multi-source stress waves in a multi-hole array. In contrast, the full-scale multi-hole model established based on engineering practice can capture the superposition of stress waves at different times under differential blasting of adjacent holes in a sector-shaped layout. The stress wave generated by the first blast hole causes pre-fracture of the rock mass, while the stress wave from the later blast hole attenuates more rapidly as it propagates through the fractured medium, requiring adjustment of charge parameters to avoid energy waste. Furthermore, in optimizing blasting parameters, the selected parameters must not only ensure effective rock fragmentation but also be based on the engineering realities of the subsequent empty-field filling method, while also considering the protection of the mine stope roof.

Claims

1. A method for double-row medium-deep hole blasting in a filling stope, characterized in that: Includes the following steps: S1. Substitute the rock mass mechanical parameters and the on-site explosive parameters into the relevant formulas to calculate the required range of hole bottom distance, explosive consumption, row spacing and plugging length for the double-row fan-shaped deep hole blasting model. Design multiple research schemes based on the parameter value range. S2. Based on the range of hole bottom distance parameters, design the charge structure and finite element method, and establish a fan-shaped medium-deep hole blasting model under different hole bottom distances; S3. Based on field data and theoretical calculations, determine the material parameters of the ore-rock model for the borehole bottom distance model and the material parameters of the explosive-air-blocking model. Set up non-reflective boundaries and free surfaces, and determine the equations of state for each material and the constitutive model of the rock. S4. Solve the fan-shaped medium-deep hole blasting model under different hole bottom distances, simulate the explosive explosion process, obtain the effective stress distribution law and stress time history diagram of the monitoring grid between holes in each section, and optimize the hole bottom distance parameters under the premise that the ore and rock in the stope are fully broken. S5. Based on the optimal hole bottom distance and existing model parameters, establish a sector-shaped medium-deep hole blasting model under different explosive consumption. Set up an effective stress monitoring grid in the edge area of ​​the roof plate, output the effective stress time history diagram of the roof plate area, compare the effective stress peak with the dynamic tensile strength of the ore and rock, and optimize the explosive consumption data without damaging the roof plate area. S6. Based on the optimal hole bottom distance and explosive consumption, establish a double-row fan-shaped medium-deep hole blasting model with different row spacing. Design an effective stress monitoring grid on the side near the free surface and between the two rows of explosives to ensure that the ore and rock are fully broken. S7. Design a blocking scheme based on theoretical formulas, and set up a blasting vibration velocity monitoring grid in the top plate and side plate areas to prevent excessive blasting vibration velocity from damaging the stability of the top plate and side plate.

2. The double-row medium-deep hole blasting method for filling a stope according to claim 1, characterized in that: The finite element model was established using ANSYS software; the parameter assignments and state equations for materials such as ore, rock, explosives, and air were all performed using LSDYNA software.

3. The double-row medium-deep hole blasting method for filling a stope according to claim 1, characterized in that: Rock mechanics parameters of the ore are obtained on-site or in the laboratory. These parameters include: ore density, uniaxial compressive strength, uniaxial tensile strength, bulk modulus, shear modulus, and Poisson's ratio.

4. The double-row medium-deep hole blasting method for filling a stope according to claim 1, characterized in that: The initial radial peak value of the stress wave during blasting and the radial and tangential stresses caused by the stress wave in each particle of the ore and rock are calculated based on theoretical formulas, providing data for subsequent analysis and calculation of the crushed zone and fracture zone of the ore and rock; Formula for calculating the initial radial peak value of a stress wave: In the formula: P0 is the initial pressure of the shock wave acting on the ore; P cj ρ0 is the detonation pressure of the explosive; ρ0 and ρ1 are the density of the explosive and the density of the ore / rock, respectively; C r γ and γ represent the wave velocity of the ore and the adiabatic index of the explosive, respectively. Radial stress calculation formula: σ r =P0(r a / r b ) -α In the formula: σ r Radial stress; r a To calculate the distance r from the point to the center of the explosive, b The radius of the borehole is μ1 = 0.8μ, where μ is the static Poisson's ratio of the ore and rock; α is the load transmission attenuation index in the ore and rock. According to relevant studies, in the shock wave zone, α = 2 + μ1 / (μ1-1), and in the stress wave zone, α = 2 - μ1 / (μ1-1). Formula for calculating tangential stress: σ θ =-fσ r σ θ Let f be the tangential stress, and f be the lateral pressure coefficient, f = μ1 / (1-μ1).

5. The double-row medium-deep hole blasting method for filling a stope according to claim 1, characterized in that: Based on the rock-breaking mechanism of explosives and the theory of columnar explosive action, the radius of the crushed zone and the radius of the fracture zone generated after rock blasting are analyzed. Combined with the calculation formulas for various parameters, the range of hole pattern parameters for medium-deep hole blasting is obtained. This confirms the hole bottom distance, charge structure, explosive consumption per unit area, and minimum resistance line for medium-deep hole blasting. Formula for calculating the radius of the crushing zone: In the formula: r b σ is the radius of the borehole; μ1 = 0.8μ, where μ is the static Poisson's ratio of the ore; α is the load transmission attenuation index in the ore. According to relevant studies, in the shock wave zone, α = 2 + μ1 / (μ1-1), and in the stress wave zone, α = 2 - μ1 / (μ1-1); σ θ Let f be the tangential stress, f be the lateral pressure coefficient, and f = μ1 / (1-μ1). B = 2 1 / 2 [f 2 +1-(μ1-μ1 2 (1-f) 2 ] 1 / 2 , where σ cd The dynamic compressive strength of the ore and rock; Formula for calculating fractured zone: Where: σ td The dynamic compressive strength of the ore and rock; Formula for calculating the unit consumption of explosives: In the formula, q represents the explosive consumption per unit volume; ρ0 represents the explosive density; r represents the borehole radius; L 药 ρ represents the total length of the explosive charge in the borehole; a represents the distance from the bottom of the borehole; m represents the density coefficient; S represents the cross-sectional area of ​​the borehole; ρ represents the density of the ore / rock. Formula for calculating the minimum resistance line: In the formula, k is the correction coefficient; d b ρ is the diameter of the medium-deep borehole; ρ0 is the explosive density; L is the depth of the underground mine borehole; m is the borehole density coefficient; q is the explosive consumption per unit volume; H is the stope segment height.

6. The double-row medium-deep hole blasting method for filling a stope according to claim 1, characterized in that: The finite element method (FEM) for deep-hole ore in a sector was established, including setting up the rock constitutive model, ore and rock mechanical parameters, explosive material parameters and equation of state, air material parameters and equation of state, plugging material parameters and constitutive model, and the model's non-reflective boundaries and free surfaces. Specifically, this included constructing a finite element method (FEM) deep-hole model in a sector based on actual field conditions; using the MAT-RHT model for ore and rock; the Jone-Wilkine-Lee (JWL) equation of state for explosives; the MAT-NULL stiffness-free material model for air; and the MAT-SOLL-AND-FOAM material constitutive model. The model was finely meshed using the Map function within the mesh module of LSDYNA, with the maximum mesh size controlled to 10 cm.

7. The double-row medium-deep hole blasting method for filling a stope according to claim 1, characterized in that: This study investigates the stress distribution characteristics of the mine stope roof under different blasting parameters, obtains the effective stress distribution law under different hole bottom distances and explosive consumption, and monitors the effective stress peak value of the grid.

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